Chapter 1 Keynote 20 Santiago Chile, 22-25 August 2004 Massmin 2004 Geomechanics: The critical engineering discip
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Chapter 1
Keynote
20
Santiago Chile, 22-25 August 2004
Massmin 2004
Geomechanics: The critical engineering discipline for mass mining Edwin T. Brown, Julius Kruttschnitt Mineral Research Centre and Golder Associates Pty Ltd, Brisbane, Australia
Abstract It is shown that the scale of the world’s largest surface and underground mass mining operations has grown at a continuing rate over the last 100 years. Indications are that both the rates of production and the depths of open pit and underground mines will continue to grow into the future. This can be expected to provide significant challenges to the engineering discipline of geomechanics which is shown to have been a major contributor to the success and growth of mass mining over the last 40 years. Important challenges for many operations remain in other engineering disciplines to improve efficiency and achieve cost reduction. Nevertheless, the importance of geomechanics to the investigation, design, construction and safe operation of profitable large-scale surface and underground mines requires that it be recognised as the basic engineering discipline for mass mining. It is essential, therefore, that the industry continues to support research into a range of geomechanics-related issues and the education and training of the future generation of mining geomechanics specialists.
1 INTRODUCTION There are apparent trends in the international mining industry towards the globalisation of company structures and operations, the mining of larger and often lower grade orebodies, mining at increasing depths, cost reduction through the use of mass mining methods, the application of caving methods of mining to more massive orebodies, and improved safety and environmental performance. These trends bring with them a range of management and engineering challenges, many of which have geomechanics bases. The purpose of this paper is to offer a wide-ranging exploration of these issues, in the course of which, it will be argued that geomechanics is the critical engineering discipline for the safe and economic use of modern mass mining methods. Following a discussion of the terminology to be used, the commonly made assertion that there is an international trend towards mass mining will be tested through a compilation and assessment of historical production and mining depth data for a number of selected open pit and underground mines. The historical development of mining geomechanics and its relation to the increased scales of both open pit and underground mining will be reviewed. The key geomechanics issues associated with the further evolution of large-scale open pit mining and of underground mining by caving and open stoping methods will then be explored and some of the requirements for research to support these developments identified. This exploration will draw on part of the work of the International Caving Study (ICS) being carried out through the Julius Kruttschnitt Mineral Research Centre, Brisbane, Australia. Finally, the difficult but critically important issue of the provision of well-educated and trained geomechanics practitioners to support current and future mass mining undertakings internationally will be addressed. 2. TERMINOLOGY At the outset it is necessary to define the term mass mining. Despite its widespread use, clear definitions of this term are not readily found. The web site for this conference (www.massmin2004.cl) indicates that the conference will address a range of mass mining problems and then says, "in this category are understood to be caving methods (block, panel and sublevel caving), large scale stoping methods Massmin 2004
(sublevel, longhole) and variations of these mentioned." The statement goes on to say that "MassMin 2004 will include not only underground mining but also open pit mining." This definition of mass mining follows that used in the Chairman’s Foreword to the Proceedings of MassMin 2000 in which the MassMin series of conferences were described as "the ‘pinnacle’ international forum for discussing and sharing both technical and operational issues and experiences associated with the application of methods such as caving (block, caving and sublevel) and large-scale stoping (sublevel and longhole), including their derivatives (underground benching, front cave, inclined footwall, etc)" (Chitombo 2000). Thus, the two recent international conferences on mass mining define their subject only in terms of the mining methods being used and not through other parameters such as production levels. Because of the nature of the mining methods identified as constituting mass mining (including open pit mining), there is an implication that the mining will be on a large or massive scale in other than narrow orebodies whose dimensions will be substantial in all three dimensions. Mass mining will also be taken to be highly mechanised and non-selective (other than in the broadest sense). In his keynote paper to MassMin 2000, Hustrulid (2000) arbitrarily took large scale undergound mines to have production rates of at least 5,000 tonnes per day (tpd) or 15 million tonnes per year (tpy) and "involving the use of panel caving, sublevel caving and sublevel stoping." For present purposes, Hustrulid’s limit may be too low and so underground mass mining will be taken to involve production rates of more than 10,000 tpd. Although underground coal mining by longwall methods has most of the characteristics of mass mining, production rates are rarely high enough for them to qualify under this criterion, even when multiple longwall panels are operated at a given mine. Accordingly, underground coal mining and its needs will not be considered here. The writer has considered elsewhere geomechanics practice in the Australian underground coal mining industry and the industry’s needs for geomechanics research, education and training (Golder Associates 2001). Large open pits provide clear examples of mass mining. For present purposes, examples will be assumed to be provided by those pits that mine 30,000 or more tonnes of ore per day (or 100 million tpy). For consistency with the
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definition adopted for underground mines, the removal of waste rock will not be taken into account. The largest open pits currently mine in excess of 200,000 tpd (see Figure 2). The other major term that requires definition in the present context is geomechanics and, more particularly, that part of the discipline known as mining geomechanics. The learned society for geomechanics in Australia, the Australian Geomechanics Society, defines its subject in the following way: "Geomechanics is the application of engineering and geological principles to the behaviour of the ground and ground water and the use of these principles in civil, mining, offshore and environmental engineering in the widest sense." This definition includes the application of geomechanics to a range of engineering purposes. In this regard, it is almost synonymous with the term geotechnical engineering which has been defined as "the application of the sciences of soil mechanics and rock mechanics, engineering geology and other related disciplines to civil engineering construction, the extractive industries and the preservation and enhancement of the environment" (Anon 1999). A more restrictive view of geomechanics is that it is that group of "sciences" referred to in this definition of geotechnical engineering, most notably soil mechanics and rock mechanics (e.g. Brown 1993). This view derives from the definition of rock mechanics developed by the US National Committee on Rock Mechanics in 1966: "Rock mechanics is the theoretical and applied science of the mechanical behaviour of rock and rock masses; it is that branch of mechanics concerned with the response of rock and rock masses to the force fields of their physical environment." The Australian Geomechanics Society’s broader definition of geomechanics which includes the applications as well as the "theoretical and applied sciences" themselves will be adopted here. As has been noted, this definition is almost synonymous with that of geotechnical engineering. Accordingly, on occasion, the adjective "geotechnical" may be used in accordance with customary usage. It follows from these definitions that mining geomechanics is that part of geomechanics that is concerned with the application of knowledge of the physical and mechanical behaviour of geological materials (soils, rocks and water) to the investigation, design and performance of mining structures including excavations (Brown 1993). Blasting mechanics is usually considered to be part of mining geomechanics as so defined. For present purposes, mining geomechanics will be taken to include the application of geomechanics to mine fill, waste and tailings disposal and the long-term stability and performance of mined and rehabilitated land forms on mine sites. It also includes applications of geomechanics in the investigation, design, excavation, stability, support and reinforcement and performance of all infrastructure, service and productive surface and underground excavations associated with the mining process. Figure 1 illustrates many of the components of mining geomechanics so defined.
Figure 1: Components of mining geomechanics (from Little and Szwedzicki 1992). 22
3. HISTORICAL TRENDS TOWARD MASS MINING It is instructive to consider the ways in which mass mining has evolved over time, in terms of mining methods, levels of production and mining depths. In these considerations, it is necessary to consider open pit and underground mining separately. Figure 2 shows a compilation of historical and projected daily production rates from the early 20th century for several of the world’s largest open pits of the day. With the notable exceptions of Bingham Canyon and Chuquicamata, none of these pits have operated throughout the full period being considered. The data plotted in Figure 2 were derived from a variety of published and unpublished sources. The available data, sometimes expressed as tonnes per month and sometimes as tonnes per year, have been converted to average tonnes per day for use in Figure 2. Although some local inaccuracies may exist because of the variety of sources of the data plotted, the general trends indicated are considered to be meaningful. Figure 2 shows clearly that production rates from the world’s largest open pits have grown progressively over the last 100 years and that they are projected to continue to grow over the next decade. This increased production owes much to increased mechanisation and to increases in the capacities of equipment such as blast hole drills, shovels, loaders and trucks (Karzulovic 2004). A similar and perhaps more significant trend from a geomechanics perspective is the increase in open pit depths with time shown in Figure 3. The open pit for which most data are available, the Chuquicamata mine, Chile, was 280 m deep in 1970, is about 850 m deep today and is projected to become 1100 m deep in the next decade (Karzulovic 2004). Also shown on Figures 2 and 3 are certain milestones in the development and application of geomechanics to open pit slope engineering. The increases in production and pit depths that took place from the 1970s coincide with the significant attention being paid to slope stability research at that time, most notably in the project undertaken by Professor Evert Hoek at Imperial College, London, in the late 1960s and reported in the seminal book, Rock Slope Engineering by Hoek and Bray (1974), and the Canadian Pit Slope Project began in 1972 and reported in the Pit Slope Manual published in 1976 and 1977. Figure 4 shows a compilation of daily production rates for some of the world’s largest underground mines of their day. The data plotted in Figure 4 were collected in a similar way to those for open pits plotted in Figure 2 and are subject to similar qualifications. Figure 2 is dominated by El Teniente, the world’s largest underground mine now producing around 100,000 tpd with significant increases in production being planned. It will be noted that the production from this mine increased steadily over the first 30 years of production to about 1940 when the rate of increase declined, possibly as a result of World War II. The rate of increase jumped sharply from the mid-1960s with the overall trend being projected to continue into the future. At several times in the past, El Teniente’s production rates have been matched and occasionally slightly exceeded by those of other block and panel caving mines such as Climax, El Salvador, Miami and San Manuel. The largest production rates achieved by other mining methods have been those of the Swedish sublevel caving iron ore mines. The production rates of the largest open stoping mines have always been below those of the caving mines. The highest current production rate from an open stoping mine known to the writer is Olympic Dam’s approximately 25,000 tpd which is planned to increase to 34,000 tpd in the near future. As a general rule, the mining costs of stoping mines may be two or three times those of caving mines.
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Figure 2: Evolution of daily production rates of selected large open pits.
Figure 3: Evolution of the depths of selected large open pits. Figure 5 shows the evolution of maximum mining depth for selected underground mines that use mass mining methods. Once again, the most complete set of data available are for El Teniente. Figure 5 clearly establishes the fact that mining depths are increasing, if anything at an increasing rate with time. The depths of mines that use mass mining methods as defined here are not as great as those mining tabular orebodies of which the deep level gold mines of South Africa provide the most notable example. However, the depths of some of these mines are now considerable in historical terms. The undercut level of the Palabora block caving mine is at a depth of 1200 m below surface (Calder et al 2000) as is the Freeport Deep Ore Massmin 2004
Zone mine. Other block and panel caving mines in the feasibility and planning stages (e.g. Bingham Canyon and Chuquicamata) may well be even deeper. Sublevel caving at depths of more than 1000 m is being practiced or planned at Kiruna (Sweden), Perseverance (Western Australia) and Ridgeway (New South Wales, Australia). The deepest mass mining open stoping mine known to the writer is the Kidd Mine in Timmins, Ontario, which is at a depth of about 2000 m and plans to reach 2990 m in about 10 years time. Another open stoping mine, the Enterprise Mine at Mount Isa, Australia, is at a depth of 1900 m below surface. The precursor of the modern block caving method of mining was developed in the iron ore mines of the
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Menominee Ranges, Michigan, USA, in the late nineteenth century (Peele 1941). Variations of the original Pewabic method were soon developed at other iron ore mines in Michigan and, from the early part of the twentieth century, in the copper mines in western USA. Before evolving to the use of full block caving, many mines initially used combined methods involving, for example, shrinkage stoping and caving methods for the subsequent mining of the pillars between the primary stopes (Peele 1941). By the 1920s and 30s, block caving methods were being used in a wide range of mines exploiting massive, weak orebodies. During this period, the method was introduced, for example, at the King Mine which mined asbestos in
Quebec, Canada, the Climax Mine mining molybdenum in Colorado, USA, and the copper mines in Chile (Peele 1941). Early block caving methods used hand, grizzlies with transfer raises, and slusher methods of loading. Under suitable geotechnical and other conditions, the productivity of block caving mines was increased significantly by the progressive introduction of mechanised loading using LHD vehicles from the late 1960s. Mechanised panel caving was used when Henderson commenced production in 1976 and was introduced at El Teniente in 1982. Peele (1941) notes that sublevel caving (SLC) was introduced as a logical development of top slicing in the Lake Superior Iron Ranges in the early part of the
Figure 4: Evolution of daily production rates at selected large underground mines.
Figure 5: Evolution of maximum mining depth for selected mines that use mass mining methods. 24
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twentieth century. It was also used in a few other parts of the USA in the same period. Perhaps the first large-scale use of SLC was that at the Kiruna iron ore mine in Sweden from 1957 (Gustafsson 1981). This marked the beginning of the development of modern SLC mining. The method was introduced into Australia, Canada and China in the 1960s. Forms of open stoping and sublevel open stoping have long been practiced, particularly in narrow and tabular orebodies. Peele (1941) notes that sublevel stoping developed in the Michigan iron ore mines in about 1902. Forms of the method were being used in disseminated copper orebodies in Canada, the USA, what is now Zambia, and at Mount Isa, Australia, in the 1930s. The introduction of LHD vehicles, advances in drilling and blasting technology and advances in mining geomechanics, helped bring about changes in mine design and increases in productivity using sublevel open stoping and blasthole stoping methods from the late 1960s and the 1970s (e.g. Goddard 1981). In many major metalliferous mining districts, these developments brought about a change to these methods from cut-andfill methods under suitable geotechnical conditions (e.g. Brown 1992, Pariseau et al 1984, Singh and Hedley 1981). A good example of the evolution of non-caving underground mass mining methods is provided by the Mount Isa Mine in northwest Queensland, Australia. Underground mining of silver-lead-zinc ore began in 1932 and of copper in 1943. Until 1963, open sublevel stoping accounted for 95% of the underground production. From 1963, two new stoping methods were introduced, mechanised cut-and-fill in the narrower silver-lead-zinc Racecourse orebodies and sublevel caving of the southern section of the 500 copper orebody (Davies 1967). Largescale mechanized sublevel open stoping with delayed filling was introduced into the wider parts of the silver-lead-zinc orebodies and the copper orebodies from the late 1960s (Goddard 1981, Grant and De Kruijff 2000). Bench and fill methods replaced what was left of cut-and-fill mining of the silver-lead-zinc (now known more simply as lead) orebodies from 1991 (Villaescusa 1996). It has been widely reported and acknowledged that geomechanics input has been centrally important to the successful development and implementation of these various mining methods (Brown 1992, Davies 1967, Mathews and Edwards 1969, Villaescusa 1996). 4. HISTORICAL DEVELOPMENT OF MINING GEOMECHANICS 4.1 Early mining geomechanics In contradistinction to many fields of modern engineering that have been preceded by, or developed in parallel with, their companion engineering sciences, mining was practiced for centuries (or even millenia) without the benefit of any formal knowledge that we would now recognise as applied mechanics, rock mechanics or geomechanics (Obert and Duvall 1967). After all, Newton’s laws were discovered only in the second half of the 17th century, and the bases of modern applied mechanics were developed mainly in the 18th and 19th centuries. Based on the practices reported by Agricola (1556), progressive improvement in the state-of-the-art of mining was made, particularly in Europe and subsequently North America. However, it was not until the 19th century that engineering societies were founded in those continents and engineers began to report their experiences in technical publications. In the area that we would now recognise as mining rock mechanics or geomechanics, the initial concerns were with the effects of underground mining on Massmin 2004
the surface (caving and subsidence) and then with the stability of the excavations themselves (Obert and Duvall 1967, Turchaninov et al 1979). Reports dating from the mid19th century were based mainly on qualitative visual observations but quite good measurements of surface subsidence and of roof and floor convergence were made in the last quarter of that century. At the same time, mechanisms of caving, arching and the development of ground pressure and deformations were postulated (e.g. Fayol 1885). Young and Stoek’s (1916) comprehensive account of mining-related subsidence lists more than 100 papers published in the preceding 60 years dealing mainly with subsidence in the European coalfields. These studies of caving and subsidence may be considered to represent an initial stage in the development of modern mining geomechanics (Hood and Brown 1999). 4.2 The establishment of rock mechanics as an engineering discipline Obert and Duvall (1967) note that in the first two decades of the 20th century, technical reports began to appear which treated rock as an engineering material. These papers addressed issues such as the mechanical properties of rock, deep mining and rockbursts. Laboratory studies using both photoelastic and material models of rock were reported (e.g. Bucky 1934) as well as theoretical and empirical studies of the states of stress around surface and underground excavations. In the late 19th and early 20th centuries, the theories developed were either empirical or applied the basic concepts of structural mechanics and the strength of materials. Subsequently, the theory of linear elasticity was used to calculate the stresses around excavations of simple shape (e.g. Terzaghi and Richart 1952, Savin 1961). Greater insight was developed using elastic-plastic and ground-support interaction concepts based on an analysis by Fenner (1938). Much of this work was carried out in Europe and then in the USA where the US Bureau of Mines was a leader in theoretical and experimental studies from the 1930s. From the 1930s, the production of scientific and engineering information about rock properties and the design and stability of structures in rock accelerated rapidly. The first issue of the first specialist journal devoted to rock mechanics, Geologie und Bauwesen, was published in Vienna in 1929. (After several name changes, the successor to this journal is now known as Rock Mechanics and Rock Engineering.) Annual colloquia on rock mechanics have been held in Austria since 1950, and in 1951 the First International Strata Control Congress was held in Liege, Belgium. In 1956, the first of what became the annual U S symposia on rock mechanics was held at the Colorado School of Mines. For some years, these symposia were sponsored by the major U S mining schools. In Eastern Europe, an International Bureau of Rock Mechanics which had a mining emphasis, was established at the Third International Strata Control Congress held in Leipzig, Germany, in 1958. Under the leadership of Leopold Müller, the International Society for Rock Mechanics (ISRM) was established in 1962. The ISRM held its First International Congress in Lisbon, Portugal, in 1964, the year in which the first volume of the International Journal of Rock Mechanics and Mining Sciences was published. It may be concluded then, that by the early 1960s, rock mechanics had become an identifiable engineering discipline with its specialist journals, books, conferences, societies, and courses and research programs in universities. It was from this time that modern rock mechanics and mining geomechanics underwent significant development. That development has been
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explored in some detail by Hood and Brown (1999). Only selected aspects of that history will be outlined here. The many seminal contributions made in respect of underground coal mining and the deep level hard rock mining of tabular orebodies, in particular, will not be considered. The development of geomechanics studies for the open pit and underground methods of mining being considered here will be dealt with separately, although there are obvious overlaps in areas such as geotechnical characterisation, rock and rock mass properties and methods of analysis. 4.3 Modern mining geomechanics By the mid-1960s, open pit mining was being carried out on an increasing scale and at increasing depths. As a consequence, the stability of open pit slopes became of major concern, not only from the obvious perspective of safety but also in terms of the overall economics of operations. By 1970, specialist conferences on open pit slope stability were being held (e.g. Van Rensberg 1970, Brawner and Milligan 1971). An example of a major open pit slope failure occurring in that period was that at the Chuquicamata mine, Chile, in 1969 (Kennedy and Niermeyer 1970). The application of the emerging discipline of rock mechanics to open pit slope design was undeveloped at the beginning of this period. Quite often, attempts were made to apply soil mechanics principles or to treat the problem as one in elastic stress analysis. In retrospect, Stacey (1993) noted that "in the early 1970s, a review of hard rock slopes around the world by height and slope angle put the majority of slopes less than 500 ft (150 m) in height, with slope angles for the higher slopes typically in the 45o range or less. At the time, slope design technology was still developing and most of the pits had been designed on the "45 degrees seems reasonable" philosophy." An important advance was made when the crucial role played by discontinuities in controlling the stability of most open pit slopes became clear. Although it had long been recognised that these structural features influenced slope behaviour (Terzaghi 1962, Müller 1964), many workers had assumed that a jointed rock mass could be regarded as an equivalent orthotropic medium. Soon, the need to describe and model the strength and deformation behaviour of key discontinuities became clear (e.g. John 1962). Pioneering work presented to the First ISRM Congress by Patton (1966) had shown that, at low normal stresses, the roughness of joint surfaces has an important effect on the shear strength of joints. In the mid- to late 1960s, research programs on open pit slope stability were being carried out, or were begun, in several parts of the world including Canada, South Africa, the UK, the USA and the USSR. One of the most significant of these programs was the industrially sponsored project led by Professor Evert Hoek at the Royal School of Mines, Imperial College, London. Hoek and other early workers recognised the value of the structural geologists’ approach of using the stereographic projection for analysing structural data. Hoek’s team, and others, developed methods of collecting and analysing discontinuity data and of using the stereographic projection to identify likely modes of rock slope failure. They also developed simple but effective instruments and procedures for the field and laboratory measurement of rock properties. From the late 1960s, significant contributions to these areas of rock mechanics were also made by several groups of international experts working as part of the ISRM Commission on Testing Methods to develop and publish a range of suggested methods of rock characterisation, testing and monitoring (Brown 1981). 26
New methods were developed for analysing single plane sliding and wedge failures, using both analytical solutions and the stereographic projection. Some of these methods had been developed initially for civil engineering applications (e.g. John 1968, Londe 1965). The influence of blasting practice and drainage on slope stability was recognised, as was the importance of monitoring the performance of slopes. A comprehensive account of this work was published by Hoek and Bray (1974) in their classic text, Rock Slope Engineering. Subsequently, revised editions were published in 1977 and 1981. Overviews of this early work were given at the 3rd ISRM Congress in Denver by Hoek and Londe (1974) and by Goodman (1974). By the mid- to late 1970s it was recognised that the solution of some types of slope stability problem required that approximations be made to the strength and deformation characteristics of large masses of rock containing several sets of discontinuities. The HoekBrown empirical rock mass strength criterion, developed originally as part of an industrially funded project on underground excavations in rock (Hoek and Brown 1980), was soon used to address this issue in slope stability analyses (Hoek and Bray 1981). At about the same time, probabilistic and risk analysis approaches to slope stability studies were being developed (e.g. Coates 1977, McMahon 1975). It had been recognised that, in some cases, the rock mass should be regarded as a discontinuum or an assembly of interacting blocks each of which may be free to translate or rotate (e.g. Trollope 1968). It was in studies of the dynamics of rock slopes in jointed rock that Cundall (1971) originally developed his distinct element approach to the numerical modelling of discontinua. It had also been recognised that, in many cases, intact bridges of rock are present between discontinuities (Jennings 1970). Although techniques for dealing with this problem were developed, they involved a number of assumptions and uncertainties. As will be discussed in Section 5 below, this remains a largely unresolved problem today. So we see that by the early 1980s, most of the basic geomechanics knowledge, methods of collecting rock mass characterisation and mechanical property data, and the techniques and framework for slope stability analyses were in place. The ability to carry out geomechanics studies of open pit rock slopes was enhanced by the numerical analysis methods that had been developed in the preceding 15 years or so and by the ever increasing computing power that became available to engineers. The use of a geomechanics based approach to rock slope design in place of the presumptive approach used previously has allowed deeper and steeper open pit slopes to be engineered, optimised and managed with positive results in terms of productivity and profitability (Karzulovic 2004, Stacey 1993). More recent developments and the current geomechanics issues associated with large scale open pit mining will be discussed in Section 5 below. From the early to mid-1960s, the emerging engineering discipline of rock mechanics was applied increasingly to large-scale underground metalliferous mining in several parts of the world. In Australia, a number of mining companies used the expertise built up on the Snowy Mountains Hydro-electric Scheme in the 1950s and 60s (May 1980). In 1963, Mount Isa Mines established what was for some time one of the strongest applied rock mechanics programs on a particular mine site anywhere in the world (Mathews and Edwards 1969). Although open stoping and caving methods of mining were well established, cut-and-fill mining of metalliferous orebodies became increasingly important from the mid1960s and the subject of basic and applied research
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programs in Australia, Canada and Sweden. Importantly, this work involved government research organisations, industry groups which often sponsored and managed the research, university research groups and individual mine sites which often contributed field test sites as well as staff expertise. The results of these various studies are reported in the specialist conference proceedings edited by Stephansson and Jones (1981). Subsequently, attention was focussed on the geomechanics problems associated with open stoping methods of mining, building on the mining geomechanics knowledge and techniques developed in the "cut-and-fill era" (Brown 1992). Up to the 1980s, advances were made in site characterisation and mine model formulation, in situ stress measurement, design analysis, numerical modelling, support and reinforcement systems including cable bolting, fill technology (hydraulic fill, rock fill and, more recently, paste fill), mine scheduling and pillar recovery, and performance monitoring and retrospective analysis (Hood and Brown 1999). From 1972 to 1976, Professor Evert Hoek and his team at the Royal School of Mines, Imperial College, London, carried out an industrially sponsored research project on the design of large underground excavations in rock, modelled on the highly successful rock slope project. The results of this project were reported by Hoek and Brown (1980). A later project carried out in Canada with the support of the Canadian mining industry built on this foundation, developing a number of new computer-based analytical tools and advancing the state-of-the-art of rock support and reinforcement for hard rock applications (Hoek et al 1995). Significant mining geomechanics research programs, associated largely with open stoping methods of mining continue in Canada to the present day. The importance of geomechanics to underground mass mining was emphasised by the series of international mass mining conferences held in Denver (Stewart 1981), Johannesburg (Glen 1992), Brisbane (Chitombo 2000) and now Santiago (Karzulovic and Alfaro 2004). In the context of the theme of this paper, an important development occurred in 1997 when the Julius Kruttschnitt Mineral Research Centre, Brisbane, Australia, and the Itasca Consulting Group, Inc, Minneapolis, USA, began work on the International Caving Study (ICS) Stage 1 (ICS I). This study was sponsored by an international consortium on mining companies engaged in, or planning, caving operations. The issues addressed by this study and its successor, the International Caving Study Stage II (ICS II) will be outlined in Sections 6.1 and 8 below. Suffice to say that, as shown by Figure 4, the need for these studies arose from the companies’ interest in mining larger, better quality and deeper orebodies by highly productive mass mining methods. Despite this apparently positive picture of the continued development and application of mining geomechanics in both open pit and underground mass mining since the early 1960s, there have been some disturbing developments (or absence of them) in the past 10-15 years. There has been a decline in the numbers of government funded research organisations working on mining geomechanics, the closure of a number of University mining programs, and an apparent drying up in the supply of well educated and trained geomechanics practitioners available in some parts of the world. These issues will be discussed in Sections 8 and 9 below. 5. GEOMECHANICS AND LARGE-SCALE OPEN PIT MINING The development of mining geomechanics for large-scale open pit mining was outlined in the preceding section where Massmin 2004
it was shown that most of the fundamental knowledge and tools were in place by the early 1980s. Since that time, geomechanics practice in large open pits has evolved, most notably in the area of slope management. This overview of geomechanics and large-scale open pit mining draws heavily on recent papers by Calderón et al (2003), Flores and Karzulovoic (2000), Hoek et al (2000a,b) and Karzulovic (2004). Hoek et al (2000a,b) emphasise the fundamental importance of a comprehensive and reliable geological model to any large open pit slope design. Geological models are now routinely developed, stored and manipulated using powerful computer-based modelling tools. The ability to interface these models with geotechnical data and stability analyses, including numerical analyses, adds to their utility in open pit slope engineering. The geotechnical and hydrogeological characterisation of the rock masses forming the current and future pit slopes is the next essential step in the process. Although the techniques for doing this have been established since the early 1980s, advances in the use of digital technologies for collecting, storing, processing and manipulating data have been made in recent years (Brown 2003). The process of collecting and evaluating geotechnical data in large open pits is never-ending. As with underground mining where the problem is even more acute, one of the major difficulties is the fact that it is possible to sample and characterise only a small proportion of the rock mass directly in advance of excavation. Although the emphasis in both geological and geotechnical data collection and model formulation for open pit mining is on structural features, the nature and properties of the rock material should also be considered. The possibility of failure through the rock material in high slopes, and the impact of variable rock strengths associated with weathering and alteration, should not be overlooked. Hoek et al (2000a,b) discuss the relative impacts of several types of alteration of the rocks at the Chuquicamata mine, Chile, for example. It is common practice, and indeed necessary for practical purposes, to divide the rock mass around the pit into structural or geotechnical domains within each of which a constant geotechnical model may be used. As mining proceeds and more data become available, it may be necessary to refine the original definition of these geotechnical domains. The next step is to carry out a series of slope stability analyses for the stage in the life of the pit being considered. Hoek et al (2000a) discuss the approach generally used in the following way: "The current state of practice tends to separate slope designs into two distinct categories. The first of these categories is for those designs that can be dealt with in terms of kinematically possible structurally controlled failures. For example, failures that involve wedges sliding along the line of intersection of two intersecting faults can be analysed using limit equilibrium models. This type of failure is commonly seen in slopes of up to 20 or 30 m high in hard jointed rock masses. The design of such slopes can sometimes be based upon an analysis of simple wedge failure. The second category is that which includes nonstructurally controlled failures in which some or all of the failure surface passes through a rock mass which has been weakened by the presence of joints or other second order structural features. An assumption commonly made is that these second order structural features are randomly or chaotically distributed and that the rock mass strength can be defined by a simple failure criterion in which ‘smeared’ or ‘average’ nondirectional strength properties are assigned to the rock mass. This approach is frequently used for the analysis of the overall stability of large slopes where it is believed that no obvious failure mode presents itself."
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Hoek et al (2000a,b) go on to point out that experience shows that these two categories alone are inadequate for open pit slope design. Many or most large scale slope failures do not appear to follow either of these relatively simple models. They are generally more complex that these models assume, often involving failure on a major structure, rock mass failure of the type discussed in the preceding paragraph and/or a step-path failure controlled by joints and intervening rock bridges as illustrated in Figure 6. In addition, these complex failures such as that shown in Figure 7 may be truly three dimensional so that they are not amenable to representation and analysis in simple two dimensional form. Clearly, these failure mechanisms are difficult, if not impossible, to foresee ahead of the event. Even when examined after the failure, the exact mechanism and the failure surfaces are not always easy to identify. The writer has had recent experience of this difficulty in the cases of two open pit failures that were only in the order of 40 m high. The development of predictive techniques for identifying and analysing these large, complex failures is seen to be perhaps the most outstanding geomechanics research need for largescale open pit mining. Any successful approach is likely to involve computer based geological and geotechnical modelling, advanced numerical analyses, probabilistic methods and the use of risk analysis techniques. At a fundamental level, further research is also required to better define the strength and deformation properties of discontinuous in situ rock masses.
It is now widely recognised that it is impossible to prevent some slope failures in large open pits, especially on the scale of benches. Indeed, it is considered that any pit in which no failures occur is less than optimally designed for profitability. Modern practice is to use a risk management approach in which the risks are identified and managed (e.g. Catalán and Calderón 2004, Calderón et al 2003, Call et al 2000, Karzulovic 2004). The monitoring of slope movements and the establishment of acceptability criteria are important elements of this approach. As well as these geomechanics issues, the general issue of materials handling and the costs involved is probably the other major challenge faced by large-scale open pit mines as their depths approach and exceed 1000 m. These issues are outside the scope of this paper and of the writer's knowledge and expertise, and so will not be discussed in any detail here. However, their significance should be acknowledged. An overall future need will be to limit the amount of waste and ore hauled from depth to the surface using conventional methods. A change away from trucks to in-pit conveyors could enable overall pit slopes to be steepened as well as saving on the capital, operating and maintenance costs of vehicles and haul roads. A further possibility into the future would be to carry out some parts of the processing cycle in underground chambers excavated below the pit floor. This concept has a number of geomechanics implications in terms of stand-off distances from the pit and the stability of the underground excavations themselves. 6. GEOMECHANICS AND CAVING METHODS OF MINING
Figure 6: Candidate failure surface involving a number of different shear failure mechanisms (Hoek et al 2000a).
Figure 7: A large-scale open pit slope failure (Hoek et al 2000a). 28
6.1 Block caving geomechanics Current practices and trends in block and panel cave mining are discussed in the paper by Flores et al (2004a). Many, if not most, of the major challenges faced in engineering the modern generation of block and panel caving mines, and those in the feasibility or planning stages, involve geomechanics issues. Many of these challenges arise because of the high rock mass qualities of the orebodies being mined (in which the critical issues of cavability and fragmentation are accentuated), the greater depths at which caving is being initiated, the associated problems with extraction level stability, and the increasing heights of caving columns. In some cases, the interaction of the cave with a pre-existing open pit becomes an additional consideration (e.g. Glazer and Hepworth 2004). Table 1 lists the major geomechanics-related issues and activities involved in the investigation, preliminary study, planning, design, construction and operating stages of block and panel caving mines. Even a cursory consideration of these issues and activities will demonstrate the vitally important role that geomechanics plays in modern cave mining, even without considering tailings management, environmental management and mine closure issues, all of which can have significant geomechanics components. Most of the issues and activities listed in Table 1 have been discussed in detail by Brown (2003) and in a range of papers presented to this and previous mass mining conferences. In order to provide some further detail, some of the geomechanics methods and processes involved in cavability assessment and in cave inducement are set out in Figure 8 which comes from the CaveRisk risk assessment methodology developed as part of ICS I (Brown 2003). Although this and the other issues listed in Table 1 have received a considerable amount of attention in recent years by both in-house groups in mining companies and in research projects such as the ICS, current levels of understanding of caving mechanics and cave engineering do not always match industrial requirements, particularly when major risks are involved in new mining ventures that "break
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Table 1: Major geomechanics-related issues and activities associated with modern block and panel caving mines PRELIMINARY STUDIES
Geotechnical model formulation Cavability assessment Fragmentation assessment Mining method selection Identification of potential major hazards Subsidence prediction Risk analysis
MINE PLANNING, DESIGN AND CONSTRUCTION
Undercut level selection and undercutting strategy Excavation sequencing Extraction level design and construction, including support and reinforcement Subsidence prediction Infrastructure location, design and construction
MINING OPERATIONS
Undercutting and caving initiation Draw management Monitoring (e.g by seismic monitoring) and managing cave propagation (e.g. by draw control, rock mass conditioning and cave inducement) Monitoring excavation (including extraction level) performance Management of major operational hazards including uncontrolled collapses, rockbursts, mud rushes and air blasts Monitoring the development of surface subsidence
new ground". The need for continuing fundamental investigation of these issues will be outlined in Section 8 below. The centrally important issues of caving mechanics and cave propagation are addressed in the paper by Flores et al (2004b). In the bench-marking exercise carried out as part of ICS II (Flores et al 2004a), it was concluded that there were no geotechnical tools available with which to define the likelihood of vertical caving propagating through ore columns of varying heights (sometimes outside the limits of precedent practice) in rock masses of varying rock mass qualities and strengths, and subjected to a range of possible in situ stress fields. The model used for caving initiation and propagation by stress caving is illustrated in Figure 9. Flores et al (2004b) carried out a series of parametric two dimensional stress analyses to assess the effects of several aspects of geometry, stress field and rock mass strength on vertical cave propagation. They defined a Caving Propagation Factor for use in estimating the likelihood of vertical cave propagation in the initial engineering stages of cave mining projects, including those involving a transition from open pit to underground cave mining.
(a) Initial undercutting defining a flat, tabular cavity where the stress release caving mechanism predominates.
(b) The upward caving propagation makes the cave back curve, increasing the importance of the stress caving mechanism.
(c) Additional upward caving propagation increases the curvature of the cave back, and makes predominant the stress caving mechanism.
Figure 8: Cavability logic tree (Brown 2003) Massmin 2004
Figure 9: Evolution of the cave back and caving mechanisms through time due to the upward propagation of caving (Flores et al 2004b). Santiago Chile, 22-25 August 2004
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Figure 10 shows a similar diagram to Figure 8 for a second risk management focus issue, excavation stability. In order to illustrate the application of geomechanics knowledge and techniques, and because it is becoming of increasing concern under the higher stress conditions being encountered in deeper mines, the general issue of extraction level excavation stability will be discussed in more detail in the following sub-section. 6.2 Extraction level excavation stability Brown (2003) lists and discusses the major factors influencing extraction level layout and design as being fragmentation, undercut strategy and design, geotechnical conditions, operational factors (relating to the efficient removal of broken ore), major operational hazards and the stability, reinforcement, wear and repair of drawpoint brows. The paramount importance of geomechanics issues is again apparent. The stability, support and reinforcement of extraction level excavations has long been a source of difficulty, production delays, cost increases and even the loss of productive sectors in caving operations. In addition to geotechnical factors, poor mining practices such as poor sequencing and poor draw control can cause excavation instabilities.
Salvador mine, Chile, with emphasis on the use of a flexible support and reinforcement system and pre-reinforcement. However, further research is required to develop reliable numerical methods and rock mass constitutive laws for use in predictive analyses (e.g. Wattimena 2003). The case involving brittle fracture or even rockbursts has been, perhaps, the more difficult of the two general cases to understand and manage. The writer suggests that advances can be made by addressing this problem through the application of concepts and techniques developed in studies of brittle rock behaviour in other engineering applications. The ideas considered to be worthy of further exploration in this context are the influence of stress path on rock mass behaviour and the adaptation of the Hoek-Brown empirical failure criterion (Hoek and Brown 1980, 1997) for brittle, slabbing conditions developed by Martin (1997) and Martin et al (1999). In laboratory and field and field studies of the behaviour of Lac du Bonnet granite, Martin (1997) found that the start of the fracture or failure process began with the initiation of damage caused by small cracks growing in the direction of the maximum applied load. For unconfined Lac du Bonnet granite, this occurred at an applied stress of 0.3 to 0.4 sc where sc is the uniaxial compressive strength of the intact rock material. As the load increased, these stable cracks continued to accumulate. Eventually, when the sample contained a sufficient density of these stable cracks, they started to interact and an unstable cracking process involving sliding was initiated. The stress level at which this unstable cracking process is initiated is referred to as the long term strength of the rock, scd. As illustrated in Figure 11, Martin (1997) first determined the laboratory peak, long term and crack initiation strengths for the Lac du Bonnet granite. He was able to fit Hoek-Brown failure envelopes to these curves, although the laboratory crack initiation curve was found to be a straight line on s1 versus s3 axes. Subsequently, in a field experiment carried out at the Underground Research Laboratory site in Manitoba, Canada, the initiation of cracks around a tunnel
Figure 10: Excavation stability logic tree (Brown 2003) Extraction level excavations can be subjected to high stresses because of the high percentage excavation on this level, the stress concentrations associated with the advancing cave front, and the increasing depths at which extraction levels are being developed. These excavations are also subjected to complex stress paths and repeated static and dynamic loading cycles during mine development and operation. Depending on the relation of the rock mass quality to the induced stresses, squeezing (or plastic deformation) or brittle fracturing (including rockburst) conditions may exist. (There is also an important third case in which instabilities are associated with structural features (e.g. Flores et al 2004a). This case will not be considered here.) The first case can be understood and managed through the use of ground-support interaction analysis and the application of the support and reinforcement principles that are well established in civil engineering tunnelling (Brown 2003). Leach et al (2000) provide an instructive example of the calculation of ground reaction curves and their application in the design of extraction level excavations in the C-cut area of the Premier Mine, South Africa. Van Sint Jan et al (1987) provide an excellent example of the successful application of modern support and reinforcement principles and practices on the extraction level of the El 30
Figure 11: Hoek-Brown failure envelope for Lac du Bonnet granite based on laboratory peak strength (Lab Peak), longterm strength (Lab scd) and in situ crack initiation stress (sci) determined by microseismic monitoring (after Martin 1997).
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excavated in the Lac du Bonnet granite was recorded using microseismic emissions. As shown in Figure 11, these data corresponded well with the laboratory crack initiation data. It was found that crack initiation at approximately constant deviatoric stress, (s1 - s3), could be well represented by the Hoek-Brown criterion with mb = 0 and s = 0.11 (Martin et al 1999), in which case, s1 - s3 = 0.33 sc. This criterion was used in conjunction with elastic stress analyses to give good predictions of the geometry of the spalled zone around the tunnel. It has since been used to predict brittle spalling or slabbing (as opposed to general shear failure) conditions in a number of underground excavations (e.g. Cai et al 2004). The suggestion is that this criterion could also be used in analyses of the likelihood of brittle rock fracture around extraction level excavations in block and panel caving mines. In addition to the uniaxial compressive strength of the rock material, the state of stress on the boundary and in the rock around particular excavations would have to be estimated. This can be done through a three dimensional elastic stress using the finite difference code, FLAC3D. Wattemina (2003) used this approach to show how elevated stresses may be produced in the extraction level excavations ahead of and near the cave front, depending on the undercutting strategy used, how the stresses at a given level in the extraction level change as the cave front approaches and passes over the point, and the effect of high horizontal in situ stresses on the stresses induced in the major and minor apices. Depending on the geometry of the problem to be investigated, two dimensional plane strain analyses may be used in some cases. It has been found that the loading path taken to the current state of stress in geological materials can influence the strength able to be developed by a soil or rock mass. This is particularly the case when plastic deformation is involved. However, when deformation is essentially elastic until "failure" and brittle fracture as opposed to general plastic deformation occurs, it has been found that the strength envelopes of "hard" rocks are essentially stress path independent (Brady and Brown 2004). It is considered likely, therefore, that as in the cases described by Martin et al (1999) and Cai et al (2004), elastic stress analyses and Martin’s representation of the Hoek-Brown criterion for crack initiation will suffice for making a first-order estimate of the occurrence and extent of brittle fracture around extraction level excavations in strong, massive rock. Clearly, this postulate will have to be tested by parametric numerical analyses and by comparison, and possible calibration, with high quality field data. This is considered to be a fertile topic for future research. In such studies, the influence of repeated loading would have to be considered.
and the frequency and orientations of the discontinuities in the rock mass. Transverse layouts generally have advantages over longitudinal layouts for orebodies of sufficient width (Bull and Page 2000) and are used for modern high production SLC operations such as the Kiruna Mine, Sweden (Quinterio et al 2001) and the Ridgeway Gold Mine, New South Wales, Australia (Trout 2002). In the last decade or so, larger sublevel and production drift spacings and drift sizes have been introduced in largescale, highly productive SLC operations. For example, the layouts used at Kiruna have been progressively scaled up since the mining method was introduced more than 40 years ago (Hustrulid 2000, Marklund and Hustrulid 1995). A major effect of the scale-up, even with increased drift sizes to accommodate larger equipment, has been to reduce the proportion of ore extracted during development from 15% to 5%. Subsequently, full-scale trials have been conducted with the sublevel interval increased from 27 m to 32 m, the drift width increased from 7 m to 11 m, and the burden increased from 3.0 m to 3.5 m (Quintero et al 2001). The transverse sublevel caving layout used at the Ridgeway Gold Mine, New South Wales, Australia, has a sublevel interval of 25 m, a cross-cut drift spacing of 14 m and a drift size of 6 m wide by 4 m high (Trout 2002, Power 2004). Despite the attention that they have received, fragmentation and the gravity flow of caved ore remain major issues for SLC operations (e.g. Bull and Page 2000, Rustan 2000, Power 2004). It appears that some of the classic concepts introduced by Janelid and Kvapil (1966) and widely used for many years, may not be applicable to the coarser fragmentation and flow of stronger orebodies in which flow may be episodic, chaotic and not interactive (Power 2004). Modern SLC designs have departed from those derived from considerations of Janelid and Kvapil’s theories. Further investigation is required to develop an improved understanding of these issues. Advances made recently suggest that realistic numerical simulations of SLC flow may soon be achievable (Power 2004). Draw control and dilution remain major concerns for some SLC operations. Well-designed and controlled drilling is required for successful sublevel caving. Blasting under confined conditions remains an issue requiring further investigation. Finally, geomechanics problems have arisen in production headings as a result of the concentration of field stresses in the lower abutment of the mining zone. Modern support and reinforcement principles and techniques derived from civil engineering and other forms of mining practice can help overcome these problems (e.g. Struthers et al 2000). However, this and the other issues identified may become accentuated as SLC is practiced at increasing depths.
6.3 Sublevel caving geomechanics Generally, sublevel caving (SLC) is suitable only for steeply dipping orebodies with reasonably strong orebody rock enclosed by weaker overlying and wall rocks. The ore must be of sufficient grade to accept dilution, perhaps exceeding 20%, arising from the entrainment of barren country rock in the ore stream. The method produces significant disturbance of the ground surface, imposing some possible limitations on its applicability. SLC may have higher production costs than the other underground mass mining methods being considered here because of the relatively high development requirement per tonne produced and the intensity of the drilling and blasting required to generate mobile, granular ore within the caving medium. Reduction of these development costs is one of the objectives of modern SLC operations. The choice between longitudinal and transverse layouts involves consideration of a range of factors such as orebody dip and plunge, orebody dimensions, the in situ stresses,
7. GEOMECHANICS AND LARGE-SCALE OPEN STOPING METHODS
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As outlined in Section 3, large-scale open or sublevel stoping of massive or steeply dipping stratiform orebodies has been practiced successfully for more than 30 years. Since open stoping requires unsupported, free-standing stope boundary surfaces, the strength of the orebody and country rock masses must be adequate to provide stable walls, faces and crowns of excavations. The orebody boundary must be fairly regular since selective mining is precluded by the requirement for regular stope outlines associated with the use of long blast holes. Blast hole penetration of stope walls due to drilling inaccuracy leads to dilution (Brady and Brown 2004). Greater control of drilling accuracy may be obtained with larger diameter blast holes. Hamrin (2001) refers to 140-165 mm holes being drilled accurately to depths of 100 m in bighole open stoping. Pillar recovery or secondary stoping is commonly
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practiced in open stoping. Backfill with a range of properties may be placed in the primary stope voids and secondary stoping or pillar mining performed by exploiting the local ground control potential of the adjacent fill. Alternatively, on the boundaries of ore blocks, pillars may be blasted into adjacent stope voids with the possibility of extensive collapse of the local country rock. Successful ore recovery would then require draw of fragmented ore from beneath less mobile, barren country rock with the potential for dilution (Brady and Brown 2004). Dilution may also occur from overbreak and the sloughing or failure of fill. On the other hand, there is also the potential for ore loss arising from insufficient ore breakage within the stope boundaries (Villaescusa 2000). The evidence shows that the development and successful implementation of these mining methods has relied heavily on advances in, and the application of, mining geomechanics knowledge and techniques in areas such as rock mass characterisation, numerical stress analysis, mine sequencing analysis, support and reinforcement including cable bolting and pre-placed reinforcement, crown pillar stability, fill technology and drilling and blasting technology (e.g. Alexander and Fabjanczyk 1981, Brady and Brown 2004, Goddard 1981, Grant and De Kruijff 2000, Potvin and Hudyma 2000, Simser and Andrieux 2000). As the depth and scale of these operations increase, problems of excavation stability and dilution can be expected to be exacerbated and the contributions made by geomechanics accentuated. In some cases, depending on the mining methods and equipment used, it may become necessary to reduce some excavation sizes. The role of fill, including the paste fill now being used increasingly, in ensuring the stability of local and overall mine structures will become ever more critical. The timely supply and placement of fill in stopes, sometimes over long distances, can also be expected to become a critical factor in some cases. Increasing productivity by reducing development and stope cycle times is likely to become an overall objective. 8. RESEARCH The summary of the development and application of mining geomechanics to mass mining given above shows that, since the 1960s, the mining industry has been active in working with university and government research groups on research programs that were identified as being relevant to the industry’s needs. Although, by the very nature of research, not all sponsored research produced significant results, there can be no doubt that the industry gained much benefit from its involvement in mining geomechanics research (e.g. May 1980, Watson 1987). At the time, there were strong government, industrial and university mining geomechanics research groups in many countries including, to the writer’s personal knowledge, Australia, Canada, France, Germany, India, South Africa, Sweden, the UK, the USA and the USSR. Not only were those programs important in terms of their contributions to knowledge, but they were also the sources of significant numbers of highly trained geomechanics specialists for the industry. A major feature Hood and Brown’s (1999) interpretation of the history of mining geomechanics is the great influence that a number of inspirational leaders had on the rapid development of the discipline and of the most influential and highly acclaimed research groups from the 1960s. As would be expected, several of these "founding fathers" are no longer living and many of those that are have either retired from their university or research positions or now work as consulting engineers. As a result, with a few exceptions, the strong research groups that they built up have either disappeared or now operate at much lower levels than they 32
did at their peaks. Furthermore, in a number of countries (e.g. Germany, UK, USA), government funded mining research organizations and mining research programs have either been closed completely or are supported at significantly reduced levels. Despite this, some mining geomechanics research is still being carried out in some of the countries identified in the opening paragraph of this section, and new programs have emerged in countries such as Chile and China. However, to the best of the writer’s knowledge, no major research programs on open pit slope stability on the scale of those carried out in the late 1960s and 1970s, are currently underway. As indicated in Section 5, there is a pressing need in this area. It is understood that plans for an industrially sponsored research project on large-scale open pit geomechanics are being developed in association with the mass mining technology initiative to be outlined below. Although many of the research programs that previously supported the development of initially cut-and-fill and then open stope mining no longer exist, the geomechanics of caving methods of mining has been studied for the last eight years by the ICS. The main topics addressed have been: • rock mass characterisation and the simulation of in situ jointing; • review, development and calibration of methods for cavability and fragmentation assessment; • undercut and extraction level design; • flow of broken rock using numerical models, a large-scale physical model and mine scale marker tests; • development of a draw control and scheduling system using linear programming and mixed integer linear programming; • pre-conditioning of strong rock masses using hydraulic fracturing methods; • development of geotechnical guidelines for the transition from open pit to underground mining by caving methods; • development of a risk assessment methodology for block and panel caving; and • collation of caving practice and knowledge, and the results of research, in a readily accessible form. Some of these topics are also being addressed on mine sites and by other research groups outside the framework of the ICS (e.g. pre-conditioning at Codelco’s Andina Division). Although useful advances have been made (e.g. Brown 2003), much remains to be done to develop the understanding and knowledge required for the engineering of the next generation of underground mass mining operations. For some time, the sponsors of the ICS have been working with the program’s Technical Director, Dr Gideon Chitombo, to establish the successor to the ICS in the area of mass mining technology. The proposed research areas and the associated research tasks are listed in Table 2 (Chitombo 2004). Clearly, many of these research tasks have geomechanics orientations. Research on these topics is considered to be essential in developing the knowledge and techniques required to successfully mine in the underground mass mining environment of the future. 9. EDUCATION AND TRAINING The importance of the engineering discipline of geomechanics to mass mining being argued here implies a requirement for a continuing supply of well-educated and trained (the writer distinguishes between the two) engineers to advance and apply the discipline. The requirement is not only for geomechanics specialists to work on mine sites, in consultancies, in research organisations and as teachers, but also for practicing mining engineers who have a good working knowledge of the discipline and its impact.
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The shortage of mining geomechanics engineers in many parts of the world, including those mainly English-speaking countries with which the writer is most familiar, is part of a perceived wider shortage of mining engineers and other professionals for the minerals industry. In some countries, the closure of mining and minerals engineering courses, of specialist postgraduate courses, and of mining research groups, has given cause for concern (Hood and Brown 1999). As a result, in several mining countries, the professional institutions, industry leaders, those universities still active in the field and the minerals industry press, have discussed and sought solutions to what is seen as being a critical future shortage of minerals industry professionals. In Australia, for example, a study carried out by a taskforce of the peak industry body, the Minerals Council of Australia (Minerals Council of Australia 1998), resulted in the establishment of the Minerals Tertiary Education Council (MTEC). MTEC aims "to build a world-class tertiary learning environment for the education of professional staff for the Australian minerals industry." The development and delivery of a flexible-format, web-based rock mechanics education program for undergraduate mining engineers forms part of MTEC’s program (Lilly et al 2003). Thankfully, as university mining and minerals education has declined in some countries, it has been maintained or developed in some others, including Chile. The problem is many-facetted and almost as old as the industry itself. Some aspects of the problem are associated with young peoples’ interests and aspirations, some may be attributed to the secondary education systems in some countries, some are the responsibility of universities and of university funding systems, while some must be visited on the mining companies themselves. This is not the place to explore the reasons for, and advance possible solutions to,
the overall problem. However, some specific comments about mining geomechanics are warranted. A research training through post-graduate research into an industry-related problem, followed (or preceded) by mine site experience is considered to provide an especially strong preparation for mining geomechanics specialists (Golder Associates 2001). The decline of mining geomechanics research groups in some parts of the world, and the failure to establish them in others, has had deleterious effects on the supply of well-trained specialists for the industry. Nevertheless, it must be recognised that some strong university mining geomechanics research groups with interests in mass mining exist in Canada and Australia, for example. In addition to contributing to advances in knowledge and understanding, industry support for projects like the ICS and the proposed mass mining technology project should serve the further important purpose of helping train some of the industry’s current and future engineering specialists to the advanced levels that will be necessary to develop solutions to the problems likely to be encountered in future mass mining operations. The writer has the highest personal regard for many of the experienced geomechanics specialists working on mine sites and in consultancies in his home country, Australia, and in several other parts of the world, including Chile. These specialists come from a wide range of initial educational and subsequent training backgrounds. In some countries, many geomechanics specialists have civil engineering, geological engineering or engineering geology, rather than mining, backgrounds. However, the living and social conditions on many remote mine sites (including the fly in-fly out system), the lack of opportunities for career advancement in specialist fields, and some company cost-
Table 2: Proposed mass mining technology research areas and tasks (Chitombo 2004) Research Area
Research Tasks
Geotechnical Characterisation
Effective use of geophysical methods for assessing fundamental geological, geomechanical and hydrological rock mass parameters.
Mechanics of Caving
Better prediction of cave initiation and propagation for the "new mining environments" including surface subsidence prediction. The mechanics of in situ to primary to secondary fragmentation for caving operations. Influence of pre-conditioning on cave initiation and propagation.
Caving Engineering
Mechanism of confined blasting and corresponding drill and blast design principles and guidelines. Gravity flow and engineering studies (large-scale physical models and field studies) and development of mine scale flow rules and design guidelines. More effective cave monitoring systems (3D) and design of appropriate instrumentation geometries including on-line analysis and interpretation. Risk approaches for major hazard identification including air blast, rock bursts, major collapses, mud rushes, and subsidence. The impact of pre-conditioning on rock mass strength reduction and change of in situ fracturing.
Systems Engineering
Application of more effective cave/draw management strategies. Process management and effective use of automation.
Enabling Technologies
Rapid horizontal development (safe and quality). Effective (and continuous) ore handling systems including mechanical gatherers. Mechanisation of secondary rock breakage.
Mine Profit Optimization
Mining sequence, scheduling and mining rates.
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cutting policies, can make it difficult to attract and retain geomechanics specialists on mine sites. Indeed, on some Australian operations, there appears to be an increasing trend to use consultants rather than company staff for geomechanics functions (Anon 2004). As indicated above, this is not the place to explore possible solutions to this critical problem. The essential point to be made in the context of this paper is that a shortage of mining geomechanics specialists could provide a major impediment to safe and profitable surface and underground mining in the challenging mass mining environment of the future.
10. CONCLUSIONS The scale of the world’s largest surface and underground mass mining operations has grown at a continuing rate over the last 100 years. Indications are that both the rates of production and the depths of open pit and underground mines will continue to grow into the future. This can be expected to provide significant challenges to the engineering discipline of geomechanics which has been shown to have been a major contributor to the success and growth of mass mining over the last 40 years. Important challenges for many operations remain in other engineering disciplines such as project management, materials handling, environmental engineering, water supply and management, information and communications technologies, and management and systems engineering to improve efficiency and achieve cost reduction. Nevertheless, the importance of geomechanics to the investigation, design, construction and safe operation of profitable large-scale surface and underground mines requires that it be recognised as the basic engineering discipline for mass mining. It is essential, therefore, that the industry continues to support research into a range of geomechanics-related issues and the education and training of the future generation of mining geomechanics specialists.
ACKNOWLEDGEMENTS The writer is most grateful to the Organising Committee of MassMin 2004 for having invited him to prepare this paper and present the associated keynote lecture. He wishes to thank the Director and staff of the Julius Kruttschnitt Mineral Research Centre and the Manager and staff of the Brisbane office of Golder Associates Pty Ltd for having provided him with facilities and assistance in the preparation of this paper. He is particularly grateful to Dr Gideon Chitombo, German Flores (who also prepared Figures 2-5), Dr Antonio Karzulovic, Neil Hepworth and Dr Gavin Power for having provided ideas and material used in the paper.
REFERENCES • Agricola, G, 1556. De Re Metallica, 1st edition. (Trans: H C Hoover and L H Hoover, 1950). Dover: New York. • Alexander, E G and Fabjanczyk, M W, 1981. Extraction design using open stopes for pillar recovery in the 1100 orebody at Mount Isa. Design and Operation of Caving and Sublevel Stoping Mines, (Ed: D R Stewart), 437-458. SME: New York. • Anon, 1999. Definition of geotechnical engineering. Ground Engineering, 32(11): 39. • Anon, 2004. Miners turn to consultancies. Australia’s Mining Monthly, March, 59. • Brady, B H G, and Brown, E T, 2004. Rock Mechanics for Underground Mining, 3rd edition. Kluwer: Dordrecht. • Brawner, C O and Milligan, V (ed), 1971. Stability in Open Pit Mining, Proceedings 1st International Conference on 34
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• Hustrulid, W, 2000. Method selection for large-scale underground mining. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 29-56. Australasian Institute of Mining and Metallurgy: Melbourne. • Janelid, I and Kvapil, R, 1966. Sublevel caving. Int J Rock Mech Min Sci, 3(2): 129-153. • Jennings, J E, 1970. A mathematical theory for the calculation of the stability of slopes in open cast mines. Planning Open Pit Mines, (Ed: P W J Van Rensberg), 87102. Balkema: Cape Town. • John, K W, 1962. An approach to rock mechanics. J Soil Mech Foundns Div, ASCE, 88(SM4): 1-30. • John, K W, 1968. Graphical stability analysis of slopes in jointed rock. J Soil Mech Foundns Div, ASCE, 94(SM2): 497-526, • Karzulovic, A, 2004. The importance of rock slope engineering in open pit mining business optimization. Proceedings IX Symposium on Landslides, Rio de Janeiro. • Karzulovic, A and Alfaro, M (ed), 2004. Proceedings MassMin 2004, Santiago. • Kennedy, B A and Niermeyer, K E, 1970. Slope monitoring systems used in the prediction of a major slope failure at the Chuquicamata mine, Chile. Planning Open Pit Mines, (Ed P W J Van Rensberg), 215-225. Balkema: Cape Town. • Leach, A R, Naidoo, K and Bartlett, P, 2000. Consideration of design of production level drawpoint layouts for a deep block cave. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 357-366. Australasian Institute of Mining and Metallurgy: Melbourne. • Lilly, P A, Thomas, S J, Hebblewhite, B K and Galvin, J M, 2003. A rock mechanics education program for undergraduate mining engineers in Australia. Technology Roadmap for Rock Mechanics, Proceedings 10th Congress, International Society for Rock Mechanics, Johannesburg, 2: 779-782. South African Institute of Mining and Metallurgy: Johannesburg. • Little, T N and Szwedzicki, T, 1992. Mining geomechanics – a Western Australian School of Mines perspective. Proceedings AusIMM Annual Conference, Broken Hill, 5-11. Australasian Institute of Mining and Metallurgy: Melbourne. • Londe, P, 1965. Une méthode d’analyse à trois dimensions de la stabilité d’une rive rocheuse. Annales des Ponts et Chaussées, No 1, Jan-Feb, 37-60. • McMahon, B K, 1975. Probability of failure and expected volume of failure in high rock slopes. Proceedings 2nd Australia New Zealand Conference on Geomechanics, Brisbane, 308-313. Institution of Engineers, Australia: Sydney. • Marklund, I and Hustrulid, W, 1995. Large-scale underground mining, new equipment and a better underground environment – result of research and development at LKAB, Sweden. Trans Instn Min Metall, Sect A: Min Industry, 104: A164-168. • Martin, C D, 1997. Seventeenth Canadian Geotechnical Colloquium: the effect of cohesion loss and stress path on brittle rock strength. Can Geotech J, 34(5): 698-725. • Martin, C D, Kaiser, P K and McCreath, D R, 1999. HoekBrown parameters for predicting the depth of brittle failure around tunnels. Can Geotech J, 36(1): 136-151. • Mathews, K E and Edwards, D B, 1969. Rock mechanics practice at Mount Isa Mines Limited, Australia. Proceedings 9th Commonwealth Mining and Metallurgical Congress, London, 1: 321-388. Institution of Mining and Metallurgy: London. • May, J R, 1980. Industry-sponsored rock mechanics research in Australia. Proceedings 13th Canadian Rock Mechanics Symposium, 219-255. • Minerals Council of Australia, 1998. Back from the Brink: Reshaping Minerals Tertiary Education. Discussion
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Paper, National Tertiary Education Taskforce. Minerals Council of Australia: Canberra. Müller, L, 1964. Application of rock mechanics in the design of rock slopes. State of Stress in the Earth’s Crust, (Ed: W R Judd), 575-598. Elsevier: New York. Obert, L and Duvall, W I, 1967. Rock Mechanics and the Design of Structures in Rock, 650 p. John Wiley & Sons: New York. Pariseau, W G, Fowler, M E and Johnson, J C, 1984. Geomechanics of the Carr Fork Mine test stope. Geomechanics Applications in Underground Hardrock Mining, (Ed: W G Pariseau), 3-38. SME: New York. Patton, F D, 1966. Multiple modes of shear failure in rock. Proceedings 1st Congress, International Society for Rock Mechanics, Lisbon, 1: 509-513. Peele, R, 1941. Mining Engineers’ Handbook, 3rd edition. John Wiley & Sons: New York. Potvin, Y and Hudyma, M, 2000. Open stoping in Canada. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 661-674. Power, G, 2004. Full scale SLC draw trials at Ridgeway Gold Mine. Proceedings MassMin 2004, Santiago, (Ed: A Karzulovic and M Alfaro). Quintero, C R, Larsson, L and Hustrulid, W A, 2001. Theory and practice of very-large-scale sublevel caving. Underground Mining Methods: Engineering Fundamentals and Engineering Case Studies, (Ed: W A Hustrulid and R L Bullock), 381-384. SME: Littleton, CO. Rustan, A, 2000. Gravity flow of broken rock – what is known and unknown. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 557-567. Australasian Institute of Mining and Metallurgy: Melbourne. Savin, G N, 1961. Stress Concentrations Around Holes. (Trans: E Gros). Pergamon Press: Oxford. Simser, B and Andrieux, P, 2002. Open stope mining strategies at Brunswick Mine. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 675-683. Australasian Institute of Mining and Metallurgy: Melbourne. Singh, K H and Hedley, D G F, 1981. Review of fill mining technology in Canada. Application of Rock Mechanics to Cut and Fill Mining, (Ed: O Stephansson and M J Jones), 11-24. Institution of Mining and Metallurgy: London. Stacey, P F, 1993. Pit slope designs for the 21st century. Innovative Mine Design for the 21st Century, (Ed: W F Bawden and J F Archibald), 3-11. Balkema: Rotterdam. Stephansson, O and Jones, M J (ed), 1981. Applications of Rock Mechanics to Cut and Fill Mining, 376 p. Institution of Mining and Metallurgy: London. Stewart, D R, (ed) 1981. Design and Operation of Caving and Sublevel Stoping Mines, 843 p. AIME: New York.
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• Struthers, M A, Turner, M H, McNabb, K and Jenkins, P A, 2000. Rock mechanics design and practice for squeezing ground and high stress conditions at Perseverance Mine. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 755-764. Australasian Institute of Mining and Metallurgy: Melbourne. • Terzaghi, K, 1962. Stability of steep slopes on hard unweathered rock. Géotechnique, 12: 251-270. • Terzaghi, K and Richart, F E, 1952. Stresses in rock about cavities. Géotechnique, 3:57-90. • Trollope, D H, 1968. The mechanics of discontinua or clastic mechanics in rock problems. Rock Mechanics in Engineering Practice, (Ed: K G Stagg and O C Zienkiewicz), 275-320. John Wiley & Sons: London. • Trout, P, 2002. Production drill and blast practices at Ridgeway Gold Mine. Growing our Underground Operations, Proceedings 8th AusIMM Underground Operators’ Conference, Townsville, 107-117. Australasian Institute of Mining and Metallurgy: Melbourne. • Turchaninov, I A, Iofis, M A and Kasparyan, E V, 1979. Principles of Rock Mechanics. (Trans: A L Peabody). Terraspace: Rockville MD. • Van Rensberg, P W J (ed), 1970. Planning Open Pit Mines, Proceedings Symposium on the Theoretical Background to the Planning of Open Pit Mines with Special Reference to Slope Stability, Johannesburg. Balkema: Cape Town. • Van Sint Jan, M L, Valenzuela, L and Morales, R, 1987. Flexible lining for underground openings in a block caving mine. Proceedings 6th Congress, International Society for Rock Mechanics, Montreal, (Ed: G Herget and S Vongpaisal), 2: 1299-1303. Balkema: Rotterdam. • Villaescusa, E, 1996. Excavation design for bench stoping at Mount Isa mine, Queensland, Australia. Trans Instn Min Metall, Sect A: Min Industry, 105:A1-A10. • Villaescusa, E, 2000. A review of sublevel stoping. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 577-590. Australasian Institute of Mining and Metallurgy: Melbourne. • Watson, B, 1987. The impact of research and development on productivity in the Australian minerals industry. Technology and Exports: the Role of the Technological Sciences in Australia’s Export Performance, Proceedings 11th Invitation Symposium, 97-108. Australian Academy of Technological Sciences and Engineering: Melbourne. • Wattimena, R, 2003. Designing Undercut and Production Level Drifts of Block Caving Mines. PhD thesis (unpublished), University of Queensland, Brisbane. • Young, L E and Stoek, H H, 1916. Subsidence resulting from mining. Bull Univ Illinois Engrg Expt Stn, 91.
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Chapter 2
Geotechnical Characterization for Mass Mining
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Mining geotechnical investigations: The need for an integrated approach Allan Haines, T. Campbell McCuaig and Esther Theron, SRK Consulting, Australia
Abstract The need for an integrated approach in mining is a key element during all phases of the evaluation of a mineable deposit. The integration of the geological, structural, hydrogeological and geotechnical disciplines, as feeds into the mine planning and scheduling process, are essential elements in the conversion of a deposit into a mine. Having the geotechnically related risks and benefits defined at an early stage may well differentiate the project from others queuing up to obtain financing, whether from internal or external sources. From the initial greenfields site, from scoping to drill out and the establishment of an advanced project, pre-feasibility, feasibility and ramp-up into operations it is beneficial to establish the links between the geotechnical characteristics and mining risks. This paper will examine the timing, functional silo mentality, the influencers and modifiers on mine design, the benefits resulting from integration and some operational examples to illustrate the benefits of this approach.
1 INTRODUCTION In the future environment of surface and underground mines, we face both a trend towards the development of ‘super pits’ and ‘super caves’, and a tendency for the underground development of ageing pits. These super projects, or long-life ore factories, sprout from the calls for increased production and more cost effective solutions for mass mining projects, and perhaps from the belief that ‘bigger is better’. The move to fewer, larger, longer life projects exposes operations to an elevated level of risk from singular events. One of the most crippling and potentially fatal events in these mass mining operations is geotechnical failures. Safe and efficient operation of large scale open pit and underground excavations thus requires an even stronger focus on geotechnical issues. As geotechnical investigations rely on information supplied by the entire range of mining related disciplines, effective integration of these services is becoming even more crucial. Having the geotechnically related risks and benefits defined at an early stage may well differentiate the project from others, all queuing up to obtain financing, whether from internal or external sources, in today’s ever-competitive marketplace. In the author’s experience, the introduction of an integrated geological – structural – hydrogeological geotechnical approach, as early as possible in the development of a project, saves time and expense throughout all phases of the project life. 2 ‘FUNCTIONAL SILO’ MENTALITY A geotechnical investigation requires the integration of data from a number of sources. The geotechnical engineer will thus be required to interact with a number of individuals. Depending on the status of the project these individuals will range from: • Exploration/resource geologists...to define the geological model and resource footprint, • Structural geologists…to assist with the interpretation of the rock mass fabric and applying geological and spatial context around the geotechnical data, • Mine geologists…to describe the performance of the encountered geological conditions, Massmin 2004
• Hydrogeologists…to provide input into the impact of ground water on the excavation stability, • Engineering geologists…to characterise the rock mass from strength and performance requirements, • Mining engineers…to develop the mining methodology (including blasting) from the geological and geotechnical interpretations, • Mine planning engineers…to plan the mine geometry, layout and scheduling in the most optimised fashion. The authors have noted that the lack of cross-discipline integration is commonplace in the industry. The industry as a whole suffers from a 'functional silo’ mentality, whereby tasks are undertaken in an assembly line fashion, often in isolation or with an incomplete understanding of how the results impact on the entire mining project or operation. Cross-discipline communication is often the critical issue. This is evident for example, when the exploration geologists often do not communicate well with mining and resource geologists, who do not communicate well with geotechnical engineers, and so on. Often, neither task team understands the others’ requirements, nor how they can help address the issues those teams face for the ultimate benefit of the study. The way consultants are used in the Australian industry often exacerbates this problem, in that companies select people for specific tasks on a perceived value for money basis. Therefore, one consulting firm may be used for geology, another for resource estimation, another for geotechnical engineering, and another for mining engineering. This presents additional challenges in ensuring the effective flow of information and knowledge across the task teams and, therefore, the entire operation. The functional silo mentality is a substantial challenge facing the industry and, therefore, focus is required on ‘cross-training’ consultants and clients so that they are aware of issues facing each of the mining disciplines and can communicate across these disciplines. This helps to effectively break down the ‘functional silos’ to allow the more effective flow of information and understanding throughout the mining process, and thus should result in more informed decisions.
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3 TIMING
4 BENEFITS FROM INTEGRATION
Geotechnical investigations should be incorporated into an exploration or mining project at a very early stage. The specific investigation can take many forms depending on the type, scale and location of the project. It should have equal status alongside the geological, resource, metallurgical, mine planning and financial modelling studies. It is equally important to know that:
The main benefits resulting from an integrated approach can be categorised into savings in time and associated cost with a better scope for communication among the professionals involved. Incorporating geotechnical investigations as early as possible in a project study can be beneficial in guiding future work, especially into feasibility. With this approach, it is possible to reduce the dependence on long and costly feasibility studies in which the geotechnical risk profile had not been adequately defined earlier on. The estimate of capital and operating cost expenditure can be better quantified early on, by following a more rigorous and integrated approach. By getting the overall slope angle for an open pit closer to its expected value at an early stage, there can be substantial cost savings in estimates of waste stripping. For example, the difference between a 45° and a 50° overall pit slope, for a pit perimeter of approximately 3000m over a slope height of 150m, can result in an additional US$ 15M of waste stripping. In this example there is also a corresponding shift in the pit crest of approximately 25m. These values are illustrated in Figure 1 and Figure 2.
• the resource exists in a certain form and location, • that it can be mined safely and economically by open pit or underground methods, • that the ore can be processed, • that it has a market that can be reached, and • that the project will produce a viable return on investment. The main benefit of an integrated approach is that the geotechnically related risks can be understood and quantified as early as possible, especially where a mine design is sensitive to these parameters. The main geotechnical risks or issues that need to be addressed from an early stage can be related to the following: • rock mass characteristics…which can be managed with empirical methodologies, but require geological context to properly establish their spatial variability • structural fabric…an evaluation of the controls on the 3D distribution of rock damage (discontinuities), and the interaction between primary, secondary and tertiary structures and the excavation, • ground water…water pressure can significantly reduce the stability of the excavation, • in-situ stresses…and their redistribution during mining needs to quantified, • seismicity…can have a detrimental impact on the stability of the rock mass, • mine excavation…and its interaction with the geotechnical environment can be modelled to determine the development of adverse tensile and shear stresses that may lead to failure of the excavation profile. All of the project scientists and engineers play a role during the development of a deposit from initial exploration or discovery through project scoping, pre-feasibility, feasibility, detailed engineering design, start-up, production and closure. The project risk profile can be changed by informed decisions taken by engineers, financiers, stakeholders and government that are involved in the emergence of a new mining project. From the initial greenfields site, from scoping to drill out and the establishment of an advanced project, prefeasibility, feasibility and ramp-up into operations, it is beneficial to establish the links between the geotechnical characteristics and mining risks. There is a real benefit in investing time and capital during early exploration to provide initial estimates of geotechnical risk. In our experience there is the potential to reduce expenditure during feasibility if the integrated geotechnicalgeological model is understood right up front. The geotechnical engineer must identify what will be the key influences on mine design and be aware of that which may subsequently modify the design. These aspects are normally assessed in relation to the risk profile for the project and whether a competitive advantage can be obtained by optimising the design process. It is the application of the influencing and modifying components that can result in significant cost saving during the project life and which may ultimately extend the operational life of a mine. This is illustrated in Table 1 for a steep slope strategy as applied to open pit design.
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5 PLANNING ON A RISK BASIS The follow extract from a 1997 paper by Oskar Steffen on the planning of open pit mines on a risk basis states the case with regard to the value that should be attached to geotechnical information during the development of a mining project. "As in the case of mineral resource estimation, the determination of slope angles is dependant on the understanding of geological and geotechnical information and the confidence of design is equally based on the degree of certainty which applies to the data available. Unlike the case of mineral resource estimation, where the exploration is specifically targeted to provide ore reserve information, the requirement for slope design only becomes necessary once a prospective mineral resource has been discovered. Exploration core has usually limited value for slope design purposes as the target areas for slope design are not necessarily the same as for the orebody. Drilling requirements for geotechnical purposes also differ considerably from that for mineral exploration. Hence a limited campaign for geotechnical purposes is usually undertaken in addition to whatever value can be obtained from the original exploration campaign. It is therefore not uncommon to have a slope design which has a much lower degree of confidence than that pertaining to the mineral resource definition. By definition, the mineable reserves within the resource are determined by applying a mine design which could economically exploit the resource." It is again emphasized that the evaluation and definition of the geotechnical risks as early as possible in the development of a project must be understood by all. The geotechnical risks described earlier in this paper must be investigated and quantified using the most appropriate techniques. In the case of caving it is almost impossible to fully define the nature of the cave without exposing the ore material in an exploration decline. The geotechnical evaluation should advance at the same rate as that for the resource model. 6 OPERATIONAL EXAMPLES Case 1 A SLOS operation in Australia has experienced difficulties with oversize in the stope drawpoints. The causes of the
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problem are varied and the overall study has had to review: • Production issues • Stope design, sequencing and performance monitoring • Geological and structural interpretation • Geotechnical aspects • Blasting practice As part of the study, SRK were contracted to focus mainly on improving the data collection, analysis and interpretation from a structural geology and geotechnical point of view. This has involved cross discipline training of personnel with varied backgrounds and establishing processes and procedures that will outlive the current staff complement. The key components of this integrated study were related to the geotechnical evaluation of the rock mass characteristics. From the results of the geotechnical database, it was shown that the Laubscher RMR/MRMR and Mathews / Potvin Q'/N' rating systems are applicable as predictive tools for the determination of rock mass stability relative to the excavation. In addition, a structural evaluation assisted in the determination the controls on the development of oversize. The initial interpretation has established the broad spectrum of rock mass and structural characteristics that influence the occurrence and frequency of oversize. The benefits from this process of integration convert to more effective data and information, with some additional effort, which may not have been possible without this concerted approach. Case 2 In Australia and West Africa there are numerous examples of open pits that have been developed through some of the most hazardous geological materials with regard to slope stability. The saprolites and saprocks within the zone of oxidation exhibit a range of geotechnical characteristics from the dry to partially saturated to the fully saturated condition that classify them as extremely challenging to say the least. They represent the geotechnical engineers "worst nightmare". Below these are commonly found the unoxidised mafic and ultramafic rocks which commonly exhibit structural complexity and weakness associated with joint coatings of dickite, brucite and tochilonite. An understanding of the spatial relationship of these rock weakening mineral coatings with regard to the mining excavation is vital to ensure that the most appropriate design parameters are selected. The value to be derived from a comprehensive geological and structural geological assessment of a property in association with the geotechnical engineer can be significant in our experience.
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Slope design work in these materials requires a fully integrated approach to determine the relic structural fabric of the saprolites, and to a lesser extent in the saprocks. This rock mass fabric is normally identifiable in the unweathered rocks. A typical weathering profile is illustrated below. • Transported • Pallid Clay • Saprolite • Saprock • Joint Oxidised Rock • Unweathered Rock Designs in this range of materials requires that all of the geological and geotechnical characteristics must be evaluated as early as possible as the open pit design parameters must be derived for both soil, soft rock and hard rock environments. The use of empirical rock mass classification systems can provide very good indicative slope angles from the first series of exploration cores. These can be supplemented when exploration declines or shafts are excavated. SRK have previously provided training in geotechnical core logging to exploration geologists and mine geologists at a similar site in Western Australia. In practice, only the orezone and a limited amount of core into the hangingwall and footwall are logged. This data would otherwise be lost, as the core is subsequently cut and sampled. Recent developments have seen both scoping and feasibility level studies required for new prospects on the same mining lease. In the first instance, no additional drilling or logging was required. The study could be completed in some detail, within a short time and at minimal cost. Similarly, the feasibility study required very little additional specific geotechnical drilling and logging. The savings here were both in direct cost and, most critically in this instance, in time. ACKNOWLEDGEMENTS SRK would like to acknowledge the assistance provided by its staff and clients in the compilation of this paper. REFERENCES • Haines, A and Terbrugge, PJ, 1991. Preliminary estimation of rock slope stability using rock mass classification systems, 7th Int. Congr. on Rock Mech, ISRM, Aachen, Germany. • Laubscher, DH, 1990. A geomechanics classification system for the rating of rock mass in mine design, Jl S Afr Inst Min and Metall, Vol 90, No 10, pp 257-273. • Steffen, OKH, 1997. Planning of Open Pit Mines on a Risk Basis, J SAIMM March/April 1997.
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Key Component
Can have significant benefits from slope depressurisation, unloading of potential failure surfaces, dry working conditions and improved blasting.
Improves slope geometry, reduces spill berm width, and reduces secondary cleaning and accidents from rock falls during bench cleaning following blasting.
For local bench crest stabilisation, ramp protection and for steepening of desirable slope sectors.
For cutting back upper weaker material to allow steepening of lower slopes in more competent rock.
Typ
Ground Water Drainage
Pit Limit Blasting
Anchoring of Bench Crests
Unloading of Upper Slope
Influencers
Overall slope angle of upper slope will be reduced to produce the required unloading
Modification of the slope geometry can benefit the influence of anchoring at bench crests.
The optimisation of bench height and face angle on the pit limits can be determined from geotechnical considerations. Can also modify the influence of postor pre- split blasting
Adjustments to slope geometry and mining sequence may be necessary to accommodate drainage requirements
Adjustments to Slope Geometry
The relative competence of the upper and lower slope material will determine the benefit of unloading
Effective for toppling failures below ramps and crucial surface infrastructure
Modifies the selection of bench face angles, local geometry, stack and overall angles. Can be a predictive tool for deeper pit sectors.
Density and orientation of structures will determine the permeability and modify the shape and gradient of dewatering surface
Structure andGDM
Implication of failure of upper slope is not crucial. Can prevent deterioration of weak upper slope and protects surface infrastructure.
Failure of anchors are not crucial following backfilling.
Special pit limit blasting techniques may be minimised should early backfilling be an option.
Permits steeper hydraulic gradients to be developed behind back filled faces
Backfilling during Mining
Modifiers
Crucial to ensure that unloading has been effective for the stability of both the upper and lower slopes.
Effectiveness of anchoring can be monitored from dis placement records
Determines the nature of displacement (acceleration/ deceleration) resulting from the improved blasting practice.
Determines the nature of displacement (acceleration/ deceleration) resulting from the influence of drainage.
Slope Displacement Monitoring
Table 1: Steep Slope Strategy Matrix: Modifiers and Influencers on Slope Geometry
To ensure that the measured groundwater profile does not exceed the design conditions.
Effective dewatering will reduce the design load on anchor tendons.
Relative benefit of dry versus wet faces will influence blast design and final face angle will determine stack angles.
Determines the nature of the hydraulic gradient and the effectiveness of drainage
Piezometric Profile Monitoring
May obtain an improvement of between 3º and 5º in the bench stack slope angle for the lower competent material. The benefit will depend on the proportion between upper and lower slope heights.
May obtain an improvement of between 3º and 5º in bench face angle, depending on orientation of most critical structure.
May obtain an improvement of between 3º and 5º in bench stack slope angle, depending on achievable bench height.
May obtain an improvement of between 5º and 10º in overall slope angle, depending on type of material and slope height.
Assessed Benefit
Figure 1: Cost saving in waste stripping (per m of pit perimeter) with slope steepening
Figure 2: Shift in pit crest with slope steepening Massmin 2004
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Estimation of geotechnical variables for mass mining Mark Howson, Principal Consultant Geologist, Rio Tinto Technical Services, Bristol, UK
Abstract In a major underground mining project, once mining starts, change can be very expensive if rock mass behaviour is found to differ from what was expected. However, issues such as ground support, caveability and fragmentation are typically major uncertainties at the planning stage. All these issues depend to a large extent on rock mass properties whose correct estimation is worth millions of dollars in either savings or avoidance of over-expenditure. A significant rock volume is rarely homogenous. Its strength will be subject to zones of weakness in otherwise stronger rock, and sampling through geotechnical logging of drill cores yields a wide scatter of values. Distributions must be estimated and used to assess rock mass behaviour, as averaged values can be misleading. A methodology is presented that is being used for several potential underground mining projects around the world. It is based on non-parametric geostatistics applied to Geotechnical Variables (GVs) such as Fracture Frequency, RMR and Q. The influences of mining, data sparseness, sources of bias, lithology, geological structure, and other spatial attributes are incorporated in the estimation of a geotechnical model. This may then be compared with reality for validation, and used in mine planning.
1 INTRODUCTION Geotechnical Variables (GVs) include Fracture Frequency, Joint Condition Rating, RQD, IRMR and Q. They may be based on drill hole logging and are used to quantify rock mass strength in various ways, so that mining attributes, including ground support, caveability and fragmentation, can be predicted. This paper is not concerned with the relative merits of different GVs, or their derivation, but with methods for their estimation as a block model. The objective is to use data collected largely from drill holes to represent the disposition of the GV in the whole mass of rock in a detailed and unbiased manner. The approach presented here is being used to investigate the underground mining potential of some major deposits in the Rio Tinto group, including two Porphyry Coppers and two diamond pipes. Geostatistical estimation and modelling of in-situ grades is a familiar process in mining. A related approach can be applied to GVs but may only be successful if their special characteristics are fully accommodated. This paper refers to methods used for grades, using geostatistical terms that are explained in standard texts such as the work by Isaaks and Srivastava (1989).
2 NON-ADDITIVE AND NON-PARAMETRIC A significant rock volume is rarely homogenous. Its strength will probably be subject to zones of weakness in otherwise stronger rock. Sampling it by geotechnical logging of drill cores yields a scattered distribution of GV values, with high and low values occurring in close proximity. How can this be represented in a model? GVs are best considered as non-additive, a quality that is illustrated by the hypothetical case of two drill holes near to a planned excavation, as sketched in figure 1. The drill holes have RMR values as shown, and between them, a planned excavation is equidistant from each hole. Assuming that development in rock of RMR less than 40 will need additional ground support, the task is to estimate the RMR at the excavation, and specify if additional support will be required. 44
Figure 1: Drill holes and a planned excavation
Using linear interpolation, one might estimate an RMR of 45 at the excavation, therefore, no additional ground support is required – but this would be incorrect! The correct estimate is that 50% of the rock is 60, and 50% is 30. Therefore additional support is expected to be required for 50% of the excavation. When estimating GVs, averages can be misleading. We must estimate the distributions of these non-additive variables. GVs are also generally "non-parametric". A parametric variable – such as some grades – can be assumed to have a characteristic distribution such as a Gaussian curve. This assumption can generally not be made for GV distributions, which may be multi-modal. 3 THE MULTIPLE-INDICATOR APPROACH The non-additive and non-parametric character suggests "Multiple Indicator Kriging" (MIK) as a geostatistical interpolation method. This is described by Barnes (1999) and is often used for estimating disseminated gold grades. However, significant adaptations must be made for GVs. In MIK, the distribution of a variable is estimated as a series of discrete "Indicators". These are proportions (between 0.0 and 1.0) of each block that are above a series of "threshold" values of the variable. For example, given some RMR thresholds selected as 30, 40, 50, 60 and 70, the indicators for a hypothetical model block were estimated as 0.95, 0.7, 0.6, 0.4 and 0.02, respectively. These describe a bi-modal distribution, whose
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median is at about RMR 55, and where 30% of the block is expected to be below RMR 40. In practice, perhaps 7 to 17 indicators are selected through the range of GV values, at regular intervals, or sometimes bunched-up at parts of the range where more detail is required. They are estimated as separate variables in their own right, and then used together to describe the distribution in each block and to determine median, other quantiles and other information.
A particular example of data bias arises where an underground mine is planned to extract the vertical continuation of an orebody below an open pit. Here, GV values near the collars of holes that were drilled in the pit must be tested for systematic reduction due to blast damage or stress release. It is likely that a zone of rock adjacent to the final open pit topography must be modelled as a separate domain (see below) to accommodate this effect.
4 INDICATORS THEN REGULARISATION
6 DOMAIN DEFINITION
Typically, geotechnical drillhole logging is in "domain intervals" of similar geotechnical values, but often of unequal lengths. Geostatistical estimation assumes that each of the input data has equal sampling support. For GVs, this means derived from equal lengths of core. Therefore it is necessary to determine equal-length data from the unequal intervals. It is normal in grade estimation to determine mean values for equal drill hole lengths in a process known as regularisation or compositing. However, because GVs are non-additive, a procedure involving mean values is inappropriate. However, indicators are similar to grades and are additive variables. Therefore they are determined first for the logging intervals. They are assigned as values of 0.0 or 1.0 depending on whether the GV value is below or above the threshold in question. These are composited into regular lengths, which may typically be a sub-multiple of a model block dimension. Each composite has a series of values from 0.0 through 1.0, which represents its distribution of the GV. Figure 2 illustrates the calculation of composited indicators from log intervals, using hypothetical Joint Spacing Rating (JSR) values as an example.
The term "domains", used in a geostatistical sense describes sub-volumes of the rock mass where samples (i.e. logging) of a domain may be used to estimate model values elsewhere in the domain. It is not usual to use data from one domain to estimate values in another. An early stage in estimation is to define the domains, and with an understanding of their geological or other origin, to characterise the behaviour of the GVs within them. Typically the basis for separation into domains includes differing lithology, faulting, shearing, folding, alteration, weathering, and, as discussed above, mining blast damage and stress release. Figure 3 shows an example of domain definition relating to faults in a project that was conducted in feet.
Figure 3: Variation of GVs adjacent to faults.
Figure 2: From logs to composited indicators The JSR log on the left has variable intervals up to 5 m in length. The indicator thresholds 15, 20 and 25 have been selected by cumulative frequency analysis of the JSR populations. The resulting indicator values, still in the variable intervals are in the middle, and the composites that were calculated in regular lengths of 5 m are on the right. 5 OPEN PIT MINING INFLUENCE Confirmation of the quality of logging and data derived from drill holes, and the assessment of any bias is important in an estimation project. Massmin 2004
In this case, the faults had been interpreted as hangingwall and footwall surfaces to define "fault envelopes". In the graph, values within the envelopes are shown at Distance = 0. The other distances show that the influence of faults extends for about 3 m (10 ft) into adjacent rocks. In this project, the domain of "faulted rock" was defined as including this extra 3 m on either side. 7 DOMAIN CHARACTERISATION GVs often exhibit strong trends within a domain. For example, rock strength in a kimberlite pipe often increases gradually but steadily with depth, due to the decreasing effect of surface weathering. It is particularly valuable to characterise the general spatial behaviour of a GV in the context of the geology, as this information will be used where data is sparse during the Kriging process, as discussed below.
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Indicators provide a useful tool for this purpose. Figure 4 shows an example where core-length-weighted mean indicators for RQD (named d1 to d8) characterise the change in RQD at a series of levels perpendicular to bedding through a sedimentary sequence.
Figure 4: Characterisation of change in GV in a domain using indicators.
8 CROSS-VALIDATION The application of geostatistical analysis to GV indicators is similar to the same process in grade estimation. It characterises spatial continuity of the indicators to provide variogram parameters that control the Kriging process for the most accurate estimation. Cross-validation is a technique that may used to assess the effectiveness of the process. However, it is not a fully accepted procedure. Cross-validation is applied by omitting a datum (e.g. a composite) from the input dataset, and then estimating its value from the other data by point Kriging. This is repeated for all the data in the set, resulting in a list of point estimates that may be compared with the corresponding true values. The closer the estimates, the better the parameters are considered to be. The best parameters are then applied to block model estimation. A problem with cross-validation as typically applied is that often the data that contribute most to each estimate are those from the same hole - typically vertically above and below, following the hole orientation, which produces a biased test. The author’s opinion is that the technique is useful if all data from the same hole are omitted from the dataset when one of its data is estimated. The effect is similar to comparing a previously estimated model with logging from new holes, as described below under Model Validation. In an estimation project it may be more appropriate to use limited time resources for better domain studies rather than cross-validation. However, it provides a useful tool for comparing estimation techniques as described below. 9 DATA SPARSENESS AND KRIGING A problem often encountered is that the GV data from drill holes may be relatively sparse compared with grade data in the same deposit. For example, many holes may have been drilled during the exploration phase, when geotechnical logging was not done, and it is impossible to re-log them. Furthermore, it may be possible to justify only limited further drilling for geotechnical purposes. Therefore, it is necessary to make the most of what data are available. 46
The following strategy utilises the domain characterisation discussed above, with a variation of the usual MIK interpolation method. When estimating grades using MIK, the variant of Kriging used (within MIK) is usually Ordinary Kriging (OK). Typically, to estimate the value of a model block, the Kriging process will include the influence of values from all data that lie within a search radius of the block. If too few values are found within the radius, then the block is not assigned a Kriged value, but a pre-determined mean value for the domain. For GVs using MIK, this would be a local mean indicator value derived from the domain characterisation. A problem with OK is that an unnatural "halo effect" is often produced with GVs. This occurs where data are sparse at the boundary where blocks estimated by Kriging are adjacent to blocks that were assigned mean values. At these locations, confidence in the Kriged block values is low. The strategy proposed to deal with this issue is to use the Simple Kriging (SK) variant within MIK for GV indicators. Unlike OK, SK always includes the influence of a predetermined local mean in its result. Where there is a high level of data in the vicinity, this influence is low, but it increases as the data become sparse, until the point at which too few values are found within the radius. Then, as in OK the block is assigned a pre-determined mean value for the domain. However, with no abrupt change in the influence from the local mean, the halo effect is avoided. Because GVs often exhibit strong trends within a domain, local means derived by well-conducted domain characterisation may themselves be close to reasonable block value estimates. Therefore, the level of confidence in a SK result that is based largely on a local domain mean may be justifiably high, even for an area with sparse data. Considering SK as an alternative to the more usual OK is suggested in the context of MIK with GVs. Graham (2003) used cross-validation to compare OK and SK estimation of indicators using a set of RQD data from a Porphyry orebody. The results support the methodology described in this paper. 10 ESTIMATION OF COMPOUND GVs Certain GVs such as the In-situ Rock Mass Rating (IRMR) described by Laubscher and Jakubec (2001) are combinations of several component values. IRMR is the sum of Ratings for Block Strength (BSR - derived from Intact Rock Strength, IRS), Joint Spacing (JSR) and Joint Condition (JCR). Values of IRMR can be determined for each logging interval, and then used to estimate model IRMR values by the straightforward MIK approach described above. However, the component ratings may have distinct distribution patterns of their own, and in places, they may behave differently with respect to one another. Variation in each component may be more regular and predictable than their sum of the three components. An alternative approach may be to separately estimate a model for each rating using MIK, as above, and then to combine the three models to produce an IRMR model. This may produce a more accurate result and has an additional advantage of making the three ratings models also available. The combination step involves combining the indicator distributions for the three ratings in each block to produce a single IRMR distribution. How best to do this is still the subject of investigation, and a full explanation is beyond the scope of this paper. However, a provisional method is being used, which works best if it is assumed that the three ratings are not correlated. Offen (2003) tested this approach to IRMR estimation, and the provisional method for combination of distributions.
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The cross-validation technique was used with data from a Porphyry Copper monzonite unit. Both the direct estimation of IRMR indicators and the estimation of ratings and then combination to produce IRMR indicators were compared with true indicator values. The results confirmed that the method described here might be better. In a recent project, this method had major advantages in dealing with sparse data. While most drillhole intervals had information for JSR, and adequate numbers had BSR and JCR, there were relatively few with all three. However, modelling each rating separately, and then combining the block models maximised the utilisation of what data were available in each case. 11 MODEL FINALISATION When the Kriging process is complete, it will yield one or more series of block model data files, or columns of data in a database. Each of these contains the values of one indicator, for all the blocks in the model. Each value is the estimate of the proportion of a block that is over the respective threshold value. A series of indicators constitutes the model of a GV, giving the distributions of that GV for each block. The block values may be utilised in their current form, with no further processing. For example, if it is assumed that an excavation in rock of RMR less than 40 will need additional support, then one would use drawings showing the RMR 40 indicator for development planning. A planned mine access may be routed through areas with high RMR 40 values – i.e. with RMR mainly above 40 - to minimise the support costs and/or the probability of ground failure. When the design is complete, the blocks intersecting the access can be summarised to determine what overall proportion is below RMR 40, and so to estimate how much of it will require additional support. Another method of displaying or utilising the models is to determine the median GV for each block. A database query may be required to select, for each block, those indicator values that are just above and just below 0.5. The GV value that represents exactly 0.5 can then be found through linear interpolation between the two. Other quantiles can be found in a similar manner. 12 MODEL VALIDATION It is desirable to compare the model with reality, to determine the accuracy of what has been estimated. Figure 5 shows an example where an RMR log from a newly drilled hole is compared with a previously estimated model. The figure shows a plot of RMR against down-hole depth. The erratic ("noisy") thick grey line is the RMR log of the new hole. This is being compared with the previously estimated RMR distribution values from those blocks in the model through which the trajectory of the hole passes. The five black lines with a more vertical nature are quantiles representing the block distributions. Left to right, these are for the 5, 25, 50, 75 and 95 percentiles respectively. In this example, the grey line should be outside the band defined by the 5 and 95 percentiles for 10% of the plot. As the accuracy and detail of the estimation is improved, one would expect this band to become narrower. It would follow the log more closely, but still with 10% outside. The hole in this example was drilled in a basement complex of granites, gneisses and schists. It is in a barren area of the deposit, and so drilling data are relatively sparse. However, there are two earlier holes with geotechnical logging that pass by several 10s of metres away. Massmin 2004
Figure 5: Comparison of model with logged data. It can be seen that a model that could predict the new hole’s log exactly would require drilling data at an impossible level of detail. However, with much sparser data, estimating distributions using the indicators can represent the ground conditions in an adequate manner for planning. Other methods of model validation may compare estimates with mining performance. One example is where an underground mine is planned to extract the vertical continuation of an orebody below an open pit. The pit slope angles as they were mined can be compared with estimated model values at the same locations, to assess the effectiveness of the model. Another example is where some underground development has already been mined. The model values can be compared with the ground conditions that were encountered. It should be possible to derive some empirical relationships that will enable mining costs to be predicted from the estimates. 13 CONCLUSIONS Estimation and modelling of geotechnical variables for predicting mining conditions is necessary in planning an underground mass mining project. It is important to estimate distributions of GVs rather than mean values, which can be misleading. In geostatistical terms, GVs are best treated as non-additive and nonparametric variables. These qualities suggest using a "Multiple Indicator" approach to estimation. Indicators provide a means of expressing the spatial distribution of a GV. They are applicable to logged data, and then through estimation, to mine development, and major excavations. Data quality and bias are important considerations in an estimation project. This includes bias introduced by mining that may not be evident in holes that were drilled earlier. The definition of geostatistical domains on the basis of lithology, faulting, shearing, folding, alteration, weathering and mining effects is an essential part of the estimation process.
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Domains may show weak or strong trends in GV variation, which must be characterised. Graphs showing indicator behaviour are recommended. The information obtained will be used to counter the problem of sparse data. Calculation of indicators from logged data, followed by regularisation or compositing is recommended prior to geostatistical processing. The geostatistical analysis stage (variograms etc) is much the same as in grade estimation and so is not described here. However, a variant of the MIK interpolation method using is proposed as a strategy for making the most of sparse data. This relies on the domain characterisation for effectiveness. Estimation of GVs such as IRMR that are combinations of several component ratings may be best achieved by estimating and then combining their components. Comparison of a GV model with reality enables its accuracy, and correlation with mining performance to be investigated. On-going investigation of techniques for estimation and utilisation of GVs to predict rock mass behaviour are vital in Rio Tinto’s strategy to develop safe and profitable worldclass underground mines.
15 REFERENCES • Barnes, T, 1999 "Practical Post-Processing of Indicator Distributions" Apcom’99, 28th International Symposium, October 20-22, ISBN 0-918062-12-8, pp 227-237. • Graham, J M, 2003. "The application of geostatistical procedures on the geotechnical variable, RQD, on data from Bingham Canyon Mine, Utah, USA." MSc dissertation, Camborne School of Mines, University of Exeter. • Isaaks, E H, & Srivastava, R M, 1989. "An Inroduction to Applied Geostatistics", Oxford University Press, ISBN 019-505012-6 or 0-19-505013-4 (pbk.) • Laubscher, D H, & Jakubec, J, 2001. "The IRMR/MRMR rock mass classification system for jointed rock masses", Underground Mining Methods: Engineering Fundamentals and International Case Studies, (Eds: W A Hustrulid and R L Bullock), 475-481. Soc. for Mining, Metallurgy and Exploration.: Littleton, Colorado. • Offen, J. (2003). "Estimation of the geotechnical variables of RMR using geostatistics, at Bingham Mine, Utah." MSc dissertation, Camborne School of Mines, University of Exeter.
14 ACKNOWLEDGEMENTS Thanks are due to staff and management of Rio Tinto Technical Services, and at business units in Rio Tinto for their support in this work, and for permission to publish this paper.
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DEESA - An approach to determine if an orebody will cave Richard Butcher, Associate, Dempers and Seymour Pty Ltd
Abstract Block caving or sub level caving rely on the ability of the rock mass to cave for production. It therefore follows that cavability assessments should be a key geotechnical aspect for all caving projects. In most cases, it is not only necessary for the geotechnical engineer to determine if an ore body will cave but at what mined dimension will caving occur. At the present time no single geotechnical approach exists that can totally simulate the complex interaction of the rock mass and the mining stresses that occur during caving. It is therefore evident that cavability assessments should be broken into a series of logical steps, using a number of tools to simulate cavability adequately. This paper critically discusses the application of existing cavability determination methods and proposes a logical step based approach to cavability determination using the most reliable tools available today.
1 INTRODUCTION In recent years, there has been renewed interest in the use of cave mining methods to extract massive low grade ore bodies. The main reason for this renewed appeal is that only cave mining methods can compete with open pit operations in terms of tonnages produced and mining costs. Since caving methods fundamentally rely on the ability of the ore body or country rock to cave in a controlled manner for ore production, cavability assessments are a prime geotechnical focus. However, despite much work having been undertaken in this field at the present time, no single accepted geotechnical approach exists to determine if and when an ore body will cave. This lack of a common approach may be attributed to the fact that a number of different approaches exist to determine if an ore body will cave (e.g. Laubscher 1994, Matab &Dixon 1976, Butcher 2002 and Brown 2003). Laubscher (1994) states that "All rock masses will cave. The manner of their caving and resultant fragmentation size distribution needs to be predicted if cave mining is to be implemented successfully". Therefore the emphasis of cavability determination changes from "if to "when and how" caving will occur. It is not only important for the geotechnical engineer to determine if an ore body will cave but at what mined dimension will caving occur. In the case of a block cave these questions are critical in determining the economic viability of the project. A six month delay in achieving the undercut dimension for sustained caving production will have serious consequences in terms of a project being able to meet required financial returns. This paper critically discusses the application of existing cavability determination methods and proposes a logical step based approach to cavability determination using the most reliable tools available today. 2 EXISTING CAVABILITY DETERMINATION METHODS
methods and research before a cavability determination approach can be formulated. Experience plays a major factor in cavability assessment internationally. In this respect the following are considered as common experiential methods of denoting the cavable size of prebreak or undercut: • Length x width of undercut (Bartlet, 1997). • Area undercut/ pre-broken or mined (Jofre et al, 2000). • Defined cavable hydraulic radii range (Owen & Guest 1994, and Butcher, 2003). • Minimum span. The important point to note about experiential cavability methods is that they are normally employed on mines with a long history of caving under known geotechnical conditions (Jofre et al, 2000). The most widely used methods of predicting the area that has to be mined to induce caving are empirical correlations between the rock quality and the undercut area. The most commonly used method is Laubscher’s stability graph (See Figure 1). Laubscher modified Bieniawski’s RMR classification system to suit caving operations for the chrysotile asbestos mines in Zimbabwe after the need to classify the ore and country rock for geotechnical purposes was identified (Heslop, 1973). Laubscher also recognized that the shape of the mined area (i.e. hydraulic radius) was an important factor for cavability determination (Laubscher, 1975). The first stability graph was published in 1987 (Diering & Laubscher, 1987). Since that time a number of variations of this graph have been published with the most recent in 1994. Over the years attempts have been made to use, integrate or modify Mathew’s stoping graphs for cavability purposes (Stewart & Fosyth, 1995, Trueman & Mawdesley, 2003). In general, these methods have not been accepted by the mining industry. In summary, empirical methods using a correlation between the quality of the rock mass and the shape of the mined area (or hydraulic radius) are frequently employed as a method to determine cavability.
Much work has been conducted into the field of cavability and it is important to review some of the most common
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Figure 1: Laubscher’s stability graph (1994)
Bartlett (1992, 1997) shows that for an ore body to cave continuously, a sufficiently large area must be undercut. The required dimensions of this area are a function of the rock mass strength (jointing, and joint shear strength characteristics) and the regional stress field that prevails before and as mining progresses. Taking this into consideration, a detailed finite difference numerical modelling exercise was commissioned for Premier Mine to predict the extent to which an undercut had to be advanced (Bartlett 1997). This exercise clearly showed that caving in all rock types was joint controlled. If accurate joint and rock mass parameters could be determined, numerical models could be used to predict the hydraulic radius to initiate caving. Bartlett’s work demonstrates that the practical application of numerical modelling as a tool for cavability determination relies on the quantum and representativeness of collected geotechnical data. This work suggests that a detailed collection of rock mass data and knowledge of the mining stress regime is fundamental for cavability determination, if numerical methods are used. Taylor (1980) detailed important geotechnical and rock mass structural aspects to be considered in the determination of cavability: • The number of joints, attitude and spacing of joints. A combination of flat and vertically spaced joints is conducive to caving. • The shear strength properties of the mentioned joints. • The magnitude and orientation of principle stresses. • The importance of rock strength increases as the number of joints decreases. The important conclusion from Taylor’s work is that the rock mass structure and it’s orientation to the mining stress regime are important considerations in terms of cavability. The importance of the variability rock mass structure and mining stress orientation in terms of cavability assessment was researched by Butcher (2002). Ehancing the concepts forwarded by Taylor (1980) and Matab & Dixon (1976) the ChasSM (Complex structural modelling) method was used to generate joint patterns from an ore body that caved. This method allowed the percentage of unstable cave back blocks to initiate and propagate caving to be determined. 50
The important point from this work is that a cavability rationale was established by simulating the variability of structural data in relation to the orientation of the stress regime. Guest and Cundell (1994) describe the use of a 3D distinct element programm (3DEC) to simulate Vertical crater retreat (VCR) caving behavior. The results of the modelling were difficult to interpret and were only fully understood after mining had been completed. In recent years, work has been conducted on numerical modelling caving behavior using the 3D Particle Flow Code ((PFC), Itasca, 1998). Using PFC, the ore body is modelled as a large number of rigid spherical particles whose behaviour is controlled by a constitutive relationship that defines the tensile and shear behavior of the particle contacts. At the present time the use of the code is in the experimental stages. Brown (2003) shows that codes like FLAC 3D are used on a regular basis for cavability determination. This is despite the normal disadvantages of using such programmes for modelling (Butcher 2004, Trueman & Mawdesley, 2003). It is the Author’s experience that numerical approaches are not normally used in isolation from other methods for caving assessment. In summary, the discussion reveals that the following methods are commonly used for cavability assessment: • Experience. • Empirical stability graphs. • Structural analysis. • Numerical analysis. • A combination of abovementioned techniques. The use of a combination of techniques is an important aspect to note since each individual method will have inherent shortcomings. From the Author’s experience, the following techniques are normally used together: • Experience. • Laubscher’s stability graph method. • FLAC 3D modelling. 3 DEESA – CAVABILITY ASSESSMENT APPROACH At the present time, a number of techniques (either individually or a combination) are used to determine cavability. The main reason for this is that there is no single geotechnical approach that can totally simulate the complex interaction of the rock mass and the mining stresses that occurs during caving. In addition, engineers may have limited confidence or experience in some of the methods listed. The critical point is to recognise that the various caving assessment methods have their own applications, advantages and disadvantages. DEESA is a method that divides cavability assessment into a series of logical steps. The results of each step should correlate with the previous steps and confirm the final mined dimension for caving after the process has been completed. The steps in the DEESA approach are: D: E: E: S: A:
Do you have a project? Experiential cavability assessment Empirical, cavability assessment Structural cavability assessment Analytical/numerical cavability assessment
Step 1: Do you have a project? Since cavability assessment is a complex and time consuming process, it is important that the project appears robust in terms of economic viability before detailed cavability work commences. The costs to undertake a cavability assessment are considerable. This situation is exacerbated
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further in cases where very competent ore bodies are being considered as potential targets for caving. In such cases, where the Mining Rock Mass Rating (MRMR) exceeds 60, it may be necessary to sink an exploration shaft or decline early in the project to gather rock mass structural data. Should this type of exploration programme be necessary, it could add significantly to the study costs. It is therefore prudent to ensure that a project appears viable before detailed geotechnical work commences. This can be done at desktop study level by assuming that the ore body will cave and by determining prebreak and undercut dimensions on a production basis only. In situations where an exploration shaft or decline has to be sunk (early in the study) to gain geotechnical information, it is important that the following aspects have been determined before excavation commences: • That the project concepts have been finalised. • That the project risks/ potential risks have been determined. • The project appears viable with strong DCF (Discounted cash flow)indicators when related to corporate hurdles. • That the resource has been geologically defined to an indicated category. • That all existing geotechnical data has been analysed, complied and the conceptual geotechnical model has been formulated. • The potential impacts of mining stresses are understood. • That the aims of the exploration programme are defined. • A geotechnical investigation and risk study have been undertaken for the exploration shaft / decline. From the Author’s experience it is becoming common to consider geotechnical exploration programmes without the above being achieved. Step 2: Experiential assessment An industrial bench marking exercise should be conducted to determine cavability characteristics of the type of deposits being caved. The experiential information is used as a "reasonableness" check for the various cavability assessment techniques’ outputs. A literature review of cavability should also be undertaken to determine what issues are of importance in terms of the type of ore body being considered and accuracies of previous studies. An example of this is the importance of joint trace lengths in cavability assessment for ore bodies where MRMR > 60. Normal industrial bench marking exercises focus on the collection of the following types of information: • Geology/ rock mass characteristics. • Mining experience during caving. • Block height. • Areas mined to induce caving, hydraulic radius and minimum span. In experiential terms ore bodies can be divided into two groups (Butcher, 2003) for caving purposes: • Hard deposits: caving is achieved with mined hydraulic radii of between 35 to 40m. • Soft deposits: caving is achieved with mined hydraulic radii of 17 to 25m. Hard deposits are considered classification values greater than 45.
to
have
MRMR
In situations where there is no previous caving experience relating to the ore body being investigated an experiential asessment has obvious limitations. The author suggests that caving experience may be limited to situations where MRMR 0.2 m Length of pieces > 0.4 m Length of pieces < 0.01 m Length of Broken Zone
Fragmentation is a major factor in an assessment of the feasibility of cave mining in large, competent orebodies (Calder et al, 2000, Carter and Russell, 2000, Chen, 1996, Nickson et al, 2000). Since the large mines depend heavily on large mechanized drilling and loading equipment, the assessment of the economic viability of caving in competent orebodies is determined by LHD productivity and the cost of breaking large fragments to a size that can be efficiently handled by the available equipment. The distribution of fragment sizes helps assess the requirement and design of material handling systems in the mine. This information can also be used for evaluating the production capability of a deposit designed for extraction using the block caving method. Fragmentation has a bearing on drawpoint spacing, dilution entry into the draw column, draw control, drawpoint productivity, secondary blasting/breaking costs and secondary blasting damage (Laubscher, 1994). In addition, the estimation of hang-ups helps in production planning and scheduling as well as equipment selection for the mitigation of hang-ups. 3 METHODS FOR ESTIMATING FRAGMENTATION
Table 1: Drill-hole piece length data collected at several projects. The authors have developed an approach for the assessment of fragmentation using the data collected from exploration and geotechnical drilling programs. This approach utilizes the drill core piece length information along with available or estimated joint set characteristics. The procedure Massmin 2004
2 IMPORTANCE OF FRAGMENTATION ESTIMATION IN BLOCK CAVE MINES
Various techniques have been developed for estimating the in-situ fragmentation and several methods for estimating the breakdown of the material within the column have been proposed. However, field-testing and calibration of these techniques is important before they can be integrated into the planning process. The development of tools for the estimation of expected fragmentation will significantly improve the process of planning and scheduling of block cave operations. Several authors have tried to address the problem of estimating block sizes in different stages of a block cave operation, and there are several schools of thought on the nature of degradation of material as it travels down a block cave column. McCormick (1968) has suggested that fragments break down within 100 meters of draw. Hustrulid (2000) has proposed that the breakdown of material should
Santiago Chile, 22-25 August 2004
55
be linked to Bond’s third theory of comminution, with gravity providing the energy to break the rock to smaller sizes within the draw column. The primary mechanism of breakdown of material is the repeated slitting of the blocks within the draw column based on the concept that blocks with high aspect ratios will split to blocks of more stable aspect ratios. Methods of estimating fragment size distributions in block cave mines have been based on joint set parameters estimated from structural mapping in available excavations or outcrops. While this is acceptable in the absence of any other means of assessing the fragmentation, the results can often be misleading since the structural mapping is often carried out in limited areas and the results are applied uniformly to the entire deposit. The Block Cave Fragmentation (BCF) program (Esterhuizen, 1999) is probably the most widely used method for estimating the fragment size distributions in block cave operations. The program uses the statistical joint set information to create primary fragments. The generation of fractures due to stresses at the cave back is also included in the primary fragmentation estimates. Secondary breakage is achieved through aspect ratio reduction, attrition and compressive strength failure of the primary blocks in the draw column. Figure 1 shows typical fragmentation curves estimated using the BCF program.
is that a certain range of core piece lengths from the drill hole is representative of the true 3-D distribution of fragments. In general, this range is from 20 to 80 centimeters. The distribution of the core piece lengths is plotted on a log-log plot and the slope of the curve in the 20 to 80 centimeter range is calculated to be the fractal dimension. The derivative of the power function is multiplied by the volume of a sphere and integrated between zero and various fragment lengths up to the longest fragment length to develop a 3-D distribution of the fragment sizes. A typical fragmentation curve based on the fractal analysis is shown in Figure 2.
Figure 2: Typical results from single parameter fractal fragmentation model. The use of single parameter fractal models for estimating fragmentation is extremely sensitive to the longest piece chosen and does not result in the standard "S" shaped fragment size distribution. The two-parameter Weibull functions give the "S" shaped curve, but is difficult to use for the fragmentation without a subjective percentile cut-off since the function is an infinite distribution.
Figure 1: Typical Results from the BCF program Programs such as BCF rely on joint set characteristics observed at the base of the deposit. This information can be misleading since the jointing characteristics may be different as the cave progresses upwards. The prediction of fragment size distributions at different times during the life of the mine helps in taking decisions regarding the size and type of mining equipment planned for use in the ore handling systems in the mine. Fractal concepts have been used for the development of fragmentation estimates from drill core data. Poulton et al (1990), Hardy et al, (1997), and Mojtabai et al, (1988), have reported that the cumulative number of pieces in a fractured rock mass plotted against the equivalent spherical dimension of the fragments exhibits a fractal relationship in the power form: y = x-b where y = number of fragments with a spherical diameter less than x, x = spherical diameter of the fragment, and b = fractal dimension. The main assumption in the use of the fractal technique for the estimation of fragment size distributions from drill hole data 56
Figure 3: Typical results from two-parameter Weibull fragmentation model. 4 THE CORE2FRAG PROGRAM The lengths of core pieces recovered from drilling in a joint rock mass represent some measure of the in-situ block sizes. Estimation of the block volumes from these core piece lengths using a cubic or spherical block shape assumption can underestimate the block sizes. In the case of a regularly jointed rock mass, the shape of the rock blocks is related to the distribution of joint set spacings. The Core2Frag program was developed for converting the drill hole core lengths to block volumes given the length of the drill hole core pieces, the orientation of the drill holes with respect to major joint sets (the orientation of the joints is estimated from cell mapping information) and some simplifying assumptions regarding the shape of the blocks and the location of the core piece within the block.
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The primary assumption in the development of primary fragment size distributions from drill core data is that each drill-hole core piece represents one in-situ rock block. The relationships between the joint spacing and lengths of the different joint sets, evaluated from the joint set characteristics gathered from available excavations, outcrops or oriented core drilling programs, can be utilized for estimating the shape of the rock blocks.
are calculated from the core lengths based on the side of the block intersected by the core piece as shown in Figure 5. The block volumes are calculated for the each piece in a drill interval and a volumetric block distribution is generated for all the drill intervals (Pratt et al, 2002).
The other assumptions used in the development of the Core2Frag program are: a) All blocks have the same aspect ratio and the block shape is defined by the joint sets b) The drill hole passes through one apex of the primary block The Core2Frag program requires the following inputs: a) Drill-hole piece length data, either in raw form or else sorted into bins of fixed sizes. b) Average dip and dip direction for up to six joint sets c) Aspect ratios calculated from the spacing of the joint sets. d) A length correction factor for drill-hole not passing through an apex of the block (Figure 4).
Figure 5: Drill hole intersecting orthogonal block The formulation given above is for orthogonal blocks and is shown for simplicity. The Core2Frag program uses vector formulations for defining the blocks and calculating the block volumes. Typical fragmentation estimates using the Core2Frag program are shown in Figure 6 for drill-hole core piece length data from different areas within the same rock type. The Core2Frag program also incorporates a secondary fragmentation module in which the primary blocks are broken mostly by reduction of the block sizes to a stable aspect ratio. The draw height is used to control the number of times the blocks are split within the draw column. Each time a block is split, a certain amount of fines is also created. The secondary fragmentation distribution is thus generated and the program stores the secondary blocks for use in the subsequent analyses related to hang-ups. Figure 4: Effect of drill-hole intercept not passing through corner of block Three joint sets are randomly selected and used to develop the shape of the rock block. The volume of the block is then calculated using the drill core piece length, as described below. Assuming that a drill hole passes through a corner of the block, the angles between the drill core piece and the axes making up the block defined by the three selected joint sets are calculated as: α = angle between core and set 1 = 1•h/L β = angle between core and set 2 = 2•h/L γ = angle between core and set 3 = 3•h/L where 1 2 3 h L
= = = = =
unit normal vector to set 1 unit normal vector to set 2 unit normal vector to set 3 vector of core piece length of core piece
Figure 6: Primary fragmentation estimates using Core2Frag in different areas in same rock type.
These angles determine which side of the block is intersected by the core piece. The volumes of the blocks Massmin 2004
A hang-up module, incorporating the logic developed by Robin Kear (as reported by Esterhuizen, 1999), is also included in the program. Five blocks are randomly selected
Santiago Chile, 22-25 August 2004
57
from the blocks created by the secondary fragmentation module. The blocks are rotated and passed through a drawbell. If the blocks do not pass through the drawbell, a hang-up is recorded. The type of hang-up depends on where in the drawbell the blocks get stuck. The blocks stored from the secondary fragmentation module are passed through the hang-up module till the actual tonnage passing through the drawpoint is reached. 5 CONCLUSIONS While an attempt has been made by the authors for the estimation of primary and secondary fragmentation based on the drill core data, the work is in no way complete. The program needs to be tested against a standard set of data to ensure robustness and consistency of the results. Data collection procedures from exploration drill holes need to be developed to provide inputs for the Core2Frag program. As with other methods of estimating fragmentation, the Core2Frag programs will need extensive field-testing and calibration before it can provide reasonably accurate quantitative results. The authors are proposing to use drawpoint fragmentation assessment as a means of direct calibration of the procedure and program. The authors are currently working to develop a standardized procedure for data collection from the drawpoints and some work in this regard has been initiated at the DOZ Block Cave operated by PT Freeport Indonesia (Srikant et al, 2004). At present, however, the Core2Frag program can be used to qualitatively compare fragmentation in different areas within block cave mines as well as open pit mines. Though the process of calibration of fragmentation estimates has a suitable feedback loop in the case of openpit design and, in many cases, stope design, this calibration process is more difficult to use in the development of design parameters in block cave design because (1) the development and equipment are put in place much in advance of actual production and (2) there is limited scope for adjustment or modification of major design parameters, especially those aspects of block cave design that are based on fragmentation. 6 ACKNOWLEDGEMENTS The authors thank David Flint and George McDonald of Freeport McMoRan Copper and Gold for supporting the development of the Core2Frag program, which was
58
developed for use on several projects in PT Freeport Indonesia. The guidance and constructive criticism from the engineers at Call & Nicholas, Inc. is also gratefully acknowledged. 7 REFERENCES • Calder, K., Townsend, P. and Russell, F., "The Palabora Underground Mine Project", Proc. MassMin 2000, AusIMM, Brisbane, 2000. • Carter, C.J., and Russell, F.M., "Modeling and Design of Block Caving at Bingham Canyon", Proc. MassMin 2000, AusIMM, Brisbane, 2000. • Chen, D., "Geotechnical assessment of block cave mining in Northparkes Mines, NSW, Australia", Rock Mechanics Tools and Techniques (ed. Aubertin, Hassani, Mitri), Balkema, Rotterdam, 1996. • Esterhuizen, G.S., BCF Version 3.0 – A Program to Predict Block Cave Fragmentation - Technical Reference and User’s Guide, Littleton, Colorado, 1999. • Hardy, A.J., Ryan, T.M. and Kemeny, J.M., "Block size distribution of in situ rock masses using digital image processing of drill core", International Journal of Rock Mechanics and Mining Sciences, v 34, n 2, Feb, 1997. • Laubscher, D.H., "Cave mining – the state of the art", Jour. South African Institute of Mining and Metallurgy, October 1994. • McCormick, "How wide does a drawpoint draw?", EM/J v 169 n 6., 1968. • Mojtabai, N., Cetintas, A., Farmer, I.W., and Savely, J.P., "In-place and excavated block size distributions", Proc. 30th US Rock Mech. Symp., West Virginia University, Morgantown, 1988. • Nickson, S., Coulson, A., and Hussey, J., "Noranda’s Approach to Evaluating a Competent Deposit for Caving", Proc. MassMin 2000, AusIMM, Brisbane, 2000. • Poulton, M.M, Mojtabai, N., and Farmer, I.W., "Scale invariant behavior of massive and fragmented rock", International Journal of Rock Mechanics and Mining Sciences & Geomechanics Abstracts, v 27 n 3, June 1990. • Pratt, R.W., Srikant, A., Nicholas, D.E. and Flint, D.C., Analysis of DOZ Fragmentation, CNI Report, May 2002. • Srikant, A., Nicholas, D.E., and Rachmad, L., "Visual Estimation of Fragment Size Distributions in the DOZ Block Cave", Proc. MassMin 2004, Santiago, August 2004.
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A review of empirical methods used to estimate rock mass compressive strength and deformability in the mining industry E. Villaescusa, J. Li, Western Australian School of Mines, Curtin University of Technology, Australia
Abstract The most widely used empirical relations, based on rock mass classifications, to estimate the rock mass compressive strength and deformation modulus are reviewed. The estimates from the relationships are compared for a wide range of general rock mass conditions and the most appropriate relationships are chosen. In addition, the selected relationships are used to quantify the scale effect by comparing the intact rock properties determined in the laboratory with the rock mass properties estimated using data from 15 mine sites. The results from the rock mass classifications are then compared with the actual estimates used at those 15 underground mines, where the geotechnical engineers simply downgrade the intact rock parameters (based on underground observations and experience) as inputs to numerical modelling. Finally, the paper compares the rock mass strength estimates using a recently developed method based on rock mass critical strain with corresponding estimates using the conventional Hoek-Brown method.
1. INTRODUCTION The deformation modulus and compressive strength of a rock mass can be determined by in situ testing. Such tests are used in civil engineering and involve among others, plate bearing, hydraulic pressure chamber, flat jack, radial jack, and direct shear tests (Lama and Vutukuri, 1978 and ISRM, 1981). These tests are extremely expensive and time consuming, and consequently, rarely carried out within the mining industry. On the other hand, over the last two decades, a number of empirical relations to estimate the deformability and compressive strength of a rock mass based on rock mass classifications have been developed (Bieniawski, 1978; Serafim and Pereira, 1983; Kalamaras et al, 1995; Hoek and Brown, 1997; Sheorey, 1997; and Hoek, 1998). Furthermore, at some mines the rock mass properties are estimated by simply downgrading some of the intact rock properties, which are then used as inputs to numerical modelling. The results from the numerical modelling are calibrated with observations of the rock mass behaviour at the mine sites and the input of the rock mass parameters are "fine tuned". 2. ROCK MASS COMPRESSIVE STRENGTH AND DEFORMABILITY The rock mass compressive strength is a measure of the peak load carrying capacity of a rock mass (i.e stress driven failure) and is likely to be a fraction of the intact rock strength due to the presence of geological discontinuities (Hoek and Brown, 1980). The actual value depends upon the geometrical nature and strength of the geological
Massmin 2004
discontinuities, as well as the physical properties of the intact rock bridges. The modulus of deformation (or deformation modulus) is the ratio of stress to corresponding strain during loading of a rock mass including elastic and inelastic behaviour. This is different to the modulus of elasticity or Young’s modulus, which is the ratio of stress to corresponding strain below the proportionality limit of a material, i. e. intact rock (ISRM, 1975). 3. EMPIRICAL RELATIONS BASED ON ROCK MASS CLASSIFICATIONS Given that large scale testing of a rock mass is cost prohibitive in mining operations, a practical alternative is to estimate the rock mass properties from rock mass classifications and intact rock core sampling testing. Consequently, a number of empirical relationships to determine the deformability and strength of rock masses using the rock mass classification indices have been proposed since the late 1970s. 3.1 Estimation of rock mass compressive strength A number of empirical relations for determining the compressive strength of a rock mass are available in the literature. The most widely used relationships are given in Table 1. In all but two cases, the empirical equations incorporate the uniaxial compressive strength of the intact rock coupled with a rock mass classification system. Table 1 also includes a procedure proposed by Hoek and Brown (1997) to calculate the rock mass compressive strength by fitting the Mohr-Coulomb failure criterion envelope to the Hoek-Brown criterion as shown in Figure 1.
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Table 2 shows uniaxial compressive strength values for typical intact rocks, as well as GSI and rock mass strength values from Hoek and Brown (1997) and the resulting estimated rock mass compressive strengths calculated using Eqs.1 to 6 from Table 1. Analysis of the data suggests that Equation 2 (Trueman, 1988) may not work well for cases in which a rock mass has a low uniaxial compressive strength of intact rock and high GSI or RMR values. For a quartz mica schist having a uniaxial compressive strength of 30MPa and a GSI (or RMR) of 65, the calculated rock mass compressive strength using equation 2 is 24.7MPa, which is much higher than the value (8.2MPa) calculated using the Mohr-Coulomb to fit the Hoek-Brown criterion (Hoek-Brown method). Similarly, it can be seen that, in most cases, the values of a rock mass compressive strength calculated using equation 3 (Singh, 1993) are higher than the values estimated using the Hoek-Brown procedure. Again, for the quartz mica schist the estimated rock mass compressive strength is 39.6MPa, which is higher than the uniaxial compressive strength of the intact rock (30MPa). In addition, equations 2 and 3 do not explicitly consider both the uniaxial compressive strength of intact rock and a
Figure 1. Hoek-Brown and Mohr-Coulomb criteria and rock mass compressive strength (Hoek and Brown, 1997).
Table 1. Empirical relations for determining compressive strength of rock masses. Empirical Relations & Procedure
Reference
Equation Number
Ramamurthy, 1986
(1)
σcm = 0.5e0.06RMR
Trueman, 1988
(2)
σcm = 0.7 γQ 1/3
Singh, 1993
(3)
Kalamaras et al, 1995
(4)
Sheorey, 1997
(5)
Hoek, 1998
(6)
σcm = σce(RMR
76
- 100)/18.75
76
RMR89 – 15
σcm = 0.5
85
σcm = σce(RMR
76
σc
- 100)/20
σcm = 0.022σce0.038GSI Mohr-Coulomb to fit Hoek-Brown criterion (Hoek-Brown method)
Hoek and Brown, 1997
Note: γ is the bulk unit weight of intact rock (kN/m3); σc is uniaxial compressive strength of intact rock; Q is the classification index of Barton et al (1974); GSI is Geological Strength Index (Hoek et al, 1995) and RMR76 and RMR89 are Rock Mass Rating of Bieniawski in 1976 and 1989.
Table 2. Rock mass compressive strength determined by empirical relations. Rock mass description
σc
σcm (MPa)
GSI or
(MPa) RMR§ Hoek & Brown (1997)# V/g quality hard rock mass
150
75
64.8
Eq.1
Eq.2
Eq.3*
Eq.4
Eq.5
Eq.6
39.5
45.0
57.4
57.4
43.0
57.0
Average quality rock mass
80
50
13.0
5.6
10.0
22.7
18.8
6.6
11.8
V/p quality rock mass
20
30
1.7
0.5
3.0
10.8
2.4
0.6
1.4
Massive gneiss
110
75
43.0
29.0
45.0
57.4
42.1
31.5
41.8
Quartz mica schist
30
65
8.2
4.6
24.7
39.6
9.7
5.2
7.8
Graphitic phyllite
15
24
1.0
0.3
2.1
8.7
1.2
0.3
0.8
* The bulk density of all rock types is assumed as 26kN/m3, and Q-index in Eq.5.12 is calculated by using RMR = 9 in Q + 44 (Bieniawski, 1984). # Compressive strength of rock mass was calculated using the Hoek-Brown method (Mohr-Coulomb to fit Hoek-Brown criterion). / In Eqs.1, 2 and 5; RMR76 = GSI and in Eq.4; RMR89 = GSI + 5. V/g is very good and V/p stands for very poor. 60
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rock mass classification system, and consequently are not considered further in this study. Eqs.1 and 5 are also excluded since these estimated rock mass compressive strength are very low compared with the results obtained using the Hoek-Brown method. This leaves Eqs. 4 and 6 coupled with the Hoek-Brown method as the recommended choices for estimation of a rock mass compressive strength using intact rock strength and rock mass classification systems. 3.2 Estimation of rock mass deformation modulus The most widely used empirical relations for determining the deformation modulus of rock masses are given in Table 3. Eqs. 7 and 8 are only valid when RMR>50 and Q>1, respectively. It is noted that in cases when the ranges from 50 to 55 or Q is less than 2, the deformation modulus values estimated using these equations appear to be too low, (see Figure 2). In addition, both equations may not provide a reasonable fit for the field measured data, since they are not applicable for a wide range of values.
Figure 2. Deformation modulus of rock mass versus rock mass rating (Hoek and Brown, 1988).
Equation 10 appears to overestimate the rock mass deformation modulus. This can be explained by the extensive literature review of Heuze (1980), which showed that the deformation modulus of a rock mass is about 20 to 60% of the intact rock Young’s modulus. When RMR ≥ 57, the ratio of deformation modulus of rock masses to Young’s modulus of intact rocks using Equation 10 is greater than 0.6. Furthermore, Hoek and Brown (1997) recommended to use Eq.9 to estimate the rock mass deformation modulus, and modified it into Eq. 11 in order to consider the intact rock strength. Hoek and Brown (1997) indicated that the deformation of better quality rock masses is controlled by the geological discontinuities; while for poorer quality rock masses the deformation of the intact rock pieces contributes to the overall deformation process. Consequently, only Eqs.9 and 11 are recommended to estimate the rock mass deformation modulus. Li (2004) has estimated the rock mass compressive strength using the Hoek-Brown method (Mohr-Coulomb to fit Hoek-Brown criterion) and the rock mass deformation modulus using Eqs. 9 and 11. The results for a number of mine sites are listed in Table 4 indicate the rock mass parameters to be a proportion of the intact rock parameters. Figure 3 shows the calculated ranges for σc /σcm, which appear to be log normally distributed. The results show that on average σcm is approximately 1/4 of σc. Figure 4 shows the calculated ranges for E/Em, which appear to be negative exponentially distributed. The results show that on average Em is approximately 1/3 of E.
Figure 3. Histogram of - Rock mass classifications.
Table 3 Empirical relations for determining rock mass deformation modulus. Empirical relationships
Reference
Equation Number
Em = 2RMR76 – 100 (RMR > 50)
(GPa)
Bieniawski, 1978
7
Em = 25 logQ (Q > 1)
(GPa)
Barton et al, 1980
8
Em = 10
(GPa)
Serafim and Pereira, 1983
9
Mitri et al, 1994
10
Hoek and Brown, 1997
11
Em E
(RMR76 – 10) /40
= 0.5 [1 – cos (π * RMR89/100)]
Em =
σc 100
10
(GSI–10)/40
(σc < 100 MPa)
Note: Em is the deformation modulus of rock mass and is Young’s modulus of intact rock. Massmin 2004
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Table 4. Intact rock and rock mass classification data with estimated rock mass parameters. Mine Site
RMR76
mb
S
a
E (GPa)
σcm (MPa)
Em (GPa)
σc/ σcm
E/E
(MPa)
σc
Ora Banda
Host rock Ore body
70 55
5.82 3.41
0.0357 0.0067
0.5 0.5
185 145
100 55
60.3 30.8
31.6 13.3
3.07 4.71
3.16 4.14
Harlequin
Quartz Porphyry Gabbro (dolerite) Basalt Granodiorite
56 57 66 62 59
5.19 3.66 8.02 4.37 6.94
0.0075 0.0084 0.0229 0.0147 0.0105
0.5 0.5 0.5 0.5 0.5
116 76 165 86 228
83 ----66 61 -----
29.6 17.1 56.2 22.1 67.5
14.1 13.0 25.1 18.5 16.8
3.92 4.52 3.17 3.94 3.79
5.89 ----2.63 3.30 -----
Quartz Porphyry Gabbro Basalt
59 61 67 70
5.32 4.22 8.31 5.42
0.0105 0.0131 0.0256 0.0286
0.5 0.5 0.5 0.5
106 221 281 246
-----------------
28.2 55.2 98.4 75.4
16.8 18.8 26.6 28.2
3.80 4.28 3.01 3.19
-----------------
Basalt (Tramways Dome) Basalt (Kambalda Dome) Basalt (Widgiemooltha Dome) Intermediate Porphyry Felsic Porphyry Ultramafic (Talc chlorite) Ultramafic (Talc magnesite) Ultramafic (Antigorite) Ultramafic (Lizardite)
78 78 78 68 79 66 66 73 62
7.75 7.75 7.75 5.42 8.03 6.53 6.53 8.39 5.66
0.0868 0.0868 0.0868 0.0286 0.0970 0.0229 0.0229 0.0498 0.0147
0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5
180 228 130 110 219 59 30 226 109
78 80 79 87 62 80 33 70 37
76.9 97.4 55.5 33.7 97.0 27.0 10.0 88.0 31.0
50.1 50.1 50.1 28.2 53.1 23.2 13.8 37.6 20.0
2.34 2.35 2.32 3.24 2.26 2.19 3.00 2.57 3.52
1.56 1.60 1.58 3.09 1.17 3.45 2.39 1.86 1.85
Mt Charlotte
Dolerite (Unit 8)
75
7.78
0.0622
0.5
177
70
70.3
42.2
2.52
1.66
Mt Isa
Pb Mine Shales
75
9.01
0.0622
0.5
185
80
76.8
42.2
2.41
1.90
New Celebration
Ultramafic Gabbro
51 72
3.95 9.94
0.0043 0.0399
0.5 0.5
40 208
25 81
7.7 82.6
7.0 33.5
8.42 2.47
3.57 2.28
Mt Marion
Ultramafic
60
7.91
0.0117
0.5
103
67
32.5
17.8
3.17
3.76
Outkumpu
Volcanite Pyrite
76 84
5.7 9.2
0.0695 0.1690
0.5 0.5
222 105
70 122
83 55.7
46.7 70.8
2.67 1.89
1.50 1.72
Centenary
Felsic Granitoid Magnetic Dolerite Magnetic Quartz Dolerite Lamprophyre
69 64 69 69 58
6.94 8.29 6.28 6.28 4.24
0.0319 0.0183 0.0319 0.0319 0.0094
0.5 0.5 0.5 0.5 0.5
220 210 231 242 123
75 81 91 89 70
75 71 76 79 30
29.9 22.4 29.9 29.9 15.8
2.93 2.96 3.04 3.06 4.08
2.51 3.62 3.05 2.97 4.40
Darlot
Basalt Felsic Granitoid Magnetic Dolerite Magnetic Quartz Dolerite Tuff Lamprophyre
55 69 64 69 69 64 58
3.41 6.94 8.29 6.28 6.28 4.15 4.24
0.0067 0.0319 0.0183 0.0319 0.0319 0.0183 0.0094
0.5 0.5 0.5 0.5 0.5 0.5 0.5
153 167 130 235 240 126 131
91 --------99 87 ----81
33 57 44 77 79 33 32
13.3 29.9 22.4 29.9 29.9 22.4 15.8
4.63 2.93 2.96 3.05 3.03 3.80 4.08
6.80 --------3.30 2.91 ----5.14
Grit Porphyry Conglomerate
82 68 82
11.57 7.97 11.57
0.1353 0.0286 0.1353
0.5 0.5 0.5
140 110 140
70 65 70
73.8 38.5 73.8
63.1 28.2 63.1
1.90 2.86 1.90
1.11 2.31 1.11
Ultramafic Ore
63 47
5.60 3.77
0.0164 0.0028
0.5 0.5
150 40
61 27
43 8
21.1 5.3
3.49 5.00
2.89 5.09
Dolerite All ore
75 68
11.06 4.78
0.0622 0.0286
0.5 0.5
335 141
92.6 74.7
148 42
42.2 28.2
2.26 3.36
2.19 2.65
Malu Quartzite Outer siltstone Footwall sandstone
78 47 58
10.94 1.36 4.23
0.0868 0.0028 0.0094
0.5 0.5 0.5
250 30 150
80 20 60
117.5 4.0 36.3
50.1 4.6 15.8
2.18 7.32 4.13
1.60 4.35 3.80
Bullen
Kambalda Nickel Mines
Kanowna Belle
Yilgarn Star
Junction
Telfer Gold
62
Rock Type
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Mine Site
Hoek and Brown (1997), Hoek (1998)
Rock Type
RMR76
Massive strong rock mass Average quality rock mass Poor quality rock mass Massive weak rock Massive strong rock mass Average quality rock mass Poor quality at shallow depth
75 65 24 80 80 70 20
mb
7.25 4.5 0.66 16.3 7.25 4.5 0.55
S
0.062 0.02 0 0.1084 0.1084 0.0357 0
Figure 4. Histogram of - Rock mass classifications. 4. NUMERICAL MODELLING PRACTICE IN UNDERGROUND MINING In underground mining practice, geotechnical engineers estimate the rock mass compressive strength and deformation modulus by either using the Hoek-Brown criterion (1988) or by simply downgrading some of the intact rock strength parameters, which are then used as input to numerical modelling. The input parameters are usually adjusted and finetuned with information from underground observations, and geotechnical instrumentation. Table 5 shows a number of input data that were used by mine site based practitioners during their numerical modelling of mining sequences. In most cases, they concluded that the stress outputs and the related excavation stability assessments from numerical modelling were acceptable and used during actual orebody sequencing. Figure 5 shows the calculated ranges for , which also appear to be log normally distributed. The results show that on average is approximately 1/5 of . Figure 6 shows the calculated ranges for , which also appear to be negative exponentially distributed. The results show that on average is approximately 1/2 of .
a
0.5 0.5 0.53 0.5 0.5 0.5 0.55
E (GPa)
σcm (MPa)
Em (GPa)
σc/ σcm
E/E
(MPa)
σc 110 30 15 51 110 30 7.5
-----------------------------
43 8.2 0.34 28.3 48.9 9.05 0.52
42 13 0.87 40.2 56.2 17.3 0.5
2.56 3.66 44.12 1.80 2.25 3.31 14.42
-----------------------------
Figure 6. Histogram of experience.
E / Em - Numerical modelling
Comparisons of the estimated parameters using rock mass classification and those used in numerical modelling are shown in Figures 7 and 8. The two methods appear to estimate different rock mass compressive strengths for the same rock mass environments. However, the deformation moduli estimated using both methods are similar.
Figure 7. Histogram of Oc / Ocm .
Figure 5. Histogram of - Numerical modelling experience. Massmin 2004
Figure 8. Histogram of E / Em.
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63
Table 5. Estimated rock mass parameters from numerical modelling experience. Mine Site
Rock Type
σc (MPa)
Em (GPa)
σc / σcm
E / Em
Ora Banda
Host rock Ore body
185 145
100 55
95 70
40 16
1.95 2.07
2.50 3.44
Harlequin
Quartz Porphyry Gabbro (dolerite) Basalt Granodiorite
116 76 165 86 228
83 ----66 61 -----
18 10 38 14 31
15 12 32 19 9
6.44 7.60 4.34 6.14 7.35
5.53 ----2.06 3.21 -----
Quartz Porphyry Gabbro Basalt
106 221 281 246
-----------------
16 28 73 56
15 12 38 32
6.63 7.89 3.85 4.39
-----------------
Mt Charlotte
Dolerite (Unit 8)
177
70
100
65
1.77
1.08
Mt Isa
Pb Mine Shales
185
80
94
60
1.96
1.33
New Celebration
Ultramafic Gabbro
40 208
25 81
---------
11 32
---------
2.27 2.53
Kambalda Nickel Mines
Basalt (Tramways Dome) Basalt (Kambalda Dome) Basalt (Widgiemooltha Dome) Intermediate Porphyry Felsic Porphyry Ultramafic (Talc chlorite) Ultramafic (Talc magnesite) Ultramafic (Antigorite) Ultramafic (Lizardite)
180 228 130 110 219 59 30 226 109
78 80 79 87 62 80 33 70 37
90 114 65 55 109 30 15 113 55
50 50 50 30 50 20 10 55 30
2.00 2.00 2.00 2.00 2.01 1.97 2.00 2.00 1.98
1.56 1.60 1.58 2.90 1.24 4.00 3.30 1.27 1.23
Mt Marion
Ultramafic
103
67
33
18
3.17
3.76
Outkumpu
Volcanite Pyrite
222 105
70 122
45 71
59 43
4.93 1.48
1.19 2.84
Grit Porphyry Conglomerate
140 110 140
70 65 70
73.8 38.5 73.8
63 28 63
1.90 2.86 1.90
1.11 2.31 1.11
Ultramafic Ore
150 40
61 27
13.1 13.1
12 12
11.45 3.05
5.08 2.25
Dolerite All ore
335 141
93 75
200 108
90 78
1.68 1.31
1.03 1.00
Malu Quartzite Outer siltstone Footwall sandstone
250 30 150
80 20 60
75 2 15
58 8 17
3.33 15.00 10.00
1.38 2.50 3.53
Massive strong rock mass Average quality rock mass Poor quality rock mass Massive weak rock Massive strong rock mass Average quality rock mass Poor quality at shallow depth
110 30 15 51 110 30 7.5
-----------------------------
43 8.2 0.34 28.3 48.9 9.05 0.52
42 13 0.87 40.2 56.2 17.3 0.5
2.56 3.66 44.12 1.80 2.25 3.31 14.42
-----------------------------
Bullen
Kanowna Belle
Yilgarn Star
Junction
Telfer
Hoek and Brown (1997), Hoek (1998)
64
E (GPa)
σcm (MPa)
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5. ROCK MASS STRENGTH FROM CRITICAL STRAIN Recent work at the Western Australian School of Mines (Li, 2004) established a correlation between critical strain for intact rock and rock masses using 81 data points collected from Western Australian underground mining operations. From this study a linear relationship between the critical strain for rock masses and intact rock can be approximately expressed as follows
cases, similar rock mass strength is estimated using both methods. The comparisons of the rock mass compressive strengths calculated using the Hoek-Brown method and Eq.15 are also shown in Figure 10.
(12)
where and are the critical strain of intact rock and rock masses expressed in % units respectively. The critical strain for intact rock is defined as follows (Sakurai, 1982, Li, 2004),
(13)
The critical strain for rock masses is defined as follows (Li, 2004),
(14)
The conceptual average axial stress versus axial strain plot for the intact rock and rock mass are shown in Figure 9.
Figure 10. Comparison of rock mass compressive strength estimated using Hoek-Brown method and Eq.15. 6. CONCLUSIONS A comprehensive review of rock mass classification and numerical modelling input data was carried out for a large number of Australian mining operations. The data were reviewed with respect to rock mass compressive strength and deformation modulus. The methodologies proposed by Kalamaras and Hoek and Brown are recommended for rock mass compressive strength, while the methods of Serafim and Pereira, as well as Hoek and Brown are recommended to estimate the rock mass deformation modulus. A comparison of data from the same mining sites showed that minebased geotechnical engineers actually use slightly lower rock mass strength values than those estimated using rock mass classifications. Finally, a recently developed methodology based on rock mass critical strain was compared to the Hoek and Brown method and the results suggest that both methods predict similar rock mass strength estimates.
7. REFERENCES
Figure 9. Comparison of intact rock and rock mass behaviour
Eq.12 can be used to estimate the rock mass compressive strength as follows (Li, 2004)
(15)
Table 6 shows a comparison of rock mass compressive strength values using Eq.15 and those calculated using the Hoek-Brown method (Mohr-Coulomb to fit HoekBrown criterion) and the rock mass deformation modulus using Eqs.9 and 11. The results suggest that, in most Massmin 2004
• Barton N. R., Lien R. and Lunde J. 1974, ‘Engineering classification of rock masses for the design of tunnel support’, Rock Mechanics, vol.6, pp189 – 236. • Barton N. R., Løset F., Lien R. and Lunde J. 1980, ‘Application of the Q-system in design decision’, Subsurface Space, (ed. M. Bergman) 2, pp553 - 561. New York: Pergamon. • Bieniawski Z. T. 1978, ‘Determining rock mass deformability – Experience from case histories’, International Journal of Rock Mechanics and Mining Sciences & Geomechanics Abstracts, vol.15, no.5, pp.237 – 247. • Bieniawski Z. T. 1984, Rock Mechanics Design in Mining and Tunnelling, A. A. Balkema, Rotterdam. • Heuze F. E 1980, ‘Scale effects in the determination of rock mass strength and deformability’, Rock Mechanics, vol.12, pp167 – 192.
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Table 6. Rock mass compressive strength calculated using Hoek-Brown method and Eq.15. Mine Site
Rock Type
RMR76
E (GPa)
εcm
(MPa)
σc
(%)
Em (GPa)
σcm (MPa) H-B method Eq. 15
Ora Banda
Host rock Ore body
70 55
185 145
100 55
0.185 0.264
31.6 13.3
60.3 30.8
46.2 27.7
Harlequin
Quartz Porphyry Gabbro (dolerite) Basalt Granodiorite
56 57 66 62 59
116 76 165 86 228
83 ----66 61 -----
0.140 ----0.250 0.141 -----
14.1 13 25.1 18.5 16.8
29.6 17.1 56.2 22.1 67.5
15.6 ----49.9 20.6 -----
Kambalda Nickel Mines
Basalt (Tramways Dome) Basalt (Kambalda Dome) Basalt (Widgiemooltha Dome) Intermed. Porphyry Felsic Porphyry Ultramafic (Talc chlorite) Ultramafic (Talc magnesite) Ultramafic (Antigorite) Ultramafic (Lizardite)
78 78 78 68 79 66 66 73 62
180 228 130 110 219 59 30 226 109
78 80 79 87 62 80 33 70 37
0.231 0.285 0.165 0.126 0.353 0.074 0.091 0.323 0.295
50.1 50.1 50.1 28.2 53.1 23.2 13.8 37.6 20
76.9 97.4 55.5 33.7 97 27 10 88 31
91.3 112.8 65.1 28.2 148.2 13.5 9.9 95.9 46.5
Mt Charlotte
Dolerite (Unit 8)
75
177
70
0.253
42.2
70.3
84.3
Mt Isa
Pb Mine Shales
75
185
80
0.231
42.2
76.8
77.1
New Celebration
Ultramafic Gabbro
51 72
40 208
25 81
0.160 0.257
7 33.5
7.7 82.6
8.8 68.0
Mt Marion
Ultramafic
65
114
78
0.146
17.8
46.0
36.3
Outkumpu
Volcanite Pyrite
76 84
222 105
70 122
0.317 0.086
46.7 70.8
83.0 55.7
117.0 48.1
Centenary
FELSIC Granitoid Magnetic Dolerite Magnetic Quartz Dolerite Lamprophyre
69 64 69 69 58
220 210 231 242 123
75 81 91 89 70
0.293 0.259 0.254 0.272 0.176
29.9 22.4 29.9 29.9 15.8
75 71 76 79 30
69.3 45.9 60.0 64.2 21.9
Darlot
Basalt FELSIC Granitoid Magnetic Dolerite Magnetic Quartz Dolerite Tuff Lamprophyre
55 69 64 69 69 64 58
153 167 130 235 240 126 131
91 --------99 87 ----81
0.168 --------0.237 0.276 ----0.162
13.3 29.9 22.4 29.9 29.9 22.4 15.8
33 57 44 77 79 33 32
17.7 --------56.1 65.2 ----20.2
Grit Porphyry Conglomerate
82 68 82
140 110 140
70 65 70
0.200 0.169 0.200
63.1 28.2 63.1
73.8 38.5 73.8
99.7 37.7 99.7
Ultramafic Ore
63 47
150 40
61 27
0.246 0.148
21.1 5.3
43 8
41.0 6.2
Dolerite All ore
75 68
335 141
92.6 74.7
0.362 0.189
42.2 28.2
148 42
120.6 42.1
Malu Quartzite Outer siltstone Footwall sandstone
78 47 58
250 30 150
80 20 60
0.313 0.150 0.250
50.1 4.6 15.8
117.5 4 36.3
123.7 5.5 31.2
Kanowna Belle
Yilgarn Star
Junction
Telfer
66
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• Hoek E. 1998, Practical Rock Engineering, [Online], Available: http://www.rocscience.com/roc/Hoek/Hoek.htm [2000, October 2]. • Hoek E. and Brown E. T. 1980, Underground Excavations in Rock, Institute of Mining and Metallurgy, London. • Hoek E. and Brown E. T. 1997, ‘Practical estimates of rock mass strength’, International Journal of Rock Mechanics & Mining Sciences, vol.34, no.8, pp.1165 – 1186. • Hoek E., Kaiser P. K and Bawden W. F. 1995, Support of Underground Excavations in Hard Rock, Balkema, Rotterdam. • International Society for Rock Mechanics (ISRM) 1975, ‘Report of the commission on Terminology’, International Society for Rock Mechanics, Lisbon. • International Society for Rock Mechanics (ISRM) 1981, Rock characterization, Testing and Monitoring: ISRM Suggested Methods, ed. Brown, E. T., Pergoman Press, Oxford. • Kalamaras G. S. and Bieniawski Z. T. 1995, ‘A rock mass strength concept for coal seams incorporating the effect of time’, in Proceedings 8th International Congress on Rock Mechanics, ISRM, ed. Fujii T., A A Balkema, Rotterdam, vol.1, pp.295 – 302. • Lama R. D. and Vutukuri V. S. 1978b, Handbook on Mechanical Properties of Rocks, vol. III, Trans Tech Publications, Clausthal.
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• Li, J. 2004, Critical Strain of Intact Rock and Rock Masses. PhD Thesis, Western Australian School of Mines, Curtin University of Technology, 186p. • Ramamurthy T. 1986, ‘Stability of rock mass’, Indian Geomechanics Journal, vol.16, no.1, pp1 – 74. • Sakurai S. 1982, ‘An evaluation technique of displacements in tunnels’, Proceedings of Japanese Society of Civil Engineering, No. 317, pp.93 – 100 (in Japanese). • Serafim L. J. and Pereira P. J. 1983, ‘Consideration on the geomechanical classification of Bieniawski’, in Proceedings of International Symposium On Engineering Geology and Underground Construction, vol.1, pp.II.33 – II.42, LNEC, Lisbon. • Sheorey P. R. 1997, Empirical Rock Failure Criteria, A A Balkema, Rotterdam. • Singh B. 1993, ‘Indian case studies of squeezing grounds and experiences of application of Barton’s Q-system’, in Workshop on Norwegian Method of Tunnelling, CSMRS, New Delhi. • Trueman R. 1988, An evaluation of strata support technique in dual life gateroads. Ph.D. Thesis, University of Wales, Cardiff.
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Chapter 3
Mass Mining Methods I: Fundamentals
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Feasibility studies for caving operations – The "State of the Art" Mike A. Struthers, Principal, AMC Consultants (UK)
Abstract Technical reviews of a number of feasibility studies for potential caving operations, and involvement in operating mines, have illustrated a number of opportunities for improvement in these investigations. This paper presents a view on the "State of the Art" of feasibility studies as applied to caving projects, and highlights some of the remaining challenges. In particular, the need for in-situ stress determinations as input to such studies is discussed, and research into a number of alternative methods should continue to be trialed in the hard rock mining industry. The value of effective risk assessment is emphasised, the process beginning during the pre-feasibility, and carried through into the feasibility study and on to operations, undergoing numerous updates and revisions. Independent review during the Study process is also invaluable.
1 INTRODUCTION Design for caving methods of mining remains one of the most significant challenges in mining geomechanics and engineering, particularly in green-field sites. During the past 2-3 decades design has been based predominantly on methods developed by Laubscher (eg. 2001), but in recent years alternative empirical schemes and numerical methods have been applied in parallel with the ‘conventional’ MRMR assessments. This has occurred in parallel with a desire by industry to exploit much more competent orebodies than were historically considered suited to caving, a trend which itself has challenged the applicability of the established design methods. This has been complemented by, and to some degree driven by, the International Caving Study (ICS). This trend has applied to both sub-level, block, and panel cave projects. At a more detailed level, knowledge of a number of specific aspects (eg. cave draw) has advanced considerably. Involvement in, and review of, a number of Feasibility Studies for caving projects has highlighted some common observations regarding the "state of the art" as it is applied to such studies. 2 TECHNICAL CHALLENGES Although the degree of sophistication used to address particular geotechnical issues in feasibility studies into caving projects does vary, in most cases the level of confidence one can have in the range of fundamental input parameters is summarised below: • Rockmass characterisation – usually good. • Major structures – moderately well-defined spatially, and associated ground conditions similarly defined, but their influence on (and reaction to) the caving process is poorly understood. • Stress regime – poorly understood. • Deposit hydrogeology – poor to fair. We are as an industry generally competent at the following tasks during the Study process (though there of course is always room for improvement): • Ore resource modelling and estimation. • Mining method selection. Massmin 2004
• Rockmass characterisation, and presentation of the results (though there remains much scope for improving modelling of geotechnical data). • Fragmentation analysis (though estimation of secondary fragmentation and size and frequency of hang-ups requires improvement). • Options analysis for materials handling systems, and simulation studies. • Ventilation modelling and design. • Cost modelling and economic analysis. Specific design tasks which remain challenging, and on which further research is required, include: • Cavability predictions in strong rockmasses, particularly where stress-induced caving is expected to play a major role. • Specifically, the state of pre-mining stress. • Assessing the likely role or impact of major structures on cavability, and excavation stability close to the cave zone. • Estimation of the rate of cave initiation and propagation. • As a consequence, estimation of surface subsidence (timing, extent, magnitude). • Estimation of draw and recovery characteristics, and factors such as dilution entry. Specifically, monitoring of draw behaviour, and draw modelling. • Assessing the degree of seismic risk • Under some circumstances, assessment of the likely water inflows into the cave, and hence assessment of inrush risks. All these aspects are of course more challenging in greenfield sites, with no prior mining experience, and little in the surrounding region. In-situ stress determination is one the most challenging problems facing studies into deep caving operations. It is beyond the scope of this paper to discuss in detail the various methods of stress measurement in use today. Overcoring techniques remain the preferred methods where access is available, but often for feasibility studies it is not. Hydraulic fracturing can obtain stress determinations from deep boreholes, but the method has it’s limitations. A range of core-based measurement methods are under development, such as DSCA (Differential Strain Curve Analysis), the ASR (Anelastic Strain Recovery) method, the AE (Acoustic Emission) method, DRA (Deformation Rate
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Analysis), and PSHA (Principal Strain Hysteresis Analysis) (Oikawa and Yamaguchi, 2000). Oikawa and Yamaguchi tested rock cores from 2200m below surface, with some success when compared to a range of other assessments. However, these are not in-situ tests, and the methods are not without their detractors. Sjöberg and Klasson (2003) presented encouraging stress results from use of a SSPB Borre probe in deep boreholes in Sweden, and significant research into a variety of methods has been undertaken in Korea, supporting a number of large civil construction projects. This has included the Compact Conical-ended Borehole Overcoring (CCBO) method (Seong-Seung Kang et al, 2003). Further research is required (and warranted) into a number of these methods. Of the above ‘alternative’ methods, only the AE technique is generally applied and supported by research in hard rock mining circles. The mining industry should support development trials of some of these other methods in hard rock mines. The need for reliable assessments of in-situ stress is widely recognised, and in some cases (eg Argyle Diamonds, WA) has been a factor in decisions to establish earlier mine access than would otherwise have been available, to provide valuable confirmatory data during the feasibility process. 3 RISK AREAS In addition to the technical risk areas listed above, there are also a number of other risks which can impact dramatically on the success of a project: • Unintentional bias in studies conducted in-house. This is a substantial risk, despite the best efforts of those involved (Bass, 1987). Rigorous independent audit and review of the Study at various stages is considered essential (McCarthy, 2002). • The pressures of time and money, which are substantial factors in large studies, and which can lead to ‘rushed’ investigations of certain issues, and a reluctance to revisit early assumptions later in the Study process. • The move from Pre-Feasibility to Feasibility Study is recognised as a risk if, as noted by McCarthy (2002), the Pre-Feasibility Study has not been sufficiently comprehensive. McCarthy argues it is the Pre-Feasibility Study which is the most important stage of investigation. • Insufficient challenging of sophisticated analyses done by consultants, such as stress modelling, or modelling of draw behaviour. This can be due in part to lack of knowledge on the part of those reviewing the results. Both these examples of complex modelling can be prone to significant errors, or at least generate misunderstandings, where not calibrated against local conditions. The latter is of course not always possible, but the results should be assessed with this clearly in mind. We suffer from two syndromes; the "Black Box", and "If it’s in colour, it’s true!". The track record of feasibility studies has generally been poor, and a number of investigations from the 1970’s onwards have analysed performances and sought explanations. In a study of production rates in US projects, Tatman (2001) found that 35% of mines did not achieve their planned rate. In a study of 105 cases, McCarthy (2003) found that three factors commonly led to underperformance; lack of impartiality of the study team, criticism of the way the study was organised or funded, and unrealistic study deadlines. 4 UNEXPECTED OUTCOMES A number of major caving projects have not performed to the expectations which existed at Feasibility. Some have 72
performed better than expectations, some not. Examples of unexpected performances include: • Unexpected seismicity in capital access development at depth, when earlier mining on higher levels gave no indication such problems may exist. The impacts were on cost and development rates. • Significantly delayed cave initiation, compared to Feasibility predictions. The impacts were on costs, and production rate (ramp up). • In another case, significantly faster cave propagation than expected, and rapid cave breakthrough. Though good for the cave and production rates, there were impacts on plans for surface preparation and security in advance of the breakthrough. • The tragedy at Northparkes, investigations of which have been well published (eg. Hebblewhite, 2003). • Marker trials and modelling of draw at a sub-level caving mine which have clearly demonstrated different widths of isolated draw compared to modern ‘conventional’ estimations. Increasingly, areas of uncertainty or risk are being recognised at feasibility, and where they cannot be designed out of the project, contingency plans are established. These rely on recognising the on-set of a unfavourable condition, and it is here that modern monitoring methods are proving very successful. One example is the Ridgeway Gold Mine in Australia. A major component of the feasibility study was identification of the major mining risks associated with using the SLC method. This process resulted in a series of Major Hazard Management Plans (MHMP), each related to a specific group of hazards (Dunstan and Power, 2003). A range of monitoring strategies and so-called ‘triggers’ were implemented, including deep open holes, and seismic monitoring (Pfitzner, 2003). 5 ARE SLC MINES OPTIMISED? The author contends that many sub-level caving (SLC) operations may not be as profitable as they could be or, at least, the industry lacks the analytical tools to make such an assessment. Some SLC mines perform better than expectations, but this can be due to a number of factors, including (i) excessively conservative or optimistic expectations during feasibility, (ii) errors in assumptions on recovery and dilution, or (iii) inaccuracies in drawpoint reconciliations, allowing the mine to draw beyond plan on the levels below, recovering material left behind. Some SLC mines do not perform as well as expected. Often SLC designs are functional, but in many cases are probably sub-optimal, largely as a function of the complexity of the problem. But optimisation requires a committed mine staff who recognise the value of controlled experimentation; robust marker field trials; and a large orebody which provides the time for incorporating the results of trials into subsequent level designs. The expanding range and application of numerical models (eg. the new draw model CaveSim (Sharrock, 2004)) has considerable potential in this regard, with proper calibration. SLC layout design is a compromise between drift and brow stability, and primary recovery (draw). Figure 1 shows a preliminary design chart for use during the early stages of design, prior to undertaking any stress analyses of proposed layouts. The proposed development extraction ratio (ER%, defined as the ratio of widths of the SLC drifts and intervening pillars) is plotted against the stress/strength ratio on the sublevel. Under difficult mining conditions, designing for an ER in excess of 40% is probably optimistic.
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Once refined, guidelines such as this can reduce the need to significantly re-design a mine layout, when stress analyses undertaken later in the design process highlight some fundamental geotechnical problems. With an estimate of the isolated draw zone width, preferably from both empirical and numerical methods, an improved preliminary SLC layout may be defined.
process, and includes features such as risk calculators, and automated reporting. 7 THE WAY FORWARD In the author’s experience, feasibility studies for caving mines should always incorporate two components: • Emphasis on independent reviews at various stages in the Study, for both Pre- and full Feasibility Studies, and • Rigorous risk assessment at all stages of investigation, and continuing into operations as a crucial management tool, with well defined triggers, responses, and accountabilities. The process must be objective, with independent input to overcome unintentional bias. The risk team must be carefully selected, comprising a mixture of Study team members and some who are independent from the Study (either within or from outside the owner organisation). Continued research is also required into certain fundamentals: • More sophisticated investigation of the impacts of major structures on cave initiation and propagation, through historical review, modelling, and sensitivity studies. • Further research into alternative methods of stress estimation or measurement from deep boreholes. • Continued increase in the use of numerical stress analysis and material flow models, but accompanied by rigorous interrogation and questioning of results, by independent specialists. • Improved design methodologies for optimising SLC layouts, balancing the opposing requirements of development stability and draw recovery / performance.
Figura 1: SLC development design chart
6 RISK ASSESSMENT Risk assessment is increasingly a feature of Health and Safety legislation around the world, and is an invaluable tool in mining feasibility studies. A major outcome of a risk assessment, beyond an understanding of the degree of risk to the operation, is a clear set of contingency plans and monitoring ‘triggers’ which, with regular updating, remain important management tools throughout the life of the project. A typical example is the risk of inrush or mudrush into the cave, in which meteoric or sub-surface waters mix with fines, from either the surface weathered zone or within the cave, to create a potential risk of inrush at the drawpoints. Various strategies have been adopted by mines at which this risk is considered sufficient to warrant specific mitigation or contingency planning. Mining consultants and project owners alike have a tendency to unwittingly mask areas of uncertainty by ‘averaging’ important design inputs, to simplify design tasks (and reporting). A statement such as "….the average MRMR of the orebody is 55, and the HR required for caving is 25m" may be an elegant summary of a complex problem, or a misleading oversimplification. If the latter, through the Study process this simplification filters out inherent variability, as the design team concentrate on other necessary ‘downstream’ tasks. Risk assessment assists in maintaining the profile of known uncertainties, and ensures those risks are incorporated into all future planning. Some mining organisations (eg. Rio Tinto plc, Newcrest Mining Ltd) have fully embraced risk assessment as a core to any feasibility investigations, and roll the results through into the operating phase, continually updating the risk assessment as conditions change. Risk assessment software such as the NSCA Electronic Risk Database (O’Donoghue and Quelch, 2001) can significantly ease and streamline the risk assessment Massmin 2004
The International Caving Study (ICS) continues to contribute to understanding of caving operations, and the development of new design methodologies and tools. The findings from these studies are only slowly disseminated to the industry at large, given the financial support provided to the project by it’s sponsors. Some of the items listed here are being addressed by the ICS, some not. 8 CONCLUSIONS Caving methods of mining have experienced a rejuvenation in the last 5-10 years, and substantial advances have been made in the range and sophistication of techniques used in feasibility studies into caving operations. Some significant challenges however remain. The trend towards caving of increasingly competent rockmasses (which were historically not considered suitable for these methods) has presented some unique challenges, and required modifications to classical design tools, and development of new alternatives. ACKNOWLEDGEMENTS The author is grateful to his colleagues in AMC Consultants for their comments and input to this article. REFERENCES • Bass, C. B, 1987. Reserve Data for Project Planning. AusIMM Reserves and Resources Symposium, Nov. 1987. • Dunstan GE and Power G, 2003. Managing technical risk at Ridgeway sublevel caving mine. Proceedings of the Mining Risk Management conference, AusIMM, Melbourne, Australia. • Hebblewhite, B K, 2003. Northparkes Findings – The implications for geotechnical professionals in the mining
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•
•
•
•
•
industry. Ground Control in Mining – Technology & Practice. AGCM Conference. Laubscher, D H, and Jakubek, J, 2001. MRMR Classification for Jointed Rockmasses. In: Underground Mining Methods Handbook: Engineering Fundamentals and International Case Studies. SME, 2001. McCarthy, P L, 2002. Feasibility Studies and Economic Models for Deep Mines. First International Seminar on Deep and High Stress Mining. Australian Centre for Geomechanics, Perth. McCarthy P L 2003. Managing Technical Risk for Mine Feasibility Studies. Mining Risk Management Conference, Sydney NSW 9-12, September 2003. O’Donoghue, L and Quelch, J. Managing Health and Safety Risks - NSCA Risk Management Tools. National Safety Council of Australia Limited, Brisbane, Queensland, 2001. Oikawa, Y and Yamaguchi, T. Stress measurement using rock core in an HDR field. Proceedings World Geothermal Congress 2000, Kyushu - Tohoku, Japan, May 28 - June 10, 2000.
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• Pfitzner, M, 2003. Monitoring a blind sub-level cave - A case study of an integrated approach at Newcrest Mining’s Ridgeway Gold Mine. Ground Control in Mining – Technology & Practice. AGCM Conference. • Seong-Seung Kang, Katsuhiko Kaneko, Jun-Mo Kim, Yuzu Obara. Clarification of the regional and local in situ stresses using the CCBO technique and numerical analysis. Island Arc, vol.12, p.p.247-255, Blackwell Science (2003). • Sharrock, G, 2004. • Sjöberg, J and Klasson, H. Stress measurements in deep boreholes using the Borre (SSPB) probe. Special Issue of the IJRMMS: Rock Stress Estimation ISRM Suggested Methods and Associated Supporting Papers, Volume 40, Issues 7-8, Pages 955-1276 (October - December 2003). • Tatman C R, 2001. Production Rate Selection for Steeply Dipping Tabular Deposits, Mining Engineering October 2001, pp. 62-64
Santiago Chile, 22-25 August 2004
Massmin 2004
Production capacity of a mass caving José Pesce, Gerente General, Alfonso Ovalle, Consultor Asociado, Metálica Consultores S.A.
Abstract This paper reviews the factors that affect the maximum production capacity of a mass caving operation (block caving, panel caving). After analysing the multiple planning, design and operating factors that traditionally are considered to affect the determination of the maximum possible production capacity, it is concluded that there are only two factors that have an influence: column height and caving rate. Surprisingly the extraction rate has no influence in determining the maximum production capacity of a mass caving. The definition of the column height and its feasible range are analysed, describing how to determine the minimum, maximum and optimal column heights. Regarding the undercutting rate, a discussion about the theoretical and practical aspects is presented. There is reference in the paper to practical examples.
1 FACTORS THAT DETERMINE THE MAXIMUM PRODUCTION CAPACITY OF A MASS CAVING OPERATION. The relevant factors normally considered in the planning, design and operation of a mass caving operation are presented in Table 1. Some of these factors are basically to define the applicability of the mass caving mining method, such as cavability, primary fragmentation, stability of mine openings, dilution, repairing of openings, water and mud inflow and the undercutting sequence. If any of these factors is unmanageable, the applicability of the method will be at stake. It can be argued that the primary fragmentation can be solved inducing the caving, which could be done if the evaluation parameters of the project justify it. This analysis supposes that the mentioned factors fall in the range that allow that applicability of the mass caving method to the ore body. The interest is to focus on those factors that affect the determination of the maximum capacity of the ore body under study. The authors, through their experience, have reduced the factors of Table 1, to only three basic independent factors affecting the maximum capacity of a mass caving: undercutting rate or development velocity (measured in square meters to undercut per year), extraction rate or draw down rate (measured as mm/day), and the panel or column height (measured in m). Therefore, the interrelation between these three factors affecting the maximum capacity of a mass caving will be analysed. 2 DISCUSSION OF COLUMN HEIGHT. The column height of a massive caving operation is firstly limited by the geometry of the ore body to exploit. There are caving operations with column heights ranging from 50 to 600 m. For very high ore bodies the definition of the column heights to select should follow the following criteria, of course with due consideration to the business aspects of the exploitation searching to maximize profits or returns: • The column height must be able provide a minimum profit. The most frequent incentives to choose a low column are: accelerate the start up time of the project (minimize the Massmin 2004
start up capital cost); restrict the exploitation to a high grade sector (improve the cash flow); adaptation to existing infrastructure (less capital expenditure), as for example the existence of a main transport level. • The selected column height must be the maximum which is compatible with technical criteria. The incentives to choose big heights are to minimize development costs and to maximize the production capacity. • The optimum column height is determined by strategic an economic criteria. Shown in figure 1 are the form of the cost curves depending on column heights, some of which increase with column heights and some of which decrease with column heights. There exists therefore, a column height or a range of column heights, with a minimum net present value of the cost. • There are four technical factors whose relationship with column height needs to be underlined: stability of openings, useful life of drawpoints, dilution, and the risk of loss of ore. o The stability of openings, especially those of the production level (which are the greater part of the permanent openings of the exploitation method) depend mostly on the type of rock on which they are on, of the stress environment and on the design of the fortification. This is one aspect that must be taken into account when considering column height. The instability of a sector can cause the loss of productive areas that can seriously affect the fulfilment of the production program. o The useful life of drawpoints is a function of the design and the construction quality of them, of the secondary blasting and of the abrasion caused by ore flow. In general, the draw point repair cost is increasing with column height. o The dilution has a behaviour related to column height, to extraction management and to the number of faces exposed to diluents. It is possible that with proper extraction management, dilution will be decreasing with increasing column height. o The risk of loosing ore because of broken and lost ground is an increasing function of column height. This depends on the extraction grid selected, on the particular geology, on the correct undercutting sequence and on the regularity of draw.
Santiago Chile, 22-25 August 2004
75
Table 1 RELEVANT FACTORS AFFECTING A MASSIVE UNDERCUTTING: DEPENDENCY ON BASIC FACTORS Overall Factors
Specific Factors
Basic Factors Undercutting rate
Design and construction
Global geometry
Interference with other productive sectors
_
Base area geometry
_
Extraction Column rate height
_
Column height
Design
Nature
Geotechnics
Geology Operations
Modality of operation
_
Overall extraction sequence
_
Quantity of independent fronts
_
Extraction grid
_
Desing of operating modulus
_
Materials handling system
_
Modality of undercut
_
Undercutting sequence
_
Primary /seconday fragmentación
_
Stability of openings and constructions
_
In-situ and induced stresses
_
Water and mud
_
Characteristics of dilutant material
_
Unit processes
_
_
Availability of resources
_
_
Utilization of resources
_
_
Secondary fragmentation
3 DISCUSSION OF THE UNDERCUT RATE The undercut rate or the velocity to incorporate new production area depends on various factors. For example, there is a great difference in the undercutting rate between the systems called "previous undercut", "post undercut" or the intermediate situation named "advanced undercut". This is due to the space and access restrictions of the previous undercut system, which limits the rate of development and constructions, basically due to the coordination difficulties of the different unit operations that compose the undercutting. In a normal or "post undercut system" it is possible to achieve an undercutting rate of around 36.000 m2/yr in one face, while in a system of "previous undercut", for the same situation, it is difficult to surpass 24.000 m2/yr.
_
For both bodies, two undercutting rates are considered: 20.000 m2/yr and 36.000 m2/yr. Also, two extraction rates are considered: 150 mm/day and 300 mm/day. These examples have a referential specific gravity of 2,6 t/m3.
The undercut rate depends ultimately on the following factors: • Design of the undercut system • Equipment availability to build that design • Work organization • Operations and sequence restrictions 4 PRACTICAL EXAMPLES OF PRODUCTION CAPACITY A simulation of the production capacity of a mass caving operation is presented in figure 2, based in a mass balance between the amount of material incorporated by undercutting and the amount of material consumed by production. Two ore bodies are considered, each with the same base area of 300.000 m2, one with a 150 m high column and the other with a 300 m high column. 76
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Massmin 2004
Figure 2 presents the production capacity for the 150 m high column ore body, with an undercutting rate of 20.000 m2/yr and with an extraction rate of 150 mm/day (120 mm/day with 80% area utilization). The maximum capacity is slightly more than 21.000 tpd, which is attained at the 4th year of production, with a life of 12 years. The steady state total active area is 62.000 m2. Table 2 summarizes the results for the 150 m and 300 m high column bodies, for all possible situations.
5 EXTRACTION RATE, RAMP UP TIME TO STADY STATE AND ACTIVE AREA It can be seen in table 2 as pointed by the arrows that the extraction rate, which is intuitively thought of as defining the maximum capacity, plays no role in this definition. It only matters in how fast the steady state is reached, but once it is reached, a fast extraction rate only means a fast depletion rate of the columns. If the column depletion rate is faster
Table 2: Production capacities of examples Area to undercut m2
Column height
Extraction rate
Undercut rate
Active area
Production capacity
Ramp up
m
mm/day
m2/yr
m2
tpd
yrs
150
300.000 300
Massmin 2004
150
20.000
61.000
21.000
4
300
20.000
31.000
21.000
3
150
36.000
120.000
38.000
4
300
36.000
56.000
38.000
3
150
20.000
120.000
42.000
7
300
20.000
60.000
42.000
4
150
36.000
220.000
76.000
7
300
36.000
110.000
76.000
4
Santiago Chile, 22-25 August 2004
77
that the undercutting rate, the production capacity will decrease. If the undercutting rate is faster than the depletion rate, the production capacity can then grow. But the determinant factor is the undercutting rate and not the extraction rate. As seen in the summarize of table 2, for a given column height and a given extraction rate, a higher extraction rate means less active area and a low extraction rate means more active area, but in both cases the maximum attainable capacity of the ore body is the same. The amount of active area to keep has two main motivations. On one side, with the least possible active area, the operations will be more concentrated and there will be lower costs in items such as ventilation, supervision, maintenance of openings. This is true up to a certain extent, because exaggerated operations concentration increases the operational interferences and therefore the operating cost, influencing even the extraction rate in a negative way. On the other hand, it is of interest to keep a large active area to minimize the risks of not complying with the production program in case of unforeseen situations that may cause large areas to go out of production. There is a compromise situation in between these two situations, to keep costs down and to keep risks down. Undoubtedly that in the short time, a variation of the extraction rate will vary the production capacity of the active area, but these are temporary effects that cannot be maintained on the long run. If in the short range, there is an increase of the extraction rate, the production capacity will momentarily increase, and this situation can be maintained for a time lapse smaller than the average life of the columns, after which the production rate will decrease to values even lower than what was being produced before the disturbance. The only way to avoid this drop in production is increasing the undercutting rate before the columns whose extraction rates were increased, are depleted. The previous discussion indicates that planning the production capacity for a given sector or project, must necessarily be complemented with strategic/operational considerations as have been mentioned.
The extraction rate has no influence in the determination of the of the maximum capacity of a mass caving, only affecting the ramp up period to achieve steady state, and once this is attained, it only determines the amount of active area available for production. An effort to increase the extraction rate past certain limits, besides the geomechanical and dilution risks that this may cause, will only decrease the total active area available for production, and will increase the risk of fulfilling the production programs. Of the two independent factors that influence the maximum production capacity of a mass caving, the column height is fixed as an initial design parameter, with an economic criteria that considers the technical restrictions imposed by the ore body, and once the floor of the production level is fixed, there is little room for change in the column height, except for changes in the cut of grade of the draw points, that will only have a marginal effect on the column height. It is the other factor that can be handled easier, the undercutting rate, which can be increased with an industrial engineering analysis orientated to the optimization of development operations. REFERENCES • Alvarez, Ricardo; Pesce, José. 1987. Blocs foudroyés. Le comportement du tirage et le phénomène de la dilution. Un modèle de prédiction. Diplome d’études approfondies, Génie Geologique et Minier, Ecole de Mines de Nancy, France. pp-54 • Araneda, Octavio. 2002. Private communications. • Laubscher, D.H. 1994. Cave mining: state of the art. Journal of the South African Institute of Mning and Metallurgy, Oct. 1994, pp 279-292 • Ovalle, Alfonso; Codoceo, Javier. Sept. 1977. Factores que inciden en la productividad de un bloque en la mina El Teniente. Revista Minerales, Instituto de Ingenieros de Minas de Chile, pp 5-29, Santiago, Chile. • Ovalle, Alfonso. Nov. 2001. Programa de simulación de capacidad productiva en panel caving. Excel.
6 CONCLUSIONS The maximum production capacity of a massive caving depends only on the column height and on the undercutting rate for a give ore body.
78
Santiago Chile, 22-25 August 2004
Massmin 2004
Production capacity of a mass caving José Pesce, Gerente General, Alfonso Ovalle, Consultor Asociado, Metálica Consultores S.A.
Abstract This paper reviews the factors that affect the maximum production capacity of a mass caving operation (block caving, panel caving). After analysing the multiple planning, design and operating factors that traditionally are considered to affect the determination of the maximum possible production capacity, it is concluded that there are only two factors that have an influence: column height and caving rate. Surprisingly the extraction rate has no influence in determining the maximum production capacity of a mass caving. The definition of the column height and its feasible range are analysed, describing how to determine the minimum, maximum and optimal column heights. Regarding the undercutting rate, a discussion about the theoretical and practical aspects is presented. There is reference in the paper to practical examples.
1 FACTORS THAT DETERMINE THE MAXIMUM PRODUCTION CAPACITY OF A MASS CAVING OPERATION. The relevant factors normally considered in the planning, design and operation of a mass caving operation are presented in Table 1. Some of these factors are basically to define the applicability of the mass caving mining method, such as cavability, primary fragmentation, stability of mine openings, dilution, repairing of openings, water and mud inflow and the undercutting sequence. If any of these factors is unmanageable, the applicability of the method will be at stake. It can be argued that the primary fragmentation can be solved inducing the caving, which could be done if the evaluation parameters of the project justify it. This analysis supposes that the mentioned factors fall in the range that allow that applicability of the mass caving method to the ore body. The interest is to focus on those factors that affect the determination of the maximum capacity of the ore body under study. The authors, through their experience, have reduced the factors of Table 1, to only three basic independent factors affecting the maximum capacity of a mass caving: undercutting rate or development velocity (measured in square meters to undercut per year), extraction rate or draw down rate (measured as mm/day), and the panel or column height (measured in m). Therefore, the interrelation between these three factors affecting the maximum capacity of a mass caving will be analysed. 2 DISCUSSION OF COLUMN HEIGHT. The column height of a massive caving operation is firstly limited by the geometry of the ore body to exploit. There are caving operations with column heights ranging from 50 to 600 m. For very high ore bodies the definition of the column heights to select should follow the following criteria, of course with due consideration to the business aspects of the exploitation searching to maximize profits or returns: • The column height must be able provide a minimum profit. The most frequent incentives to choose a low column are: accelerate the start up time of the project (minimize the Massmin 2004
start up capital cost); restrict the exploitation to a high grade sector (improve the cash flow); adaptation to existing infrastructure (less capital expenditure), as for example the existence of a main transport level. • The selected column height must be the maximum which is compatible with technical criteria. The incentives to choose big heights are to minimize development costs and to maximize the production capacity. • The optimum column height is determined by strategic an economic criteria. Shown in figure 1 are the form of the cost curves depending on column heights, some of which increase with column heights and some of which decrease with column heights. There exists therefore, a column height or a range of column heights, with a minimum net present value of the cost. • There are four technical factors whose relationship with column height needs to be underlined: stability of openings, useful life of drawpoints, dilution, and the risk of loss of ore. o The stability of openings, especially those of the production level (which are the greater part of the permanent openings of the exploitation method) depend mostly on the type of rock on which they are on, of the stress environment and on the design of the fortification. This is one aspect that must be taken into account when considering column height. The instability of a sector can cause the loss of productive areas that can seriously affect the fulfilment of the production program. o The useful life of drawpoints is a function of the design and the construction quality of them, of the secondary blasting and of the abrasion caused by ore flow. In general, the draw point repair cost is increasing with column height. o The dilution has a behaviour related to column height, to extraction management and to the number of faces exposed to diluents. It is possible that with proper extraction management, dilution will be decreasing with increasing column height. o The risk of loosing ore because of broken and lost ground is an increasing function of column height. This depends on the extraction grid selected, on the particular geology, on the correct undercutting sequence and on the regularity of draw.
Santiago Chile, 22-25 August 2004
75
Table 1 RELEVANT FACTORS AFFECTING A MASSIVE UNDERCUTTING: DEPENDENCY ON BASIC FACTORS Overall Factors
Specific Factors
Basic Factors Undercutting rate
Design and construction
Global geometry
Interference with other productive sectors
_
Base area geometry
_
Extraction Column rate height
_
Column height
Design
Nature
Geotechnics
Geology Operations
Modality of operation
_
Overall extraction sequence
_
Quantity of independent fronts
_
Extraction grid
_
Desing of operating modulus
_
Materials handling system
_
Modality of undercut
_
Undercutting sequence
_
Primary /seconday fragmentación
_
Stability of openings and constructions
_
In-situ and induced stresses
_
Water and mud
_
Characteristics of dilutant material
_
Unit processes
_
_
Availability of resources
_
_
Utilization of resources
_
_
Secondary fragmentation
3 DISCUSSION OF THE UNDERCUT RATE The undercut rate or the velocity to incorporate new production area depends on various factors. For example, there is a great difference in the undercutting rate between the systems called "previous undercut", "post undercut" or the intermediate situation named "advanced undercut". This is due to the space and access restrictions of the previous undercut system, which limits the rate of development and constructions, basically due to the coordination difficulties of the different unit operations that compose the undercutting. In a normal or "post undercut system" it is possible to achieve an undercutting rate of around 36.000 m2/yr in one face, while in a system of "previous undercut", for the same situation, it is difficult to surpass 24.000 m2/yr.
_
For both bodies, two undercutting rates are considered: 20.000 m2/yr and 36.000 m2/yr. Also, two extraction rates are considered: 150 mm/day and 300 mm/day. These examples have a referential specific gravity of 2,6 t/m3.
The undercut rate depends ultimately on the following factors: • Design of the undercut system • Equipment availability to build that design • Work organization • Operations and sequence restrictions 4 PRACTICAL EXAMPLES OF PRODUCTION CAPACITY A simulation of the production capacity of a mass caving operation is presented in figure 2, based in a mass balance between the amount of material incorporated by undercutting and the amount of material consumed by production. Two ore bodies are considered, each with the same base area of 300.000 m2, one with a 150 m high column and the other with a 300 m high column. 76
Santiago Chile, 22-25 August 2004
Massmin 2004
Figure 2 presents the production capacity for the 150 m high column ore body, with an undercutting rate of 20.000 m2/yr and with an extraction rate of 150 mm/day (120 mm/day with 80% area utilization). The maximum capacity is slightly more than 21.000 tpd, which is attained at the 4th year of production, with a life of 12 years. The steady state total active area is 62.000 m2. Table 2 summarizes the results for the 150 m and 300 m high column bodies, for all possible situations.
5 EXTRACTION RATE, RAMP UP TIME TO STADY STATE AND ACTIVE AREA It can be seen in table 2 as pointed by the arrows that the extraction rate, which is intuitively thought of as defining the maximum capacity, plays no role in this definition. It only matters in how fast the steady state is reached, but once it is reached, a fast extraction rate only means a fast depletion rate of the columns. If the column depletion rate is faster
Table 2: Production capacities of examples Area to undercut m2
Column height
Extraction rate
Undercut rate
Active area
Production capacity
Ramp up
m
mm/day
m2/yr
m2
tpd
yrs
150
300.000 300
Massmin 2004
150
20.000
61.000
21.000
4
300
20.000
31.000
21.000
3
150
36.000
120.000
38.000
4
300
36.000
56.000
38.000
3
150
20.000
120.000
42.000
7
300
20.000
60.000
42.000
4
150
36.000
220.000
76.000
7
300
36.000
110.000
76.000
4
Santiago Chile, 22-25 August 2004
77
that the undercutting rate, the production capacity will decrease. If the undercutting rate is faster than the depletion rate, the production capacity can then grow. But the determinant factor is the undercutting rate and not the extraction rate. As seen in the summarize of table 2, for a given column height and a given extraction rate, a higher extraction rate means less active area and a low extraction rate means more active area, but in both cases the maximum attainable capacity of the ore body is the same. The amount of active area to keep has two main motivations. On one side, with the least possible active area, the operations will be more concentrated and there will be lower costs in items such as ventilation, supervision, maintenance of openings. This is true up to a certain extent, because exaggerated operations concentration increases the operational interferences and therefore the operating cost, influencing even the extraction rate in a negative way. On the other hand, it is of interest to keep a large active area to minimize the risks of not complying with the production program in case of unforeseen situations that may cause large areas to go out of production. There is a compromise situation in between these two situations, to keep costs down and to keep risks down. Undoubtedly that in the short time, a variation of the extraction rate will vary the production capacity of the active area, but these are temporary effects that cannot be maintained on the long run. If in the short range, there is an increase of the extraction rate, the production capacity will momentarily increase, and this situation can be maintained for a time lapse smaller than the average life of the columns, after which the production rate will decrease to values even lower than what was being produced before the disturbance. The only way to avoid this drop in production is increasing the undercutting rate before the columns whose extraction rates were increased, are depleted. The previous discussion indicates that planning the production capacity for a given sector or project, must necessarily be complemented with strategic/operational considerations as have been mentioned.
The extraction rate has no influence in the determination of the of the maximum capacity of a mass caving, only affecting the ramp up period to achieve steady state, and once this is attained, it only determines the amount of active area available for production. An effort to increase the extraction rate past certain limits, besides the geomechanical and dilution risks that this may cause, will only decrease the total active area available for production, and will increase the risk of fulfilling the production programs. Of the two independent factors that influence the maximum production capacity of a mass caving, the column height is fixed as an initial design parameter, with an economic criteria that considers the technical restrictions imposed by the ore body, and once the floor of the production level is fixed, there is little room for change in the column height, except for changes in the cut of grade of the draw points, that will only have a marginal effect on the column height. It is the other factor that can be handled easier, the undercutting rate, which can be increased with an industrial engineering analysis orientated to the optimization of development operations. REFERENCES • Alvarez, Ricardo; Pesce, José. 1987. Blocs foudroyés. Le comportement du tirage et le phénomène de la dilution. Un modèle de prédiction. Diplome d’études approfondies, Génie Geologique et Minier, Ecole de Mines de Nancy, France. pp-54 • Araneda, Octavio. 2002. Private communications. • Laubscher, D.H. 1994. Cave mining: state of the art. Journal of the South African Institute of Mning and Metallurgy, Oct. 1994, pp 279-292 • Ovalle, Alfonso; Codoceo, Javier. Sept. 1977. Factores que inciden en la productividad de un bloque en la mina El Teniente. Revista Minerales, Instituto de Ingenieros de Minas de Chile, pp 5-29, Santiago, Chile. • Ovalle, Alfonso. Nov. 2001. Programa de simulación de capacidad productiva en panel caving. Excel.
6 CONCLUSIONS The maximum production capacity of a massive caving depends only on the column height and on the undercutting rate for a give ore body.
78
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Massmin 2004
Continuous mining for caving method Francisco Carrasco J., Víctor Encina M., René Le-Féaux C. Instituto de Innovación en Minería y Metalurgia S.A., IM2, Fernando Geister B., Dirección Investigación y Desarrollo, Codelco Chile
Abstract Continuous mining system is an alternative mine design to perform block caving method. This paper describes the whole configuration of the system and its generic parameters, showing technical and economical convenience of its application. The system is based on simultaneous ore extraction from all active draw points. To do that, it is proposed to use a set of stationary extractor devices in each drawpoint. Extractors feed a conveying system trough transfer points, the conveyor system hauls the ore to an underground crusher or sizer after which a conventional belt conveyor system transports the crushed material to surface. The proposed system is able to get high rates of extraction, allowing the owner to set great production capacity even if the extension of the footprint is no so large. Cost of production will be less than conventional LHD system; full automated and remote operation can be done from a control room based on standard central command technology of stationary equipment in industrial plants.
1 INTRODUCTION Since 1998, Codelco Chile has been developing its Technological Programs for Underground Mining with its Institute for Innovation in Mining and Metallurgy, IM2, which has introduced the Continuous Mining concept as one commanding objective, aimed to surpass the limit of the dominant critical technological dimension of underground mining methods, which in present time tends to be less than 0.5 t/m2-day when mining hard rock by caving mining method. After six years, those research projects have given some results and in present time Codelco-Chile is running a specific project to validate the new Continuous Mining for Block Caving Method in an industrial scale. Nowadays Continuous Mining concept is much more pertinent, as we consider that long term plans of CodelcoChile, are demanding 550 ktpd of underground production for next decade, what means more than 3 times than current underground mines production. Such a challenge cannot be face following continuous improvements, but it requires a technological break out, especially if we see that those mines will be located in low grade, harder rock and in deeper stressed zones. The strategy followed was to think the process first and after that, to demand the technologies required to achieve the requested output. That means work in a process driven innovation instead of a technological one. In such a strategy, it was set that the new mining method will be based on the following concepts: • Method will be not adapted to rock characteristics, but rock will be conditioned to best apply block caving method. • Draw will be done simultaneously and continually, in time and in ore flow, from all active drawpoints. • Mine has to be operated remotely and fully assisted by automated tasks. The target under those concepts are to be able to get a greater rate of extraction and to reduce the mining costs while assures a safe and reliable production process independently of ore body rock and ore body location. This paper describes the fundamentals of Continuous Massmin 2004
Mining for Block Caving Method, its whole configuration and its generic parameters to show its technical and economical convenience. 2 FUNDAMENTALS Mining process is composed by only two sub processes: Fragmentation and Transportation. Fragmentation is the process by means the rock in place is transformed in broken ore and transportation consist on remove those fragments and transfer them to a mineral processing plant. Fragmentation can be done by blasting or caving, and different material handling systems can be used to transport the ore depending of the size distribution of bulk material. In underground mining of large ore bodies, the most economical mining method applied is the one called Block Caving, consisting in get the fragmentation by caving and extract the ore, like "milking" it, from a drawpoints infrastructure arranged below the caved material. Initially this method was developed to mine ore bodies whose rock was well fractured. As its costs are very attractive, nowadays it is also applied in rock without or with very few open discontinuities, what we call "hard rock", where under stress conditions, fracturing can be induced during the caving process. Main aspects limiting block caving application in hard rock under stress conditions are, large size of material resulting of caving and seismic effect of caving propagation. Seismic events caused by caving propagation may eventually cause damage on infrastructure or equipment or even personnel injuries. Coarse material had push the designer to introduce large loaders and secondary reduction equipment underground with the consequence of enlarging drawpoint spacing and drifts size and to have a batch process extracting the ore from few drawpoints at a time. Although many improvements have been introduced to maintain acceptable safety and effectiveness as, large size material handling equipment, mine design
Santiago Chile, 22-25 August 2004
79
adjustments and some production restrictions to avoid seismic effects, mines productivity have became stuck and limited to a mean rate of extraction less than 0.5 t/m2-day. This stagnation can be due to improvement have been oriented not to causes but to consequences of caving in hard rock. That is, assuming as an unavoidable fact that broken ore have to bring large pieces so, extraction process had to be adapted to that condition. When IM2 looked for new mining process able to increase underground production, they realize they could choose to continue with block caving if they look for complementary technologies to aid its sub process to be more effective. That means to have a sustainable caving process producing well sized broken ore, to assure expeditious confined gravity flow to drawpoints and to define a continuous material handling process. These aiding technologies are based on following concepts:
Continuous and simultaneous extraction cannot be practiced using loaders, but it requires having special kind of feeders to be installed on the drawpoints and a continuous conveying system. Both feeders and conveyor are now under designing and testing process taking care of following conditions: • Material flow has to occur only when the feeder is on and stops when it is off. • In case of blasting hangups, as it is expected to occur, the feeder has not to be damaged. • Feeder and conveyor design has to allow performing maintenance and repairing tasks free at all of any risk to workers caused by material flow. • Continuous conveying system should be able to be loaded simultaneously at different points along the conveyor and should be able to stop and start up fully loaded. • Both feeders and conveyor should be operated in remote and automated mode.
2.1 Preconditioning As natural caving was not producing a suitable product when applied on not well fractured hard rock, it was define a target in looking for a treatment of ore mass before caving with the purpose of: • To assure caving propagation: It is known that the more fractured is the ore mass, the easier and faster is the caving propagation process. • To improve the fragmentation: As in caving, fragments are finally formed trough fractures, the more fractured the ore mass, the smaller the size of fragments, and smaller the size of larger pieces.
Then, by doing both continuous and simultaneous extraction, a combined effect of easier and greater drawing is expected to be achieved from the same active area. That effect results in a greater rate of extraction and in a greater utilization of time, in such a way that no stops of production will be need for changing operation shifts.
Then, an important work was done in developing preconditioning technology adapting hidrofracturing petroleum industry experience to mining, as part of the research program of second stage of International Caving Study, and in other internal research about confined blasting method, both to be applied on rock mass before caving. This is not the paper to give a detailed description of those preconditioning technologies, but we can say that great progress has been obtained in experimental applications performed at Andina and Salvador mines. So for the purpose of designing the new mining process, preconditioning is an enabler technology, as it is the base to assure the ore will be caved in a reasonable time and the product of caving will be well fragmented for allowing an expeditious flow of broken material trough drawpoints. 2.2 Continuous and Simultaneous Drawing The main requirement to this block caving reengineering is to be able to increase the production coming from same or similar extension of footprint mines. The target defined for ore extraction was to design a system of continuous and simultaneous ore drawing from all active drawpoints with the following purpose: • To increase the rate of extraction: Drawpoints utilization when using batch LHD system is very low, so a way to increase the rate of extraction it to increase such utilization by drawing them simultaneously. An important increase of rate of extraction will result of simultaneous extraction even extracting much less than actually does LHD equipment when loading from a drawpoint. • To assure the ore flow: Gravity flow can be seen as a succession of loosening effect due to the extraction of some material from the drawpoint. Based on experience on gravity flow, it is easier to draw material when it is loose than when it is settled or compacted, so continuous drawing will aid to maintain a loosened status of broken ore favoring ore flow. 80
2.3 Automation and Remote Control Considering that this new method will be applied to new projects during the next decade, when labor requirements will be more exigent. Then, the target was to look for a system of work that could assure a better comfort and life quality for workers than current standards, avoiding as much as possible, the personnel working on night shifts and living far from family. It was also looked for a centralized command of feeders as a mean of draw control, for driving depletion strategies on line and in real time. Using stationary equipments, it is easy to have a centralized command and control room governing the process as a "rock factory", based on digital instrumentation and information technology as commonly used in modern industry. This command room could be located very far from the mine, for example in a well developed city or even in a different country. Undoubtedly, this vision of the mine as a rock factory, remotely commanded will contribute to improve mining work condition and workers perception. So better possibilities will be open to hire and keep personnel for mine development and that mentioned day shift of production resetting and maintenance. 3 MINE DESIGN The following is a brief description of a Continuous Mine design. As previously mentioned, preconditioning is the technology that traces the path to continuous mining system described as follow. The ore body preconditioning take place from undercut level, previous to the operation and using technologies like hydraulic fragmentation and massive blasting without a free face. The undercut for caving can be done in any standard way as it is done in every block caving mine of Codelco-Chile. This operation marks the beginning of the extraction from the drawpoints. As it said before, each drawpoint is equipped with a stationary feeder to extract the ore from the drawpoints for feeding a continuous conveyor. Continuous and simultaneous operation is required to increase the rate of extraction. In fact, considering a very conservative case, where the feeder has one half of the
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extraction capacity of LHD equipment (6 t) and if it takes twice the time to produce it, feeder productivity will be approximately 46 t/h per drawpoint. Then the combined effect of all drawpoint working simultaneously at that productivity gives a rate of extraction of approximately 3 t/m2-d that is, 6 to 8 times the current rates. Conveying system and an appropriate arrangement of ore passes are to be set to transfer the ore to a crusher for reducing the size of ore up to be compatible with belt conveyor transporting, which has to convey the ore to the mineral processing plant. An isometric sketch of the whole system is shown as follow:
• For Conventional LHD system propagation rate of extraction is 0.40 t/m2-d but in Continuous Mining with preconditioning it will increase to 0.6 t/m2-d. After 30% of ore extraction, no restriction in extraction rate is assumed except material handling capacity. • Material handling capacity for Conventional LHD System is 0.7 t/m2-d and for Continuous Mining is 2.8 t/m2-d. • Ore body footprint is as follows: - Width: 270 m. - Length: 800 m. - Panel High: 200 m. • Two fronts of 270 m long (equal to footprint width) production can be run from central part to extremes of the ore body. For Conventional LHD System, production in one front is limited by the amount of production drifts, in this case 9 (30 m spaced) and the extension of a drift is defined by some optimum distance of LHD hauling, usually 120 m. So the maximum area in full production in such a front for Conventional LHD System would be 32,400 m2 (270 m x 120 m) plus other 21,600 m2 in propagation giving a total of 54,000 m2 of active area per front. On the other hand, for Continuous Mining System, whose production module is composed by 36 drawpoints of 225 m2 each, one module is 90 x 90 m. In order to balance the production in a stable output, it is required to have 2 modules in propagation per each module in full production. So we can say that production in CM will come from blocks composed by 3 modules (1 full + 2 in propagation) giving block dimensions of 270 m long by 90 m width. Therefore, for the same front, it is possible to operate with 3 blocks giving a total active area of 72,900 m2. On Table 1, it is presented a comparison of technical performance of both systems applied to the same front of production:
Figure 1: Isometric view unit MC (1): UCL; (2): Drawbell; (3): Service Level (drawpoints); (4): Production Level (secondary transport); (5): ore pass; (6): Reduction Level; (7): Crusher equipment; (8): Ore pass to main conveyor. In this scheme, unhangup process and secondary blasting will be done with conventional technology or with new technologies nowadays under development. The material handling configuration, based in automated stationary equipments, allows an expert control system for commanding the system functioning in each one of its stages. Continuous Mining is thought to operate continuously during 18 hours, since 3 PM to 9 AM in the next day, letting a day shift of 6 hours per day to maintenance, repairing and ore flow resetting. If any drawpoint get hanged up during the operation time, it will remains in that status till next day shift when production crew go into the mine to reset the flow, the same if a feeder or other equipment brakes down. Development and preparing of new area can operate all day in 4 shifts of 6 hours, for assuring opportune replacement of exhausted area. 4 TECHNICAL ECONOMICS INDEXS To emphasize the impact of production performance, it is hypothetical example with following • Drawpoint spacing arranged in 15x15 m, corresponding to of drawpoint. Massmin 2004
the new method on presented ahead a design criteria: an equilateral grid of 225 m2 of area per
Table 1: Technical performance CM Conventional Continuous LHD mining mining
Extraction rate Propagation Full Mean Production Active area width length Active drawpoint Ramp up time
t/m2-d t/m2-d t/m2-d tpd m2 m m unit year
current
future
0.40 0.70 0.57 30,700 54,000 270 200 240 3.8
0.60 2.80 1.33 97,000 72,900 270 270 324 1.6
A Table 1 result suggests a great potential to increase the production of Codelco-Chile’s underground mining due to: • It can be seen an important increase in production per front is got with CM without a proportional increase of active area. • Table 1 refers to only one front of production, if more fronts are available, CM has a potential of getting much more production than any calculation done applying Conventional LHD system. • Also it can be seen an important decrease of time required to get full production (ramp up) in CM System
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compared to Conventional LHD System, This will have a great effect in NPV when applied in a new mine. All these result leads the great conclusion of CM will give a significant contribution to the business when applying this new mining concept in new mining projects. 5 CONCLUSIONS A new approach for innovation following the strategy of thinking first the process, and after that, to demand the required technologies to achieve the requested output was applied to study mining process. That strategy has been very fruitful leading the researches of Underground Mining Technological Program started in 1998, trough unexpected paths and getting surprising results, that can be interpreted as a validation of the strategy of having a process driven innovation focus instead of a technological one. One product of Underground Mining Technological Program is Continuous Mining for Block Caving Method, which is a mine design and a mining concept able to face future challenge of Codelco Chile’s mines. Continuous Mining for Block Caving Method concept is based on three main fundaments: • Preconditioning as an enabler technology, to assure the caving propagation and an expeditious flow of broken material trough drawpoints. • Continuous and simultaneous drawing for increasing the rate of extraction and day time utilization. • Remote commanding of stationary equipment to have a draw control on line and in real time governing the process as a "rock factory". Although both preconditioning and continuous and simultaneous drawing are nowadays under its final steps of development, there are not fatal failures threatening the research program, so in the near future an industrial validation test of the whole system will be performed. ACKNOWLEDGEMENTS The authors are grateful to all their colleagues of CodelcoChile operations and IM2 that helped them during the development of this research and to ICS II team of
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researchers that have been involve in hidrofracturing for preconditioning. Also, the authors want to acknowledge the permission given by Codelco Chile to publish this technical paper. REFERENCES • Chacón, E., Quiñones, L., Gonzalez, J., Barrera, V., 2002. Pre-acondicionamiento de macizos rocosos competentes para la explotación por métodos de hundimiento, pp.1934. IIMCH magazine, Santiago, Chile Vol. 57, N° 245, may-june 2002. • Encina, V. and Correa, L., 2001. Minería Continua: Un quiebre tecnológico ad portas, in 52th IIMCH Convention, 7-10 November, La Serena, Chile. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2001. Técnicas de manejo de materiales en el nivel de producción, División Salvador, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2002. Tecnologías de transporte continuo, División Salvador, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Diseño Conceptual para Minería Continua, División Salvador, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Estudio de metodología de acondicionamiento de macizo rocoso para hundimiento, División Andina, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Estudio de minería alternativa en Tercer Panel, División Andina, Codelco Chile. Internal Report. • Maass, S, 2003. Análisis de competitividad de métodos de explotación de minas mediante indicadores tecnológicos, 118 p. Final Thesis to obtain Degree in Mining Engineer, Universidad de Chile. • Minería Chilena. Nº 275, May 2004, Santiago, Chile. • Pérez, V, 2001. La investigación e innovación tecnológica en la minería del cobre, 54 p. Cochilco research, Santiago, Chile. • Villarzú, J., 2003. Industria del Cobre: Los desafíos del Siglo XXI. Codelco Chile CEO presentation in Cooper2003, Santiago, Chile.
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Current practices and trends in cave mining Germán Flores, Codelco Norte Division, Codelco Chile Antonio Karzulovic, A. Karzulovic & Assoc. Ltd., Santiago, Chile Edwin T. Brown, Julius Kruttschnitt Mineral Research Centre, University of Queensland, Australia
Abstract Although many caving mines are now exploiting deeper and more massive orebodies, and several large open pits are planning a transition to underground cave mining, there is no single and convenient reference on current practices and trends in cave mining. Hence, as part of the International Caving Study Stage II (ICS-II) funded by nine major international mining companies, a comprehensive benchmarking study was undertaken. The goal of this study was to establish current practices and trends, with emphasis on the key areas that have large impacts on the economics of a mining project, geomechanics, mine design, mine operation, and geotechnical hazards. Data were collected and interviews conducted during visits made to 17 mines in Australia, Canada, Chile, Indonesia, South Africa, Sweden and the USA. A comprehensive review of the technical literature provided information on a further 88 mines. This paper outlines the current trends and practices in caving mining, and summarizes them in a table containing data that may be used during scoping and pre-feasibility studies of cave mining projects, including transitions from open pit to underground mining. 1 INTRODUCTION Table 1: Mines Visited Several mines are planning a transition from open pit to underground cave mining in the medium or long term. Accordingly, the International Caving Study Stage II (ICS-II), managed by the Julius Kruttschnitt Mineral Research Centre, Brisbane, Australia, included a task of providing practical geotechnical guidelines for the transition from open pit to underground cave mining. As part of this task, a benchmarking study was carried out to collect data from mines that have developed, or are planning to develop such transitions, and from other mines (open pits or underground) that could provide relevant data. The collected data were supplemented by a comprehensive review of the technical literature. This study which was planned and executed so as to optimize data collection, included the following elements: Survey Design: To facilitate data collection, an Excel© spreadsheet was designed, and e-mailed to the mines included in the study. Mine Visits: The mines listed in Table 1 were visited, and relevant information was collected. Additional Data Collection: A comprehensive survey of the technical literature was undertaken to collect supplementary data from 88 additional mines. Data Processing: The collected data were analysed in order to develop histograms and, where possible, correlations showing current practice and trends in underground mining by caving methods. When enough data were available the relative frequency of selected parameters was computed, and when the available data were limited, the relative importance of the parameters was assessed. Benchmarking Report: The data, conclusions and recommendations resulting from this study were presented in a Final Report submitted to the sponsors of ICS-II (Flores and Karzulovic, 2002). The results of this benchmarking study reflecting current practices and trends in cave mining, are summarized in this paper.
Country
Mine
Mining Method(s)
Australia
Cadia Hill
OP.
Mount Keith
OP considering a transition to UG.
Northparkes
Mine that developed a transition from OP to UG by BC.
Ridgeway
UG by SLC.
Canada
Kidd Creek
Mine that developed a transition from OP to UG by SOP.
Chile
Andina
OP and UG by BC/PC.
Chuquicamata
OP considering a transition to UG by PC.
El Teniente
UG by PC.
Salvador
UG by PC.
Grasberg
OP.
Indonesia
Grasberg DOZ
UG by PC.
Finsch
Mine that developed a transition from OP to UG by SOP.
Koffiefontein
Mine that developed a transition from OP to UG by SOP/SLC/FC.
Palabora
Mine developing a transition from OP to UG by PC.
Sweden
Kiruna
Mine that developed a transition from OP to UG by SLC.
USA
Bingham Canyon Henderson
OP considering a transition to UG by PC. UG by PC.
OP UG SLC / SOP BC / PC
Open pit mine/mining Underground mine/mining Sublevel caving / Sublevel open stoping Block caving / Panel caving
South Africa
2 GEOMECHANICS The analysis of the data on Geological Structures indicates that: Massmin 2004
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a. In most underground caving mines, subvertical structures (dips > 60o) predominate over subhorizontal structures (dips < 30o). b. In underground and open pit mines the data on the orientations of structures are typically much better than the data on their lengths, spacings, and apertures. c. The geotechnical characterization of structures generally is poorer for underground mines than for open pit mines. d. In spite of the increasing use of numerical models, the quality of input data on the mechanical properties of structures is, in most of cases, poor. The analysis of the data on Intact Rock and Rock Mass Properties indicates that: a. Usually the unit weight (g), and the uniaxial compressive strength (UCS) of the intact rock are known, but information on shear strength and deformability is much poorer. b. Typically UCS values are smaller for open pit mine rocks (average: 80 MPa) that for underground mine rocks (average: 130 MPa). c. Typically RQD values are smaller for open pit mine rocks (average: 60% to 70%) than for underground mine rocks (average: 70% to 85%). d. The most commonly used method of rock mass classification in the underground mines studied is Laubscher’s RMR (53%), followed by Barton’s Q (26%), and Bieniawski’s RMR (15%). e. Typical Laubscher’s RMR ranges found in different mining conditions are shown in Table 2.
Table 2: Typical Laubscher’s RMR Range Mining Method
Laubscher’s RMR Range
Open Pit
20 to 40
Open Stoping
40 to 80
Sublevel Caving
40 to 70
Block Caving
30 to 70
Panel Caving
40 to 80
(1)
g. The geotechnical characterization of rock masses seems to be poorer in underground than in open pit mines. Indeed, in spite of the increasing use of numerical models the quality of input data on rock mass properties is, in most cases, poor to fair. The analysis of the data on In Situ Stresses indicates that: a. Currently the CSIRO Hollow Inclusion Cell is the most commonly used method for in situ stress measurements. b. Typically in underground caving mines the in situ major principal stress, S1, varies from 30 to 40 MPa; and the in situ minimum principal stress, S3, varies from 10 to 20 MPa. c. In underground mines the mean value of the stress ratio, KMEAN, is bounded as proposed by Hoek and Brown (1980), where the depth z is in meters:
0.5 +
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1500 100 ≥ K MEAN ≥ 0.3 + z z
0.6 +
1250 100 ≥ K Min ≥ 0.2 + z z
1.0 +
1500 90 ≥ K Max ≥ 0.3 + z z
(3)
(4)
The analysis of the data on Hydrogeology indicates that: a. Most mines do not consider hydrogeological characterization to be a high priority. b. The most commonly used drainage systems are sumps (78%), subhorizontal drains (14%) and drainage tunnels (8%). c. The most typical monitoring systems are observation wells (open holes), piezometers and flow rate measurement devices. The analysis of the data on Geotechnical Software indicates that: a. For two-dimensional numerical analyses the most commonly used codes are FLAC (50%), UDEC (33%), and EXAMINE (10%). b. For three-dimensional numerical analyses the most commonly used codes are FLAC3D (44%), 3DEC (26%), and MAP3D (18%). 3 MINE DESIGN
f. The central trend relating Laubscher’s RMR and MRMR is: MRMR = 0,9 ↔ RMR
d. Similar relationships were derived for the minimum and maximum values of the stress ratio, KMIN and KMAX:
(2)
The analysis of the data on Mine Accesses indicates that: a. The use of shafts as the only access has decreased since 1970. b. The use of declines as the only access has increased since 1970. c. Before 1970 in 70% of the case surveyed, shafts were used as accesses and in 30% declines were used. d. In the period 1970 to 1990, in 46% of the cases shafts were used as accesses, in 42% declines were used, and in 13% both shafts and declines were used. e. After 1990, in 36% of the cases shafts were used as accesses, in 50% declines were used, and in 14% both shafts and declines, were used. The analysis of the data on Block Heights and Footprints indicates that: a. As shown in Figure 1, since 1970 block heights have increased in block and panel caving mines. Before 1970, the typical block height was 100 m; for the period 19701990 it was 160 m, and since 1990 it has been 240 m. b. In block and panel caving mines the footprint area varies widely, but in 80% of the cases, it is smaller than 250,000 m2 with an average of 165,000 m2. c. As shown in Figure 2, the ratio between footprint length (L) and width (B) rarely exceeds 3, and in almost 60% of the cases it is smaller than 2. Nevertheless, in many large open pit mines currently considering a transition to underground cave mining this ratio will be larger than 3. d. It seems that most block and panel caving mines have ignored a possible relationship between block height (H) and footprint geometry (defined by its width B). As a preliminary conclusion, and as shown in Figure 3, the data collected suggested that: If H/B ≤ 1 the cave will easily connect to surface (or to an upper mined out level).
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a. As shown in Figure 4, the shape of the initial area for caving is predominantly square or rectangular, but in a few cases other shapes (e.g. triangular) have been used. b. As shown in Figure 4, the available data indicates that the area for caving initiation has an average value of 10,000 m2, and typically varies from 5,000 to 15,000 m2.
Figure 1: Evolution with time of block height in block and panel caving mines.
Figure 4: Relative frequency of initial caving areas and their shapes in block and panel caving mines. c. As shown in Figure 5, the hydraulic radius, HR, of the initial caving area varies from 15 to 45 m, with an average of 20 to 30 m.
Figure 2: Values of the footprint ratio L/B in block and panel caving mines.
Figure 5: Relative frequency of hydraulic radii of the initial caving area in block and panel caving mines. d. In 60% of the cases some measures were taken to facilitate cave initiation (slots: 53%, artificial chimneys: 7%). The analysis of the data on the Undercut Level, UCL, indicates that:
Figure 3: Relation between H and B in block and panel caving mines.
If 2 ≥ H/B > 1 the cave probably will connect to surface (or to an upper mined out level). If H/B > 2 the cave could have problems to connect to surface (or to an upper mined out level). The analysis of the data on Caving Initiation indicates that: Massmin 2004
a. The distance between UCL drifts varies from 10 to 35 m, with an average of 20 to 25 m. b. The width of UCL drifts has increased with time. Typically, before 1970 it was 2 to 3 m, during 1970-1990 it was 3 m, and since 1990 it has been 4 m. c. The height of UCL drifts has increased with time. Typically, before 1970 it was 2 to 2.5 m, during 1970-1990 it was 3 to 3.5 m, and since 1990 it has been 3.5 to 4 m. d. The undercut height shows no time-dependent trend. It varies from 3 to 20 m, with an average of 8 to 12 m. e. As shown in Figure 6, the average undercutting rate, RUC, varies from 500 to 5,000 m2/month, with an average of 2,000 to 3,000 m2/month.
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The analysis of the data on Support indicates that: a. In most underground caving mines support on the UCL consists of only bolts. In some mines this support also includes mesh and shotcrete in the UCL access drifts. b. In most underground caving mines support on the EXL includes bolts (typically 1.8 to 2.4 m long, at spacings of 1.0 to 1.3 m), mesh and shotcrete (typically 50 mm). In many cases it also includes cables, typically 5 to 8 m long, at intersections). Also some mines use straps and osro-straps, as illustrated in Figure 8. c. Bolt lengths range from 1.2 to 3.8 m, with typical values of 2 to 2.3 m on the UCL, and 2 to 2.5 m on the EXL. d. Bolt spacings range from 0.6 to 1.4 m, being typically 1.0 m for both the UCL and the EXL. e. There is no clear difference in the ratio of the bolt length, L, to the drift width, W, used on the UCL and EXL. In most cases the ratio L/W ranges from 1.5 to 3. Figure 6: Relative frequencies of average undercutting rates, RUC, in block and panel caving mines.
The analysis of the data on the Extraction Level, EXL, indicates that: a. The crown-pillar thickness (from the EXL floor to the UCL floor) shows an increasing trend with time. Typically, before 1970 it was 7.5 to 10 m, from 1970-1990 it was 12.5 m, and since 1990 it has been 15 to 17.5 m. b. The spacing between EXL drifts shows an increasing trend with time. Typically, before 1970 it was 12 to 16 m, from 1970-1990 it was 20 to 24 m, and since 1990 it has been 26 to 28 m. c. The width of EXL drifts shows an increasing trend with time. Typically, before 1970 it was 2.5 m, from 1970-1990 it was 3 to 3.5 m, and since 1990 it has been 4 to 4.5 m. d. The height of EXL drifts shows an increasing trend with time. Typically, before 1970 it was 2 to 2.5 m, from 19701990 it was 3 to 3.5 m, and since 1990 it has been 3.5 to 4.5 m. e. The draw point spacing shows an increasing trend with time. Typically, before 1970 it was 8 m, from 1970-1990 it was 12 m, and since 1990 it has been 15 m. f. The influence area of draw points shows an increasing trend with time. Typically, before 1970 it was 50 m2, from 1970-1990 it was 125 m2, and since 1990 it has been 200 to 225 m2. g. The most commonly used extraction level geometry is the herringbone layout (54% of the cases), followed by the El Teniente layout (40% of cases). h. As shown in Figure 7, the average draw rate, RDW, varies from 0.05 to 0.7 m/day, with an average of 0.2 to 0.25 m/day. Figure 8: EXL support by bolts, mesh and osro-straps at a South African underground mine. f. For preliminary estimations of bolt length, the following relations are suggested: Poor quality rock masses (20 ≤ RMR ≤ 40): L (m) = 0.60 x W + 0.60
(5)
Fair quality rock masses (40 ≤ RMR ≤ 60): L (m) = 0.45 x W + 0.45
(6)
Good quality rock masses (60 ≤ RMR ≤ 80): L (m) = 0.30 x W + 0.30
(7)
Figure 7: Relative frequencies of average draw rates, RDW, in block and panel caving mines. 86
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g. Underground caving mines in rockburst prone ground have also used mesh and lacing, as a complementary support for the EXL drifts. h. The support of the draw points differs from one mine to another, but in most cases it includes steel arches, cable bolts, and concrete and/or shotcrete. The number of steel arches previously varied from 2 to 7, but currently most mines used 2 to 3 steel arches.
Collapses correspond to the relatively slow failure of UCL and/or EXL pillars, which triggers crown pillar failure, and drift closure. Figures 10 and 11 illustrate this type of geotechnical hazard.
The analysis of the data on Material Handling Systems indicates that in 57% of the cases examined, caving mines use production shafts; in 27% of cases they use conveyor belts; in 12% they use trains; and in 4% they use trucks. 4 MINE OPERATION The analysis of the data on Mine Operations indicates that: a. The powder factor for undercut blasting varies widely, from 200 to 1000 g/tonne; with an average of 400 to 500 g/tonne. b. LHD capacity varies from 7 to 19 tonnes, with an average of 11 tonnes. c. LHD tramming distance varies widely, from 25 to 300 m, with an average of 125 to 150 m. d. In almost 50% of cases, the oversize limit varies from 1.8 to 2 m3 but has a wide range from 0.4 to 2.4 m3. The average oversize limit is 1.6 m3.
Figure 10: Collapse of an UCL drift at Te 4 Sur, El Teniente mine, 1989.
5 GEOTECHNICAL MONITORING The analysis of the data on Geotechnical Monitoring in caving mines indicates that: a. As shown in Figure 9, the most commonly used monitoring systems are field inspections (100%), local displacement measurements (82%), seismic monitoring (64%), time domain reflectometry (64%), convergence measurements (36%) and observation boreholes (36%). b. As shown in Figure 9, the geotechnical monitoring systems that have generally produced the most satisfactory results are convergence measurements, field inspections, seismic monitoring, time domain reflectometry, aerial photography and water flow measurements. Figure 11: Collapse of an EXL drift at Ten 4 Sur, El Teniente mine, 1989.
The analysis of the data on Collapses indicates that: a. The area affected by a single collapse has varied from 140 to 17,500 m2, with an average of 3,700 m2. b. The most common causes of collapses are irregular draw rate and/or poor draw management, presence of major geological structures, deficient mine planning and/or poor mining sequences, and lack of communication between the geotechnical, planning and mine operations groups. c. The most common remedial measures for collapses are regularization of draw rates and/or improved draw management, improved mining sequences, additional support, and improved communication between the geotechnical, planning and mine operations groups. Figure 9: Relative frequency of use and degree of satisfaction of geotechnical monitoring systems used in underground caving mines.
Rockbursts arise from mining-induced seismic events that cause sudden and violent failure of a volume of rock, which may or may not be defined by structures. Figures 12 and 13 illustrate this type of geotechnical hazard.
6 GEOTECHNICAL HAZARDS The main geotechnical hazards affecting underground caving mines are collapses, rockbursts, subsidence, water inflows and mudrushes, and hangups. Massmin 2004
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Figure 14 illustrates a subsidence crater associated with cave mining.
‘
Figure 12: Typical major rockburst damage without structural control, affecting a UCL drift on Ten Sub 6, El Teniente mine, 1991.
Figure 14: Subsidence crater caused by block and panel caving at Salvador mine, Chile, 2003.
Figure 13: Typical major rockburst damage with structural control, affecting a drift on the Ventilation Level of Ten Sub 6, El Teniente mine, 1990.
The analysis of the data on Subsidence indicates that: a. The depth of a subsidence crater may vary from less than 100 to 1400 m, with the average being about 450 m. b. The angle of break, a, varies from 40° to 90°, depending on the rock mass quality and the structures in the rock mass (in few cases there are even overhanging crater walls). c. The most common factors that may exacerbate subsidence are the presence of major unfavourably oriented geological structures, a poorer rock mass quality than expected, poor draw management and deficient mine planning. d. The most common remedial measures for dealing with subsidence are geotechnical monitoring, geotechnically improved mine planning and mining sequence and, in the final analysis, relocating infrastructure that could be affected by subsidence. Water Inflows / Mudrushes correspond to sudden inflows of water and/or mud from drawpoints or other underground openings. Due to their nature these phenomena propagate rapidly, endangering people, equipment and infrastructure. Figure 15 illustrates the damage that can be caused.
The analysis of the data on Rockbursts indicates that: a. The most common causes of rockbursts are the presence of major geological structures, high stresses, a nonfavourable mining sequence, an excessive undercutting rate, and an excessive draw rate. b. The most common remedial measures for rockbursts are seismic monitoring, improved mining sequences from geotechnical considerations, reduced draw rates, and the implementation of contingency plans (from access restrictions to the closure of a mining sector). Subsidence corresponds to the lowering of the ground surface due to mining causing the development of a crater. Within the cave’s zone of influence, the rock mass undergoes deformations that could damage the mine’s surface and/or underground infrastructure. Commonly, the geometry of a subsidence crater is defined by its depth, H, and the angle of break, a, defined as the inclination with the horizontal of an imaginary line connecting the UCL edge with the perimeter of the crater at surface (defined by the limit of the zone where the rock mass shows large discontinuous deformations). 88
Figure 15: Damage caused by a mudrush in an underground caving mine.
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The analysis of the data on Water Inflows and Mudrushes indicates that: a. The most common causes of water inflows and mudrushes are the presence of wet/clayey ore, fine fragmentation, a crater acting as a water collector and deficient draw management. b. The most common remedial measures for dealing with these hazards are surface stabilization, improved draw management, development and implementation of contingency plans, implementation of drainage systems and instrumentation and monitoring.
c. The most common causes of hangups are unexpected geological-geotechnical conditions, stresses too low to induce stress caving, stresses high enough to develop clamping forces arresting stress release caving, reductions in the undercut height, non-standard undercut geometry, no measures taken to facilitate cave initiation and poor draw extraction/management. d. The most common remedial measures for dealing with hangups are increasing the undercut area, weakening of the hangup boundaries, conditioning the rock mass in the cave back and improved draw management.
Hangups arise from the arrest of cave propagation due to the formation of a metastable geometry that could eventually fail suddenly, generating air blasts and causing significant damage in the underground mine’s excavations. Figure 16 shows an example of this type of hazard. The analysis of the data on Hangups indicates that:
7 CONCLUSIONS The interpretation of the data collected in this study has allowed current trends and practices in underground mining by caving methods to be identified. Table 3 summarizes the currently used values of the most relevant parameters. This compilation may be of value in scoping and pre-feasibility studies. ACKNOWLEDGEMENTS The authors wish to acknowledge the support of the JKMRC and the sponsors of ICS-II: Codelco, De Beers, LKAB, Newcrest Mining, Northparkes, Rio Tinto, Sandvik-Tamrock and WMC. They also wish to express their gratitude to all those who helped them to collect data at the mines visited: Andina, Bingham Canyon, Cadia, Chuquicamata, El Teniente, Finsch, Grasberg Open Pit, Grasberg Underground (DOZ), Henderson, Kidd Creek, Kiruna, Koffiefontein, Mount Keith, Northparkes, Palabora, Ridgeway and Salvador. Finally, the authors wish to especially acknowledge the support provided by Division Codelco Norte for this research. REFERENCES
Figure 16: Formation of chimney craters immediately after the failure of the hangup at Inca West Sector, Salvador mine, Chile (December 5, 1999). a. The area of hangups has varied widely, from 1,000 m2 to more than 35,000 m2, with an average of 12,000 m2. b. Reported data on air blasts are related to hangups having areas larger than 10,000 m2.
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• Flores, G and Karzulovic, A, 2002. Benchmarking Report, prepared for ICS-II, JKMRC and Itasca Consulting Group, Inc.: Brisbane. • Hoek, E and Brown, E T, 1980. Underground Excavations in Rock, 527 p., Institution of Mining and Metallurgy: London.
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Table 3: Typical Parameters for Block and Panel Caving Mines Parameter
Typical Value
Rock Mass Quality
50 ≤ RMR < 60
Accesses
Decline
Currently 70% of mines prefer declines, and 20% use both declines and shafts as mine access.
Block Height
210 m
This typical block height could vary by ± 20%.
< 50,000 m2 50,000 to 100,000 m2 > 100,000 m2
These typical areas could vary +20%. It is recommended to use equal or larger areas, but not smaller than the typical values. Also, square areas are better than the rectangular ones.
Area
10,000 m2
Shape Measures to Facilitate Hydraulic Radius
Square Slot 20 to 30 m
Smaller areas are not recommended, especially in massive rock masses. Internal corners must be avoided (eg an L shaped area). Highly recommended to facilitate cave initiation. Avoid being close to the limit in Laubscher’s chart.
Caving Initiation
UCL
Drifts
Spacing Height Width Undercut Height Undercutting Rate
EXL
Drifts
Draw Points
If RMR > 60 rock mass cavability must be evaluated carefully.
30,000 m2 75,000 m2 170,000 m2
Footprint Area
15 m 4m 4m 8m 2100 m2/month
Crown-Pillar Spacing Height Width
17 m 30 m 4m 4m
Spacing Influence Area
15 m 225 m2
Draw Rates
LHD Equipment
Capacity Tramming Distance
This is the current practice. Could be increased but not decreased. Could vary, but be careful if using small undercutting heights. Could be increased but be careful with induced seismicity, especially if in a high stress environment. Could vary by ± 20% (measured from UCL floor to EXT floor). Could vary from 26 to 36 m. Could be increased but not decreased.
Could vary from 13 to 18 m. Could vary from 169 to 324 m2.
0.20 m/day
This is an average value. Typically lower values are used at the beginning of caving, and higher values are used when more than ‘ 30% of the block height has been extracted.
11 tonnes 140 m
Could vary by ± 20%. Smaller tramming distances are preferable.
Powder Factor
400 g/tonne
For undercutting blasting. It could vary ± 20%.
Oversize Limit
1.8 to 2.0 m3
Could vary by ± 20%.
Subsidence
RMR < 70 RMR > 70
α α
> 45° > 60°
a is the angle of break.
Geotechnical Hazards
The project must take account that collapses, rockbursts, subsidence, water inflows and mudrushes, and hangups could occur.
Instrumentation & Monitoring
The most common monitoring systems include displacements and seismicity. It is recommended to include a seismic monitoring system, especially in massive hard rock and/or high stress environments.
(1) (2)
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Comments
These typical values are intended only for the scoping and pre-feasibility stages of a mining project. RMR values are for Laubscher’s 1990 system.
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Current practices and trends in cave mining Germán Flores, Codelco Norte Division, Codelco Chile Antonio Karzulovic, A. Karzulovic & Assoc. Ltd., Santiago, Chile Edwin T. Brown, Julius Kruttschnitt Mineral Research Centre, University of Queensland, Australia
Abstract Although many caving mines are now exploiting deeper and more massive orebodies, and several large open pits are planning a transition to underground cave mining, there is no single and convenient reference on current practices and trends in cave mining. Hence, as part of the International Caving Study Stage II (ICS-II) funded by nine major international mining companies, a comprehensive benchmarking study was undertaken. The goal of this study was to establish current practices and trends, with emphasis on the key areas that have large impacts on the economics of a mining project, geomechanics, mine design, mine operation, and geotechnical hazards. Data were collected and interviews conducted during visits made to 17 mines in Australia, Canada, Chile, Indonesia, South Africa, Sweden and the USA. A comprehensive review of the technical literature provided information on a further 88 mines. This paper outlines the current trends and practices in caving mining, and summarizes them in a table containing data that may be used during scoping and pre-feasibility studies of cave mining projects, including transitions from open pit to underground mining. 1 INTRODUCTION Table 1: Mines Visited Several mines are planning a transition from open pit to underground cave mining in the medium or long term. Accordingly, the International Caving Study Stage II (ICS-II), managed by the Julius Kruttschnitt Mineral Research Centre, Brisbane, Australia, included a task of providing practical geotechnical guidelines for the transition from open pit to underground cave mining. As part of this task, a benchmarking study was carried out to collect data from mines that have developed, or are planning to develop such transitions, and from other mines (open pits or underground) that could provide relevant data. The collected data were supplemented by a comprehensive review of the technical literature. This study which was planned and executed so as to optimize data collection, included the following elements: Survey Design: To facilitate data collection, an Excel© spreadsheet was designed, and e-mailed to the mines included in the study. Mine Visits: The mines listed in Table 1 were visited, and relevant information was collected. Additional Data Collection: A comprehensive survey of the technical literature was undertaken to collect supplementary data from 88 additional mines. Data Processing: The collected data were analysed in order to develop histograms and, where possible, correlations showing current practice and trends in underground mining by caving methods. When enough data were available the relative frequency of selected parameters was computed, and when the available data were limited, the relative importance of the parameters was assessed. Benchmarking Report: The data, conclusions and recommendations resulting from this study were presented in a Final Report submitted to the sponsors of ICS-II (Flores and Karzulovic, 2002). The results of this benchmarking study reflecting current practices and trends in cave mining, are summarized in this paper.
Country
Mine
Mining Method(s)
Australia
Cadia Hill
OP.
Mount Keith
OP considering a transition to UG.
Northparkes
Mine that developed a transition from OP to UG by BC.
Ridgeway
UG by SLC.
Canada
Kidd Creek
Mine that developed a transition from OP to UG by SOP.
Chile
Andina
OP and UG by BC/PC.
Chuquicamata
OP considering a transition to UG by PC.
El Teniente
UG by PC.
Salvador
UG by PC.
Grasberg
OP.
Indonesia
Grasberg DOZ
UG by PC.
Finsch
Mine that developed a transition from OP to UG by SOP.
Koffiefontein
Mine that developed a transition from OP to UG by SOP/SLC/FC.
Palabora
Mine developing a transition from OP to UG by PC.
Sweden
Kiruna
Mine that developed a transition from OP to UG by SLC.
USA
Bingham Canyon Henderson
OP considering a transition to UG by PC. UG by PC.
OP UG SLC / SOP BC / PC
Open pit mine/mining Underground mine/mining Sublevel caving / Sublevel open stoping Block caving / Panel caving
South Africa
2 GEOMECHANICS The analysis of the data on Geological Structures indicates that: Massmin 2004
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a. In most underground caving mines, subvertical structures (dips > 60o) predominate over subhorizontal structures (dips < 30o). b. In underground and open pit mines the data on the orientations of structures are typically much better than the data on their lengths, spacings, and apertures. c. The geotechnical characterization of structures generally is poorer for underground mines than for open pit mines. d. In spite of the increasing use of numerical models, the quality of input data on the mechanical properties of structures is, in most of cases, poor. The analysis of the data on Intact Rock and Rock Mass Properties indicates that: a. Usually the unit weight (g), and the uniaxial compressive strength (UCS) of the intact rock are known, but information on shear strength and deformability is much poorer. b. Typically UCS values are smaller for open pit mine rocks (average: 80 MPa) that for underground mine rocks (average: 130 MPa). c. Typically RQD values are smaller for open pit mine rocks (average: 60% to 70%) than for underground mine rocks (average: 70% to 85%). d. The most commonly used method of rock mass classification in the underground mines studied is Laubscher’s RMR (53%), followed by Barton’s Q (26%), and Bieniawski’s RMR (15%). e. Typical Laubscher’s RMR ranges found in different mining conditions are shown in Table 2.
Table 2: Typical Laubscher’s RMR Range Mining Method
Laubscher’s RMR Range
Open Pit
20 to 40
Open Stoping
40 to 80
Sublevel Caving
40 to 70
Block Caving
30 to 70
Panel Caving
40 to 80
(1)
g. The geotechnical characterization of rock masses seems to be poorer in underground than in open pit mines. Indeed, in spite of the increasing use of numerical models the quality of input data on rock mass properties is, in most cases, poor to fair. The analysis of the data on In Situ Stresses indicates that: a. Currently the CSIRO Hollow Inclusion Cell is the most commonly used method for in situ stress measurements. b. Typically in underground caving mines the in situ major principal stress, S1, varies from 30 to 40 MPa; and the in situ minimum principal stress, S3, varies from 10 to 20 MPa. c. In underground mines the mean value of the stress ratio, KMEAN, is bounded as proposed by Hoek and Brown (1980), where the depth z is in meters:
0.5 +
84
1500 100 ≥ K MEAN ≥ 0.3 + z z
0.6 +
1250 100 ≥ K Min ≥ 0.2 + z z
1.0 +
1500 90 ≥ K Max ≥ 0.3 + z z
(3)
(4)
The analysis of the data on Hydrogeology indicates that: a. Most mines do not consider hydrogeological characterization to be a high priority. b. The most commonly used drainage systems are sumps (78%), subhorizontal drains (14%) and drainage tunnels (8%). c. The most typical monitoring systems are observation wells (open holes), piezometers and flow rate measurement devices. The analysis of the data on Geotechnical Software indicates that: a. For two-dimensional numerical analyses the most commonly used codes are FLAC (50%), UDEC (33%), and EXAMINE (10%). b. For three-dimensional numerical analyses the most commonly used codes are FLAC3D (44%), 3DEC (26%), and MAP3D (18%). 3 MINE DESIGN
f. The central trend relating Laubscher’s RMR and MRMR is: MRMR = 0,9 ↔ RMR
d. Similar relationships were derived for the minimum and maximum values of the stress ratio, KMIN and KMAX:
(2)
The analysis of the data on Mine Accesses indicates that: a. The use of shafts as the only access has decreased since 1970. b. The use of declines as the only access has increased since 1970. c. Before 1970 in 70% of the case surveyed, shafts were used as accesses and in 30% declines were used. d. In the period 1970 to 1990, in 46% of the cases shafts were used as accesses, in 42% declines were used, and in 13% both shafts and declines were used. e. After 1990, in 36% of the cases shafts were used as accesses, in 50% declines were used, and in 14% both shafts and declines, were used. The analysis of the data on Block Heights and Footprints indicates that: a. As shown in Figure 1, since 1970 block heights have increased in block and panel caving mines. Before 1970, the typical block height was 100 m; for the period 19701990 it was 160 m, and since 1990 it has been 240 m. b. In block and panel caving mines the footprint area varies widely, but in 80% of the cases, it is smaller than 250,000 m2 with an average of 165,000 m2. c. As shown in Figure 2, the ratio between footprint length (L) and width (B) rarely exceeds 3, and in almost 60% of the cases it is smaller than 2. Nevertheless, in many large open pit mines currently considering a transition to underground cave mining this ratio will be larger than 3. d. It seems that most block and panel caving mines have ignored a possible relationship between block height (H) and footprint geometry (defined by its width B). As a preliminary conclusion, and as shown in Figure 3, the data collected suggested that: If H/B ≤ 1 the cave will easily connect to surface (or to an upper mined out level).
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a. As shown in Figure 4, the shape of the initial area for caving is predominantly square or rectangular, but in a few cases other shapes (e.g. triangular) have been used. b. As shown in Figure 4, the available data indicates that the area for caving initiation has an average value of 10,000 m2, and typically varies from 5,000 to 15,000 m2.
Figure 1: Evolution with time of block height in block and panel caving mines.
Figure 4: Relative frequency of initial caving areas and their shapes in block and panel caving mines. c. As shown in Figure 5, the hydraulic radius, HR, of the initial caving area varies from 15 to 45 m, with an average of 20 to 30 m.
Figure 2: Values of the footprint ratio L/B in block and panel caving mines.
Figure 5: Relative frequency of hydraulic radii of the initial caving area in block and panel caving mines. d. In 60% of the cases some measures were taken to facilitate cave initiation (slots: 53%, artificial chimneys: 7%). The analysis of the data on the Undercut Level, UCL, indicates that:
Figure 3: Relation between H and B in block and panel caving mines.
If 2 ≥ H/B > 1 the cave probably will connect to surface (or to an upper mined out level). If H/B > 2 the cave could have problems to connect to surface (or to an upper mined out level). The analysis of the data on Caving Initiation indicates that: Massmin 2004
a. The distance between UCL drifts varies from 10 to 35 m, with an average of 20 to 25 m. b. The width of UCL drifts has increased with time. Typically, before 1970 it was 2 to 3 m, during 1970-1990 it was 3 m, and since 1990 it has been 4 m. c. The height of UCL drifts has increased with time. Typically, before 1970 it was 2 to 2.5 m, during 1970-1990 it was 3 to 3.5 m, and since 1990 it has been 3.5 to 4 m. d. The undercut height shows no time-dependent trend. It varies from 3 to 20 m, with an average of 8 to 12 m. e. As shown in Figure 6, the average undercutting rate, RUC, varies from 500 to 5,000 m2/month, with an average of 2,000 to 3,000 m2/month.
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The analysis of the data on Support indicates that: a. In most underground caving mines support on the UCL consists of only bolts. In some mines this support also includes mesh and shotcrete in the UCL access drifts. b. In most underground caving mines support on the EXL includes bolts (typically 1.8 to 2.4 m long, at spacings of 1.0 to 1.3 m), mesh and shotcrete (typically 50 mm). In many cases it also includes cables, typically 5 to 8 m long, at intersections). Also some mines use straps and osro-straps, as illustrated in Figure 8. c. Bolt lengths range from 1.2 to 3.8 m, with typical values of 2 to 2.3 m on the UCL, and 2 to 2.5 m on the EXL. d. Bolt spacings range from 0.6 to 1.4 m, being typically 1.0 m for both the UCL and the EXL. e. There is no clear difference in the ratio of the bolt length, L, to the drift width, W, used on the UCL and EXL. In most cases the ratio L/W ranges from 1.5 to 3. Figure 6: Relative frequencies of average undercutting rates, RUC, in block and panel caving mines.
The analysis of the data on the Extraction Level, EXL, indicates that: a. The crown-pillar thickness (from the EXL floor to the UCL floor) shows an increasing trend with time. Typically, before 1970 it was 7.5 to 10 m, from 1970-1990 it was 12.5 m, and since 1990 it has been 15 to 17.5 m. b. The spacing between EXL drifts shows an increasing trend with time. Typically, before 1970 it was 12 to 16 m, from 1970-1990 it was 20 to 24 m, and since 1990 it has been 26 to 28 m. c. The width of EXL drifts shows an increasing trend with time. Typically, before 1970 it was 2.5 m, from 1970-1990 it was 3 to 3.5 m, and since 1990 it has been 4 to 4.5 m. d. The height of EXL drifts shows an increasing trend with time. Typically, before 1970 it was 2 to 2.5 m, from 19701990 it was 3 to 3.5 m, and since 1990 it has been 3.5 to 4.5 m. e. The draw point spacing shows an increasing trend with time. Typically, before 1970 it was 8 m, from 1970-1990 it was 12 m, and since 1990 it has been 15 m. f. The influence area of draw points shows an increasing trend with time. Typically, before 1970 it was 50 m2, from 1970-1990 it was 125 m2, and since 1990 it has been 200 to 225 m2. g. The most commonly used extraction level geometry is the herringbone layout (54% of the cases), followed by the El Teniente layout (40% of cases). h. As shown in Figure 7, the average draw rate, RDW, varies from 0.05 to 0.7 m/day, with an average of 0.2 to 0.25 m/day. Figure 8: EXL support by bolts, mesh and osro-straps at a South African underground mine. f. For preliminary estimations of bolt length, the following relations are suggested: Poor quality rock masses (20 ≤ RMR ≤ 40): L (m) = 0.60 x W + 0.60
(5)
Fair quality rock masses (40 ≤ RMR ≤ 60): L (m) = 0.45 x W + 0.45
(6)
Good quality rock masses (60 ≤ RMR ≤ 80): L (m) = 0.30 x W + 0.30
(7)
Figure 7: Relative frequencies of average draw rates, RDW, in block and panel caving mines. 86
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g. Underground caving mines in rockburst prone ground have also used mesh and lacing, as a complementary support for the EXL drifts. h. The support of the draw points differs from one mine to another, but in most cases it includes steel arches, cable bolts, and concrete and/or shotcrete. The number of steel arches previously varied from 2 to 7, but currently most mines used 2 to 3 steel arches.
Collapses correspond to the relatively slow failure of UCL and/or EXL pillars, which triggers crown pillar failure, and drift closure. Figures 10 and 11 illustrate this type of geotechnical hazard.
The analysis of the data on Material Handling Systems indicates that in 57% of the cases examined, caving mines use production shafts; in 27% of cases they use conveyor belts; in 12% they use trains; and in 4% they use trucks. 4 MINE OPERATION The analysis of the data on Mine Operations indicates that: a. The powder factor for undercut blasting varies widely, from 200 to 1000 g/tonne; with an average of 400 to 500 g/tonne. b. LHD capacity varies from 7 to 19 tonnes, with an average of 11 tonnes. c. LHD tramming distance varies widely, from 25 to 300 m, with an average of 125 to 150 m. d. In almost 50% of cases, the oversize limit varies from 1.8 to 2 m3 but has a wide range from 0.4 to 2.4 m3. The average oversize limit is 1.6 m3.
Figure 10: Collapse of an UCL drift at Te 4 Sur, El Teniente mine, 1989.
5 GEOTECHNICAL MONITORING The analysis of the data on Geotechnical Monitoring in caving mines indicates that: a. As shown in Figure 9, the most commonly used monitoring systems are field inspections (100%), local displacement measurements (82%), seismic monitoring (64%), time domain reflectometry (64%), convergence measurements (36%) and observation boreholes (36%). b. As shown in Figure 9, the geotechnical monitoring systems that have generally produced the most satisfactory results are convergence measurements, field inspections, seismic monitoring, time domain reflectometry, aerial photography and water flow measurements. Figure 11: Collapse of an EXL drift at Ten 4 Sur, El Teniente mine, 1989.
The analysis of the data on Collapses indicates that: a. The area affected by a single collapse has varied from 140 to 17,500 m2, with an average of 3,700 m2. b. The most common causes of collapses are irregular draw rate and/or poor draw management, presence of major geological structures, deficient mine planning and/or poor mining sequences, and lack of communication between the geotechnical, planning and mine operations groups. c. The most common remedial measures for collapses are regularization of draw rates and/or improved draw management, improved mining sequences, additional support, and improved communication between the geotechnical, planning and mine operations groups. Figure 9: Relative frequency of use and degree of satisfaction of geotechnical monitoring systems used in underground caving mines.
Rockbursts arise from mining-induced seismic events that cause sudden and violent failure of a volume of rock, which may or may not be defined by structures. Figures 12 and 13 illustrate this type of geotechnical hazard.
6 GEOTECHNICAL HAZARDS The main geotechnical hazards affecting underground caving mines are collapses, rockbursts, subsidence, water inflows and mudrushes, and hangups. Massmin 2004
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Figure 14 illustrates a subsidence crater associated with cave mining.
‘
Figure 12: Typical major rockburst damage without structural control, affecting a UCL drift on Ten Sub 6, El Teniente mine, 1991.
Figure 14: Subsidence crater caused by block and panel caving at Salvador mine, Chile, 2003.
Figure 13: Typical major rockburst damage with structural control, affecting a drift on the Ventilation Level of Ten Sub 6, El Teniente mine, 1990.
The analysis of the data on Subsidence indicates that: a. The depth of a subsidence crater may vary from less than 100 to 1400 m, with the average being about 450 m. b. The angle of break, a, varies from 40° to 90°, depending on the rock mass quality and the structures in the rock mass (in few cases there are even overhanging crater walls). c. The most common factors that may exacerbate subsidence are the presence of major unfavourably oriented geological structures, a poorer rock mass quality than expected, poor draw management and deficient mine planning. d. The most common remedial measures for dealing with subsidence are geotechnical monitoring, geotechnically improved mine planning and mining sequence and, in the final analysis, relocating infrastructure that could be affected by subsidence. Water Inflows / Mudrushes correspond to sudden inflows of water and/or mud from drawpoints or other underground openings. Due to their nature these phenomena propagate rapidly, endangering people, equipment and infrastructure. Figure 15 illustrates the damage that can be caused.
The analysis of the data on Rockbursts indicates that: a. The most common causes of rockbursts are the presence of major geological structures, high stresses, a nonfavourable mining sequence, an excessive undercutting rate, and an excessive draw rate. b. The most common remedial measures for rockbursts are seismic monitoring, improved mining sequences from geotechnical considerations, reduced draw rates, and the implementation of contingency plans (from access restrictions to the closure of a mining sector). Subsidence corresponds to the lowering of the ground surface due to mining causing the development of a crater. Within the cave’s zone of influence, the rock mass undergoes deformations that could damage the mine’s surface and/or underground infrastructure. Commonly, the geometry of a subsidence crater is defined by its depth, H, and the angle of break, a, defined as the inclination with the horizontal of an imaginary line connecting the UCL edge with the perimeter of the crater at surface (defined by the limit of the zone where the rock mass shows large discontinuous deformations). 88
Figure 15: Damage caused by a mudrush in an underground caving mine.
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The analysis of the data on Water Inflows and Mudrushes indicates that: a. The most common causes of water inflows and mudrushes are the presence of wet/clayey ore, fine fragmentation, a crater acting as a water collector and deficient draw management. b. The most common remedial measures for dealing with these hazards are surface stabilization, improved draw management, development and implementation of contingency plans, implementation of drainage systems and instrumentation and monitoring.
c. The most common causes of hangups are unexpected geological-geotechnical conditions, stresses too low to induce stress caving, stresses high enough to develop clamping forces arresting stress release caving, reductions in the undercut height, non-standard undercut geometry, no measures taken to facilitate cave initiation and poor draw extraction/management. d. The most common remedial measures for dealing with hangups are increasing the undercut area, weakening of the hangup boundaries, conditioning the rock mass in the cave back and improved draw management.
Hangups arise from the arrest of cave propagation due to the formation of a metastable geometry that could eventually fail suddenly, generating air blasts and causing significant damage in the underground mine’s excavations. Figure 16 shows an example of this type of hazard. The analysis of the data on Hangups indicates that:
7 CONCLUSIONS The interpretation of the data collected in this study has allowed current trends and practices in underground mining by caving methods to be identified. Table 3 summarizes the currently used values of the most relevant parameters. This compilation may be of value in scoping and pre-feasibility studies. ACKNOWLEDGEMENTS The authors wish to acknowledge the support of the JKMRC and the sponsors of ICS-II: Codelco, De Beers, LKAB, Newcrest Mining, Northparkes, Rio Tinto, Sandvik-Tamrock and WMC. They also wish to express their gratitude to all those who helped them to collect data at the mines visited: Andina, Bingham Canyon, Cadia, Chuquicamata, El Teniente, Finsch, Grasberg Open Pit, Grasberg Underground (DOZ), Henderson, Kidd Creek, Kiruna, Koffiefontein, Mount Keith, Northparkes, Palabora, Ridgeway and Salvador. Finally, the authors wish to especially acknowledge the support provided by Division Codelco Norte for this research. REFERENCES
Figure 16: Formation of chimney craters immediately after the failure of the hangup at Inca West Sector, Salvador mine, Chile (December 5, 1999). a. The area of hangups has varied widely, from 1,000 m2 to more than 35,000 m2, with an average of 12,000 m2. b. Reported data on air blasts are related to hangups having areas larger than 10,000 m2.
Massmin 2004
• Flores, G and Karzulovic, A, 2002. Benchmarking Report, prepared for ICS-II, JKMRC and Itasca Consulting Group, Inc.: Brisbane. • Hoek, E and Brown, E T, 1980. Underground Excavations in Rock, 527 p., Institution of Mining and Metallurgy: London.
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Table 3: Typical Parameters for Block and Panel Caving Mines Parameter
Typical Value
Rock Mass Quality
50 ≤ RMR < 60
Accesses
Decline
Currently 70% of mines prefer declines, and 20% use both declines and shafts as mine access.
Block Height
210 m
This typical block height could vary by ± 20%.
< 50,000 m2 50,000 to 100,000 m2 > 100,000 m2
These typical areas could vary +20%. It is recommended to use equal or larger areas, but not smaller than the typical values. Also, square areas are better than the rectangular ones.
Area
10,000 m2
Shape Measures to Facilitate Hydraulic Radius
Square Slot 20 to 30 m
Smaller areas are not recommended, especially in massive rock masses. Internal corners must be avoided (eg an L shaped area). Highly recommended to facilitate cave initiation. Avoid being close to the limit in Laubscher’s chart.
Caving Initiation
UCL
Drifts
Spacing Height Width Undercut Height Undercutting Rate
EXL
Drifts
Draw Points
If RMR > 60 rock mass cavability must be evaluated carefully.
30,000 m2 75,000 m2 170,000 m2
Footprint Area
15 m 4m 4m 8m 2100 m2/month
Crown-Pillar Spacing Height Width
17 m 30 m 4m 4m
Spacing Influence Area
15 m 225 m2
Draw Rates
LHD Equipment
Capacity Tramming Distance
This is the current practice. Could be increased but not decreased. Could vary, but be careful if using small undercutting heights. Could be increased but be careful with induced seismicity, especially if in a high stress environment. Could vary by ± 20% (measured from UCL floor to EXT floor). Could vary from 26 to 36 m. Could be increased but not decreased.
Could vary from 13 to 18 m. Could vary from 169 to 324 m2.
0.20 m/day
This is an average value. Typically lower values are used at the beginning of caving, and higher values are used when more than ‘ 30% of the block height has been extracted.
11 tonnes 140 m
Could vary by ± 20%. Smaller tramming distances are preferable.
Powder Factor
400 g/tonne
For undercutting blasting. It could vary ± 20%.
Oversize Limit
1.8 to 2.0 m3
Could vary by ± 20%.
Subsidence
RMR < 70 RMR > 70
α α
> 45° > 60°
a is the angle of break.
Geotechnical Hazards
The project must take account that collapses, rockbursts, subsidence, water inflows and mudrushes, and hangups could occur.
Instrumentation & Monitoring
The most common monitoring systems include displacements and seismicity. It is recommended to include a seismic monitoring system, especially in massive hard rock and/or high stress environments.
(1) (2)
90
Comments
These typical values are intended only for the scoping and pre-feasibility stages of a mining project. RMR values are for Laubscher’s 1990 system.
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Incline cave: A technical alternative method to mine kimberlite deposits at depth. Miguel Paucar and Collin Mthombeni, Finsch Mine, De Beers, South Africa
Abstract De Beers underground operations have mined kimberlitic pipes utilizing the Block caving as the preferred mining method since the late 1950’s. These pipes are been mined at shallow depth, with large ore body footprint size, and with low block heights. The requirement to extend the life of the existing operations has resulted in investigations for mining at greater depth and higher block heights due to the decreasing footprint and volume of the orebody at depth. Therefore, it is strategic for De Beers to design and propose alternative mining methods, which will reduce the technical risks at similar or lower costs to ensure a safe and efficient ore extraction. This paper describes the experience and challenges of the Block cave method to meet the technical requirements of mining at depth and higher block heights. An incline cave mining method is proposed as an alternative method to meet some of the challenges of mining kimberlite at depth.
1 INTRODUCTION De Beers has been exploiting the Finsch Mine diamondbearing kimberlite pipe since 1966. Open-pit methods were first used to exploit the diamond pipe. After the open-pit operation the mine changed to underground operations and the pipe was divided into series of blocks, as shown in Figure 1. Block 1 and 2 were mined by a combination of open pit and blasthole open stoping methods. Block 3 is exclusively blasthole open stoping while Block 4 will employ block caving (Preece, 1998). Block 5 is currently being delineated and as part of the pre-feasibility study alternative mining methods are being investigated. Due to the pipes decreasing geometry (tapering at an angles between 76 -86 degrees), and footprint size with depth, the Block 5 drop down has been set at 250 metres. This is more from a financial stand point than anything else. This will result in a considerable change in the mining environment, since no kimberlite pipe has at yet been mined at these block heights. As part of the investigations into determining the most suitable mining method for the block, an inclined drawpoint caving layout was identified as one. An incline drawpoint layout was first introduced at Gaths Mine, Zimbabwe, as it was not possible to maintain a horizontal layout (Laubscher, 2000). This paper describes the proposed mining method, compared to the block caving method, using the Finsch pipe geometry as a case study. Although the incline cave is mostly applicable to dipping orebodies, there is no reason why this method should not be applied to other mining situations.
shape with steeply dipping (80o to 85o) walls, as shown in Figure 2. This shape is consistent, although it reflects the hardness of the country rock that the kimberlite broke through on its way to surface. The kimberlite within the pipe can usually be sub-divided into three broad facies: • The hypabyssal facies, which is plug-like, dyke intrusion, comprising of a harder, more competent kimberlite. • The diatreme facies, which constitutes the mid portion of the pipe, consist of various types of Tuffistic Kimberlite Breccias (TKB). • The crater facies, which is the upper portion, is influenced by the surface venting of the volcano. Most of the pipes have inclusions of the country rock through which they have intruded, being mostly basalt at Finsch Mine or quartzite at Cullinan Diamond Mine. These inclusions create a complex rock mass structures which are exacerbated by the hygroscopic nature of kimberlite. On exposure kimberlite deteriorates rapidly, leading to relatively low design rockmass strength, being between 18 to 25 MPa for TKB kimberlite.
2 GEOLOGY AND GEOMECHANICS Kimberlite, the rock in which diamonds are generally found, consists of a mixture of different rock materials whose most important constituents are: (1) fragments of peridotite, eclogite, (2) large crystals of peridotite or eclogite altered by fluids in the kimberlite and by ground water and (3) the matrix, which is the crystallized kimberlite magma. The original kimberlitic volcanoes are now known as "pipes" because of the typically carrot or inverted-cone Massmin 2004
Figure 1: Diagrammatic section of Finsch Mine.
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Rockmass classification techniques, as well as numerical modeling techniques have been applied to the mine design and planning process. Due to the complexity of the kimberlite, these models generally require extensive validation by back analysis.
Figure 2: Cross section of a typical kimberlite pipe in South Africa (Harlow, 1997).
3 BLOCK CAVE MINING CHALLENGES IN KIMBERLITE PIPES The challenges of mining kimberlitic pipes at depth are that: • The decreasing footprint results in increasing the block heights in order to optimize the required economic life from the block being mined.
• Higher in-situ stresses, poor rock conditions (structures) and water resulting in an increased risk of not achieving mine production. Common challenges for mining kimberlite pipes are the difficulties being experienced in maintaining the extraction level. The drawpoint horizon is only capable of standing for a limited time and in the worst cases completely collapsed tunnels may be required to be re-established and resupported several times at high cost during their productive life (Laubscher and Esterhuizen, 1994). This is because the extraction level is developed in relatively weak rock and also as much as 50% of the rock on the extraction level is extracted to create production tunnels, crosscuts and drawbells required for mining. The rock strength capacity starts being challenged, and also the rock is deteriorating due the blasting/mining activities which cause serious damage. In an attempt to alleviate the situation, various modifications are being made based on past experiences. These modifications vary from various support techniques and extraction level layouts (distance between drawpoints, shapes of drawpoints, El Teniente v/s herringbone layouts) to undercutting techniques (pre and post undercut, advance undercut, undercut height, etc). Detailed information on these modifications is to be found in Bartlett, 1998. Although some benefit can be derived from these measures, it becomes increasingly apparent that the horizontal extraction level could be becoming totally unsuited to the ground conditions that are being experienced. Anticipating that the problems encountered would compound with depth, the need to investigate an alternative mining method for the kimberlite pipes at depth and higher block heights is becoming more and more apparent. Alternative caving mining methods previously being applied in kimberlite pipes such as the Sub Level Caving (previously applied in Kimberley Mines) and the Front Cave (applied in Koffiefontein Mine) have been investigated. The SLC was found to be not economically attractive due to the
Figure 3: Layout and design of Incline cave - 3D model of levels and infraestructure required. 92
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amount of development, support and the drill and blast running cost associated with the method. The Front Cave had the limitation of not being suitable for higher block heights and the limited production output due to the geometry of the pipe.
4 DESIGN AND LAYOUT OF THE INCLINE CAVE METHOD
will be blasted above the top extraction level to extend the undercut to the contact of the orebody. Rings will be drilled from the crosscuts, and blasted into the slot for a distance of 11m to form the draw cones. A permanent drawpoint brow will then be established at that position. Caving is initiated once the slot and draw cones are connected between the different levels, forming a continuous undercut area inclined at 45 degrees, as shown in Figure 5.
4.1 General description The incline cave method proposed here is based on the false footwall incline cave method being applied in Gaths Mine, Zimbabwe. The preliminary design and layout presented in this paper is just a step in the iterative process and will be revisited in the cycle of planning and design before a final layout is produced. All the design parameters for a horizontal block caving layouts have been assumed to be the same for the incline caving layout. The method entails the development of a number of extraction levels at the lower part of the block. For the block height of about 250m, seven extraction levels, at 15m vertical intervals, will be developed in the bottom of the block (Figure 3). The extraction tunnels will be 15m apart, and will be staggered from level to level as in sub-level caving. The ends of the tunnels will be located along a plane inclined at 45 degrees as shown in Figure 4. The eighth level will be the goundhandling level, where the ore will be crushed and then conveyed to the shaft station.
1) Slot opening Scale 1:1000
4.2 Block size and height At 250m below the current mining level, the pipe has a footprint of approximately 4 hectares (max 250m length by 150m wide). A ‘mirror image’ (see Figure 4) incline layout has been applied in order to limit the production working area to the least number of levels. This layout has resulted in 7 production (extraction) levels, with the bottom level drawing the highest columns (220m) and the top level drawing shortest column height (175m).
2) Undercut opening Scale 1:1000
Figure 4: Section view of the pipe and level positions.
4.3 Undercutting Continuous slots will be blasted, 4m wide and 15m high, to connect the ends of the crosscuts on each level, as shown in Figure 5 (slot opening). An additional wider slot Massmin 2004
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3) Caving and Loading Scale 1:1000
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in conjunction with Blumhm Burton consultants. The findings recommended a force exhaust system for the development phase and an exhaust system for the undercutting and production stages. The volume of air required is 760 m3/sec. compared to 680 m3/sec. for the block caving method. 5 STABILITY AND SUPPORT
4) Retreat using SLC method Scale 1:1000
Figure 5: Mining sequence of the Incline cave method. The undercut initiation position was influenced by: • Geotechnical factors to ensure caving propagation, • High grade area, • The amount of development before caving ocurrs. The selected option was to initiate undercut from the south position; from the bottom level up, as shown in Figure 6.
From the experience at Gaths and Cassiar mines it has been shown that if the inclined cave design is correctly orientated in relation to the stresses and structures, the general or global stability is much better than the traditional horizontal block caving method. The pillars formed in a conventional Block cave; by the undercut and extraction level tunnel development (major apex) and by drawbell opening (minor apex) are freestanding pillars that are inherently weak and require extensive support and major expense. The advantage of the incline cave layout is that the drawpoints and production tunnels are located in a narrow zone backed by solid ground. Support has always been an issue in Block caving; the extraction level has to undergo various stress changes. During undercutting operations it is under a high compressive load, and once the undercut is completed the extraction level tunnel goes under tension or stress relaxation. Therefore for these severe stress changes a increased level of support is required for the extraction level. In the case of the incline cave the stress distribution is different since it has production tunnels in different levels. The amount of support required on the production tunnels and drawpoints will be less (lower cost) than in the block caving due to the better stability. The incline cave design provides more drawpoints than the block cave method and the strengh of them is greater too, meaning the drawpoint availability is higher, and hopefully a lower repair cost. Support is designed on the basis of experience, rock mass characterisation, and expected stress changes associated with cave mining as predicted by numerical stress modelling. The support requirements for the various areas in the mine excavations are: Dolomite Access ramps and rim tunnels – 2.9m X 25mm fully grouted rock bolts at a 1.5m square pattern installed during primary development. In areas where additional support is necessary, shotcrete and wire mesh will supplement the rock bolt support. Rim tunnel/loading drive break-aways – 3.8m X 25mm rock bolts, 6.0m fully grouted pre-tensioned rope anchors and meshing/shotcrete if required.
Figure 6: Method of undercutting – From corner bottom up.
4.4 Remnants and wedge mining The ore in the final triangular wedge of ground containing the tunnels will be recovered by sub-level caving of the tunnels, as shown in Figure 5. The wedge that could not be mined after the sub level cave (SLC) mining will be allowed to cave down to the next mining block. 4.5 Ventilation It was first thought that the Incline Cave method would have a much higher air requirements than a Block Cave. A ventilation study was carried out by Anglo Technical Division 94
Kimberlite Loading drives in kimberlite – sealant, 25mm shotcrete, 3.8m X 25mm rock bolts at 1.0m square pattern, meshing and strapping and a final 100mm layer of shotcrete. Slot tunnels – As per the loading drives above except at the intersections with the loading tunnels where additional 6.0m rope anchors will be installed to support the large spans. Loading drive contact support – where necessary steel arches will be installed in highly unstable contact areas to support the key blocks. Drawpoint support will be as per current mine practices, these being the Koffiefontein type semi-permanent
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drawpoint (SPD) support design for the front cave. Consisting of 6m long anchors, RSJ sets and concrete. 6 PRODUCTION AND DRAW CONTROL Ore is removed from the drawpoint by diesel LHD machines, which will transport the ore to orepasses. All the levels feed a common orepass system on both sides of the pipe. Unlike in panel caving if an access tunnels get damaged, the column can be drawn from the adjacent tunnel. As Dr. D. Laubscher defines draw control: "Practice of controlling tonnages drawn from individual drawpoints with the object of: Minimize dilution, ensure maximum ore recovery, avoid damaging load concentration on the extraction level and avoid creation of conditions leading to air blast, mud rushes, etc". (Laubscher, 2000). The grade and fragmentation in the dilution zone must be known in order to practice good draw control. A poor draw control means drawing of fine material at the expense of coarse material. Draw control discipline must be applied so that coarse material must be drilled and blasted at the end of the shift. During the production phase, brow wear will occur, and conditions may deteriorate because of weak rock, but the instalattion of brow support as semi permanent drawpoint (SPD) will reduce it and enhance stability in this area. Unlike a conventional block-cave layout, a particular brow position can be abandoned and mining can fall back along the drawpoint to a more stable location. This is made possible by the greater length of drawpoint drifts and the absence of extraction pillars with limited dimensions. In falling back, the effect on the draw control and interaction with other levels should be considered.
Advantages • High productivity due to close drawpoint spacing. • Covering levels from 790m down to 880m, makes easier to recover ore from upper levels. • In the event of damage to an access tunnel, columns can be drawn from the adjacent tunnel because they are placed in different levels. • More drawpoints covering larger footprint area – more productivity. • Stress distribution is improved, especially during undercutting and reduction in tunnels support cost. • One drawpoint is dedicated to a tunnel on a level, makes automation more applicable. Disadvantages • Never applied in kimberlite mining. • Long lead time is needed to install most of the infrastructure in place before the start of undercutting. • More metres to be developed and capital investment are relatively higher. • Multi-tipping level is seen as major challenge due to ventilation. • Supervision not concentrated in one level. ACKNOWLEDGEMENTS The authors would like to thank Alan Guest, General Manager, Geotechnical Engineering for his encouragement and comments during the preparation of this paper. The authors acknowledge the permission of Management, Finsch Mine and the Director Operations, De Beers Consolidated Mines to publish this paper. REFERENCES
7 CONCLUSIONS The incline cave method is a combination of sub-level caving and block caving (low cost - high production in weak rock) but it has never been applied in the mining of kimberlite pipes, although successes have been noted at King section, Gaths Mine in Zimbabwe and Cassiar Mine in Canada (Laubscher and Esterhuizen, 1994). However, due to advantages mentioned before, we believe that this method warrants further investigation in order to minimize risks inherent to unproven technology. The risks can be reduced by fragmentation analysis, draw control, dilution model through PCBC studies, scheduling of activities using a software tool, and simulation of mining equipment requirements and the overall ore extraction system. The advantages and disadvantages of the method are as follows:
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• Bartlett, P, 1998. Planning a mechanised cave with coarse fragmentation in Kimberlite, PhD thesis. • Diaz, G and Tobar, P, 2000. Panel caving experiences and macrotrench – An alternative exploitation method at the El Teniente mine, Codelco – Chile. Massmin 2000, Brisbane, Australia, pp. 235 – 247. • Harlow, G, 1997. The Nature of Diamonds – Cambridge University press in association with the American Museum of Natural History. • Laubscher, D, 2000. A practical manual on Block Caving – For the International Caving Study. • Laubscher, D and Esterhuizen, G, 1994. Inclineddrawpoint caving – a cave mining method. XVth Congress, Johannesburg, SAIMM. Volume 1 pp. 247-250. • Owen, K and Guest, A, 1994. Underground mining of kimberlite pipes. XVth Congress, Johannesburg, SAIMM. Volume 1 pp. 207-218. • Preece, C, 1998. Finsch Mine; Open pit to open stoping to Block caving, Underground mining methods, pp. 439 – 453.
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Quantifying open stope performance E. Villaescusa, Professor of Geomechanics, Western Australian School of Mines
Abstract Stope performance is reviewed with respect to the overall stope design process. Global and detailed design issues are identified along the way, and the stope design note is described in detail. Stope performance is quantified based on depth of failure measurements, which are calculated using block models of Cavity Monitoring System wireframes and tested against the stope design boundaries. Finally a stope performance assessment summary data sheet is also provided.
1 INTRODUCTION The sublevel open stoping method (SLOS) are used to extract large massive or tabular, steeply-dipping competent orebodies surrounded by competent host rocks which in general have few constraints regarding the shape, size and continuity of the mineralization. In general, open stopes are relatively large excavations in which ring drilling is the main method of rock breakage (Villaescusa, 2000). The SLOS method offers several advantages including, low cost and efficient non-entry production operations, utilization of highly mechanized, mobile drilling and loading production equipment, high production rates with a minimum level of personnel. Furthermore, production operations are concentrated into few locations such as ring drilling, blasting and drawpoint mucking. The disadvantages include a requirement for a significant level of development infrastructure before production starts, thus incurring a high initial capital investment. However, most of the development occurs within the orebody. In addition, the stopes must be designed with regular boundaries and internal waste pockets can not be separated within the broken ore. Similarly, delineated ore can not be recovered beyond a designed stope boundary. Consequently, ore dilution, consisting of low-grade, waste rock or minefill materials, may occur at the stope boundaries. Furthermore, ore loss due to insufficient breakage can also occur within the stope boundaries. The stope performance is measured by the ability to achieve maximum extraction with minimal dilution. Hence, the success of the method relies on the stability of large (mainly un-reinforced) stope walls and crowns as well as the stability of any fill masses exposed 2 STOPE DESIGN PROCESS Stope design for dilution control requires interactions among geology, mine planning, rock mechanics and operating personnel (Villaescusa, 1998). The overall rational methodology for the stope design process is shown in Figure 1. Six key stages (and key personnel) are identified, with the orebody delineation and rock mass characterization stages as the basic input. The tasks consists of an early determination of rockmass properties on a block scale, followed by an estimate of the likely loading conditions from the mining sequences. The process requires a global and a detailed design stage, where global design issues are relevant and applicable within entire areas of a mine, such as an extension of an existing orebody, while detailed design issues are applicable to the 96
extraction of individual stopes. Finally, a monitoring and back analysis strategy is required to allow a documented closure of the mine design loop. 3 GEOLOGICAL AND GEOTECHNICAL CHARACTERIZATION The stope design process starts with an initial orebody delineation process to provide an interpolated outline of the grade contours. This information is critical and is initially used to locate the required drilling and mucking drives along the orebody in question. The accuracy of the delineated grade boundaries is a function of the nature of the orebody, the amount of drilling information and the mining access through the orebody. For narrow orebodies, development is carried out under strict geological control, a process that requires geological mapping of drives and crosscuts through the orebodies (and sometimes additional in-fill drilling) in order to define the stope ore-waste contacts. The suggested approach is to obtain representative (mine-wide) rock mass properties required during the global excavation design and stability analysis stages. In most cases, this information is obtained from diamond drill holes (core logging of non-oriented holes, as well as geotechnical holes) and direct mapping of underground openings. Geophysical tools can also used for orebody delineation and rock mass characterization. The confidence in the geological information must be sufficient to establish the nature and irregularities of the orebody, the nature and location of major controlling geological structures, the general rock mass characteristics as well as to perform an economic evaluation to determine whether a particular block should be mined. This type of information requires that the sampling process extend beyond the orebody boundaries in order to determine the likelihood of failure from orebody hangingwalls, footwalls or stope crowns. Experience has shown that the interpolated grade may define the economics of a stope, but the geological structures and the location (and alignment) of the drives up-dip may define the final shape. Lithology and the presence of major faults and joints relative to the stope wall orientations need to be anticipated in order to control dilution (See Figure 2). A need exists for routine geological mapping and timely interpretations to keep the geology current and to determine areas of low rock mass strength, associated with clay fracture filling and moderate to complete wall alteration. Interpreted geology maps across all stope levels, on a stope composite basis, are essential tools for evaluating the likely influence of rock type
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and major geological discontinuities on the actual stope performance.
of mine access and infrastructure, dimensions of sublevel intervals, backfill requirements and infrastructure, equipment and ventilation considerations, etc. Stress analysis of the global production schedules is critical to determine the loading conditions (stress and displacement) likely to result from a proposed mine-wide stoping sequence. Table 1. Global (block) design issues. Exploration drilling requirements for orebody delineation for the designed area Area wide rock mass characterization from borehole data and direct access Overall mining method selection Quantity and grade of ore required with respect to scheduled metal targets Access and infrastructure development requirements ore handling systems, workshops, etc. Production scheduling, details and timing Induced stresses from scheduled sequences, including extraction directions Primary and secondary stope dimensions (including regional access pillars) Backfill system requirements Equipment requirements Ventilation Global economic assessment
5 DETAILED DESIGN Figure 1. A formalized stope design methodology from data collection to stope reconciliation.
Figure 2: Massive stope hangingwall failure controlled by large scale faults.
4 GLOBAL DESIGN Global design issues are related to the design and stability of large sections of a mine, such as a new orebody extension at depth or at the abutment for an existing deposit. Global design issues are schematically represented on Figure 1, and listed in detail in Table 1. The issues involved include global orebody delineation, design Massmin 2004
Detailed design is related to the extraction of individual stopes within a global area and it represents the process of establishing an optimum extraction method for an individual stope, subject to a number of variables and constraints. Blasthole geometry, firing sequence, ground support, ventilation and economics are some of the key variables considered. The constraints include the orebody boundaries, the geological structures, any existing development, and in some cases, any adjacent backfill masses (See Figure 3). Detailed design is achieved by means of a stope design note issued to the planning and operating personnel. Such a document includes detail on the overall extraction philosophy, plans of sublevel development, sections showing blasthole design concepts and drilling and blasting parameters, ventilation, geology, rock mechanics and overall firing sequence. All the topics included on a stope design document are inter-related. The extraction philosophy provides a general overview of the design, safety and production issues for a particular section of an orebody. Properly reinforced stope development is required to allow access for drilling, blasting and mucking. Development size is a function of the stoping method and the equipment utilised. Knowledge of the nature and stability of the adjacent backfill masses is needed to design cleaner rings or to avoid toeing of blastholes into the backfill. Structural geology considerations such as the presence of major geological discontinuities often influence the blasting sequences. Other factors considered are the stress redistributions within and around a stope and likely to control fall-off behaviour on the exposed walls. In addition, the
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retreat direction of the blasthole rings must take into account the stope ventilation network, with a retreat direction into fresh air. 6 STOPE DESIGN NOTE A stope design note covers many aspects involved in the development and production of a stope (See Table 2). Technical presentations are required to encourage technical input from all the members of the design team (geology, rock mechanics, planning, operations and management). They usually occur twice within the design process: at the conceptual design stages and prior to the issue of the final drill and blast design. Feedback from both meetings should be incorporated into the final stope design.
Once a final stope design status has been achieved, the blasthole design is undertaken by considering the production rigs that will be used, the ore limits, the survey pick-up of the access development, the extent and sublevels of the stope, as well as the ring burden and toe spacing (See Figure 4). The ore limits are usually updated in accordance with the completed stope development. A scaled floor plan showing details of the latest survey information including any vertical openings and status of surrounding stopes will be provided to assist the drillers. Locations of hangingwall, footwall, cut-off detail and location of the main rings are also included. A long section that includes a schematic view of the stope cut-off raise, the cutoff, the production rings and the trough undercuts, is also completed. This section helps to explain the stope design philosophy, and becomes a useful tool during drilling and blasting of the stope.
Table 2. Stope design presentation issues 7 STOPE EXTRACTION
Geological structures Stope access and development requirements Ore passes, loading bays, etc. Stope cut-off location Production blast directions Stability issues, ground support requirements Stress re-distributions assessment Backfill or permanent pillar demands Production schedule Ventilation requirements Detailed economic analysis
The actual firing sequence used to extract individual stopes is likely to influence stress re-distribution as well as blast induced damage within a stope. Stress and blast induced fall-off within a stope boundary may lead to poor mucking performance during extraction. Although fall-off resulting from stope firing is not the only source of poor fragmentation, it can be minimized by avoiding excessive undercutting of the stope walls (See Figure 5). A number of design options can be used to reduce stope undercut; including for example, firing the cut-off slot to the full height of the stope before blasting of the main rings commences. This can be followed by the sequential blasting of the main rings to the full stope height (Villaescusa, 2000). The objective is to reduce the number of stope faces exposed, thereby reducing the potential for time related
Figure 3. Multiple lift stope showing main ring and diaphragm ring details. 98
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Figure 4. Cross section view showing drilling details in multiple lift open stoping. A stope firing sequence also determines the rate of exposure of the main geological discontinuities intersecting a stope (See Figure 6). A rapid exposure of a large fault may occur after mass blasting or after progressive firing to a fault (See Figure 6a). Such exposures may not allow sufficient time for a gradual stress relief. If the orientation of the stress field is unfavorable, large shear stresses may result and induce local and regional fault movements leading to stope fall-off. In order to optimize stability, stope firings should proceed across a structure to allow progressive stress relief of the shear stresses (See Figure 6b).
Figure 5. Stope wall undercutting within a stope firing sequence.
structurally controlled fall-off. Undercut of the main rings can be avoided by designing the troughs to be blasted with coinciding faces. Massmin 2004
Figure 6. Exposure of large geological features during stoping operations.
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Table 3. Example of potential problems and solutions in open stoping. Open Stope Activity
Potential Problem
Potential Solution
Rock mass characterization
Design may not be stable
Back analyze previous extracted stopes
Different domain for design Within stope boundaries
Geological engineering judgement
Insufficient information
More geological mapping
Major discontinuities intersect stope walls
Consider firing sequences and cablebolt reinforcement
Design by default
Better preparation job – use databases of stope performance
Tonnage and grade do not match the design
Better geological interpretation needed
Stope access is not in the appropriate location
Better planning
Orebody delineation do not match the geological interpretation
More definition drilling, use geophysical techniques
Excessive development in waste
Optimize the block design
Not following the design
Spot check and quality control, better communication with production
Excessive hole deviation
Down hole surveys, better operator skills, laser alignment
Not following design
Efficient supervision
Not drilling to required depth
Efficient supervision
Poor workmanship due to bonus driven
There may not be a short term solution
Explosive malfunctioning
Review pattern
Area of low or high powder factor
Use modelling blasting software
Fall–off
Less aggressive design?
Inability to establish failure triggering mechanism
Use information from seismic system
Orepass hang-up
Limit intake size (use screen)
Large fragmentation/fall-off
Optimize drilling and blasting
Long tramming distances
Improve block design
Poor ventilation
Review ventilation system
Poor reporting practices
More personnel training
Poor drawpoint condition
Support and reinforcement
Continuous fall-off inside the stope
Exclusion periods
Ability to survey as stope is extracted
Communication with survey department
Limited access
Establish stope access doors
Poor ventilation, laser beam can not shoot through
Improve ventilation
Fall-off may damage equipment
Wait until ground stabilizes
Stope design
Drilling and blasting
Mucking
Stope survey
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8 STOPE PERFORMANCE A stope performance review is undertaken as a technical audit of a stope design process. The review is performed during the stope extraction (after each firing) to monitor the conditions at the exposed stope walls, including backbreak, underbreak and broken ore fragmentation. The purpose of the review is to determine any variations from a planned stope design extraction strategy. To achieve this, a series of stope surveys can be carried out after each significant firing, and also following the completion of all firings (See Figure 7).
Figure 8. Production profile from a high lift bench stope
Figure 9. Cumulative plot of time vs volume for fired and mucked volumes. Figure 7. Longitudinal section view of a large scale bench stope showing consecutive surveys indicating minimal backbreak.
The performance review provides a mechanism to record the observations from operators and technical personnel in order to indicate problems and successes during stope extraction. A database that highlights lessons to be learnt and improvements to be made can be adopted for each stope. Table 3 shows some of the typical problems and possible solutions (by no means exhaustive) encountered in open stoping. In addition to those problems, stopes left open over long periods of time may be influenced by time-dependent regional fault behaviour. Stress re-distribution, production blasting and backfill drainage from adjacent stopes are likely to influence stope stability over a period of time. Blast damage and the effects of water from backfill can be transmitted along common fault structures intersecting a number of stopes. Instability may create difficult remote mucking conditions due to large material falling off into the stope. These delays (stope production tails) actually extend the stope life, which in turn may contribute to more overbreak and more mucking delays. Production profiles are usually shown as histograms of mucked volume on a daily basis. The data in Figure 8 show that long-hole winzing (or any re-slotting) actually slow down productivity. Since dilution is defined as any material that is extracted beyond the boundaries of a designed orebody outline, a comparison of mucked versus designed volume can be used to estimate dilution as shown in Figure 9.
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With the advent of the Cavity Monitoring System (CMS) stope survey technique (Miller et al., 1992), information about the actual variations from a designed stope shape can be routinely obtained and used analytically to calculate dilution, depth of failure and to determine structural control by large faults at the stope boundaries. Contours of depth of failure can be determined by filling the CMS wireframes with blocks and using the stope orientation information to orient the block model such that Y direction of the blocks is perpendicular to the hangingwall, the X direction parallel to the strike and Z direction parallel to the dip of the stopes as shown in Figure 10.
Figure 10. A CMS wireframe filled with 0.25m x 0.25m x 0.25m blocks. The block model can then be interrogated using the lode hangingwall and footwall and the CMS wireframes. The blocks inside the CMS wireframe, yet outside the lode hangingwall boundaries (depth of failure) need to be determined. Once the thickness for each column of blocks in the Y direction is calculated, the information can then be contoured using 0.5 metre intervals as shown in Figure 11.
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Figure 11. Longitudinal view of hangingwall depth of failure contours showing structurally controlled failure
Information from failure depths can be used to compare stope performance between double and single lift stopes for a similar range of strike lengths and rock mass conditions (See Figure 12). Back analysis of CMS data can be used as a diagnostic tool to identify stopes where blast damage may be causing early failures as shown in Figure 12(a). The stope highlighted by a large circle shows a depth of failure that is not in accordance with the other stopes of similar size and shape at this particular mining operation. The data in Figure 12(b) show that for this site similar depths of failure were experienced within the short stopes (25m high down dip) compared with the large stopes (50m down dip). A stronger geometrical control on the behaviour was experienced within the large stopes, where a range of failure depths can be established for stopes having a similar strike length. The depth of failure within the short stopes was controlled by factors other then geometry, such as blast damage or time dependency. The depth of failure increases sharply when the hydraulic radius exceeds 8 as shown in Figure 13. The depth of failure in the stope footwalls is not controlled by stope geometry. The stope performance can also be quantified by plotting depth of failure versus critical span as shown in Figure 14. The economical impact of dilution can readily be linked to depth of failure. The larger the critical span for this particular
Figure 12. Depth of failure for different hangingwall stope geometries.
Figure 13. Depth of failure for different hydraulic radius 102
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Figure 14. Depth of failure for different spans within a shallow dipping tabular open stope operation. operation, the larger the failure depth. A reduction on the critical spans may require additional pillars (hence ore loss). The balance between additional pillars versus the
detrimental effects of failures (See Figure 15) can only be established using an economic model of dilution. In order to ensure that the actual stope performance information is used to the best advantage, and to improve future designs, the details of stope design and its underlying assumptions can be documented in a Stope Atlas, where the history of the stope performance is recorded from the initial firing through to final stope completion. The information contained varies depending upon the stoping practices at a particular mine site. The following issues may be included: • Geology Geological orebody model and interpretations, geological structural and rock mass properties. • Stope design Initial stope design geometry, documentation of changes to design parameters, the reason and the results.
Table 4. Suggested stope performance assessment summary STOPE PERFORMANCE REVIEW Stope name: Material
Designed
By:
Date:
Actual
Tonnes mucked
Ore (t) Grade (%) Internal dilution (%) External dilution (%) Underbreak (%)
-
Fill dilution (%)
-
Geology: The effects of major geological structures, rock types and properties Reasons for any difference between design and actual grade and tonnes Development: Problems and concerns regarding ground conditions Performance of ground support Drilling: Whether any holes or ring section could not be drilled as planned, set-up or deviation problems. Reasons for variation from design. Blasting: Any problems encountered with charging, firing or design sequence. The results of the blast, eg. Fragmentation, misfires, freezing of holes, induced failures Production mucking: Ventilation problems or otherwise with chosen circuit. Drawpoint and orepass conditions. Broken ore left in base of stope? Backfill: Condition of fill passes, filling times and cement ratios used, any problems encountered. Rock mechanics: Stope and adjacent development stability. Timing of failures, and features that contributed to dilution, effects of blasting, structure and stress. Exposure and stability of adjacent fill masses. Planning and design: General comments on original vs. actual extraction. Recommended changes to design procedure. Financial analysis of stope extraction. Massmin 2004
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• Stope summary A one page stope performance review for easy reference (See Table 4). 9 CONCLUSIONS The Stope performance is measured by the ability to achieve maximum extraction with minimal dilution. The key variable used to compare performance is depth of failure, which is calculated using CMS wireframes and the designed stope boundaries. The data show that depth of failure can be used to identify blast damage and other factors controlling stability such as time dependency. Depth of failure increases significantly when the stope size exceeds a critical value and can be readily used to develop economic models of dilution. Figure 15. The detrimental effects of stope back failure following stope blasting leading to ore contamination and ore loss. • Stope extraction Drilling and blasting practice, in-the-hole survey data and comparison to design, fragmentation assessment. • Stope performance Maximum spans achieved, stope survey (CMS) data, back analysis of failures, geotechnical information that contributed to understanding the failures, ground support performance.
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REFERENCES • Miller, F., Jacob, D. and Y. Potvin, 1992. Cavity Monitoring System: Update and applications. 94th Annual General Meeting. Canadian Institute of Mining and Metallurgy, Montreal. • Villaescusa, E., 1998. Geotechnical design for dilution control in underground mining. Mine Planning and Equipment Selection. Singhal R. (ed), Balkema, Rotterdam, 141-149. • Villaescusa, E. 2000. A review of sublevel stoping. MassMin2000, Chitombo G. (ed), The AusIMM: Melbourne, 577-590.
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Mining method selection for diamond mines - Challenges in the Arctic Jaroslav Jakubec, Chris Page, SRK Consulting, Canada Paul Harvey, BHP Billiton Diamonds Inc, Canada
Abstract Following South Africa and Russia, Canada has become the latest country to open an underground diamond mine. The Koala North underground mine, North America’s first underground diamond mine, formally opened in November 2002. The Canadian Arctic poses several challenges that are not common on mines located in warmer climates and the mining method selection process must take such challenges into consideration. This paper illustrates the decision making process used at EKATI Diamond Mine to select an underground mining method that is applicable for kimberlite pipes in Arctic conditions.
1 INTRODUCTION EKATI Diamond Mine was the first diamond mine in Canada. Since its opening in 1998, the mine has produced more than 23 million carats. Currently its annual output contributes to approximately 6% of world diamond production by value. EKATI Diamond Mine has another "first". Mass underground mining has found a new frontier in the Arctic with the successful commissioning and operation of the Koala North mine (see Figure 1).
Figure 1: Open benching method was introduced at Koala North kimberlite pipe. The photograph shows upper extraction levels The work carried out for the choice and design of the mining method at Koala North identified a number of significant differences between underground operations in kimberlite and those in other hard rocks. In addition to the specifics of mining kimberlite, the challenges posed by the Arctic conditions also had to be considered. These differences are due to the specific context of kimberlite rock mass properties, the typical evolution of mining in kimberlites and, also, the consideration of the very harsh weather and often the presence of permafrost The Koala North pipe has been selected as a trial underground mine for the purpose of testing mining Massmin 2004
methods and to provide access to the lower elevations of the Panda and Koala pipes which will also be developed as underground operations once the open pit mining is completed. This paper does not aim at detailed mining method description but rather tries to comment on the specifics of kimberlite mining in the Arctic context. It is restricted to the kimberlite pipes not to dikes or sills. 2 CHOICE OF MINING METHOD - BASIC CONSIDERATIONS The context for a mining operation is usually something that man cannot change: the in situ rock mass and geometrical characteristics of the economic mineralization, the weather, the topography, geographical location - are things that must be taken into account and the planned operations must be appropriate to this context and not continually in conflict. A mining method must be appropriate to the context of rock mass, orebody geometry and geology. Those are the conditions which cannot be changed and in which the mining method must perform. If the mining method is in conflict with the context then it will not perform as expected (and expectations usually come from other mines where the method is applied). In summary the choice of a mining method depends on the following: • Geology • Orebody size and geometry • Grade distribution • Rock mass competency • Disturbances • External constraints Geology - includes continuity and predictability of the lithology that hosts the economic material; whether the ore can be visually identified or whether cut-off is to a grade boundary. Geometry - actual mining shapes, size, attitude, dip; variability of short and long range geometry that might affect planned dilution Grade distribution - values contoured to different cut-offs; geometry and continuity of value shapes; value of dilution
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Rock mass - strength and competence of the rock mass for both orebody and host rock and affect on excavation stability (especially unplanned dilution) and fragmentation from blasting or natural caving Geological structure - major through-going, significantly weakening, structures that could affect very large excavations or mining directions and sequences Disturbances - in situ stresses, magnitude and orientation; presence of water that could decrease stability or cause wet muck rushes and/or protection of ground water; high rock temperatures or presence of permafrost External Constraints - production rate required versus the TVM (tonnes per vertical metre), rock mass conditions and mining method; protection of surface and avoidance of caving or limiting of subsidence.
Supported Methods These methods use back fill, either cemented or uncemented, to modify the size and shape of the open excavation or to provide a working platform. Open stoping - usually large excavations drilled with rings and filled with cemented fill. Although this method is widely used in the world, it is not used in diamond mining. Cut-and-Fill - there are several varieties of cut and fill which generally use un-cemented fill with varying amounts of back support; it is notable as a "man-entry" method and is sensitive to the quality of the ore rock mass. The varieties of this method include: • Conventional cut and fill • Post-Pillar cut and fill • Bench-and-Fill
3 BASIC MINING METHODS - REVIEW
Drift-and-Fill - usually applied where the rock mass is too weak for conventional Post-Pillar cut and fill. Excavations are mined and then filled before starting the next excavation. Depending if the mining progresses up or down, overhand or underhand cut and fill could be developed.
Most of the underground mining methods can be grouped into three basic categories: • Caving methods • Open stope methods (unsupported) • Supported methods 3.1 Caving Methods Natural caving This includes a group of methods that rely on the rock mass that will cave under gravity and/or stress. The method has several variations depending on orebody size and geometry and on the rock mass quality. Block Caving - applies usually to smaller ore bodies where the whole footprint is caved at once. Panel Caving - all the major caving mines use panel with a caving face that progresses across the footprint of the orebody. Inclined Caving - where there is a significant dip or plunge and the production level can be fitted to the plunging geometry with a "Sub-Level Caving" type layout. Front Caving - applied where the rock mass is too weak to maintain the production development over a large area or in steeply dipping ore bodies. A single line of draw points is developed, caved and exhausted before starting the next line. Induced Caving This includes methods where the ore might be too strong to cave (but the hanging wall will cave, or where extraction is developed directly from surface or from an open pit floor); or where the fragmentation from natural caving is too coarse. Sub-Level Caving (SLC) - very regular pattern of diamond shaped blast hole rings. This method is typically used in competent rock mass. Blast or Hydrofractured Assisted Caving – this method had limited use in the past on some of the De Beers mines. It was also used at Northparkes mine but as a remedial measure not as a primary mining method. New advances are currently being made in this field in South America and Australia. Open Stoping with mass blasting of pillars - typical open stoping geometry with an array of pillars left and then massblasted and most of the draw is under choke conditions. 3.2 Open Stope Methods These include methods where the stoping excavations are un-caved and unfilled; excavations must remain reasonably stable usually for the life of the mine. Room-and-pillar - regular or random pillar layout in flat lying geometry usually less than 15º to 30º; method very sensitive to ground conditions Open stoping with permanent pillars - applied in steeper deposits where room and pillar cannot be operated.3.3 106
4 KIMBERLITE CONTEXT Someone once said "kimberlite is not a rock"! And there is some substance to this statement. Kimberlites, although they often appear strong and competent, can weather and lose strength when exposed to moisture. They can also be very absorbent to blasting and therefore difficult to break. Kimberlites generally behave very differently to other hard rocks and the most significant kimberlite issues that could influence mining method selection are discussed in following paragraphs. 4.1 Country Rock Most of the known kimberlite pipes in Canada’s Slave Province have intruded into strong and competent granites. This high contrast in rock mass properties between granite and kimberlite could be both a bonus and a possible problem. When the kimberlite is excavated, the stress balance is upset and kimberlite can yield to re-distribute the stresses back into the country rock. The bonus is that rock mass dilation is minimal due to the stiffness of the material and once the stress is re-distributed the pipe walls can be very stable allowing maximum recovery of the kimberlite ore with relatively little dilution. 4.2 Pipe Geometry The geometry of kimberlite pipes is relatively unique with their carrot shaped bodies. There is often increased complexity of the root of the pipe and a single pipe can break into separate zones with quite variable geometry. There is always a risk of assuming that the pipe geometry is regular where there is not enough pierce points. Often there are significant overhangs of country rock or gouges where the kimberlite has protruded into the pipe walls; as well, extremely large fragments of country rock can be slightly detached from the pipe wall and be surrounded by kimberlite. 4.3 Pipe Contact Zones The kimberlite pipe contact zones in the mining context are disturbed rocks adjacent to the pipe. Two types of contact zones can be recognized; the internal pipe contact zone within the kimberlite body and external pipe contact zones located in the country rocks (Jakubec, 2003). Both contact zones could have significant impact on a mining operation from the point of view of stability, cavability, wet muck flow and dilution. It is not always easy to recognize the contact zones in the drill core and as a result they could be
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overlooked when considering mining method choice and designs. 4.4 Kimberlite facies and Inclusions There can often be several facies within the same pipe with very contrasting rock mass properties and susceptibility to weathering. This could pose difficulties for selecting a single mining method. Kimberlites often contain varying amounts of country rock as a result of emplacement and formation of the pipe. These waste xenoliths can sometimes be extremely large in size and can have significant affects on excavation behavior and could pose problems for caving or material handling. 4.5 Kimberlite Weathering The susceptibility to weathering can impose constraints on operations and mine design. Where rapidly weathering kimberlite is present, no water should be allowed to come into contact with kimberlite. Exposed kimberlite surfaces might have to be kept completely dry with the application of sealants and the use of dry drilling. This could pose a significant problem in an Arctic environment since there are no commercially available sealants for wet conditions or very cold climate.
Freezing muckpile - broken rock (ore or waste) can freeze after blasting and may require re-blasting before mucking. Also, snow could accumulate in open drawpoints during the storms. Freezing of the cave material - free caving material can subsequently be wetted during thaw and then be re-frozen sufficiently solid that it causes caving problems or bridging that interrupts the natural flow of broken material. Icing-up of blast holes: - ice build up in the drillholes was experienced at Koala North during the freshet and freeze up season and requires re-drill. Stabilizing effect of frozen pipe walls - during the winter months exposed pipe walls will be frozen and very stable. De-stabilizing effect of ice jacking – freezing of water in the open joints will cause volumetric expansion and movements of the loose blocks. Subsequent melting could destabilize the blocks resulting in a rockfall.
Figure 3: Snow accumulation outside the drawpoint could pose operational problems. Figure 2: Photo illustrates rapid weathering and strength deterioration of clay rich kimberlite. Another issue could be floor deterioration resulting in very poor tramming conditions for mechanized equipment. However, excellent running surfaces could be obtained when the kimberlite drawpoint roadways freeze with impacted ice in winter. Lastly, the kimberlite susceptible to weathering could form mud resulting in conditions that give rise to mud or wet muck rushes into the underground excavations. 5 ARCTIC CONTEXT The Artic imposes another layer of complexity when considering the mining method selection. The main issues are associated with: • The frigid subzero temperatures and extremely challenging weather conditions • Permafrost which can extend to over 400m below surface in certain parts of the Arctic In the context of EKATI, the permafrost is usually located only in the country rock as the softer, more erodable kimberlite often results in a shallow lake below which the ground is not frozen – a talik zone. The permafrost can effectively prevent the flow of ground water and significantly improve the rock mass competency The impact of the freezing temperatures and permafrost can impose the following design and operating constraints: Massmin 2004
Fogging - affect of very cold air coming into contact with warmer and moist conditions, especially in a draw point or in a decline. This could have a significant impact on haulage and people and material movement. Trafficability - build-up of ice on roadways, especially at the portals to declines and in the open drawpoints, requires special and continual attention to prevent slippery conditions. (At exhaust shafts, to such an extent that could constrict the opening and flow of air.) For example, at Polaris Mine, located in the Canadian high Arctic, damp passes could freeze completely in a matter of weeks. Effect of underground cold on productivity - the effectiveness of both men and/or equipment can be affected if air must be kept below freezing to preserve the strengthening effect of the permafrost. Shotcrete mix for cold climate - special additives are often required to ensure that shotcrete will set. Constituents have to be stored above freezing before use and heated water or heated aggregates and brine is necessary. Similar considerations are required for grouting of rock reinforcement. Brine drilling - if rapid weathering rocks are not present and dry drilling is not possible then brine is necessary to combat freezing. 6 DECISION PROCESS FOR MINING METHOD SELECTION The decision process that was used for identifying the most suitable mining method was based on the following fundamentals:
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• Must be safe - no compromise on safety • Technically viable - work more or less as planned - one can never be sure but the method must be appropriate to the in situ context • Economically viable - this might be obvious but it is important to take into account what might go wrong rather than assume that everything will go as planned (often "as hoped") • Should have sufficient flexibility within the method to be able to manage problems (in underground mining results seldom go entirely according to plan and problems need to be anticipated and solutions in place) The decision process can be divided into two main steps: • First, ensure that the concepts on which a method depends are technically viable; the method must work approximately as designed. • Second, that the mining rate and grade deliver a value well in excess of the cost. There must be sufficient "room" to operate at the desired production rates and that dilution can be managed. In most methods the dilution can be decreased at the expense of recovery and recovery does not usually have a significant impact on the net present value. It is very important to keep a close watch on the risk/reward relationship; a formal risk analysis cannot be left until the end of the study; it is a continuing process. Technical Considerations The dominant technical considerations that have to be tested during the process are: Caving - If caving is necessary then it must cave: there are several recognized methods for estimating caving and empirical information database. Stability - If a stable face is needed (open benching), or stable open stopes are necessary, or the economics are sensitive to the cost of ground support then stability is an issue. There are again several recognized approaches that can be used for stability assessment but the overriding consideration for kimberlites is the question of weathering on exposure and, in the Arctic conditions, the potentially strengthening effect of permafrost. Fragmentation - If the orebody will cave and is sufficiently competent for a caving layout then what will be the fragmentation? The estimation of fragmentation is not nearly as well understood as rock mass stability and cavability but methods are being developed to get better estimates Mud rush - If a natural or pre-break cave is feasible but the rock mass in the cave or in the country rock is susceptible to weathering and degradation, is there a potential for mud rush and can the mud rush potential be managed? Rock masses that are insufficiently competent for caving layouts are unlikely to be sufficiently competent for open stoping with cemented back fill either; in this case the methods like drift-and-fill have to be considered. Howeve,r such methods have extremely high cost and very low production rates and are unlikely to be economical in the Arctic. If either open benching or caving is an option then choosing between the options is now a question of operational achievements. Operating Considerations As we said before: safe and economical. The economics are mostly a question of managing grade (dilution) and production rate (having the mining room and understanding the things that might go wrong and reduce the planned rate). These are the operating considerations as outlined previously. Dilution - In most kimberlite pipes the dilution comes from the wall of the pipe (in larger pipes there are occasions 108
when some of the kimberlite facies are also low value and form dilution); in the Canadian Arctic the waste is usually a very competent granite type rock; the stability and, therefore, the amount of waste material that will fall into the space left as the kimberlite is extracted will depend on: • The thickness of the contact zone within the wall rock; the contact zone will almost definitely fail and is the minimum amount of waste material that will fall into the excavation • Overall slope stability of the exposed wall rock beyond the contact zone; a function of rock mass competence, through-going weakening structures and geometry standard slope stability considerations • The enhanced stability initially given by permafrost must be considered especially if the freezing conditions mean there is slope pressurization from ground water • The action of near-surface ice-jacking from freeze/thaw cycle Operating Issues - The operating factors are nearly all a question of the challenging cold: • Freezing muck piles: muck-piles that are moist and are not immediately moved can freeze and require re-blasting • Icing-up of blast holes • Fogging: mixing of very cold and heated and moist air usually at the draw points: a question of direction of flow and pressures • Ice build-up: at the entrance of declines and ventilation shafts: simply a question of operating procedures • Freezing of cave: limited or extensive freezing of the cave could interrupt the flow of material and could result in major, high hang-ups • Productivity of men and equipment: there appears to be adequate information from mines like Polaris, Nanisivik and Raglan, and now Koala North for reasonable assessment of this issue 7 MINING METHOD BASICS – EKATI CASE STUDY As a result of the decision process four basic methods can be identified as candidates in a possible order of increasing costs: • Natural caving • Open benching • Pre-break caving • Supported methods with cemented fill It should be noted that currently the only significant underground diamond mining experiences are from De Beers operations, mainly in South Africa. The methods that are, or were, used in any significant way include natural caving, VCR assisted caving, open benching, and sublevel caving. Besides the International kimberlite pipe in Russia which uses the drift and fill method to combat poor ground conditions and some trial developments at Aikhal kimberlite pipe, there are no operations using back fill in kimberlite. Some of the operational issues specific to kimberlite mining in Arctic conditions for the four selected methods are discussed below. 7.1 Natural Caving The natural caving is low cost (although up-front capital can be high) and is the best for managing dilution through good draw control. If the method is feasible (can achieve a sustainable cave, can maintain a stable production level and oversize and hang-ups are manageable) then operating problems specific to the Arctic environment are possibly limited due to the considerable height of caved material overlying the production level for much of the cave’s life. • Mud: many kimberlites can form mud; the build-up of snow and ice in the crater and the thawing of this material could cause problems
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• Freezing of cave: if the ice in the base of the cave forms the equivalent of a cirque (glacier in a depression) and does not thaw during summer this might interfere with the cave. 7.2 Open Benching Open benching, if properly operated, is a cheaper method than SLC as it does not operate under choke conditions and therefore the drilling and blasting and development layout does not have to be as tight. As the name implies, all the blasting faces are open and a certain level of stability is required: • Remote control loading: much of the draw would be in the open and operators would have to stand in a safety bay to guard against rocks falling from the pipe walls; but with the advances in tele-remote this is no longer an issue. • Blast hole loading: the open draw points pose a risk; can be modified by having a number of rings pre-loaded or as in case of Koala North a scoop-movable, heavy duty steel barricade was used in the drawpoint before charging (see Figure 4). • Open bench stability: individual face heights might need to be limited in order to maintain overall face mining front stability. • Loss of drill holes: damage due to blasting, freezing closed or potentially weak ground • Freezing of muck-piles: very cold temperatures combined with wet rock • Coarse fragmentation: small blasts in potentially damaged and unstable faces; need for remote secondary blasting • Ventilation: open draw points; extreme cold in the work place impacting on productivity, equipment selection and requiring regular breaks and dedicated warming facilities u/g. Also need to regulate air onto levels and install air dryers on compressors to ensure no moisture that can freeze in compressed air lines • Loss of brows: over-break of brows and need for charging and hole cleaning in an open draw point; precharging is probably advisable if explosives can resist freezing temperatures • Pipe wall stability: the pipe walls and contact zones may be a source of loose material and/or massive failures; dilution and possible risk at open draw points If open benching does not proceed as planned then SLC might very quickly become a more predictable and therefore more effective method. The good results that have been achieved by Koala North are encouraging.
7.3 Pre-break Caving Sub-level caving Sub-level caving is a choke method and it relies on broken material to confine the blast and keep the broken ore near to the draw point. Although it is considered a high dilution method in the context of narrow kimberlite ore bodies with strong country rocks that will not cave this may not be a consideration unless the grade carrying kimberlite does not go all the way to the surface. The use of SLC may overcome some of the concerns with open benching, such as: • No requirement for remote control loading • No need for remote secondary breakage (both these two could be an issue if kimberlite is "sticky" and doesn’t flow well) • Draw points always full of muck so less problems with ventilation • No open benches thus no risk of rock fall into the drawpoint SLC has some of the same problems as open benching such as concerns with brow stability and loss of drill holes. There are also some additional concerns: • Mud: because the SLC is a top down method the risk of wet muck flow is higher than in block cave. This is also supported by De Beers’ experiences in South Africa. • Freezing of cave: Wet broken material could freeze, bridge and interfere with the draw. Blast assisted caving VCR type (or other types of layouts) caving assistance might be best considered under the management of cavability for natural caving methods. The method was successfully introduced at De Beers and Bultfontein mines in Kimberley (Granger, 1992). In the context of kimberlite the VCR can "suffer" from the same problem as any method relying on blasting, and that is drillhole stability in poor quality ground. Since the method does not communicate with the surface there are no Arctic specific issues for kimberlites that are in talik zones. Kimberlites located in permafrost could have potential challenges associated with stability and effectiveness of explosives. 7.4 Supported Methods with Cemented Fill In a situation where caving is not feasible and where the economic material does not go through to surface, then open stoping with post fill is a possibility. This is probably the most widely used mining method outside diamond mining and needs no further explanation here. Apart from the same operating issues that would affect all methods, the one particular issue affecting the backfill method in the Arctic context would be risk of freezing the fill. If the fill material is below freezing and the kimberlite orebody is in permafrost or – more likely – the material has to be cemented; back fill would have to be kept in storage to be above freezing and possibly special additives might be required. Polaris successfully operated with cemented fill and therefore there are solutions to these issues. 8 CONCLUSIONS
Figure 4: Protective steel barricade was used in the drawpoints during the blast hole charging. Massmin 2004
The mining method selection process for kimberlite pipes located in the Arctic poses numerous challenges that have to be considered. Without taking kimberlite characteristics and severe weather conditions into context the mine design could fail. The Koala North project using "open benching" is a successful outcome of a decision process that BHP Billiton Diamonds went through to reach a compromise between
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the certainty of operational outcomes and economics whilst ensuring no compromise on safety. ACKNOWLEDGEMENTS The authors would like to thank BHP Billiton Diamonds for permission to publish this paper. Also, the help of Tyla Hay of SRK in making this paper a reality is greatly appreciated.
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REFERENCES • Granger, Q.P., 1992. VCR - assisted Block Caving – a Viable Future for Kimberley Mines. MASSMIN 92, SAIMM, Johannesburg, South Africa, pp.11-19 • Jakubec, J., 2003. Role of Geology in Diamond Project Development. 8th International Kimberlite Conference Abstracts. Victoria, Canada
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In Situ Leaching as a mass mining method Steve Hildebrand and Dan Ramey, CRCMining, UQ Experimental Mine, Isles Road, Indooroopilly, QLD 4068, Australia And Mining Solutions, Inc., Tucson, AZ USA
Abstract In situ leaching is a proven mining method that has been used for over forty years within the USA. It can be considered as the primary method for mining or it can be used as the final stage in the extraction history from an ore body, much like salvage mining. When an ore body is appraised for mining, generally open pit mining and underground mining techniques are considered. The historic evolution of the minerals industry has led to the mining of the highest grade and most accessible deposits first. As the world continues to mine out its shallow and high-grade ores, deeper and lower –grade ores are being accessed at ever increasing expense. More energy at an ever increasing cost is required to process deeper and lower grade ores. In situ mining represents an opportunity to successfully employ mass mining methods that can achieve the goal of low cost production from low grade and/or deep ore bodies. In addition in situ mining can increase the consumptive efficiency of ore bodies that are currently using conventional mining methods during their life cycle. The mining operations at San Manuel, Arizona are an example of using in situ technology for mass mining. The San Manuel Mine was comprised of an underground sulphide block caving mine, an oxide open pit mine and an in situ operation located between these two larger production entities. The combination of these three mining methods increased the consumptive efficiency of the ore body. In situ mass mining allowed additional mineral resources to be placed into production and transformed a mineral resource into a mineral reserve.
INTRODUCTION
THE CONCEPT OF MASS MINING
In situ mining technology has been used in the metals, energy and industrial mineral segments of the mining industry for many years. In this discussion of in situ leaching as a mass mining technology, the discussion will confine itself to comments about resources within the copper industry. The copper industry has been, and remains, an increasingly competitive environment and the world price is a dominate influence on the industry. The ability for a mining company to remain economically competitive in the current copper market depends on its ability to generate cash flow and returns on investments, even during periods of significantly depressed copper prices. Due to the influence of the world copper market, the copper industry is forced to have a flexible cost profile, which allows greater control over the major cost components used to produce copper. Escalating mining and recovery costs require methods that are more efficient and responsive in order to remain economically competitive within the world market. Historically, most successful companies have relied on applying traditional technology to rich ore bodies. As time progresses, the largest, richest and most accessible of the known copper ore bodies become depleted. However, a great deal of copper mineralization known to exist is too deeply-buried and/or low-grade to be mined economically using conventional techniques, which involve massive material handling and processing. However, the leaching potential of a particular resource is often an under exploited resource, to which most copper companies need to examine and develop to its fullest potential. The ability to access these resources depends on the development of alternative mining technologies. The application of in situ mining technology is one method of accessing the full potential of these resources.
It would be presumptuous to offer a definition for a concept that is in common usage today in the mining industry. However a few comments are worth while. A distinction of mass mining is that it must be a continuous process of mining which provides a consistent feed to a reduction plant, for example a concentrator. The concept of mass mining applies to underground production whose costs profiles must be comparable to those obtained in large open pit mines. Gideon Chitombo1 states that "The holy grail of mass mining is to produce large tonnages of ore at a similar cost to a large open pit mine," Within the copper industry, mass mining can apply equally to sulphide or oxide resources. The methods of mass mining underground are known widely and include block caving, sublevel caving, panel caving, and others. The mass mining process flow is coupled to a reduction unit such as a heap leach or a concentrator. In most mines the end product of production is measured in tonnes of ore. For instance, the Magma Copper Company’s San Manuel Mine in Arizona had an underground block caving sulphide operation that was rated at greater than 55,000 tonnes per day (tpd). If this was an oxide operation, it would have delivered tonnes of ore to a heap leach operation, not a concentrator. However if we are to consider in situ leaching as a method of mass mining, the end product of mass mining must be presented differently. With in situ leaching (ISL) the end product is a pregnant leach solution (PLS) with a particular copper tenor that is considered on a mass or load basis. With a sulphide or oxide underground operation, a mass or load basis can be calculated enabling a true comparison with ISL mass mining. A comparison of mass mining techniques can accurately be accomplished using daily production rates and unit cost of production. The following
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two tables using hypothetical tonnages and costs are presented to strengthen this comparison. A comparison of mass mining rates described on a daily basis is shown in Table 1. The information demonstrates that although the descriptive metrics for the mining methods are different they can all be compared on a daily mass basis.
Table 1 Mass Mining
Daily tonnes Or M3/hr
Copper Tenor
Daily Copper (kg)
Sulphide UG
55,000
0.65%
358,000
Oxide
90,000
0.35%
315,000
In situ
2,270
2.00gpl
109,000
If we assume typical costs for an underground sulphide block caving operation, an oxide open pit operation and an in situ leach operation on a daily cost basis, the values presented in Table 2 can be generated.
Table 2 Mass Mining
Total Daily Costs
Daily Copper lbs.
Cost per lb.
Sulphide UG
$550,000
780,000
$0.705
Oxide
$350,000
700,000
$0.500
In situ
$75,000
240,000
$0.313
Table 2 illustrates the cost relationship between 3 mining situations. Although the comparison is theoretical, the values presented in the two tables are based on the experience of the authors. Generally, the underground sulphide mine will have a larger work force and a physically larger reduction plant. An oxide mine will also have a larger work force, use heap leaching or vat leaching for dissolution of copper and higher costs. An ISL operation generally has a much smaller work force and the reduction plant is simplified to an SX-EW operation. The ISL mine will usually have a lower unit cost for a pound of copper than underground operations.
situ leaching techniques follows. Classifying ISL operations can be accomplished by determining whether the resource has previously been mined or not. If the ISL mine is virgin resource, it is called a "green-field" application, while if the ISL mine is following previous mining efforts; it is called a "brown-field" application. A brown-field application is a situation in which the remnant resource is mined with ISL techniques following open pit or underground mining activity. "Brown-field" applications have occurred in numerous sites within the copper industry. Magma Copper Company (and later BHP Copper) operated a successful in situ leaching operation at Miami, Arizona for over forty-five years. Block caving at Miami was curtailed over fifty years ago and the initial in situ solutions were treated in cementation cells for direct smelter feed. The San Manuel Mine, also in Arizona, operated a very successful in situ leaching operation in the remnant block caving ore body and the oxide resource left after the completion of open pit mining. The green-field applications of ISL technology have not been advanced into operations yet because most mining companies have not developed their ISL technology. Examinations of virgin resources have been primarily completed by evaluating the potential of open pit or underground techniques. However, recently there has been much greater interest within the mining community for examining new resources, that have never been mined, for the potential of open pit, underground and in situ leaching. A recent green-field property that evaluated underground, open pit and ISL mining technology was Poston Butte porphyry copper deposit at Florence, Arizona. Magma Copper Company and later BHP Copper examined the property and initiated a proof of concept test for in situ leaching. The results of the evaluations can be summed up by the abstract from another paper which states, "Low costs associated with in situ copper leaching provide the opportunity for developing copper resources which are uneconomic by conventional mining methods." In the future, new copper resources will be located deeper underground and generally have lower tenors. All three major categories of mining, underground, open pit and in situ leaching will have to be considered in order to make a thorough examination of a resource. These resources will be evaluated using a sequential methodology moving from very well known methods and costs to less well known methods. There is also the possibility to mix the major technologies if the resource is conducive to simultaneous development. "Target orebodies for mass mining can be technically challenging and require the application of ingenious operational practices to maximise extraction rates and achieve lowest possible operating costs while maintaining high levels of safety." This is a challenge for mass mining and ISL is a technique that meets the demands of modern mining.
UNDERGROUND ACCESS IN SITU LEACHING; A NEW DIRECTION There is another important distinction involving in situ leaching operations versus other underground mass mining techniques. All mass mining methods require physical underground access to the ore body. Access is also required by an ISL mine however, in most in situ operations, access to the underground ore body is obtained by drilling holes from the surface, which act as injection and recovery wells. There is no requirement for the work force to have physical access to the underground in an ISL mine. The elimination of the necessity to place personnel underground is an improvement from a safety point of view. CLASSIFYING IN SITU MINES If it is accepted that in situ leaching is a legitimate mass mining method, then a discussion of the types of possible in 112
Each in situ leaching opportunity requires a logical approach to the investigation of its potential, evolving from simple to more complex techniques. A quick and relatively inexpensive analysis of a resource, termed a portfolio analysis, examines the resource in a general way to quickly eliminate those resources which clearly demonstrate their unsuitability to ISL mining. If a resource is acceptable after the first analysis, the next investigation, termed a fatal flaw analysis, delves into the resource in more depth. The fatal flaw analysis considers those characteristics which are favourable to ISL or any which would be fatal and preclude any additional project development. The third examination, which is technically termed the critical element analysis, completes the formal examination of a resource’s potential for in situ technology. This investigation is
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performed with a thorough look at all major disciplines required for a successful ISL mine and includes a detailed financial examination. The owner will be able to make knowledgeable decisions concerning the potential of an ISL mine, balancing potential risk with potential gain. Whether to proceed with the development can be made at this time. The final phase in an in situ investigation involves a large scaled test which becomes the proof of concept examination. ISL MINES RESEARCH POTENTIAL There have been many successful in situ leaching mines, mining not just copper but also uranium and other industrial minerals. There is the potential to investigate other mineral leaching in the future. These different minerals will require some "proof of concept" testing either done through the auspices of a research centre or by interested mining companies. The equipment used typically with ISL mining operations is commonly accessible and has an extensive history of use. However, as this technology develops new methods and new technology will be developed such as the use of hydraulic pumps. Sensing probes are being developed that help move ISL mining from the status of an art, into a recognized technology. Drilling is a major cost component for an ISL mine, whether on the surface or underground, and research into methods and new approaches is ongoing.
CONSUMPTIVE EFFICIENCY In comparing various mining methods, there is an economic tool called consumptive efficiency that is useful. Consumptive efficiency can be defined as the measure of resources contributing to production and is determined as the percent of an ore body that contributes to production. All mines have economic limits that continuously change with time and market conditions. These economic conditions provide boundaries to the amount of ore that can be placed into production at any point in time. These boundaries separate economic ore from uneconomic ore, and as the resource is mined, move ore from a reserve to resource status. The highest consumptive efficiency of an ore body would involve the use of a single or multiple mining techniques which allows mining of the greatest amount of the ore body.
Figure 2 Aerial View of San Manuel Mine Massmin 2004
THE SAN MANUEL IN SITU MINE – A CASE STUDY OF AN ISL MINE The San Manuel Mine was developed initially in the 1950’s as an underground block caving mine of a massive low grade porphyry copper deposit. The deposit was sliced in half by a normal fault that displaced the down-thrown block a thousand metres to the West. Figure 2 is an aerial view of the Open Pit, production shafts and ISL mining location. Weathering of the upper half (San Manuel), which was closer to the surface, occurred over time eventually creating a copper oxide cap comprised primarily of chrysocolla. The down-thrown block, called the Kalamazoo, is primarily chalcopyrite. Average sulphide tenor in the San Manuel portion of the orebody was 0.65% copper with an acid soluble tenor of 0.45%. The underground block caving mined the lower portion of the San Manuel orebody for over 40 years before production was shifted to the Kalamazoo exclusively. In the early 1980’s during a slump in the world copper price, Magma Copper Company began an extensive investigation of the potential of exploiting the oxide resource. In 1985 several reverse circulation injection holes were drilled in order to verify collection points underground on 4 different levels. Drill holes were cased and a mild acidic solution injected. The location of PLS collected underground was documented as well as ground water dilution. Acid and copper fronts in the collection solution were verified. In 1986, a large scale test was directed by an organization called the San Manuel Oxide Committee (SMOX), composed of metallurgists, analytical chemists, and oxide personnel from the Magma mine. The constant overview by this committee was instrumental in determining the economic feasibility of the project. In addition the determination to initiate open pit mining of the upper portion of the cap originated with this committee. In 1987 and 1988, engineering of the underground in situ PLS pumping system was begun and constructed in a partnership between contractual construction workers and Magma miners. This project involved construction of three major underground pump stations, an independent electrical supply and a 100% backup source. A SCADA system was designed with each pump station and operated with a programmable logic controller. Two 12" 316L stainless steel pipelines were installed in an operating shaft at the mine site. This PLS handling system was designed for 1,817m3/hr This PLS handling system ran flawlessly for more than twelve years with only minor equipment maintenance requirements.
Figure 3 ISL Mine at San Manuel Santiago Chile, 22-25 August 2004
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In January 1995, open pit mining of oxide was terminated. A comparison of the unit cost of production between open pit mining and in situ leaching showed that in situ had greater economic potential. In late 1993 engineering for a surface PLS pumping system was initiated with construction and commissioning completed just prior to the termination of the open pit mining. Figure 3 shows the ISL mine within the San Manuel ore body. Much of the construction was completed in an active subsidence zone from the underground block caving activity. The ISL system was designed to handle 2,725m3/hr. Over 25,000 metres of HDPE pipe in various sizes and SDR’s was installed during this period. Figure 4 has a view of the in situ wells fields and the top two pump stations.
Figure 6 PLS Pumping Well
Figure 4 ISL well fields Figure 5 is a typical well field with both injection and recovery wells and using a production mode known as wellto-well leaching. All wells were developed in a similar manner so that changes in operation for hydrological reasons could be easily accomplished.
Figure 7 Copper Cathode for Export
The consumptive efficiency of the ore body was maximized by utilizing three unique mining techniques in the same ore body at the same time. The competiveness of the current copper market makes it important for companies to place consumptive efficiency as a goal in their strategic planning process. The ability to utilize more than one mining technique at an ore body is a proven method to attain higher efficiencies. In situ leaching can be, and will be, a mass mining technology that will allow many companies to extract additional value from their resources. In situ leaching is also an opportunity for research and development as the technique is utilized in the copper industry and with other minerals.
Figure 5 ISL Well Field Figure 6 shows a typical production well surface arrangement. The finished product from the SX-EW plant is shown in Figure 7 and was five nine copper (99.999%) pure, meeting all LME and Comex standards easily. CONCLUSIONS The experiences at San Manuel demonstrate that in situ leaching is a mass mining method. The daily production rates for the ISL mine, although lower than the other underground and surface circuits, were sufficient to influence the overall finances of the entire mining operation. 114
REFERENCE • Chitombo, Gideon, "Mass Mining – Challenge, Opportunties and Trends – Part 1, MiningNews.Net, 31 July 2000 • Ramey, Dan and Beane, Richard, "In situ Project Evaluation: Magma Copper’s Approach, COPPER 95 – COBRE 95 • Chitombo, Gideon, "Mass Mining – Challenge, Opportunties and Trends – Part 1, MiningNews.Net, 31 July 2000
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Chapter 4
Mine Planning I: Fundamentals
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Mine project life cycle Robin M Kear, Independent Mining Advisor
Abstract Significant projects are often undertaken in the mining industry. These projects can extend mine life by several years and involve considerable resources. Such projects are normally developed through several stages from conception to implementation and each of these stages has a specific function with a particular environment thus requiring different management and technical skills. In order to correctly manage and staff a project during these stages and also to enable an orderly progression through the project it is helpful to recognise the requirements of these stages. This paper examines a mining project from the delineation of a resource to implementation and attempts to identify the individual stages, the broad concepts that define them and the management issues that need to be considered. The views are based on the author’s experience on several mining projects. It is hoped that the paper may assist a reader in ascertaining their requirements when commencing or during a project execution.
1 INTRODUCTION Mining projects are normally highly expensive, have a long lead time and extend mine life for many years. The actual planning of a project can take considerable time and resources. Because of these factors it is important that a project moves through various distinct phases prior to construction. These phases should ensure that if the project is not viable, expenditure is curtailed as soon as possible or if viable, then the best possible return is obtained with the minimum risk. These project phases require different mind sets and consequently have different intellectual environments. This and the purpose of each phase are often misunderstood which can cause significant problems in bringing a project to a successful conclusion. For the purposes of this paper, a mining project covers the phases after the ore resource has been defined, either as an extension of an existing mine or as new ore body. This paper examines the phases through which a project moves and, should the mine prove viable, to the final construction and implementation. The paper is more a discussion of the various principles and concepts involved rather than a definitive study. 2 PROJECT CONCEPTION Most major mining projects will be one, or a combination of the following • A new ore body has been discovered. • An extension to an existing mine. • An expansion of, or a change to, an existing mine.
3.1 Strategic Environment This environment is required when the development of a strategy is required. The strategy is the broad plan required to achieve an objective. This requires "free" or lateral thinking and all possible scenarios, which could lead to the objective, need to be identified. Normally, for a mining project, the objective is to obtain the best economics from a particular resource. In summary "Define the Goal". 3.2 Tactical Environment In this environment the objective is to develop and implement the tactics required to achieve a strategic objective. In the mining sense this would be the procurement and utilisation of resources such as capital and labour to achieve the defined strategic plan. In summary "Achieve the Goal". The mining industry operates mainly in a tactical environment. A strategy is set and the organisation and management effort is concentrated on the best employment of the available resources to achieve the strategic objectives, such as production and construction targets. However, when a new project commences it is important for the strategy to be first determined and requires that the planning team shifts from the tactical to the strategic environment. This can present various problems especially as not all people are strategic thinkers, and even those who are will require some time to adjust to this environment. During the progress of a project the environment changes from strategic to tactical as shown in the following figure.
Within most mining groups the objective of a project is to increase the value of the company. In some situations other factors may enter the equation. It is therefore important to define what parameter or parameters the project must satisfy. Also the correct environment should be set up to ensure that project phases are correctly undertaken. 3 INTELLECTUAL ENVIRONMENT For the purposes of this paper, two environments will be considered. The first being strategic and the second tactical. The following broadly defines the differences between them. Massmin 2004
Figure 1 Environment over Time
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Strategic issues are most prominent at the start of a project falling off as criteria are defined. Tactical issues grow over time. Expenditure on a project normally follows the tactical curve, that is, increases over time. The following paragraphs broadly indicate the requirements of each phase and the transition between these environments. 4 PHASE 1 - CONCEPTUAL The purpose of the conceptual study is to determine if there is a potential project and if so justify the expenditure for a pre-feasibility study. A potential project would be one which meets the required selection criteria. Normally the selection is based on economics but could be on other parameters. This phase will indicate some of the major design criteria. Unfortunately, it is this conceptual phase that is often inadequately undertaken. With a rush for production, major criteria can be arbitrarily set to continue with present practice or something that has been observed on a visit to another property. Often a junior engineer is requested to do some basic calculations and if positive the study leaps into the detail. What is required is to stand back from the present situation and critically examine all facets of the situation. To do this adequately requires properly experienced engineers who have had the necessary exposure to the mining industry. Unfortunately engineers of this type are not too common and are often thought to be "too valuable" to "waste" on a study. However, this stage of the project starts to lay the foundations for subsequent work, has the most impact on the outcome and is therefore of the utmost importance. In order to successfully complete a conceptual study it should comprise of the following components. 4.1 Identification of Major Design Parameters All potential options must be looked at such as the mining method, production rates, major design boundaries and mine elevations. Where possible, constraints should be left out of the equation. This will enable the costs of these constraints to be evaluated when they are later applied. Assumptions should be positive and tests done to evaluate the effects of varying the parameters between the perceived confidence limits. These iterations should identify the major design criteria for the project. Various techniques are available to assist in these evaluations, one of which is the evaluation or economic surface analysis of the project. 4.2 Develop the Economics A cash flow model is required to assist in the evaluation of the various options. The model should be simple and easily understood. The purpose of the model is to allow for the ranking of alternatives rather than for absolute values. Checks should be made on the effect of different selection criteria to determine if these affect the choice of an option. If so then this should be discussed with the management to ensure that the correct selection criteria have been chosen. If the selection criteria are to be economic then it is normally best, at this stage, to use the Net Present Value (NPV) at the required hurdle rate rather than say the Internal Rate of Return (IRR). There are many views on this and the team must be sure that they are using the correct parameter for ranking of the alternatives. 4.3 List the Assumptions In order to complete this stage of the investigation various assumptions will be made. These will need to be confirmed during later phases and as such will be required to be included in the further schedule and budget.
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4.4 Identify Trade off Studies From the assumptions and other parts of the investigation various alternatives will be identified. These will require to be further investigated and incorporated into the on going schedule and budget. 4.5 Develop the Pre Feasibility Scope, Schedule and Budget If a potential project is identified then the conceptual study is required to justify the expenditure for the subsequent phase, normally a pre-feasibility study. A schedule and cost for this study must be produced for approval. The schedule should show the various trade off studies and data gathering exercises identified during the conceptual study. This conceptual phase should not be rushed and the design team should be small and comprise of senior staff. The staff should not have routine responsibilities as well as the project. Ideally the team should vary between the following numbers:Table 1 Phase One Manning Position
Maximum Number
Minimum Number
Project Manager
1
0
Mining Engineer
1
1
Mech/Elect Engineer
1
0
Metallurgy
1
0
Geotech/Geology
1
0
Accountant
1
0
Secretary
1
0
Total
7
1
Some of the disciplines may not be required to be full time but the team should have access to these skills for particular areas of the investigation. Routine responsibilities for the team should be as minimum as possible as these are normally tactical whereas the study is strategic. These modes require different mindsets and changing from one to the other is difficult. In the end neither is completed satisfactorily. A small team is required, as a conceptual study is extremely fluid with major changes occurring regularly. Each member must be aware of the latest status of these changes. After being involved in several projects the author has devised the following formula for strategic planning Kears’ Rule of Planning -- C proportional to n^2 Where n is the number of components and C is confusion. Although this may seem facetious there is some truth in an inverse relationship between numbers and the ability to be innovative, especially during conceptual planning. Obviously there should be some form of progress management during the study and the best would appear to be management by milestones. It is pointless to attempt to identify every task that must be undertaken, as the tasks in this stage will constantly be changing. To be continuously updating the schedule to reflect this will require enormous resources and probably all the team’s time. Rather identify the broad areas that need to be accomplished and when they need to be completed. This keeps the schedule simple, preferably on one page or at most two. The individual(s) will
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then know at what point these sections must be completed and adjust the investigations accordingly. Once a conceptual study has been correctly undertaken it is often observed that the study has raised more questions than it has answered. In fact, this is as it should be as various assumptions must be tested and trade off studies will be required in the following stage. The conceptual stage is strategic in nature and as such requires lateral thinking. Very little is routine and many changes will take place during this stage. Should the conceptual phase indicate that there is a suitable prize, under various assumptions it should justify the following study phase. However, if the project does not appear positive even under optimistic assumptions the project should be stopped. In this case it is important to document the work completed and also the major drivers which could change the outcome. This will make it relatively easy to recommence should one of the drivers change, for example the mineral price.
the work completed during this phase show that the project does not meet the requirements it should be stopped with suitable documentation to revisit the study should any of the main drivers change over time. Table 2 Pre-FeasibilityManning Position
Project Manager Mining Engineer Mech/Elect Engineer Metallurgy Geotech/Geology Accountant Secretary Total
Maximum Number
Minimum Number
1 2 2 2 1 1 1 10
0 1 1 1 0 0 0 3
5 PHASE 2 - PRE FEASIBILITY This phase of the project is to firm up and or confirm the issues identified in the conceptual phase and to justify the expenditure for the feasibility and detailed design stage of the project. The major design criteria should have been decided during the conceptual study. However, it is often the case that some parameters have not been settled due to lack of data, or further investigation is required. Therefore the pre feasibility study should be done in two stages. 5.1 Pre Feasibility Initial Stage This stage is required to obtain the required data and/or complete the investigations highlighted in the conceptual study. The assumptions made during the conceptual study should be tested and updated if required. There are often trade off studies to be completed and visits to other operations are useful at this point. A major problem, which may be experienced, is the lack of reliable non mining costs. It is sometimes supposed that a project is a mining project and therefore all the detail and planning is mining related. However, the design criteria can be hugely affected by the non mining costs and these need to be properly identified and quantified for the correct determination of the design criteria. This first stage remains strategic with a few tactical issues being required. 5.2 Pre Feasibility Second Stage Once the assumptions and trade off studies have been competed it is then required to commence detailing the design criteria, develop a schedule of resources and budget for the Feasibility stage. More detailed cash flow models are required and a detailed justification for the feasibility should be developed. Up to this point study expenditure and manpower requirements are normally not high. The exception would be data capture, for example a drilling program. Once the project moves into the following feasibility and detailed design phase considerable resources, both manpower and money, are required. For this reason a detailed and reasonable accurate justification is required prior to committing these resources. It is important that all major issues have been resolved at the completion of this phase. This second stage has a reduced strategic component and the tactical issues are beginning to become the predominant ones. At this point the team will possibly expand and typical numbers are shown in the following table Control of this phase can still be on milestones, but these will be more specific than in the previous phase. Basically this phase, collects data that is required, completes trade off studies, confirms that there is a viable project and if so details and justifies the next phase. Should Massmin 2004
6 PHASE 3 - FEASIBILITY AND DETAILED DESIGN The purpose of this phase is to provide an estimate of the required accuracy to justify the major capital required for construction and implementation. Provided that the previous phases have been satisfactorily completed, the project at this stage should have a comprehensive design criteria document. All major decisions should have been made. Decisions, which are required in this phase of the project, should relate to the specific system under design with no or almost no impact on other systems. This phase will require considerable manpower, either in the team or contractors. If major changes are made during this stage there is a very real danger that team members will be working with obsolete design criteria with the possibility of serious flaws in the design and the associated money and timing consequences. If the previous phases have been properly completed then this stage is largely tactical. The design criteria having been set now need to be incorporated into the design. Major decisions should not be made during this stage but rather the effort should be directed at producing the required detail to give an estimate of the desired accuracy. This stage is largely tactical, with the management concentrating on using the available resources to produce a design to the specifications of the required accuracy. Continuity between this and the previous phase is required and a properly convened steering committee is one technique to ensure continuity by the inclusion of the conceptual and prefeasibility team members. The high costs of this phase are mainly due to the number of persons required with typical numbers being shown below
Table 3 Feasibility Manning Position
Project Manager Mining Engineer Mech/Elect Engineer Metallurgy Geotech/Geology Accountant Secretary Technicians
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Maximum Number
Minimum Number
1 3 4 3 2 1 1 300
0 1 2 1 1 0 0 100
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7 APPROVAL Whilst each phase of the project requires approval, the move from the feasibility and detailed design to the construction phase is a major event in the project development. Considerable sums of money are involved, not to mention that the mines’ future for many years is probably decided at this point. The time required to obtain approval should be included in the schedule as it can take several months to obtain this approval. 8 PHASE 4 - CONSTRUCTION The construction phase should have a dedicated management team which should be different from the design team. This phase is definitely tactical and requires good control and discipline. It is essential that both schedule and cost tracking systems are in place before commencing construction. These systems should be able to show the original schedule, possibly a revised schedule and the actual to date with variances highlighted. These systems should be installed and demonstrated before any activities or expenditure take place. The design team is still required to continue with details. When approval has been granted it is normal that only 30%40% of the detailed design has been completed. Therefore an excess of 60% of the design still has to be completed. This should not be left to the construction team to manage and hence the design team will still function well into the construction phase. Control of this phase is important firstly to ensure that what has been designed is constructed and secondly the quality of work must not be an impediment to the future production. The Steering Committee becomes much more of a control issue and should have members from both the design team and the future production team. The committee should not be under the control of the Project Manager. 9 SUB PROJECTS Often, several possible improvements to the approved design have been identified but rely on new technology or systems still under development. Whilst the approved design should be based on practical and achievable systems there is often a window of time available to prove up these newer systems and include them in the project. The cut off times for proving these systems should be determined and separate project teams be convened to undertake this work with their own schedules and budgets. The management of these teams should not fall under the
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construction team as they will be focussed on the construction. Rather it is suggested that either these fall under the design team or preferably under their own management structure. 10 IMPLEMENTATION - COMMISSIONING Major projects most often do not just switch on. Systems require time to bed down, people require time to learn new techniques and, especially with mining methods, time is required to build up production. If sufficient time is not allowed for these activities the pressures on the team will often dictate the use of "short-cuts". These "short-cuts" can be extremely detrimental to the long term viability of the mine and should be avoided. The best avoidance strategy is to allow a realistic time for these activities and ramp-up. 11 SUMMARY This paper has emphasised the conceptual phase of a project. This is because the concepts set the foundation for all further work. Properly done the other phases flow from this stage and the team will be confident that the correct criteria have been set. In practice and for various reasons it is often the pre feasibility or feasibility study that is used as the start point with the major criteria being arbitrarily set, this does not often lead to the best utilisation of the resource. A mine has only one resource and can mine this resource once. There is only one method that will produce the best economics for this resource, others may be economic but of lesser value. In view of the magnitude of the capital normally required for a mining project and the usual life of a project of many years the effort in determining the best project is well justified. ACKNOWLEDGEMENTS There are doubtless many papers and textbooks which cover the management of a mining project. The author readily admits that these will probably give the reader much greater insight into the issues involved however they tend to deal with the management of a project whereas this paper has attempted to highlight the issues to be managed. This paper has been the result of the observations made on many mines and the author would like to thank his many clients for the ideas and discussions over the years. The author stresses that this paper is based on his observations and conclusions but hopes that these will form a basis for discussion and may assist in the formulation of a suitable strategy for a mining project.
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Block caving planning under metal price uncertainty Sergio Fuentes S., Mining Civil Engineer, Universidad de Chile, M.Sc. (Eng), Queen’s University, 2003, Director, Metálica Consultores S.A., Chile
Abstract The common practice in mine planning, irrespective of mining method, is to make decisions based on deterministic future metal prices. Later, during the economic evaluation of the project, stochastic tools are employed to identify and quantify risk factors such as metal prices when mine design and planning has already been completed. Given the capital-intensive nature and long preproduction periods of mining projects, especially in block caving, incorrect price assumptions present a significant source of uncertainty to project feasibility. This proposal outlines the introduction of a stochastic simulation method in the primary phase of mine planning, assigning metal prices based on probabilistic distributions. Utilizing this method combined with standard procedures for defining mining areas in block caving permits characterizing and prioritizing ore reserves based on price uncertainty, giving an additional decision tool to mine planners.
1. INTRODUCTION
metal scenario, including identification and quantification of associated risks at an early stage of the planning process.
Conventional mine planning encompasses all or part of the development of a mine, including feasibility, profitability, detail development, scheduling of extraction sequences, operation of equipment and transportation. Mine planners usually estimate the different parameters used for the planning process from standard procedures available in the engineering field such as cost information databases from similar operations, experience and engineering judgement. This standard approach is usually based on the ‘state of the art’ associated with a given mining method. Nevertheless, these procedures, databases and experience are usually incomplete or have a large degree of uncertainty related to the nature of basic estimations on such as the heterogeneity of the rock mass, geological environment and etc. These uncertainties are present in every stage of the mine planning and design process and include: • Definition of the geological model. • Ore resource or mineral inventory estimation. • Rock mechanics behaviour of a design. • Metallurgical performance. • Operational performance. • Cost and economic estimations. • Local political conditions and regulations. • Market product price forecasting. • Environmental issues. From the above abbreviated list, it is readily apparent that mining, and especially block caving, is inherently risky. Traditionally, the decision making process under uncertainty has been managed using standard risk and sensitivity analyses for the cash flows of projects. However, sensitivity analyses alone are not sufficient, and to be rigorous it is necessary to check and adjust major assumptions used to define a project such as the final exploitation limits and cut-off grade applied to ore, and further re-evaluate cash flows. The risk is greater with block caving as opposed to conventional mining because the response time required for adjustments of cut-off grades and mine costs, even if these are possible, is large. This paper proposes a method for managing uncertainty during the block cave mine planning process under a multiMassmin 2004
2. RISK RELATED TO COSTS, GRADES AND METAL PRICES The mining business is intrinsically more risky than most other common business areas. Figure 1 shows a scheme of uncertainties for open pit planning associated with the stage of the development of the pit (Blackwell,1993), and this figure defines the uncertain information and conditions that apply to mining generally.
Figure 1: Uncertainty of Factors in Mine Planning versus Pit Development (Blackwell, 1993) Traditionally mine planners have been managing uncertainty in the planning and design process by using iterative estimations and sensitivity analyses of project cash flows. This process demands substantial and intensive use of resources to identify the most important parameters influencing the results of mine planning and design.
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Currently, many tools are available to produce data, conduct sensitivity analyses, and produce ‘final numbers’ from a base scenario. These tools include geostatistical approaches for ore grade estimations, probabilistic risk and sensitivity analyses for cash flows, and economic indexes. In most cases, operating and capital costs and similar technical variables are defined within reasonable limits with limited impact on project economics. This is in contrast to the risk associated with external variables such as metal prices and market forecasts that have much greater impact. Although several tools are being used in the mining industry for risk management, low metal price trends of important commodities, along with the lack of availability of high grade deposits, are pressing the mining industry to produce, and improve on, more reliable designs and production schedules. This is especially true in low grade (marginal) and impurity rich (penalty producing) ore bodies where profit margins are small, if they exist at all. 3. RISK AND UNCERTAINTY Decision theory literature indicates two distinct types of uncertainty (Rose, 1976): • Risk is a concept used to characterize situations where past data can be useful in predicting an occurrence in the future such as the probability of a road accident or the tossing of a coin. • Uncertainty is a concept referring to situations where there is no suitably supported past data or experience. Metal prices fall more reasonably under the ‘uncertainty concept’ as it is not known when the metal price will return to some past value. The main questions here are, firstly how could we characterize this uncertainty factor, and secondly, how could we manage this characterization in block caving mine planning procedures. 4. METAL PRICES UNCERTAINTY Metal price uncertainty and its impact have been under discussion for many years. The effects on mine planning are critical, hence the continuing recurrence in mining industry publications. Metal prices are critical inputs in defining ore and waste, and the foundations of the mining business, including ‘ore reserves’, are defined based on commodity (metal) prices. In 1903, Williams (in Rickard et al, 1907, page 1) wrote "It has been estimated that 95 per cent of the commercial and industrial enterprises which are started every year ultimately prove unprofitable. Such business failures are primarily due to incorrect estimation of the trade conditions which obtain in every field of commercial operation". This quotation shows that this problem has been of concern for at least a century. During the 1930’s, other authors affirmed that forecasting metal prices is perhaps the most doubtful and speculative of any of the elements in mineral valuation. At that time some of the important elements affecting price forecasts were identified, including relative rate of growth of supply and demand, competition, substitution of commodities, technology trends, and international market conditions being the most relevant. More recently, organizations, such as the Chamber of Mines of South Africa have stated (Storrad, 1981); "…the uncertainties associated with risk factors such as selling prices can not be defined in a completely objective way. Human judgement unavoidably plays the major part in assessing these uncertainties, and the quality of the final answer obtained therefore depends critically on the quality of the judgement involved at this stage". 122
The future is never certain and so metal price forecasts cannot be expected to be accurate. Indeed, it therefore follows that any economic evaluation will never be exact. 5. STOCHASTIC SIMULATION "The fundamental rationale for using simulation in any discipline (whether or not this is economics or operations research) is man’s unceasing quest for knowledge about the future. This search for knowledge and the desire to predict the future are as old as the history of mankind." (Naylor, Balintfy, Burdick & Chu, 1968). Although simulation has been applied to some extremely diverse forms of model building, ranging from Renaissance painting and sculpture to the spacing and computing of neurological processes, it has come to mean something quite specific. Modern use of the word traces its origin to the late 1940’s with the work of Neumann and Ulman (Naylor, Balintfy, Burdick & Chu, 1968). They used an original mathematical technique termed "Monte Carlo Analysis" to solve certain nuclear-shielding problems (Monte Carlo simulation was named after Monte Carlo, Monaco, where the primary attractions are casinos and games of chance). At the beginning the term "Monte Carlo" applied only to the use of stochastic simulation methods for solving strictly deterministic problems. Later it was generalized and became "a popularized synonym for ‘simulation of stochastic processes’ " (Naylor, Balintfy, Burdick & Chu, 1968). Generally speaking Monte Carlo simulation is a form of simulation where randomly generated values for uncertain variables are used over and over again to simulate a model (Goldman, 2000). Stochastic or Monte Carlo simulations are based on the generation of pseudo random numbers (usually between 0 and 1), which are introduced into the cumulative probability functions for predefined variables. As a result of this probabilistic definition of values, the model is fed a new state permitting the updating of the internal database describing the model. Consequently, this procedure is repeated 100 to 1000 times or more, each time prompting a random choice of one or more values for variables. A counter controls the simulation, and the counter is increased until a predefined condition or imposed constraint is met, as Figure 2 shows in a very simplistic scheme.
Figure 2: Stochastic Simulations, Simplistic and Schematic Representation
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The number of times the simulation must be carried out depends on the confidence limits required for the final result. 6. GENERAL PROPOSAL FOR MINE PLANNING & METAL PRICE UNCERTAINTY Figure 3 shows the traditional procedure used to evaluate metal price uncertainty in a project. In this example a Monte Carlo Simulation (MCS) of the cash flow varying under a random evaluation of a metal price distribution is used for the entire life of the project.
Figure 4: Simple Open Pit Cross Section, Defining Price Certainty Parameter (PCP) Assuming a metal price probability distribution was provided for the economic model, and each block was evaluated independently N times (with N large), then 100% of the time block W1 will be waste, but will always be included in the design because it leads to the mining of the highly profitable block O9. The Price Certainty Parameter for this waste block is 100%, and the block has a 100% economic certainty of being included in the design envelope for any given metal price and should be mined. On the other hand, if the peripheral block O6 were included only 50 % of the time in a series of open pit optimizations, then its PCP value is 50%. Figure 5 below shows a flow sheet diagram describing the proposed assignment and management of a stochastic certainty index to each ‘block of resources’ in the block model as a result of a mine planning evaluation. Figure 3: Traditional Mine Planning and Decision Process During this standard deterministic mine planning process, no probabilities of outcomes are specified, so the decision maker (mine planner) must rely on personal judgement and intuition to put the sensitivities of the variables into perspective, and to assess what might happen with many inter-related sources of price uncertainty, especially in the case of a multiple metals ore body. In this paper it is proposed to introduce probability metal price distributions for the different valuable metals present in the ore body in the primary economic evaluation of the resource. By providing a measure of this uncertainty will allow the mine planner to make decisions during the many stages of the planning process. As this measure represents the certainty generated from metal price variability, the author will name it the ‘Price Certainty Parameter’ (PCP), which will provide a value for the probability that any block has a chance to participate in a profitable design, i.e. be considered for mining in the planning process, even though it may not be ore. This Price Certainty Parameter (PCP) will not reflect the exact final economic value of the mine plan and design, because detailed estimation is carried out downstream of this process. Even if the reserves are well scheduled, the final certainty of the project should be less than the average of the PCP when the discount rate and dilution are taken into account. Using a simple open pit ‘moving cone’ example, it is possible to deduce that the PCP is the probability of a resource block being considered inside a profitable design. That does not mean the block is profitable by itself, but, by leading to ore, it might increase the value of the design as a whole. Cleary we could observe this concept if we analyse a block of waste located above a high grade ore block in an open pit planning process as shown in Figure 4. Massmin 2004
Figure 5: Proposed Mine Planning Flow Sheet
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The Price Certainty Parameter of a ‘resource block’ could be defined as "The probability of that resource block being included in a specific design, taking into account metal price probability distributions". This probability could be estimated as a result of several random metal price evaluations of the mining limits for a given mining plan or method. For block caving the PCP parameter can be estimated by evaluating each column of ore using randomly obtained metal prices for each block in the column several times. In the case of an open pit those parameters could be obtained by carrying out several optimization evaluations (e.g. Lerchs-Grossmann (1965) or moving cone, Lizotte(1988)), where the PCP for each block is re-calculated several times. Using this characterization of "Certainty" as associated with metal prices, it is possible to generate alternative designs, envelopes and sequences, and make decisions by referencing PCP values as a numerical parameter. This parameter quantifies, in some way, the economic advantage of mining a defined volume of material (ore or waste) considered in the design before any economical evaluation is carried out. This characterization, with a numerical qualification of resources in the entire block model, will help support the making of decisions during the mine planning process. The assumptions used to assign a certainty ‘index’ (PCP) to each block in the block model are: • Unknown Metal Price in the Future The first assumption is that the market price of any given metal in the future is unknown; intuitions, opinions, shortterm trends, etc. all produce a rough expected value range. • Metal Prices as Probability Distributions It is assumed that we can represent each metal price with a probability distribution describing possible future price behaviour. There are so many factors affecting metal price behaviour that it is impossible to demonstrate that metal prices are deterministic variables. It is also impossible to demonstrate that metal prices are truly random variables taking values in accordance with a probability distribution.
Figure 6: PCP Footprint Output for a Block Cave Column Valuation The ore reserves included as part as mineable envelop can be characterized using standard tonnage-"certainty" curves, as is shown in Figure 7. This "tonnage certainty" distribution (or certainty/tonnage curve, comparable to a grade/tonnage curve) can be obtained and should provide a good general numerical measure of the risk in any particular design. This distribution can give a global indicator of the resources already considered in the mining envelope, including all non-profitable columns added to the design due to operational or geotechnical reasons.
The author considers the characterization of the uncertain behaviour of metal prices as a random variable, and this is far more reasonable than the assumption of a deterministic model. The same assumption is made by all risk analysis software packages. There is a direct relation between the profitability of a column and PCP because a high PCP value is an indication of the profitability of the column for a wide range of metal prices as used in the metal price simulation. 7. MANAGING PRICE CERTAINTY PARAMETER IN BLOCK CAVING Specifically in block caving, using the "Price Certainty Parameter" it is possible to characterize each column in the entire area, representing the probability of each column being included in a mine design for a given metal prices distributions. Additionally, as a result of the several evaluations for the columns, an expected economic value is available for each column and should to be used, together with PCP values, to make a single informed decision in the mine planning process without waiting for several economic evaluations to be completed. A footprint representation for PCP values associated to each column can be observed graphically in Figure 6. 124
Figure 7: Tonnage – Certainty Curves This information provides the planning engineer some confirmation that the operational mine envelope developed is adequately mining with the available resource. For the definition of an operational area to mine, each project board should select the minimum "cut off certainty", taking into account PCP distribution resulting from the area under analysis, "risk guidelines" for the project, etc.
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At this stage, some of the most important definitions in block cave planning are achieved, only using mine planning outputs Once mineable envelope and final column heights to be exploited are defined, the mining sequence definition is made. Again, this situation could be solved using PCP values as a good guideline for profitability of the ore resources. Scheduling columns that show highest PCP values at the beginning of the mining sequence will imply highest profit resulting in the initial production periods of the project. 8. CASE STUDY A case study was developed using information from Codelco-Chile, El Salvador Mine. The A- Norte Sector, located in the Northern part of the main ore body as shows Figure 8, was evaluated using the standard process (fixed metal prices) and also using this proposed and complementary methodology (stochastic metal price distributions).
The stochastic scenario considered trapezoidal probabilistic price distributions for Cu, Mo, Ag and Au. The specific distributions are shown in Figure 9. This probabilistic distribution is usually used when the behaviour of the variable is unknown, maximum and minimum values are defined, clearly highlighting best and worst commodity price scenarios. As well, a range with equal probability (uniform) in between indicates the most probable expectation, guess or feeling, etc., about the future trend of mean prices. In consequence, this probabilistic approach for our lack of knowledge appears as a reasonable representation of this ignorance and actual metal price forecasting. If we consider that this sector will not be in production before 2010, then the question is what metal price must be used for ore reserve definition. Generally speaking, to make the decision, we can use several metal price valuations, an economical envelope using our best "engineering judgement". Standard methodology used in block caving planning, permitted to obtain outputs associated with metal production, ore production rate and copper ore grade profiles as shown in Figure 10 to Figure 12.
Figure 8: General Plan View of El Salvador Mine This marginal area was studied using five deterministic copper prices (75, 80, 85, 90 and 95¢US$/lb), keeping constant Mo, Ag and Au prices.
Figure 10: Fine Copper Production
Figure 9: Trapezoidal Distributions Used
Figure 11: Mine Production Rate
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Figure 12: Mine Production Rate
Figure 14: Expected NPV Values using MCS
The main differences can be observed in 95 cents scenario with longer exploitable reserves due to the grater area to mine. Indeed, this scenario has approximately three more times of ore reserves compared with the 90 cents or stochastic scenario. If we develop a standard economic evaluation, means deterministic metal prices for each of them, we obtain outputs such are shown in Figure 13.
Using the certainty characterization made in the Stochastic Scenario, it is possible to observe a direct relation between the mean PCP value of each scenario and NPV variability by metal prices and discount rate.
Figure 15: NPV Variability & Certainty by Scenario and Discount Rate 9. CONCLUSIONS Figure 13: Deterministic NPV Evaluation Same cash flows were evaluating using MCS, obtaining an expected NPV for several discount rates for each scenario, as is shown in Figure 14.
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The introduction and estimation of a certainty parameter during the primary definition of the (ore) block model in the block cave planning process, based on metal price probabilistic distributions and Monte Carlo simulation techniques, shows that: • There is a direct relationship between values of the price certainty parameter (PCP) and the profitability of a given resource block. • Independent of the cut-off certainty value used by the mine planner, it is possible to obtain in one step a complete picture of the economic potential of the
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resources under analysis, and consequently an advanced measure of the economic risk associated with any given design. • This proposed procedure provides mine planners with a safe and profitable exploitation envelope and mining sequence far more quickly than the standard (deterministic plus sensitivity) methodology. • The estimation of the certainty value for each resource block permits the concept of risk management to be applied during the entire block caving planning process, obtaining quantifiable support for many decisions instead of using "engineering judgement". Further, the identification of material in the (say) 40-50% and 50-60% certainty range by location and volume would provide an equivalent sensitivity analysis to several deterministic scenarios. • In a similar manner to the development of the metal price certainty concept, it would be advantageous to introduce the same concepts to other uncertain variables such as mineral inventory estimates (grades) and geotechnical parameters to provide a more complete characterization of the certainty of any given resource block. The methodology developed can still be improved upon, especially the fixed ‘metal price probability distribution’ which will continue to be a subject for discussion. The basic objective of providing block caving planners with an additional tool for making decisions prior to reaching the stage of the final cash flow evaluation has been met. Without this tool, and without substantial investment of time and human resources, a single mine plan could lead to the rejection of the decision to start a mine, or the starting of a mine with no prospects of making a return on capital employed.
• Fuentes, S., (2003). Planning Block Caving Operations with Metal Price Uncertainty, M Sc Thesis, Queen’s University, Kingston, Canada. • Goldman, L., (2000). Risk Analysis and Monte Carlo Simulation, Decisioneering Inc., h t t p : / / w w w. d e c i s i o n e e r i n g . c o m / m o n t e - c a r l o simulation.html. • Lane, K.F., (1991). The Economic Definition of Ore, Mining Journal Books Ltd., London, England, ISBN 0 900117 45 1, pp 6 – 33 • Leith, C.K., (1938). Mineral Valuation of the Future, The Maple Press Co., New York. • Lerchs, H. and I. Grossmann, (1965), Optimum Design of Open-pit Mines, Transactions, CIM, Vol LXVIII, pp. 17-24 • Lizotte, Y., (1988). The Economics of Computerized Open Pit Design, International Journal of Surface Mining, A.A. Balkema, Rotterdam, Netherlands,Vol 2, 59-78. • Metálica S.A., (2001). Ingeniería Básica A-Norte El Salvador, Internal Report. • Metálica S.A., (1998). Manual de uso Módulo de Planificación Block Caving, MineSight, Internal Report. • Naylor, T., J. Balintfy, D. Burdick and K. Chu, (1968). Computer Simulation Techniques, John Wiley & Sons, U.S.A., pp 1 – 22; 68-71. • Rickard, T.A. and H.C. Hoover, (1907). The Economics of Mining, Ingalls, Gilman & Others (eds.), 2nd Edition, New York, Hill Publishing Co, pp 1. • Rose, L.M., (1976). Engineering Investment Decisions, Planning under Uncertainty, Elsevier Scientific Publishing Co., Amsterdam, The Netherlands, ISBN-0 444 41522 X, pp 42 – 136. • Storrad, C.D. (1981). South African Mine Valuation, Chamber of Mines of South Africa, Johannesburg, ISBN 0 620 02155 1, pp 448 – 456.
REFERENCES • Blackwell, G. H., (1993). Computerized Mine Planning for Medium-Size Open-Pits, Trans. Instn. Min. Metall. (Sect. A: Min, Industry), 102, May- August 1993, A83 – A88.
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Towards an integrated approach to block cave planning Enrique Rubio, PhD, Candidate, Cristián Cáceres, Graduate Student, Malcolm Scoble, Department Head, University of British Columbia
Abstract Block cave planning is a challenging task that is dependent upon effective predictive modeling of the rock mass and the mining system. In reviewing the planning methodology of several operations worldwide it seems evident that such models are not fully integrated within the planning process. The lack of integration challenges realistic production plans and potentially results in conservatively using more resources than needed to achieve desired production targets. This paper presents a methodology to develop a mine planning process for block caving that integrates geomechanical and fragmentation models within the production schedule algorithms. This aims to demonstrate a more robust and reliable approach to block cave planning. Case studies are presented to demonstrate the applicability of the proposed approach compared to some current practice.
1 INTRODUCTION
more resources than needed to achieve a desired production target.
Current mine planning practices at block caving operations tend to be based upon a set of heuristic rules that have been learned throughout the life of the mine and similar deposits. It is well accepted by the underground mining community that geotechnical-geomechanical aspects of the rock mass and the mining method need to be included as part of the planning process. Nevertheless, only a few attempts have been made in order to actually integrate ground-related problems into underground mine planning (Kazakidis and Scoble 2002). In block caving their influence affects the definition of the mining sequence, draw rate and draw point failure rate among others. This paper relates to ongoing research whose main objective is to embed the fundamental geotechnical-geomechanical models into the production scheduling tools so that the algorithms can respond to variation in the rock mass behaviour. 2 THE CONCEPT Several decisions related to mine design and mine planning in block caving are based upon initial modeling that holds a high degree of uncertainty related to the behavior of the rock mass. Consequently, a fair amount of modeling has to take place in order to achieve a comprehensive view of the rock mass and the mining system. The relationships between these initial models is shown in figure 1. The modeling is normally used to estimate parameters such as: stress distribution at the front cave to decide upon the mining sequence; stress re-distribution on the cave back to estimate ultimate fragmentation; fragmentation models to estimate draw point productivity. Even though there may be a fair amount of modeling at the beginning of a block cave project, very little ouput tends to be carried forward into the ongoing mine planning activities. For example, how often do we see a stress model supporting any changes to the undercut sequence within a yearly plan? The same could be asked about the angle of draw or draw rate. At the moment these rock mass models are not fully integrated into the mine planning systems. This challenges the ability to generate realistic production plans and often leads to using 128
Figure 1: Rock mass interaction within the mine planning process There are four main models identified in this research as that are needed in order to sustain the regular mine planning activities. These models are fragmentation, geomechanical, geological and reconciliation models. Figure 2 shows how these fundamental models should be supporting the mine planning parameters, such as draw rate, undercut sequence, development rate, tonnage, draw method (Diering, 2004) and production targets. The fragmentation model estimates the ultimate fragmentation that leads to the estimation of mine design, mixing parameters, mining equipment and draw point productivity. The geomechanical model inducts the mine design into a three- dimensional stress analysis computer program such as FLAC 3D, MAP3D that can simulate the effect from a stresses point of view of different mining strategies in conjunction with the mining plan. The main output of this model will be the stress distribution on the cave back, front cave (abutment stress zone), and induced stresses due to differential draw across the active layout.
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The geological model links data relating to structure, lithology and mineralogy with the ultimate metallurgical recovery. This model aims to build the information towards a geometallurgical model that can provide a reasonable estimate of the metallurgical recovery based on the combination of the composite lithologies. The reconciliation model is one of the most important models supporting the mine planning system. It provides the tools to analyze the historical behavior of the mine. It can capture information on the underground mine and will provide a set of reliability measures regarding the compliance with different production plans. This model also provides the information to feed the fragmentation and geomechanical models to calibrate and reconcile the initial estimates.
ground support and draw point repair. Therefore the important relationship to be used in production planning is the production drift availability versus maximum draw rate per draw point. Figure 3 shows an example of maximum draw rate per draw point as a function of the production drift availability based on a given layout configuration (ore pass spacing), equipment size, cycle time, and number of loaders per drift.
Figure 3: Maximum draw rate per draw point as a function of production drift availability.
Figure 2: Block caving fundamental models This paper concentrates mainly on the fragmentation and geomechanical models.
3 FRAGMENTATION MODELS Fragmentation models are essential to the estimation of the productivity of a draw point. They consist of an estimation of the primary and secondary fragmentation based on initial rock mass jointing, joint conditions and an estimation of the stress behaviour on the cave back (Esterhuizen, 1994). This model should also estimate the amount of rock over 2m3 in volume which seems to be a good indicator of the medium height hang up frequency. Stress behaviour plays an important role in the estimation of the ultimate fragmentation. Therefore the fragmentation models need to be linked to the geomechanical model. Using a recently developed algorithm the fragmentation models could eventually integrate estimates of erosion as part of the secondary fragmentation process (Jensen, 1999). There are a few models that have succeeded in representing this fragmentation models such as Joints (Villaescusa, 1991), BCF (Esterhuizen, 1994). Further analysis and calibration need to be done to assess the quality of these models. Little information has been found on the relationships supporting draw point productivity as a function of fragmentation. In fact, individual draw point productivity as a function of fragmentation alone is meaningless, since the productivity ultimately will be determined by the availability of the production drift. The drift availability is also a function of the secondary breakage and other activities related to Massmin 2004
A more comprehensive simulation model is under construction to model the relationships shown in figure 3, including draw point oversize and hang ups versus different secondary blasting strategies. In an attempt to model the effect of fragmentation on a production schedule, an initial fragmentation model was built for different lithological settings, that were categorized according to the quantity of rock over 2m3. This estimation decreased in height in taking into account the effect of secondary fragmentation. A block model was then constructed using as an attribute the percentage of the block over 2m3. This attribute was built into the draw column and depleted using PC-BC software (Diering, 2000) and average of the attribute >2m3 per draw column is shown in figure 4.
Figure 4: Plan of the percentage of rock >2m3 The next step consisted of simulating a long term production schedule and reporting the amount of rock over >2m3. Within the production schedule the draw rate was modified depending on the amount of rock over >2m3.
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Figure 5 shows the typical adjustment factor used based on the amount of rock over >2m3. In this model there was no consideration regarding production drift availability.
effect of the stresses on the front cave. In this paper the abutment stress has been the starting point to build a full scale model. The initial modeling was performed using the MAP3D boundary element method (MAP 3D, 2003). The rock mass parameters used in the modeling are shown in table 1 (Karzulovic ,1999). The stress state is also shown in table 2. Table 1: Rock mass parameters Parameter
Figure 5: Draw rate adjustment as a function of percentage of rock >2m3 Finally the production plan and the forecast of the tonnage of coarse fragmentation are shown in figure 6. The fragmentation forecast needs to be calibrated against field observations. This step is considered to be crucial in order to understand and modify several assumptions that are made at the beginning of the modeling process.
Value
Young Modulus [GPa]
40.0
Poisson’s ratio
0.2
Cohesion [MPa]
6.2
Friction angle [o]
43.0
Tensile Strength [MPa]
1.0
UCS [MPa]
30.0
Table 2: Far field stress state Stress component
Value (MPa)
O1
80
O2
50
O3
30
The stress state showed in table 2 represents the premining stress conditions at 1000 m depth. The orientation of the principal stress is East West due to tectonics. The spacing of undercut drifts is 15m. The initial model is shown in figure 7. At this stage it was interesting to measure the change in the stress conditions in the center of the pillars supporting the undercut level. Points spaced 5m apart were placed along the undercut pillars to measure the change in the stress conditions, see figure 8. Figure 6: Long term production plan including the fragmentation forecast.
4 GEOMECHANICS ASPECTS OF BLOCK CAVING MINE PLANNING In this research the main aspects of rock mechanics under study are as follows: • Stress distribution at the cave back which affects primary fragmentation and alters the risk associated with air gaps • Seismic activity induced at the cave back by mining activity • Stress re-distribution at the draw points induced by uneven draw • Abutment stress acting on the front cave as a result of stress re-distribution, causing early damage to production and undercut drifts and other related production areas. 4.1. Abutment stress and angle of draw The abutment stress zones have been examined closely in the literature Lorig et al (1995) and McKinnon and Lorig (1999). The abutment stress area is the result of the stress re-distribution at the front cave due to the large extension of the cave. Usually linear models are sufficient to capture the 130
Figure 7: Conceptual model to verify abutment stress condition at the front cave.
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back these two are coupled together through the caving rate. A full scaled model shown in figure 10 was constructed in order to determine the relationship between the angle of draw and the changes in geometry and magnitude of the abutment stress zone. For the model in figure 10 the deviatoric stress (σ1 – σ2) pre- and post-mining were computed and compared for different angles of draw. There are two main results derived from figure 11. The first is that as the angle of draw becomes shallower then the deviatoric stress decreases. The second is that the extent of the abutment stress area increases as the angle of draw becomes steeper. Figure 8: Abutment stress zone using an elastic rock mass model. Figure 8 shows that the abutment stress zone extends up to 30 m ahead of the front cave. The magnitude of the vertical stress increases up to two times the initial premining stress. It is generally accepted that the geometry of the cave back controls the stress re-distribution on the front cave. The angle of draw is a function of the amount of new production area incorporated in a given period and the draw performance in the area of old draw points. A low new opened area and incremental height of draw ratio leads to a steeper angle of draw than a higher ratio. An example of the evolution of the angle of draw is shown in figure 9. Usually this angle is measured in the direction of the undercut sequence. Even though the angle of draw is not the same as the angle of the cave
Figure 11: Change of deviatoric stress as a function of the angle of draw.
A relationship similar to that shown in figure 11 is useful to quantify the potential damage and loss of undercut area as a result of a given production strategy. Several mines around the world will mature and face the question as to whether to keep drawing from old "theoretically exhausted" draw points or to keep undercutting. It is clear that the decision to stop undercutting is economically correct since it leads to capital cost savings. However, the geomechanical implications of this decision may perhaps end up jeopardizing the mine’s operational life.
Figure 9: Evolution of the angle of draw throughout the mine life.
4.2. Frictional forces as a result of uneven draw The second aspect of the geomechanical model supporting mine planning is related to the induced stresses due to uneven draw. It is well known among the mining community that isolated draw produces two effects: early dilution and induced stresses on the production drifts. Recently, Freeport has conducted extensive convergence monitoring, Febrian et al (2004), that raises the issue of induced stresses due to differential draw. The flow of broken rock surrounded by compacted or semi-compacted, broken rock induces frictional forces that will develop on the flow boundary impacting the overall load of the pillars sustaining the production tunnels (Jenike, 1962). Also Kvapil (1965) developed a gravity flow model in which the components are illustrated in figure 13. The components of the gravity flow model developed by Kvapil are as follows: 1 Ellipsoid of motion 2 Funnel of loose material produced by the flow 3 Boundary of motion zone 4 Zone without motion d deviation angle due to frictional properties of boundary wall
Figure 10: Full scale model to assess angle of draw on the stress configuration at the front cave area using MAP3D
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Kvapil also recognized the fact that the flow is altered by an angle due to the existence of a frictional force acting on
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the wall. Angle will increase as the friction of the granular material and the wall increases.
Figure 12: Induced stress as a result of uneven draw
resolution of the stresses on the area under study. There were 8 columns modeled using interfaces to enable the differential movement of them. Vertical right and left boundaries were fixed to prevent horizontal movement. The bottom boundary was fixed to prevent vertical movement. Gravity was applied with large strain coordinate update to show the deformation of the elements. The bottom boundary of the size of a column was freed in the middle of the caved zone to represent extraction of material. The properties used in the model are as follows Density Cohesion Friction Angle Dilation Tensile Strength Young’s Modulus Poisson’s ratio Cohesion Friction Angle Dilation Tensile Strength Stiffness
= = = = = = = = = = = =
2.3 0.0 40.0 0.0 0.0 500 0.3 0.0 40.0 0.0 0.0 500
t/m MPa deg deg MPa MPa MPa deg deg MPa MPa/m
In order to prevent free falling material, a velocity was assigned at the bottom nodes to represent the extraction rate. In this case, 0.0001 meters per time step was modeled. The model was cycle for 60000 cycles until the vertical stress histories converged to a stable situation. Six history points were located 15 m from the edge of the opening to measure the evolution of the vertical stress. Figure 14 shows the location of the history points and velocity vectors. Figure 15 shows the vertical stress history of points located 15m from the edge of the opening. Vertical stress increased from 18 to 28 MPa due to frictional forces between the drawing and static columns. A profile of the vertical stress across the production level is shown in figure 16. It is possible to see that the induced stress area due to frictional forces extends up to 25 m from the edge of the opening. Beyond this point the vertical stress stays at the same pre mining condition.
Figure 13: Gravity flow model (Kvapil 1964) Differential draw develops a frictional force that works against the motion of the flowing rock mass. As result of this behavior, a surface of contact between rock in motion and static is developed. This frictional force is proportional to the friction angle of the surface of contact, horizontal stress and the length of the surface of contact. A model using FLAC 2D (Board, 1989) was constructed to analyze the behaviour of the frictional forces and the impact of this frictional force on the stresses acting on the major apex pillar. The number of grid elements was 50 x 50, representing a 100m wide by 600m height muck pile. The grid density increases with depth to achieve a better
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Figure 14: Velocity vectors and position of the history points.
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Figure 15: Vertical stress measured at the history points Figure 17: PFC model simulating even draw
Figure 16: Profile of vertical stress across the production level. Figure 18: PRC model simulating isolated draw A second model was developed using PFC 2D to test the effect of the frictional forces on the behaviour of the gravity flow process. The rock mass was modeled using an assembly of 1000 particles of a radius varying between 1.8 and 2.2 m, a multiplier factor of 1.5 was used to reach the desired initial stress condition. The block dimensions were 200 height and 150 m wide. The initial stress state was 6 MPa vertical stress at top of the undercut level. The initial stress equilibrium was reached after 8000 cycles. The friction coefficient assigned to the particles was 0.5 according to the friction angle under study. There were two models in order to simulate even draw and isolated draw condition. The even draw condition was simulated by creating two opening at the bottom wall and leave the model running for 25000 cycles. The second model consisted of running 10000 cycles as the first model, in which both opening were drawing particles, and after one of the openings was closed simulating idle draw point condition for 15000 cycles. Figures 17 and 18 are showing both of the models described above, being 17 the even draw simulation and 18 the isolated draw simulation
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From figure 18 is seeing that as a result of performing isolated draw there was formed an stable arc. This is consistent with the theory that a higher frictional forces acting on the edges will tend to produce stable arcs; Egger (1983), Morgan (1999), have presented similar behaviour in soil mechanics. Therefore it is possible to conclude that isolated draw does increase the frictional forces which induce the rotation of the principal stresses tensor approaching a state of equilibrium. The even draw scenario shows that the pillar of rock overlying on the major apex pillar continuously fails without allowing the frictional forces to induce a consistent rotation of the principal stress tensor. There were several measurement points defined above the major apex pillar at 5, 30 and 60 meters above the top of the pillar. Figure 19 shows the change of porosity at 5 meters above the major apex pillar. It is possible to see from figure 19 that porosity consistently decreases as a result of performing even draw. Figure 20 shows the evolution of porosity using the isolated draw model. It is seeing that by performing isolated draw the major apex becomes fully loaded of body forces that make the muck pile sited on top of the major apex to compact and consequently failure.
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More research needs to be done to better understand the constitutive relationship between stress and primary fragmentation, as well as dilution behavior. Thus the stresses will not only be linked to the undercutting sequence but also to the entire production system. REFERENCES
Figure 19: Porosity above major apex pillar measuring on the even draw model
Figure 20: Porosity above major apex pillar measuring on the isolated draw model There is on going research to construct a more comprehensive model to correlate the differential draw versus the activity of body forces on the crown pillar supporting the production level of a block cave operation. 5 CONCLUSIONS Integrating geomechanical rock mass behavior into underground production scheduling provides a more robust mine planning model that will lead to a more realistic financial plan. This should avoid hidden costs that may impact the economics of the project and improve safety in the workplace. The integration of stress behavior into production planning aims to better understand the behaviour of the caving back, enabling mine planners to anticipate eventual air gaps and rock bursting at the caving back. It has been demonstrated throughout numerical modeling that differential draw induces frictional forces that affect the orientation of the principal stress tensor. This change on the stresses configuration leads to higher probability of experiencing of hang ups and compaction of the material located above the major apex pillar inducing higher vertical stress.
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• Board, M. FLAC (Fast Lagrangian Analysis of Continua) Version 2.20. U.S. NRC, NUREG/CR-5430, October 1989 • Cundall, P., and M. Board, 1988. A Microcomputer Program for Modeling Large- Strain Plasticity Problems. In Numerical Methods in Geomechanics (Proceedings of the 6th International Conference, Innsbruck, Austria, April 1988), pp. 2101-2108. Rotterdam: A. A. Balkema. • Diering T, (2000). PC-BC: A Block Cave Design and Draw Control System. Proceedings Massmin 2000. Brisbane, AusIMM, pp301-335. • Diering T, (2004). Computational considerations for production scheduling of block cave mines. Proceedings MassMin 2004, Santiago, Chile. • Egger 1983 • Esterhuizen, G S, 1994. A program to predict block cave fragmentation. Technical reference and user guide, version 2.1. • Febrian, I, Yudanto, W, Rubio E. (2004) Application of Convergence monitoring to Manage Induced Stress by Mining Activities at PT Freeport Indonesia Deep Ore Zone Mine. Proceedings MassMin 2004, Santiago, Chile. • Itasca, 2000. FLAC3D, Fast Lagrangian Analysis of Continua in 3 Dimensions. Version 2.0. Itasca consulting group, Inc: Minneapolis • Jenike, A W, (1961). Gravity flow of solids. Trans. Inst. Chem. Engrs. Vol.40, pp 264-271 • Karzulovic, A,1999. Comentarios relativos a las propiedades del macizo rocoso que conforman los pilares del sector Teniente Sub 6 N FW1. Geotechnical note DTCG-99-016 from A Karzulovic & Asoc to El Teniente Division, Codelco- Chile. • Kazakidis, V.N. and M. Scoble, 2002. Accounting for Ground-related Problems in Planning Mine Production Systems. Int. Jnl. Mineral Resources Engineering, Imperial College Press, London, 11, 1, pp. 35-57. • Kvapil R. (1965). Gravity Flow of Granular Materials in Hoppers and Bins. Part I and II, Int. J. Rock Mech. Mining Sci. Vol 2, pp.35-41 • Lorig, J L, Board, M P, Potyondy, D O and Coetzee, M J, (1995). Numerical modelling of caving using continuum and micro-mechanical models. In CAMAI’95: 3rd Canadian Conference on Computer Applications in the Mineral Industry, pp 416-425. • McKinnon, S D and Lorig, L J, (1999). Considerations for three dimensional modelling in analysis of underground excavations. In Distinct Element Modelling in Geomechanics (Ed: V W Sharma, X R Saxena and R D Woods), pp 145-166. Oxford and IBH Publishing: New Delhi. • MAP3D, 2003. MAP3D, Three-Dimensional Boundary Element Formulation, versión 49. Mine Modelling Pty Ltd. • Morgan, J. K., and M. S. Boettcher. (1999). Numerical Simulations of Granular Shear Zones Using the Distinct Element Method: 1. Shear Zone Kinematics and the Micromechanics of Localization, J. Geophys. Res. 104(B2), 2703-2719 • Villaescusa, E, 1991. A three dimensional modelo f rock jointing. PhD thesis (unpublished), University of Queensland, Brisbane.
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Computational considerations for production scheduling of block cave mines Tony Diering, Principal Consultant, Gemcom Software International Inc.
Abstract Production scheduling for block cave operations can be complex. The factors to be considered include geotechnical constraints, cave shape, draw point development sequence, draw point productivity, production block limits such as loader capacity and/or ore pass capacity and variable shut-off grade or mining costs. In addition, for some caves, the material flow can be very non-linear, especially when material moves between the draw point during the depletion process. Advantages and disadvantages of treating this as a single multi-period problem or as a number of smaller problems for each time period are presented. The nature of the problem also changes during the life of a cave from initial production build up to final closure. Over the years, a number of different scheduling methods have been introduced into the PC-BC program to handle these different situations. Methods include cave surface following, mining oldest draw points first so as to move a cave front, NPV optimization or depletion of historical tonnages. LP techniques may be used in some, but not all cases. In addition, special tools are required for updating long term plans due to short term draw order variations. The logic, advantages and disadvantages of these different methods are described in this paper.
1 DESCRIPTION OF THE PROBLEM The basic problem is to try to predict or schedule the "best" tonnages to extract from a number of draw points for various periods of time. The time period can be very short, such as a day or it can extend over the life of mine. This is an old problem in mining which has been solved (or not solved) to various extents for different mining methods. In the case of block caving, much work has already been done by A. Guest of De Beers, the International Cave Study and also by others such as NCL in Chile and T. Diering with Gemcom’s PC-BC package. In a typical instance, one will start off with a number of draw points (di) from which it is desired to extract tons tij for a number of time periods j. In doing this, the intent is to maximize revenue R or some time discounted equivalent of R (i.e. NPV). The allowable tonnages tij are heavily constrained, to the extent that the constraints will often determine the allowable values of tij. In many cases, the objective may be to achieve a certain "draw profile" which is in fact not primarily dependent on NPV or revenues. Even in this case, NPV is still an important objective, since, if a bad profile is achieved, then this is adverse from a mining perspective which in turn leads to higher mining costs and more dilution and lower grades which reduces the NPV. This then becomes a linear or more likely non-linear optimization problem typical in operations research. However, the value of R can be a very non-linear function of the tonnages tij, since the revenue is dependent on the draw point grades gij which in turn are variable with tij and also on the interaction of the tij from one draw point with all other ones. Of course, the extent to which the problem can be solved depends on making various assumptions about the process of solving the problem. 2 OVERALL OBJECTIVES OF THE OPTIMIZATION PROCESS As mentioned above the overall objective is to maximize NPV subject to various constraints. Some of the constraints which can be applied are as follows: Massmin 2004
• • • • •
Minimum tonnage per period Maximum tonnage per period Maximum total tonnage per draw point Maximum total tonnage per period Ratio of tonnage from current draw point compared with neighbours. • Height of draw of current draw point with respect to neighbours • Percentage drawn for current draw point with respect to neighbours • Maximum tonnage from selected groups of draw points in a period (Usually the groups of draw points are referred to as production blocks or panels) Many of these constraints are themselves quite complex to estimate and may well be non-linearly dependent on tij. A key factor in all of this is the maximum allowable tonnage per draw points per period. This is typically limited by what is called a PRC (Production Rate Curve). This will specify the maximum tonnage per square metre per day which is allowed to be extracted from a draw point. The computation of this curve is in itself complex. In trying to solve the various aspects of the problem, one needs to consider two main scenarios, solving for tij for a single time period j, and solving for tij for multiple time steps j=1,M If there are many time steps and many draw points, then the total number of variables becomes large if one is using non-linear programming solvers and it then becomes desirable to linearize the problem by various assumptions. Another way to tackle the problem is to apply common sense rules or heuristics which, perhaps, could provide common sense solutions and also give an indication as to what extent the problem is flexible enough to warrant optimization procedures. For example, if the tonnage requested from draw points always exceeds what is able to be produced, then the problem is self limiting and no optimization is necessary
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3 NOTATION The following notation is used in this paper. N M i j tij Ti Tj cij bij
Number of Draw points Number of time periods The ith draw point The jth time period Tons for draw point i for time period j Total tonnage allowed for draw point i Total tonnage allowed for time period j PRC tonnage limit for draw point i for period j Minimum tonnage allowed for draw point i in period j. bij ≥ 0 Allowable difference in tonnage between draw eij point and its neighbours mij Mean tonnage of neighbours HOD for draw point i in period j hij F Objective Function to be maximized Grade of draw point i in time period j gij Revenue from draw point i in period j rij (adjusted for mining costs) Discount fraction in period j dj NPV Net present value HOD Height of draw LP Linear programming problem NLP Non-linear programming problem 4 A SAMPLE PROBLEM Consider the sample problem of maximizing NPV for M periods:
F
= =
ΣΣ rij dj ΣΣ gij * tij *
dj
Constraints tij tij
Σtij Σtij
≤ ≥ ≤ =
cij bij Ti Tj
≤ ≥
PRC limits per draw point Lower limit per draw point Total tonnage limits per draw point Tonnage limits for each period
mij + eij mij - eij
In general, there are considerable benefits to be gained if the problem can be formulated as an LP instead of NLP. Solutions are faster and the solver routines can generally handle larger problems. So the problem will often be how to linearize a problem or to convert it to a quadratic programming problem. In the above example, the grades are variable and are functions of tij and tij (for previous periods). Thus the objective is actually non-linear. This can be accommodated by having a look up table to get the grades, but this will slow down solution time as well as probably resulting in multiple local maxima in the objective function. 5 SINGLE PERIOD vs. MULTI-PERIOD There are several advantages and disadvantages in each approach. These are listed below:
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Disadvantages of single step approach • Decisions made now can potentially lead to a "bottle neck" down the road. There is no forward looking capability Advantages of multi step approach • Solves the overall problem in a single step • Short term decisions are guided by future considerations Disadvantages of multi step approach • Leads to large problems (number of draw points * number of time periods) • Grades are more complex to model and likely to be nonlinear. Grade profiles may have to be simplified • Care has to be taken to input the development sequence • Convergence to a unique "best" solution is not assured • Likely to require commercial solver engines • In some cases, the extent of non-linear behaviour makes numeric solution virtually impossible. 6 THE DEVELOPMENT CYCLE OF DRAW POINTS AND PRODUCTION BUILD UP
If we do this, the program will simply take the maximum allowable tons from the highest grade draw points as soon as possible. Therefore, we need to add in extra constraints which limit how tonnages are related to neighbouring draw points. tij tij
Advantages of single step approach • Grades can be assumed constant (and known) for each run • Less equations to solve • Objective function can be set up to maximize local draw control constraints or some local dollar value or grade target • Sequencing of draw points does not arise, since it is known which draw points are active in each step • If some intermediate HOD surfaces can be established by some other means (e.g. Upside down pit optimizer) then these can be used as intermediate control surfaces for the step by step approach. In this manner, the overall problem of NPV optimization is removed from the problem and it becomes one of draw control (or draw management as it is sometimes referred to)
In producing schedules, it is useful to consider what happens to draw points during their "life" in a schedule. Consider the following steps: • The draw point is Planned status. During this time tij=0. • Draw point has just been developed or undercut. During this time, the PRC limits cij are likely to be quite low and the tonnage from each draw point must be kept in close alignment with the tonnages of its neighbours. • The time until the cave limit breaks to surface or other previous cave zone (or gets "far away"). In general if a swell factor is 20%, then the broken zone will extend 5 times as high as the height of draw. During this time, the shape of the cave back is important. Thus the hij values need to be constrained or monitored. • In the final stages of draw, Height of draw is of less concern and percent draw is of more relevance. It is desirable to have draw points close in an "orderly" manner. Tonnage from one draw point compared with its neighbours is still important. • The draw point is closed. This usually happens when the limiting tons for the draw point has been reached. In some circumstances, when variable shut-off grades are being used, then the limiting tons are also variable. During this overall process, the ability of the system to meet various production demands varies significantly as well. In the early stages, the production capacity (C) falls well short of the tonnage demand (D) and basically, draw points are drawn as fast as is allowed (via the PRC curve). As more draw points are developed, the draw area increases and the production rate for individual draw points
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increases, then production capacity increases until C > D and results in some production flexibility. This is where the scope for NLP optimization arises. Later in the life, there are no more new draw points available and the active draw area starts to reduce resulting again in a stage where C < D. Here again, there is little or no need for NLP optimizers. This is studied in more detail later. For block cave mines (including panel caving), there are a number of different draw strategies which can be applied. It is useful to represent these graphically to better understand their characteristics and behaviour. The graph below (termed Draw Method Graph) shows a list of draw points on the x-axis and the tonnage per draw point on the y-axis. Draw points are ordered on the X-axis by sequence with oldest to the left and newest to the right. The number of active draw points defines the "active area". Maximum tons per draw point are constrained by the PRC limit and minimum tons per draw point are also shown (often set equal to zero). The PRC limit and active zone define the capacity of the draw points for a given period. If D < C, then there is some flexibility as to how tons are taken from each draw point. For a moving front or panel caving method, maximum tons are taken from the oldest draw points and minimum tons from the newer ones. In PC-BC, this method is referred to as AUTO as shown in the figure below.
The use of AUTO method often results in "idle" draw points where no tons are extracted in a given period. This is a very useful technique to study the minimum development rate (commissioning of new draw points) to achieve a desired production rate (reduced rate of capital expenditure). Other methods can also be represented in a similar manner. See the EVEN, SMOOTH and COMBO methods shown below. SMOOTH is similar to AUTO, but ensures that no draw points are idle by having a transition from the PRC limited tonnages of older to the minimum tons for the newest draw points. EVEN method tries to take tons from each draw point proportional to what is remaining in the column. This is the opposite of AUTO and takes the highest tonnages from the newest draw points. "COMBO" is similar to EVEN, but constrained by the maximum PRC. 6.1. Differences between daily and longer term schedules There are various differences between the two situations which are listed below: • Usually in the short term, the target tonnages are aimed to reach a given tonnage over the next few days (e.g. daily tonnages aim to reach a monthly target). Massmin 2004
• Draw points are usually assumed to be of known and constant status during the short range runs. • Draw tends to be "Even" in the short term (i.e. proportional in that draw points which are behind are drawn faster and draw points which are ahead are drawn less.) • Draw can be "Even" or in the form of a moving front or even a "segmented moving front" for the long term. In this case, the shape of the draw profile is more important. • In the short term, it is quite likely that tonnage limits will also be set for groups of draw points. For example, various production blocks or panels (or drifts) might each have their own tonnage target (or often referred to as a "call").
6.2. A Panel caving example This example demonstrates two aspects of the scheduling process: • If actual historical tonnages result in an uneven HOD profile, then the schedule needs to be corrective in putting the HOD profile back on track • For a moving front, the HOD surface is inclined and moves across the block cave footprint. The top line below represents a target cave surface (or shape). The line "Past" might represent actual tons drawn after a given period of time. We want to evolve the shape to the top line.
In step 1, the program is able to correct the tonnages (New) for most of the draw points except where maximum
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current tons per draw point (2500t) per step is reached. (New = Past + Current). After a few more steps, we can see how the shape is restored to the target, except for draw points to the right, which have closed due to reaching their maximum allowable tonnages.
It should be noted that if the PRC curves are changing dynamically, then the program will continually need to make these adjustments. Also, in the shorter term, hang ups will restrict actual tonnages so that the correction process will always be ongoing.
second (or subsequent draw points are developed). Thereafter, in a simple case, equal tons will be drawn from each draw point until one or other of the draw points closes due to the maximum tonnage for that draw point being reached. This defines the "target tonnage line" for the schedule. Next, we know that we cannot mine negative tonnages from any draw point. Thus, we can only move upwards and to the right. We cannot move down or to the left on this graph. The next constraint is the PRC limit for each draw point. This is a limit on the maximum tonnage for each draw point in a given period. These show up as vertical and horizontal lines on the above graph. Points outside this region are not feasible. In addition, there may be constraints which limit the tonnage from one draw point depending on the tonnages from neighbours. These are referred to as "neighbour constraints". They are shown as short diagonal lines on the above figure. Next we define the production target for a given period. This is the total tonnage from all draw points in a given period. In this example, with only two draw points, we get a straight line (of slope = -1). The ideal target would be where the production line meets the target line. However, this lies outside the feasible domain, so we have to look only at feasible solutions. This is represented by the heavy line where the production line is inside the domain defined by the other tonnage constraints. At this stage, it is easy to see that the required solution is where the heavy line comes closest to the target line. See "best solution" above.
7 SCHEDULING CONSIDERATIONS FOR A TWO DRAW POINT EXAMPLE 7.1. Single time step It is common in the literature to study complex problems by reducing the situation to a simple example with only a few unknowns. The example here for block cave scheduling provides a very useful tool to help study different aspects of the overall problem. The next figures show this concept. Each figure has various comments by way of explanation.
The example above shows a situation with two draw points and the two axes represent the cumulative tons drawn for each draw point. At any point in time, the current state of draw is represented by a single point on this graph (see "Start" above. For a moving front example, we will always draw some tons from the first draw points before the 138
In the second example above, the production capacity of the draw points has been reduced. Thus the ability of the system to meet the required production tons is also reduced. In fact, there is no feasible solution as the feasible domain (based on tonnage constraints) does not intersect the production line. We thus have one of two possible alternatives. One is to take the maximum tonnage from each draw point so as to get as close to the production target as possible. (This is referred to as a PRC limited tonnage). The problem with this approach is that it does not move us any closer to the target tonnage. The second is to sacrifice some tonnage production so as to be able to move closer to the target tonnage line. Of course, there are a large number of similar alternatives which could arise. But inspection of the graphical representation is always useful in helping to understand the different possibilities. One can imagine that with hundreds or thousands of draw points, the solution is quite complex. 7.2. Multi-time steps Next we look at a multi-step scenario. The production tons for each period is represented by a diagonal line. The
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distance between the lines depends on the production target. In addition to the short term PRC limits for draw point tonnages, we also have the maximum total tons per draw point which define the overall "box" in which we need to work. The example below shows a typical production solution in which tons are drawn in an unequal manner from each draw point over the life of the draw points and the "solution" moves towards the top right corner.
broken. This represents a more realistic scenario. Draw point 1 is mined faster than draw point 2 initially, since it has higher grades, then draw point 2 is mined faster, then draw point 1 again and so on. Doing this, instead of following a single trend line, results in better NPV, while still obeying the applied constraints for each step. 8 SOME PRACTICAL CONSIDERATIONS Experience has taught us some basic factors to be aware of: • You won’t get the final solution the first time. • There is no "optimal" solution. You need to study and compare different alternatives to get a better understanding of the important factors and constraints in a schedule. • Engineering judgement is still very important in this process. • Aspect ratio (length to height) of the cave has a significant effect on the production strategy. • The grade of dilution (and its effect on ore treatment) can also be significant. In some block caves, the grade of "dilution material" is often close to ore grade. This is obviously a lot less "harmful" than dilution material of zero grade, such as gabbro material in a kimberlite pipe. However situation like this can induce operators to disregard draw control practices that could damage the geomechnical stability of the mine.
In this example, every change in direction of the tonnage line represents a change in draw strategy driven by the set of constraints and the target defined in that particular period. In general, big changes would be considered undesirable as would near vertical or near horizontal lines Now (above), we superimpose on the production graph the expected grades from each draw point. For draw point 1, the grades are (3.0, 4.0, 2.0, 1.0, 0.5). For draw point number 2, the grades are (2.5, 3.5, 2.2, 0.8, 0.6, 0.1). By "grade" we mean any variable which can be linked to the dollar value of the draw points.
8.1. A "Northparkes" type deposit • Height to width ratio is high • Dilution is less significant • Usually has higher grade central core • Caveability and hydraulic radius are problematic • Usually all draw points active for most of cave life. Strategy is to pull draw points at equal rate initially for first 20%-30% of HOD to get cave established and then pull central core (high grade zone) faster than outer draw points, while maintaining relatively smooth draw profile. 8.2. A "Kimberlite" type deposit • Height to width ratio ≈ 1 • Dilution is very significant • Grades generally less variable • Need to pull new draw points as fast as possible (to PRC limit) to try to get cave profile as flat as possible. • Have to close older draw points to limit age of draw points (repair costs) and total number of active draw points. 8.3. Large Cu/Au type deposit • Height to width ratio < 0.8 • Dilution is often, but not always problematic • Grades are variable, so sequence and cave shape affect value significantly • New draw point sequence (undercut shape) is very important. 9 CONCLUSIONS
In this example, we now constrain the tonnages using PRC and neighbour limits. It is seen that the overall trend of the schedule line is similar to before, but cannot deviate too far from the diagonal line connecting the bottom left with the top right otherwise neighbourhood constraints are
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Much of the logic around the scheduling of large open pits applies to scheduling of block caves. In both cases, geotechnical constraints are very important. In both cases, the potential to add value to the overall project, through careful scheduling is significant. Using tools, such as PCBC allows you to understand the scheduling characteristics and achieve a sound engineering balance between geotechnical risk and the reward of improved Net Present Value.
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ACKNOWLEDGEMENTS
REFERENCES
The author is grateful to National Research Council of Canada who sponsored some of this work via the Industrial Research Assistance Program. Thanks are also due to Gemcom Software International Inc. for time and funding to complete the work and submit this paper.
• Diering, T, 2000. PC-BC: A block cave design and draw control system, Proceedings MassMin 2000, Brisbane, pp. 469-484. • Guest, A., van Houd, G.J. van Johannides, A. and Scheepers, L.F., An Application of Linear Programming for Block Cave Draw Control. Proceedings MassMin 2000, Brisbane, pp. 461-488.
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Block cave production planning using operation research tools Enrique Rubio, PhD Candidate, University of British Columbia Mining Consultant, Tony Diering, Principal Consultant, Gemcom Software International
Abstract In the pas, manual methods have been used to plan and schedule the extraction of ore from different block cave operations worldwide. The basic assumption of these methods has been the validity of a set of heuristics, traditionally, used to plan and schedule production coming out of an active panel. Currently, however, there are several operations research tools (previously used in the manufacturing sector) that could be used in block cave mine planning. This paper describes the application of mathematical programming to formulate optimization problems whose solution may perhaps drive the production strategy of a block cave mine. Some of these strategies such as net present value optimization, draw profile optimization and minimization of long – short term gap have been formulated. The construction of the optimization problems has required a rational study of which mining constraints are applicable in each case. In doing so it has been found that the formulation of the objective function as well as the set of constraints that define the feasible space of solutions are both critical to effective mine planning solutions. At the moment the full scale algorithms have been incorporated into the PC-BC block caving commercial package. One of the results of this research has been the integration of the opportunity cost into PC-BC to compute best height of draw in a dynamic manner. The second result has been the development of draw method called NPV which maximizes the net earnings per period. Another result has been the introduction of a new draw method called SURF, which aims to minimize the difference between actual height of draw and the target represented by a surface. Different mathematical techniques have been used to solve the optimization problems such as direct iterative methods, linear programming, golden section search technique and integer programming. The results of applying optimization to different operations worldwide will be presented and outlined in this paper. Finally a discussion about the role of optimization in block caving will be presented
1 INTRODUCTION
2 OPERATIONS RESEARCH IN PRODUCTION SCHEDULING
The planning of a block cave mine poses considerable difficulties in the areas of safety, environment, ground control and production scheduling. As the industry is faced with more marginal resources, it is becoming imperative to generate production schedules which will provide optimal operating strategies and make the industry more competitive (Chanda, 1990). Production scheduling of any mining system has a profound effect on the economics of the operation. In a marginal deposit the application of the correct scheduling mechanism might affect the life of the mine. Usually the scheduling problems are complex due to the nature and variety of the constraints acting upon the system (Denby, 1994). Although several authors such as Caccetta and Giannini (1988), Wilke et al (1984), Gershon (1987) have attempted to develop methodologies to optimize production schedules, none has satisfactorily produced a robust technique which has an acceptable level of success. One of the main reasons for this unsuccessful history has been the failure in defining the objective function in relation to the mine planning horizons. In this research two main planning strategies will be formulated as potential goals to be optimised as part of the long term planning process. The first one is the maximization of net present value, which has been a traditional interest of mining companies to optimise in such a way that all the mining, metallurgic and environmental constraints are fulfilled. The second strategy developed in this research is maximization of mine life, which often has been associated with a societal goal to maintain employment levels. Massmin 2004
The problem of computing a production schedule in an underground mine can be understood as an operations research problem in which there is an objective function subject to operational constraints. Trout in 1994 developed a model to optimize the cycle time of the unit operations related to a long-hole mining method. Also Chanda in 1990 developed a model to optimize production from a slusher block cave method using scrapers as production machines. Both of these authors concentrated on a short term planning problems that cover a time horizon of a few weeks to a few months. Neither of these algorithms have recognized the fact that the set of constraints is a function of the planning horizon under study, for example, a long term production schedule should contain much less detail than a short term plan. However the long term plan includes clear definitions related to mining reserves, production sequence, and production rate. More sophisticated algorithms have been developed by Guest (2000) and Matthews (2001) to analyse and compute long term plans. Guest in 2000 postulated that by following a set of surfaces that conceptually define a draw control strategy dilution can be minimized and therefore NPV maximized. Matthew also presented an algorithm which could be used to define the optimum opening and closure sequence in a cut an fill mine. Both of these algorithms recognized the fact that by using integer variables in their formulation the computation time often is inadequate. Also both authors described that the solution for the computing time is relaxing the integer variable to reach a feasible solution in a reasonable time. It has been proven (Terlaky, 1996) that by relaxing the integer variables
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in a mixed integer algorithm the optimum solution can differ dangerously from the solution provided by the optimizer. One of the problems found in the current literature is that there has been very little analysis of the adequate set of constraints applicable for different planning horizons. Also none of these algorithms have shown a case study in which a large scale model had been computed. Before stating the mathematical problem of computing a production schedule in a block cave mine, it is important to describe the operational constraints applicable to block cave as a mining method, the following list presented by Rubio, 2000 summarizes a few of them: • Development rate states the maximum feasible number of draw points to be opened at any given time within the schedule horizon. This constraint is usually based on the geometry and geotechnics of the ore body and the existent infrastructure of the mine, which typically will define the number of accesses available to the mining faces • Undercut sequence defines the order in which the draw points will be open. This constraint usually acts on the draw point status activating those that are at the front of the production face. This component assumes that the layout has been previously computed and it is fixed in the optimization. • Maximum opened production area at any given time within the production schedule has to be constrained according to the size of the ore body, available infrastructure and equipment availability. A large number of active draw points might lead into serious operational problems such as exceeding optimum haulage distance or problematic maintenance of draw points. • Draw rate; the draw rate will control flow of muck at the draw point. The draw ratio is a function of the fragmentation and the caveability model. Ultimately the draw ratio will define the capacity of the draw point and it needs to be fast enough to avoid compaction and slow enough to avoid air gaps. • Draw ratio defines a temporarily relationship in tonnage between one draw point and its neighbors. It is believed that this parameter will control the dilution entry point and the damage of the production level due to induced stresses. • Period Constraints; the period constraints forces the mining system to produce the desired production usually keeping it within a range that allows flexibility for potential operational variations. Note that in this formulation mining reserves are not part of the set of constraints. This breaks the traditional paradigm of computing mining reserves in advance of computing a production schedule. In this case the mining reserves will be computed as a result of the optimal production schedule. 3 NPV OPTIMIZATION IN BLOCK CAVE PRODUCTION SCHEDULING According to the theory of non renewable natural resources the problem of optimizing net present value can be written as follows (Conrad, J M 1999): Suppose a utility function πt = ptqt – C (qt), where is the rate of depletion of the natural resources in period t, pt is the spot price of the underlying resource, C (qt) is the production cost as a function of the depletion rate. The objective is to maximize the discounted utility function subject to the limited amount of resources R0. Let’s call the discount factor pt which is a function of the discount rate and the period of the depletion to be discounted. The problem to solve can be formulated as follows:
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Subject to the depletion rate Rt+1 – Rt = –qt R0 given and optimization
T a variable in the
Maximization of the discounted utility function subject to the exhaustion constraint leads to the Lagrangian:
(1)
The first order optimality condition is presented by:
(2)
Note that represents the first derivative of the production cost with respect to the depletion rate in period t. The derivative of the Lagrangian with respect to the remaining resources in period t is presented as follows:
(3)
Finally the derivative with respect to the Lagrange constant is as follows:
(4) By replacing (3) on (2) pt = Cq + λt
(5)
Equation (5) means that the marginal profit should equal the marginal production cost plus a variable cost . This variable cost is called the shadow price of initial reserves, or the value of having an extra initial ton to be extracted optimally in period t. From an optimization point of view λt represents the opportunity cost of depleting a particular unit of resource in period t instead of saving it for the next period of production. In summary the mining interpretation for equation 5 is that the result of applying the marginal cost plus an extra "artificial" cost per period to define the economic outline per period will lead to the strategy {qt }T0 that will deplete optimally the mineral deposit. In the context of an arithmetic example Gray (1914) was the first author recognizing an additional cost to marginal extraction calling it opportunity cost. The second author that developed a formulation for non renewable natural resources optimization was Hotelling, 1931 who introduced the concept of depletion strategy. Hotelling stated that the optimum depletion strategy is the one that depletes the natural resources at such a rate that the growth of the rent generated for depleting natural resources is similar to the rate of return. Nevertheless it was Lane in 1988 who first introduced formally a methodology to compute the opportunity cost in mining. Lane postulates that the optimum strategy should be optimal at the whole resource depletion path, the optimum strategy should not only
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maximize the individual cash flows but also include the effect of these cash flows in the value of the remaining reserves. The last could be summarized in the following formula (6).
(6)
From (6) represents the maximum net present value for the deposit, is the total amount of mineral resource to be depleted, represents the marginal cash flow of depleting one ton of ore in period t following the shut off grade w, with w being part of the overall optimal depletion strategy. τ is the time to deplete and process one ton of ore. δ is the period discount rate. dV* is the gain or the loss of depleting dT one ton of ore in period t upon variable economic and market conditions. Note from 6 that the optimum marginal reserves for depletion (left hand side of the equation 6), will be reached when the contribution to the optimal NPV (V* ) of the last ton depleted is equal to zero. Therefore the last ton of ore depleted in period t should fulfill the following relationship:
(7)
The above formulation has tremendous implications from the perspective of the economics of natural resources, because it provides a mechanism to compute the Langrage constants presented before in equation 5. Thus combining 5 and 7:
(8)
In summary the economy of natural resource theory says that the optimum depletion strategy is the one that covers the marginal cost plus the opportunity cost of depleting the actual resources instead of leaving them in the ground for the next depletion period. One of the problems with the above formulation applied to Block Caving would be to find the set of shut off grades (shut off grade policy) that leads to an optimum solution. The following section will show how the concept of opportunity cost has been implemented in order to derive an optimum depletion strategy. 3.1. Application of opportunity cost in block caving production schedules The following algorithm has been introduced into the PCBC software from Gemcom Software International (Diering, 2000) as part of their routines to optimize the NPV of a production schedule. Before stating the algorithm used to introduce the opportunity cost in the production schedule it is important to draw graphically all the constraints related to the production schedule, figure 1 shows the feasible area for one period of the production schedule. Figure 1 the X axis represents the dollar value per draw point, this value is computed by integrating vertically the value of the draw column until it declines due to a decrease in the metal content. The Y axis represents the tonnage to be mined per draw point; this is one of the optimization variables. The chart is divided into 4 areas by closed draw points (C) which are draw points already exhausted, active draw points (A) which are draw points in production, new draw points (N) which is also an optimization variable and represents new draw points commissioned to production in Massmin 2004
Figure1: Schematic representation of the feasible area of production at any given period of the production schedule
the current period and planned draw points that are draw points located next in the sequence. Also all the constraints of the optimization have been graphically represented to define the feasible area. The solution of the optimization will provide a draw method (Diering, 2004) that defines the tonnage to be drawn from everyone of the active and new draw points. Note that the feasible area should be large enough to contain the draw method that would fulfill the total tons production target. Otherwise the problem is considered to be fully constrained and sub-optimal solution will be found. The optimal draw method will be such that draw points containing higher dollar value will be drawn more and draw points with lower will be drawn less or not at all. The draw points that are not drawn in a period are shut down moving the boundary (C), shown in figure 1, to the right. Consequently the next period in the schedule the active area would be reduced and more new draw points will be needed. Thus the chart in figure 1 is re-drawn for this new period and the draw method re-computed. This process is repeated until reaching the end of the life of the mine. The mechanism to incorporate the opportunity cost in the production schedule consists of computing the dollar value per draw point in every period of the production schedule. The dollar value per draw point calculation will be used for two purposes, the first one will shut down those draw points that do not have enough remaining value and second will be used to plot the chart shown in figure 1. Consequently the active draw points will be drawn according to the draw method described above to drive the NPV to its maximum point. The algorithm to add the opportunity cost in the production schedule is as follows: 1. Set the initial boundary conditions TMINi ∀ i ∈ I sets the minimum tonnage to be mined per draw point with I draw points across the layout OCto = ∀ t ∈ T sets the initial value of opportunity cost for the time horizon of the production schedule 2. Incorporate new draw points according to the given undercut sequence. 3. Compute dollar value per draw point DVit ∀ i ∈ Activet +Newt, ∀ t ∈ T. Note that already contains the marginal cost. If draw point i contains less value than the opportunity cost (DVti < OCkt ) and the tonnage drawn from draw point i exceeds the minimum allowable (TMINi) then draw point i is shut down. Otherwise draw point i will still be in production and will be drawn according to the draw method.
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If there is an extra capacity the newest draw points are flagged as idle status. If there is not enough capacity the tonnage target constraint is broken. 4. Deplete assigned tonnages from the draw column and update the model. Then move to the next period t=t +1, return to 2 After computing the production schedule in this first iteration the opportunity cost per period is updated as follows: • Compute revenue per period Rt = Valvuet *dt where Valvuet is the average dollar value in period t and dt is the total tonnage sent to the mill in period t • Development cost per period Dt = nt * DPC where nt is the number of new draw points incorporated and used in period t and DPC is the construction cost of a draw point. • Profit per period Pt = Rt – Dt • Remaining deposit value per period Pk
T
Vt =
∑ k=t+1
( 1 + δ)
k-1
, where δ is the discount rate per period. Note that this value is computed at every period of the production schedule. • Opportunity cost per period is computed according δ • Vt equation 6 OCt = C , where C represents the average mill capacity. This equation does not integrate the term _ dV dt because it assumes that the economic conditions as well as the market conditions stay steady along the life of the mine. Table 1 shows an example of how the system computes one set of opportunity costs after one iteration.
Every iteration would produce a different set of opportunity costs per period. Thus OCkt would represent the opportunity cost of period t after iteration k. Calling the optimum iteration z, OCzt would represent the optimum opportunity cost policy that would drive the optimum production strategy. The total tonnage drawn per draw point would represent the optimum mining reserves. Thus this algorithm integrates the production schedule and the mining reserves optimisation in a single algorithm. This result is fairly significant considering the fact that traditionally these two processes are computed independently of each other. Ultimately the above algorithm produces a variable shut off grade policy that drives the production schedule to its maximum net present value. 3.2. Opportunity Cost varying economic conditions In the presented algorithm the term _ dV has not been dt included as the scenario under analysis has been steady economic and market conditions. However in the real world the metal prices change as well as the supply and demand for metals. Therefore it would be meaningless to optimize a production schedule without considering metal price changes. However the problem would be to forecast how the prices will behave in the future. The following algorithm assumes that the vector of metal prices along the life of the mine is known and is part of the evaluation variables. For example if the market is facing a rise in metal prices, then it may be more appropriate to wait for prices to recover. Alternative, while waiting for increased prices the deposit is losing value in delaying its operation. Yet there is a trade off between the incremental value gained by economical external changes and by opportunity cost. There are a few parameters that need to be defined to formulate the integration of – dV / dT into the optimization algorithm such as: RFt is the revenue factor per period. MCt is the mining cost per period including the processing cost. Rt is the revenue earned in period t.
Table 1: Opportunity cost calculation after iterating once the production schedule nt
Rt
Dt
3 0 2 1 0 1 1 1 0 1 0 0 0 0 0
2,919,365 2,921,512 2,343,786 1,451,375 1,374,377 1,375,122 1,874,756 2,543,706 1,892,513 1,901,043 1,294,969 208,914 57,387 0 0
360,000 0 240,000 120,000 0 120,000 120,000 120,000 0 120,000 0 0 0 0 0
DPC C NPV
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Pt
Vt
OCt
2,559,365 11,576,159 6,32 2,921,512 9,812,263 5,36 2,103,786 8,689,703 4,75 1,331,375 8,227,298 4,49 1,374,377 7,675,651 4,19 1,255,122 7,188,094 3,93 1,754,756 6,152,147 3,36 2,423,706 4,343,656 2,37 1,892,513 2,885,509 1,58 1,781,043 1,393,016 0,76 1,294,969 237,349 0,13 208,914 52,170 0,03 57,387 0 0,00 0 0 0,00 0 0 0,00
is the revenue that would be earned if the project is delayed in one period of time. It takes economic parameters from t +1 period. Pt is the profit earned in period t. is the profit using . Vt is the remaining value of the mine at the end of year t. Wt is the future value of the mine at the end of year t using economic parameters of year t +1. Gt is the head grade simulated by the production schedule. The calculations proceed as follows:
0,1 120,000 183,050 12,850,476 Santiago Chile, 22-25 August 2004
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Table 2: Comparison of two different schedules without and with –dV/dT Draw Point Name
dV = Wt - Vt is known and the opportunity dt cost per period can be computed as follows: Then the factor
Figure 2 shows the effect of –dV/dT on the resulting opportunity costs. Note that the opportunity cost without incorporating –dV/dT does not have any relation to the revenue factor or metal price. In contrast, the opportunity cost including –dV/dT has a direct correlation with the revenue factor. Yet the inclusion of
on the calculation
of the opportunity cost could lead to a totally different production schedule and therefore the set of opportunity costs that does not include the term accounting for variable economic conditions would drive to a sub optimal production schedule.
HOD_OC
HOD_OC-dV/dT
E1N1
75
75
E2N1
75
75
E3N1
75
75
E4N1
105
216
E5N1
105
191
E6N1
105
205
E7N1
105
186
E8N1
105
172
E9N1
105
126
E10N1
105
122
NPV 000$
14,965
15,047
Reserves
2,029,028
3,093,550
objective function is avoided. Clearly the problem is 3D because the decisions variables are, for example, when to shut off a draw point, which represents the vertical dimension of the problem and also when to open a new draw point, which represents the horizontal dimension of the problem. The following representation of 4 draw points that contain 3 slices each is used to represent the problem in two dimensions. This representation shows profit per slice scaled by 1000. The order in which the draw points have been sorted depends on the opening sequence previously defined.
Table 3: 2D representation of the slice file. 2,24
1,12
1,12
11,23
2,24
4,49
2,24
-3,38
9,98
3,24
5,48
6,61
Sllice #, j Draw point #, i
Figure 2: Opportunity cost with and without –dV/dT Table 2 shows the resulting optimal height of draws for two different scenarios HOD_OC and HOD_OC-dV/dT. The first scenario (HOD_OC) does not include the term –dV/dT, alternative HOD_OC-dV/dT does it. It is clear that the impact of the change in value due to economic change is significant. The algorithm without the differential of value with time does not reproduce a realistic scenario because it does not account for the relation between shut off grade and metal price returning lower NPV and mining reserves. On the contrary, the algorithm with the differential of value on time does couple price and grade, reporting higher NPV and mining reserves. 3.3. Integer programming approach to optimize NPV in a block cave production schuedule The problem of NPV maximization in a block caving operation can be described as a large scale, multi-period, mixed-integer linear programming problem. The development of the model to solve this problem began with translating the optimization problem in two dimensions, so that the non linearity between tonnage and grade in the Massmin 2004
The next step of the optimization process is to find a combination of the above blocks to extract in every period of the schedule that will optimize the overall NPV. Therefore the problem is to find a set of binary matrixes that will tell when to mine every one of the blocks making up the slice file. A representation of these matrices is shown in Table 4 with the extraction of the first year of the production schedule. The meaning of this matrix is that the first 2 slices of draw point 1 and the 1 slice of draw point 2 are drawn in the first year of the production schedule.
Table 4: Binary matrix representing the first period of the production schedule. 0
0
0
0
1
0
0
0
1
1
0
0
Clearly the binary variables should fulfill all the rules related to block cave mining. For example the slices as well as the draw point need to be mined following the matrix’s sequence.
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The above algorithm was written on AMPL 1999 which is a system for writing the optimization problem in mathematical language. Once the problem is written in mathematical language AMPL translates the problem and passes it to CPLEX which is the engine used to search for the solution to the optimization problem. The algorithm is presented as follows: Problem dimensions I , total number of draw points across the layout J , total number of slices within a draw point T , time horizon for the production schedule Problem Parameters
The above algorithm has been tested using up to 100 draw points with 20 slices each and the time horizon for the production schedule has been set to be 14 years. The time to solve the problem has been around 2.5 hours using a Pentium 4 computer, 2.1 GHz of speed and 520 MB of ram memory. Table 5 shows the resolution of a 7 draw point theoretical problem. Every draw point contains 3 slices per column. The economic values per slice are presented in the first matrix, the following matrixes show how the depletion of the slices will be performed in every of the production schedule. Note that the draw rate used in this case constrains the extraction of 1 slice per period, per draw point. Also the maximum number of new draw points per period was set up to be 2.
valvueij , dollar value for draw point i slice number j. This parameter is similar to the matrix presented in Table 3 Ton_t arg et , production target for period t dri , maximum draw rate for draw point i newt , maximum development rate per period. Mblocki, minimum number of blocks to be drawn from each draw point i Decision Variables
{
dijt =
1, If draw point i, slice j is mined in period t 0, Otherwise
This set of variables represents the binary matrixes showed in figure 4 Objective Function i , j ,T d * valueij Maxdijt ∑ ijt ; where is the period discount rate t i , j ,t (1+ a ) Constraints • Draw point sequence within a draw point
Table 5: Resolution of NPV optimization using integer programming.
k
∑d
ijt
≥ dij +1k ;j=1..J,K=1..T,i=1..I
t =1
• Draw point sequence across the layout k
∑d
ijt
≥ di +1 jk ;j=1,i=1..II,k=1..T
t =1
The final reserves outline of the above problem is presented in Table 6. Note that the algorithm does not smooth the final reserves outline ("hair cut") because this process is believed to be part of a second optimization, which perhaps specifically may not be a task of the strategic long term planning.
• Every slice can be mined just once T
∑d
ijt
≤ 1;i=1..I,j=1..j
t =1
• Maximum development rate I
∑d
ilt
≤ newt ;t=1..T
i =1
• Maximum draw rate Table 6: Final reserves outline using integer programming.
j
∑d
ijt
≤ dri ;i=1..I,t=1..T
j =1
• Maximum production rate I, j
∑d
ijt
≤ ton _ t arg et1
i, j
• Minimum number of slices per draw point j ,T
∑d
ijt
This algorithm lacks several constraints that may apply in the planning of a block cave mine such as reserves outline smoothing, draw ratio between a draw point and its neighbours. However the intend of this algorithm is to operate in conjunction with a system such as PC-BC that could introduce the level of detail desired for the planning of the block cave.
≥ Mblocki ; i = 1..I
j ,t
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4 MINIMIZE THE DIFFERENCE BETWEEN TWO SURFACES USING QUADRATIC PROGRAMMING There are two applications of this objective function; the first one is the application of an angle of draw as the desire draw profile, the second one is the minimization of differences between actual height of draw versus a desired target. The first application is related with having a "good" draw performance which leads to retarding the dilution entry point. If dilution is delayed the life of the mine is prolonged, since draw points can still be opened for a longer period of time. The second application of this objective function is more related with the link between long term plans and the short term plans. The long term plan provides the target height of draw for a certain period of time and the short term plan provides the current height of draw situation. The first application to minimize the difference between the current draw profile and a desired draw profile has already been developed by Rahal, 2003. However this author uses a linear function consisting of two main deviations: current profile with respect to a target and current production with respect to a target. The problem with this formulation is that the minimum deviation could be achieved by having a large deviation at the beginning of the schedule and a small deviation at the end of the schedule, eventually resulting in a total deviation equal 0. The proposed algorithm in this paper is the minimization of the square of the deviation which will produce a much efficient search mechanism using quadratic programming and also a better decision from a scheduling point of view. Therefore all deviations either happening at the beginning of the schedule or at the end of it will count the same for the objective function. Objective function
inimum draw rate per draw point per period. In this case mining reserves represent a constraint in the optimization unlike NPV optimization where the mining reserves represent a variable in the optimization process. The way of introducing the reserves as a constraint is by using a binary status variable called "closed" which indicates whether the draw point has been depleted or not. Since the status variable "active" is affected by the closed status variable, reserves affect the status of the active draw points. Therefore if a draw point has been depleted the status variable "closed" would be 1 and the respective "active" variable would be 0. Some assumptions to solve the problem are as follows: • Integer variables were relaxed • Draw points shut down when they reach their Best Height of Draw • Draw rates used were constant along the production schedule The algorithm used to solve the problem was the basic Linear and Quadratic Solver commercialized by Frontline Technology. A model of 10 draw points with 10 slices each was set up in order to solve the original problem. This optimization also fits into the category of multi period problem, in this case 10 period optimization. A graph showing the overall cumulated production drawn from draw points is shown in Figure 3. In this example the main objective of the optimization is to minimize the variance of the tonnages being drawn at any given period of the production schedule. Note from Figure 3 that by applying this algorithm the height of draws per period per draw point describes an overall angle of extraction or draw. It is also possible to demonstrate that the angle of draw is directly related to the draw ratio ( ). Therefore the angle of draw can be easily planned and evaluated by modifying the draw ratio constraint.
I 2 T min ∑ at * ∑ {^ dit − dt− } i =1 t =1
at, number of active draw points in period t dt, average tonnage drawn in period "t" from the active draw points or any desired target dt, tonnage to be drawn from draw point i in period t. This is the main variable in the optimization process, which ultimately leads to the production schedule Constraints • Development rate
vt ≤ Newt ∀t = 1..T
. Note that this is an integer variable.
• Tonnage target I
∑d
it
≤ TTU t
it
≥ TTLt
i =1
∀
t=1..T
being TTU
I
∑d i =1
and TTL the maximum and the minimum production rate in period t respectively. • Draw rate
dit ≤ TU it * ait
∀ i=1 to I and t=1 to T
dit ≥ TLit * ait Note that draw ratio dcf is part of the above constraint by TUit l, being TU and the maximum and the TLit it dcft Massmin 2004
Figure 3: Angle of draw as a result of minimize the variance of the tonnages drawn per period. A different way of approaching the problem of optimizing draw performance is to impose a desired draw surface that is believed to follow a particular cave behavior which ultimately will minimize dilution. Figure 4 shows a draw profile in which the first 2 years of the schedule were drawn without constraints and the following years a particular angle was imposed to be followed as the main objective of the optimization. The above methodology needs to be carefully constrained with the minimum tonnage to be drawn per draw point otherwise isolated draw (which is undesidable) could be an optimal solution for the algorithm. Therefore the upper and lower bound for the draw rate should be carefully studied and controlled by the draw ratio parameter. It is clear that a more relaxed draw ratio constraint will produce a more productive schedule however this may induce the entrance of early dilution and point load on the major apex pillar.
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Table 8: NPV optimization using variable shut off grade approach
Figure 4: Minimize the difference between the current and the desired draw profile.
Note from table 8 that the increase in the NPV is due to the optimization of the blended grade and the reduction of the development rate. The evolution of opportunity cost throughout the optimization process is shown in figure 6.
5 CASE STUDY The following case study presents a mine that contains 1219 draw points in the current layout. The mining method is panel caving with traditional undercutting. The pre computed mining reserves corresponds to 237 Mt with 0.98%Cu. The mill capacity has been set up to be a maximum of 11 Mt/year. Figure 5 shows a 3D display of the layout using PC-BC software.
Figure 6: Opportunity cost for different iterations throughout the optimization process
Figure 5: 3D display of a 1219 draw points layout.
The current mining cost structure used in the optimization is presented in table 7. These costs do not include fixed costs which are added separately in the evaluation of the production schedule.
The difference between shut off grades across the economic layout for the base case and the optimized scenario is presented in figure 7. It is interesting to note that the last draw points in the optimized sequence and the base case shut off grade is similar. Between sequence number 200 and 300 the shut off grade for the optimized case is lower than the base case because there is no enough flexibility or rather active draw points to fulfill the production target. Then the optimization algorithm decides to keep low grade draw points active to achieve the desired production rate.
Table 7: Cost structure used in the optimization. The maximum development rate per year was set up to be 120 draw points. The draw rate varies per draw point and it moves in range 0.5 to 0.65 t/m2/day. Table 8 shows the result of the optimization using variable shut off grade approach. It is possible to see that the increase in the NPV is about 19% while the mining reserves are reduced by 12% with respect to the base case.
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Figure 7: Resulting shut off grade throughout the life of the mine.
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6 CONCLUSIONS Operations research tools can be used to plan and schedule block cave mines. The level at which these tools are applied would define the success of the resultant production schedules. The process of identifying the constraints that apply to the corresponding planning horizon is a critical step in defining the operations research problem. A wrong decision about a set of constraints could lead to a good answer for the wrong problem. The process of establishing the adequate constraints enables mine planners to better understand the mining problem. The use of opportunity cost in production scheduling can lead to improvement of the NPV of the operations by several million dollars. The reserves as well as the development rate are in this case rather an output of the optimization process than an input. Further research needs to be done in order to develop new technologies that could perhaps have the ability to integrate new constraints that would better forecast the reaction of the rock mass to different mining strategies. In particular the addition of uncertainty based upon actual performance will be a key parameter to be incorporated in the future generation of production schedulers in block caving operations. ACKNOWLEDGEMENTS Funding for this research project was provided by Gemcom Software International and the National Research Council of Canada. Also University of British Columbia to provide guidance along the research presented in this paper. In particular Dr. Scott Dunbar, Dr. Malcolm Scoble for their contribution to finish this paper. REFERENCES • Cacceta, L and Giannini, L M, 1988. The generation of minimum search patterns in the optimum design of open pit mines. Bull Proc Australasian Inst Min Metall, 293(5): 57-61. • Chanda E C K, (1990). An Application of Integer Programming and Simulation to Production Planning for a Stratiform Ore Body. Mining Science and Technology, 11:165/172
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• Conrad J M, (1999). Resource Economics. Cambridge University Press; 0 edition (October 28, 1999). ISBN: 0521649749 • Denby, B and Schofield, D, 1995. The Use of Genetic Algorithms in Underground Mine Scheduling. Proceedings XXVth APCOM, pp 389-394 (AusIMM:Brisbane) • Diering T, (2000). PC-BC: A Block Cave Design and Draw Control System. Proceedings Massmin 2000. Brisbane, AusIMM, pp301-335. • Diering T, (2004). Computational considerations for production scheduling of block cave mines . Proceedings MassMin 2004, Santiago, Chile. • Gershon, M, 1987. Heuristic approaches for mine planning and production scheduling. Int J of Min and Geol Eng. 5:1-13 • Gray L C, (1914). Rent Under the Assumption of Exhaustibility. The Journal of Political Economy, Vol. 28, No.3, May 1914 466-489 • Guest A R et al, (2000). An Application of Linear Programming for Block Cave Draw Control. Proceeding Massmin 2000, Brisbane. • Hotelling H, (1931). The Economics of Exhaustible Resources. The Journal of Political Economy, Vol. 39, No.2, Apr. 1931 137-175 • Lane K L (1988). The Economic Definition of Ore. Mining Journal Books LTD, London. • Rahal, D., Smith, M., van Hout, G., and von Johannedis, A., 2003. The use of mixed integer linear programming for long-term scheduling in block caving mines. Proceedings APCOM 2003, Cape Town, South Africa. • Rubio, E., Scoble, M and Dunbar, W. S., 2001. Scheduling in block caving operations using operational research tools. In Proceedings Minespace 2001. Proceedings Annual General Meeting of the Canadian Institute of Mining, Metallurgy, and Petroleum, Quebec City, Canada. • Terlaky T (1996). Interior Point Methods of Mathematical Programming. Kluwer, Dordrecht, The Netherlands. • Trout L P, (1995). Underground Mine Production Scheduling Using Mixed Integer Programming. Proceeding APCOM XXV, Brisbane, pp 395-400.
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Reliability theory applied to block cave production scheduling Enrique Rubio, PhD, Candidate, W Scott Dunbar, Associate Professor; Malcolm Scoble, Department Head, Robert Hall, Assistant Professor, University of British Columbia
Abstract The long term plan in a block cave mine is based upon a number of assumptions about the behaviour of the rock mass. Production forecasts will rely on these assumptions even when data are available to suggest modifications to those assumptions. This can compromise not only the economics of the project but also the global geomechanical stability of the mine. Even though there might be several goals that a production schedule of a block cave mine could follow, at the moment, there is no tool to measure how precisely those goals are met. Reliability theory introduces a new metric to production schedules which ultimately will measure the ability of different production strategies to achieve production targets. In this approach the reliability of a draw point can be computed using historical forecasts versus historical production data. The individual draw point reliabilities can then be linked though a set of equations to compute the overall block cave reliability. This aims to provide a different means to schedule block cave mines adding an index of uncertainty to the overall production schedule as well as the factors that contribute to it. Several examples will be presented as a proof of concept. 1 INTRODUCTION Historically, production schedules in Block Caving operations have been computed using heuristic methods learned over the years during the operation of the mine. Some research on planning or scheduling block cave operations has focused on the use of operations research tools to allocate resources such as equipment and labor (Rubio, Scoble and Dunbar, 2001) or to minimize costs or maximize net present value (Smith and Rahal, 2001; Rahal et al, 2003). There is no published record of the regular application of such methods in a block caving operation. Recently, operations research tools have been introduced to facilitate the planning methodology either in the short term planning (Chanda, 1990) and long term planning (Guest et al, 2000). Nevertheless, there is still a need to introduce algorithms that are able to adapt to the dynamic conditions of an underground mine. One of the main problems that block cave mines are facing at the moment is the lack of integration of operational information into the construction of production schedules. There have been a few attempts to integrate uncertainty models into operations research models stochastical programming methods (Smith, 1999) have not been able to provide an easy way of integrating the variance of the rock mass models into the mine planning system. It is clear that the success of a production plan will depend upon the ability of the mine planning system to incorporate uncertainty found in rock mass behaviour and in the mining system. The understanding of what constitutes a successful mine plan in this paper is the means to be able to forecast the correct amount of resources needed to achieve a desired production target. Operations research tools in mine planning tend to be limited to strategic mine planning or long term production schedules that do not incorporate the degree of detail needed at the operational levels. Reliability systems theory has been extensively used in mechanical engineering to compute maintenance plans and derive operational decisions. One of the advantages of this methodology is that it can integrate all the operational components of a production system, incorporating the failure rate of the components as part of the forecasted 150
availability of the production system as a whole. As an analogy, this could be implemented in a mine that is composed out of different processes in which every one of the components has a different failure rate. This paper aims to demonstrate the ability of reliability models to integrate operational information in the estimation of the best production strategy. 2 CAVING METHODS AND UNCERTAINTY MODELS Block caving has gained increased popularity in recent year due to its ability to produce large tonnages at low operating cost. However, there are several issues that add considerable uncertainty to the mining method such as: caveability in competent and highly stressed ore bodies (De Nicola and Fishwick, 2000); seismicity due the presence of high stresses that could adversely affect the mining method (Dunlop and Gaete, 1995); stress redistribution due to a particular draw strategy; ultimate fragmentation that may have been poorly estimated for the rock mass; lack of precision in estimating the grade distribution within the ore body; and dilution or the manner in which waste is included in the caved rock mass as it moves toward the draw points (Dolipas R, 2000). Inadequate recognition and understanding of these issues can lead to disruption of production performance. However, it has been observed that some block cave operations perform better than others when facing these uncertainties. It seems that the amount of planning and its ability to integrate the above mentioned issues plays a significant role in the success of a block caving operation. 3 PLANNING AND SCHEDULING METHODS APPLIED TO CAVING METHODS In order to ensure that the ore production rate meets requirements and to efficiently allocate resources such as capital, equipment and labor, a block cave mine plan and schedule must be defined. The aspects of mine planning that need to be fully considered to properly plan a block cave mine are as follows: • Draw point sequence:. i.e. the order and timing by which the draw points should be incorporated in production.
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• Active area: i.e. the number of draw points that should be developed per period • Draw rate: i.e. how fast can material be extracted from these draw points to provide the best value to the operations. • Draw constraints: i.e. identification of the main operational constraints that limit the productivity of a draw point. • Draw profile: i.e. what should be the distribution of tonnages within an active panel to guaranty the global stability of the mine • Geotechnical constraints: i.e. how does the draw profile affect the geomechanical response of the rock mass The above factors are linked through several production rules that traditionally have been derived from heuristics and experience at different operations. Figure 1 shows an operating mine that has successfully forecast its production ramp-up as a result of using the factors above and the appropriate heuristic rules.
the caved zone; amount of secondary blasting activity (Dessureault, Scoble, Rubio, 2000). There is a lack of published work dealing with the relationship of fragmentation to production scheduling. Fragmentation tends to vary across the active area due to factors such as discontinuity frequency, rock mass strength, and other geomechanical factors such as the stresses acting in the rock mass. In turn, the stresses are related to rock mass properties, the rate of draw, the draw pattern and the location of draw points as the operation proceeds. These relationships are complex and are likely to be sitedependent. Thus it seems clear that a robust production planning tool should be empirically based and should integrate available production data with measured geomechanical data, such as: deformation, stress indicators, fragmentation data, and rock mass properties. 4 UNCERTAINTY IN BLOCK CAVE PRODUCTION PLANNING The difference between forecast and actual production of a draw point could be used as a measure of the reliability of the draw point to produce ore. Figure 2 compares one month of production between forecast and actual tonnage taken from the same operation as shown in figure 1. Note that even though the global tonnage forecasted has been fulfilled, the distribution of tonnage per period across the active area has not been achieved. This tonnage variance per draw point induces two well known operational behaviors: Under Pulling and Over Pulling. Under pulling means that the actual tonnage is less than the forecast for the draw point and over pulling means that the draw point has exceeded its planned tonnage.
Figure 1: Production back-analysis of an existent operation Currently, when a production plan is computed then all draw points have the same chance of being part of the schedule. However, in every block cave operation there are draw points that tend to produce more easily than others or the productivity of draw points varies across the active area. This variance in draw point productivity clearly reflects the uncertainty in predicting rock mass behavior to plan the production associated with a particular schedule. Production from a draw point depends on several rock mass and design parameters such as: equipment size, layout configuration, stresses on the production tunnels, haulage infrastructure, seismic activity. One of the most relevant parameters, however, appears to be the ultimate fragmentation of the rock mass. Fragmentation models such as BCF developed by Esterhuinzen (1994), Brown (2003) and Wang et al (2003) could be used to estimate the fragmentation curve of a given rock type and thus forecast the frequency of oversize and hang ups occurring at draw points. These fragmentations models will finally affect the draw point productivity. However, current practice is to employ a trial and error process until full production is achieved without introducing the interruptions that the secondary blasting activity adds to the production system. Generally by adding the secondary blasting activity then the productivity of a block or a production unit decreases and therefore the time to achieve full production is usually longer than planned. The impact of this situation on the economics of the mine is significant. Fragmentation models are also used to define other aspects of the design and planning of a block cave mine, such as: draw point layouts (Laubscher, 1994); mixing within Massmin 2004
Figure 2: One month of tonnage reconciliation per draw point
It is generally accepted that under pull and over pull behaviour leads to early dilution entry, over induced stresses, and increased discrepancies between planned and operational performance. Therefore it seems clear that production plans need to be based on a draw point by draw point basis; otherwise imprecise analysis can place at risk the life and economic return of the mine. Figure 3 shows the monthly average relative deviation between forecast and actual tonnages drawn from 40 draw points during a 36 month period at an actual mine. It demonstrates that draw points can vary considerably in the precision of estimating production performance. It also demonstrates the need for production planning to integrate a new way of quantifying the historical production variance between forecast and actual. Then this variance could be used to correct the future forecast or even better help to study the relationship between this deviation and rock mass
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properties. This paper now will introduce this concept through a method that computes a reliability parameter as a measure of production variance
Table 2. Actual tonnage drawn planned tons 09 01H
Mar-98
Apr-98
1800
3500
09 02F
May-98 Jun-98 2500
3000
500
4500
5500
09 02H
2500
2400
11 01F
1350
2150
Table 3. Relative tonnage deviation: actual versus forecast planned tons 09 01H
Mar-98
Apr-98
May-98
Jun-98
0.20
0.40
0.29
0.83
1.00
0.50
0.10
09 02F Figure 3: Monthly average relative deviations between forecast and actual tonnages The reliability of a production plan is computed as a function of the individual draw point reliability. The reliability of a draw point is computed as another property of a draw point, alongside grade, dilution, draw rate, as follows: Draw points i = 1..l, Periods j = 1..J
09 02H
1.50
0.04
11 01F
0.13
0.13
The reliability of a draw point will be evaluated by assuming that a 50% or less average monthly relative tonnage deviation is acceptable. Then the reliability index is computed as the percentage of the time that a draw point has been reliable during its life. Thus the reliability index of draw point 09 01H would be _=75%, since the deviations in March, April and May were less than 50%. The reliability of the selected four draw points is shown in Table 4. Table 4. Draw point reliability
, actual tonnage drawn from draw point i in period j
Draw Point
Reliability
09 01H
75%
09 02F
67%
09 02H
50%
11 01F
100%
, planned tonnage to be drawn from draw point i in period j
ri j =
{
1 if di j ≤ K 0 otherwise
is an indicator of the compliance over a deviation K of the plan with respect to the actual tonnage drawn. Finally the reliability of the draw point is computed as: j
∑r
j
i
Ri =
j =1
Applying the same concept to the data shown in figure 3 it is possible to see in figure 4 that there is a variation of reliability within the active panel. Consequently for the next monthly forecast there will be more confidence in predicting draw point 13 02F than 01 01F in achieving the production target.
j
Table 1 and Table 2 show a production plan and the actual tonnages drawn from four of the 40 draw points showed in figure 3. Table 3 shows the relative deviation between forecast and actual tonnages. Table 1. Production forecast for four draw points planned tons
Mar-98
Apr-98
May-98
Jun-98
1500
2500
3500
3000
1500
3000
5000
09 02H
1000
2500
11 01F
1200
1900
09 01H 09 02F
Figure 4. Reliability index of draw points in figure 1. As more observations are made, then the reliability index can be recomputed. Figure 5 shows the change of the reliability index with time for four draw points chosen from the 40 draw points of figure 3. This shows that the reliability index cannot be assumed to be independent of time and
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therefore an adjustment of the index will be necessary to develop long term schedules.
Figure 5. Change of reliability index with time for different draw points The next step in developing the application of the reliability theory to block cave planning is to assemble the individual draw point reliabilities into the block cave reliability 5 APPLICATION OF RELIABILITY THEORY TO PRODUCTION PLANNING The reliability of the system relates to the probability of an entire system failing, based on the knowledge of the failure distribution associated with the system’s component (Kaufmann et al, 1977). Reliability theory has been used extensively to analyze mechanical systems, even the process of ants foraging (Herbers, 1981). In order to compute the reliability of the system, it is first necessary to map its component processes. The relationships between these processes also need to be established. Then an analogy with electric circuits is used to compute the system reliability. For example, the reliability of two dependant processes is computed as if these two processes were connected in series. The same applies for two independent processes in which the reliability is computed as if the processes were connected in parallel. Figure 6 shows a few different kinds of process connections, Herbers (1981), as follows: a) in series b) in parallel c) parallel - series d) in series - parallel Each of the above would have a different formula to compute the overall system reliability.
Figure 7. Representation of a block cave mine The ore passes, numbered from 1 to 3, and draw points are indexed according to position with respect to ore pass and draw point, e.g. would be the jth draw point in ore pass i. Draw points in the same panel will be in parallel and in series with the ore pass. Therefore the reliability of a panel is given by:
j (i ) RPi = ROPi 1 − C (1 − RDik ) k =1 where is the reliability of panel i, is the reliability of ore pass i, and is the reliability of the jth draw point located on panel i. J(i) is the number of draw points in panel i. Since the three panels in figure 7 operate in series, then the reliability of the system shown is given by: R = RP1 • RP2 • RP3 The base case scenario will consist of having the same reliability for all elements equal to 0.8, i.e. ROP1 = RDij = 0.8 for all i and j. Then the system reliability would be R = 0.487. Now consider two cases: Case 1: Suppose that the reliability of any draw point in panel 1 decreases to 0.5 due to poor draw control and operational factors. Then the reliability of the production area would be R = 0.486 (See Table 5), not a significant difference from the base case. Case 2: If instead the reliability of any draw point in panel 3 (the cave front) decreases to 0.5, then the reliability of the entire system would be R = 0.456. This is an interesting result; since the reliability model suggests that it is more important to keep draw points at the cave front in proper operation rather than old draw points to enhance the performance of the production area.
Figure 6: Processes connections presented by Herbers (1981) Consider a mine with three production panels. Each panel has an ore pass and several draw points linked to it, as shown in figure 7. Each of the panels has to produce an equal tonnage at any given period to maintain the uniform draw pattern in order to avoid early dilution as well as high stresses in the production area.
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Table 5: Production area reliability by panel for different cases Panel 1
Panel 2 Panel 3
System R
Base case
0.799
0.794
0.768
0.487
Case 1
0.797
0.794
0.768
0.486
Case 2
0.799
0.794
0.72
0.456
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The reliability model could also be used to help decide where to open a new draw point by showing in which way the system becomes more reliable. For example, consider Case 2 above where there is a draw point in the cave front with low reliability and it is desired to open a new draw point (with reliability 0.8) so that the entire production area becomes more reliable. Suppose also that operational constraints in panel 2 mean that the new draw point can only be located in panels 1 or 3. The calculation of the system reliability for both these scenarios is shown in Tables 6 and 7 respectively.
become so large that a "hang-up" occurs. Production from the draw points may still proceed, however, at a lower reliability. If production stops, then caving still occurs but this induces potentially damaging stresses onto the adjacent draw points, thus affecting production from the panel in the long term. Development of a realistic and robust production planning model for block caving operations is a challenging task. The issues and factors described above should also be included in such a model. 6 CONCLUSIONS
Table 6: System reliability for a new draw point in panel 1 (shown in bold) Panel 1 Panel 2 Panel 3 ROPi
0.8
0.8
0.8
RDil
0.8
0.8
0.8
RDi2
0.8
0.8
0.5
RDi3
0.8
0.8
RDi4
0.8
RDi5
0.8
RPi
0.8
0.79
0.72
System R = 0,46
Table 7: System reliability for a new draw point in panel 3 (shown in bold)
Even though current production scheduling methods may forecast the global tonnage to be mined per period accurately, there is often exists a significant tonnage variance between forecast and actual production on a draw point by draw point basis. This variance generally results from inadequate integration of the fundamental models that sustain the planning of a block cave operation and also the inability to deal with operational data. Reliability theory has been studied as a means to allow mine planners to account for the actual production performance as part of the mine planning system. Also, reliability theory provides a way to analyze the weakest link within a mine plan from a process point of view. There is a need to integrate the more fundamental process models and field monitoring data to improve the precision of production scheduling ACKNOWLEDGEMENTS
ROPi
0.8
0.8
0.8
The authors are grateful to Gemcom Software International Inc. for time and funding to complete the work and submit this paper.
RDil
0.8
0.8
0.8
REFERENCES
RDi2
0.8
0.8
0.5
RDi3
0.8
0.8
0.8
RDi4
0.8 0.79
0.78
Panel 1 Panel 2 Panel 3
RDi5 RPi
0.8
System R = 0,50
The results in Tables 6 and 7 show that the effect of opening a new draw point at the cave front is much greater than opening it at the back of the active area. This is because the contribution of the extra draw point in panel 3 to the overall reliability is much larger than in panel 1. However, a new draw point in the cave front will lead to increased stresses at the corners of the active area, possibly damaging the rock mass and affecting fragmentation at other draw points. This could render production from other draw points difficult or impossible, even though the reliability model suggests otherwise. This clearly illustrates the need to integrate the reliability model with geomechanical factors. Moreover, since geomechanical data are measured during operations, these relationships should account for the observed time variation in production reliability, thus providing a physical basis for adjustment of the draw point reliability. In addition to constraints imposed by geomechanical factors, there may also be operational constraints on the development and operation of draw points such as equipment availability and logistics associated with draw point development. These issues are more related to short term planning but could also affect the reliability of the entire production system. During production from a panel, it is common for the fragments in one or more draw points to 154
• Brown, E. T., (2003). Block Caving Geomechanics. Julius Kruttschnitt Mineral Research Centre, JKTech Pty Ltd. • Chanda E C K, (1990). An Application of Integer Programming and Simulation to Production Planning for a Stratiform Ore Body. Mining Science and Technology, 11:165/172 • De Nicola, R., Fishwick, M., (2000). An Underground Air Blast- Codelco Chile Division El Salvador. Proceeding Massmin 2000, Brisbane. • Dessureault, S., Scoble, M., Rubio, E, (2000). Simulating Block Cave Secondary Breakage - An Application of Information and Operations Management Tools in Mass Mining Systems. Proc. of Massmin 2000. (AusIMM: Carlton Victoria Australia) Brisbane, Aust. Oct. 29 - Nov. 22 2000, pp 893-896 • Dolipas R S, (2000). Rock Mechanics as Applied in Philex Block Cave Operations. Proceeding Massmin 2000, Brisbane. • Dunlop R and Gaete S, (1995). Seismicity at El Teniente mine: a mining process approach. 4th International Symposium on Mine Planning and Equipment Selection. October 31 –November 03, Calgary, Canada. • A R Guest, G J Van Hout, A Von Johannides and L F Scheepers, (2000). An Application of Linear Programming for Block Cave Draw Control. Proceeding Massmin 2000, Brisbane. • Herbers, J.M (1981). "Reliability Theory and Foraging by Ants," Journal of Theoretical Biology, 89 (1981) 175-189. • Kaufmann, A., Grouchko, D. & Cruon, R (1977). Mathematical Models for the Study of the Reliability of System. New York: Academic Press (1977). • Laubscher D H, (1994). Cave Mining- The State of the Art, J. S. Afr. Inst. Min. Metall. , Vol. 94, pp. 279-292.
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• Rahal, D., Smith, M., van Hout, G., and von Johannedis, A., (2003). The use of mixed integer linear programming for long-term scheduling in block caving mines. In Proceedings APCOM 2003: 31st International Symposium on the Application of Computers and Operations research in the Minerals Industries, May 1416, 2003, Cape Town, South Africa. • Rubio, E., Scoble, M and Dunbar, W. S., (2001). Scheduling in block caving operations using operational research tools. In Proceedings Minespace 2001. Annual General Meeting of the Canadian Institute of Mining, Metallurgy, and Petroleum, Quebec City, Canada.
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• Smith, M. L. and Rahal, D., (2001). Draw control optimization in the context of production scheduling. In Proceedings 17th International Congress and Exhibition of Turkey, Chamber of Mining Engineers of Turkey, Ankara, Turkey, pp 831-838. • Smith M L, (1999). The influence Deposit Uncertainty on Mine Production Scheduling. International Journal of Surface Mining, 13:173-178 • Wang, L. G., Yamashita, S., Sugimoto, F., Pan, C., and Tan, G., (2003). A methodology for predicting the in situ size and shape distribution of rock blocks. Rock Mechanics and Rock Engineering, 36, 121-142.
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Integrating Work Index into mine planning at large scale mining operations Jose A. Caceres S., MSc(Eng), PhD Candidate Charles W. Pelley, PhD, P.Eng P.D. (Takis) Katsabanis, PhD, P.Eng. Shadan Kelebek, PhD, Associate Professors, Department of Mining Engineering, Queen’s University
Abstract Mine planning and mineral processing optimization are usually treated as two unconnected problems, especially at open pit and panel caving mines where the cut-off grade is practically the only variable optimized and analyzed for in the planning proposes. Using existing planning tool, the output is a reserve consumption strategy privileging higher grades. With the earlier consumption of the highest grades, every year produces revenues which are closer to the cost and potentially even below cost. However, there are other ore characteristics that can also affect the profitability of the operation. A new methodology has been developed, incorporating the grinding and flotation relationship into the mining economic models for the simultaneous analysis and optimization of the throughput-work index-recovery relationship through changes in the mining limits, sequence and redefinition of the reserve consumption strategy. A mine-site that implements this new planning strategy, can expect to increase the NPV from 5 to 15% depending on the actual operational settings. Two case studies have been developed, showing how the inclusion of the work index in the economic model changes the phase and mining sequence in an open pit mine and the caving sequence and the optimum column height in a panel caving mine.
1 INTRODUCTION The most recent studies integrating mining and mineral processing have been named "Mine to Mill". These studies try to maximize the net present value by optimizing the recovery-throughput relation for an on-going operation. The concept has been previously explored. Carlisle(1954), suggested the advantage of increased throughput over the mill design capacity, Steine(1978) shows the potential improvement of cash flow by the throughput increments of increasing grinding size, without considering the impact on the mine planning. Camus(2002) suggests the advantage of increasing throughput by sacrificing recovery, considering the recovery as a lineal function of throughput. These studies assume a maximum capacity in the mill, accept the actual mine design and mining schedules, and optimize only the cut-off grades. Initial apprehension from mineral processing professionals for these kinds of studies is understood as it is risky to assume a maximum capacity without having information about the rock properties. It is not valid to assume a potential recovery without having certainty about product size as mineral processing performances are evaluated by throughput and recovery goals. Considering these factors, the "initial" apprehension is largely justified and explained by the uncertainty in one of the important variables for the performance and design of conventional grinding circuits, the Work Index (Wi). The research presented here, developed by the Queen’s University Mine to Mill project, incorporates previous researcher’s findings but goes a step forward with the inclusion of Work Index into the mining planning. 1.1 Wi Planning Methodology The methodology is a recursive algorithm for long term planning for operating mines. The method starts by assuming a known mining method, an initial mining 156
schedule, an operational plant with its behavior studied by laboratory testing or developed from historic data. Better models or better optimization engines can improve every step of the methodology. The target function is the NPV for the long-term planning, and changing that target would not change the structure of the methodology, just the parameter and potentially the results. 1.1.1 Concentrator Elements Considering an operating plant, the main work must be centered in grinding and flotation. In conventional grinding circuits, the first task is to build throughput-Wi-product size relationships, usually based on the Bond approach, calibrated and modified for real performance data. The analysis will generate a set of equations or surfaces placing a potential throughput in function of product size (Usually P80) and Wi. The achievable production rate will be indicated by the grinding analysis and must be constrained by a maximum capacity analysis. In flotation, the task is based on statistical analysis using ANOVA, Stepwise Regression or Structural Equation Modeling to build a predictive model for the Metallurgical Recovery (Cu, Nickel….etc). The model can be checked, corrected even replaced by flotation batch tests, where the recovery is well correlated with control variables (grinding size, average grade, lithology), the support must be accompanied by a sensitivity analysis where the predictive capacity can be checked. Back analysis and/or cross validations are highly recommended. The flotation analysis must include the time/ residence effect on the recovery, by the statistical analysis or the flotation kinetics made on the plant. The economical analysis in the concentrator must be able to split the fixed and variable costs (considered fixed in this marginal scenario where no new investments are considered). The main items to be analyzed generally must be the energy and iron consumption of grinding.
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The final elements must be the analysis of the maximum capacity reachable for the concentrator; the analysis must review all the plant stages and identify bottle necks. Summarizing, the elements required in the Concentrator to apply to Mine Plant Simultaneous Planning are the following: • The P80-Tonnage-Wi Relationship • A Recovery (P8,Wi, Grade, …, other), predictive model. • Maximum Capacity Analysis • The Fixed and Variable Cost Structure. 1.1.2 Mine Elements. Assuming a mining method is already selected, a mining limit and sequence has been defined, a cut-off optimization has been carried out based on the traditional throughput definition (Probably defined by the concentrator); the mine must produce an initial reserve schedule and mining plan with a average grade and Wi by period (if the Wi optimization is going to be included). All these initial approaches are useful to build a Base Case to use an initial evaluation. The same structure’s cost analysis carried out on the concentrator must be repeated on the mine, splitting again the fixed and the variable costs. Finally, just like the concentrator, a maximum capacity analysis must be included. For an open pit mine, the capacity is almost unconstrained and is directly related with the equipment. For underground mines upper limits exist and must be identified. Summarizing, the elements required to apply to mine and plant simultaneous planning for mining professionals are the following: • An Initial Reserves Schedule • A Maximum Production Rate Analysis • The Fixed and Variable Cost Structure. 1.1.3 Evaluation Methodology including a Variable Recovery A critical step is to accept the fact that recovery cannot be estimated as a constant value for the next 25 years. All the previous and following steps, and the extra value gained by carrying out better planning, will be wasted if at the end, somebody declares "The recovery for the next 25 years will be 89.99%". The most critical step is to modify the actual evaluation system in every mine site by including the recovery as a reliable function of input data, like throughputs, work index, grades or rock types. The evaluation of any system depends on every operating characteristic. 1.1.4 Base Case Economic Evaluation Considering the data generated for the mine and the concentrator, an economic evaluation estimating the initial NPV has to be built. The economic evaluation must be the result of the compromises assumed by the mine in the longterm period, accepted by management, and define the most relevant parameter of the mine. In most cases this will be the NPV of the actual operation. 1.1.5 The Optimization Module. The optimization model must recreate the economic value of the base case, (based on the economic evaluation already made) and must be able to calculate directly the target function, net present value, cash flow or rate of return as a function of changes in tonnage, recovery, grades, and Wi. This module, core of the Mine to Mill, must be able to calculate average grades facing increases or decreases in the production rate, and simultaneously must calculate the new recovery caused by new throughput and/or new grades and Wi. The way to solve the problem in an open pit is using the incremental technique provided by software package like Massmin 2004
Opticut. In underground mines, an incremental technique can be set up. Once all the variables are set up and related to each other, it is possible to estimate the effect of changes in throughput and Wi on recovery and liberation size to find the optimum production rate. The optimization engine could be an iterative method like dynamic programming, or multivariable methods for non lineal problems starting with the generalized reduced gradient, considering an important number of starting solutions to increase the chance of finding the global optimal solution. 1.1.6 Wi inclusion. The Wi has to be included in the base case, the production plan, the recovery-production rate functions and into the maximum capacity analysis. The Wi must be included in the economic model affecting simultaneously production rate, capacity and recovery, in the block model or mining units, (Blocks, Benches, Stopes, Panel and Areas) and reevaluate these mining units and apply the method used to define the sequence depending of every mining method. For example for an open pit, it is necessary to build the profit matrix and apply the Lerch & Grossman method or Moving Cone with the decreasing price series or increasing cost series and define the new potential phases and sequence and optimize the cut-off grade including the variable recovery and re-evaluate the production schedule. For Panel Caving mines, the method is similar, accepting the caving level, including the effects of Wi on the caving limits, the caving sequences and Lane’s method applied to optimize the column height. Figure 1 shows this concept. 1.1.7 Stop Criteria. The global methodology stops when the difference between the NPV for the solution immediately posterior and the actual solution is lower than a tolerance value.
Figure 1 Wi Inclusion Concept
2 INDUSTRIAL CASE The following is a simplified industrial case where grades, work index, mineral processing data and correlations belong to an actual mine site in central of Chile. With the intention of simplifying and enhancing the general character of the exercise and protecting the mine information, the analysis is based only on copper as the main valuable component. The analysis considers a conventional circuit and applies the methodology previously described for an underground Panel Caving.
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2.1 Concentrator Analysis The grinding circuit is composed of a single open-circuit mill plus a conventional mill initially designed for 33 [Kton/day], based on an average Wi of 15 [Kwh/ston]. The capacity analysis shows that without a constraint in the product size the technical maximum capacity will be 38 Kton/day, limited by the behavior of the rod mill. Additionally there is a product size constraint in the range 260-300 um because of pulp transportation problems. The throughput, work index, and product size relations, calibrated with actual data, are shown in Figure 2. Statistical Analysis shows that the P80 explains more than 80% of the variance in the recovery data and flotation kinetics ensures adequate residence time for a scenario of 38 Kton/day. This justifies no inclusion of a penalty in the recovery associated with a lack of residence time. Figure 3 shows the P80-Recovery relation considered. These figures are from Kelebek (2000).
Table 1 Economic Parameters Underground Mine Parameters Cu Price Discount Interest Rate Total Mining Cost Processing Cost Mining Cost Processing Variable Cost
130.00 35.00 10.00 3.26 4.30 0.98 2.15
cUS$/lb cUS$/lb % US$/Ton US$/Ton US$/Ton US$/Ton
Furthermore, it seems logical to assume that the "soft rock" is associated with the higher grades which assumes when planning using opportunity cost only one should get the best grades and softest rocks in the first years of the mine life. Under this assumption it is enough to consider the average Wi from the annual reports to equal 15 Kwh/st.
Figure 2 Actual Conventional Grinding Circuit Calibration
Figure 4 Optimal Sequence, Average Wi Considering a production rate of 33 Kton/day, average, an average Wi (15 Kwh) and the economic parameters from table 1, the optimal caving sequence, is highly orientated to the south part of the mine as shown in Figure 4. This follows higher grades and higher column heights due to the copper price considered. Table 2 shows the production plan and economic evaluation resulting after smooth the caving sequences and considering the concentrator information shown in Fgures 2 and 3. Table 2 Production Plan Average Wi NPV = 1164.8 MUS$
Figure 3 Recovery-P80 Relation
2.2 Underground Mine The underground mine is a Block Caving operation with a maximum capacity of 48 Kton/Day because of area and caving constraints. The analysis is based on copper for envelope limits and caving sequences design. Lane’s methodology was used to define the optimum column height. A copper price of 130 US$/Ton was used with an interest rate of 10%. Table 1 shows the economic parameters considered. The base case considers the same economic parameters from Table 1 and a production rate of 33 Kton/day. An important first step is to examine what happens if work index information is simply omitted because, for example, the high copper price suggests that grade is far more important or as in the recent past with a price of $US0.70 /lb, there is neither the time or economic resource to get Wi information. 158
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The production plan is one from a mine site which does consider mine to mill relationships and assumes a flat recovery and a flat throughput. But, what happens if the previous plan does not assume an average work index but instead measures and includes those values in the block model? Because of work index inclusion, recovery is no longer flat but changes according to liberation size, reflecting the actual behavior of a flotation plant. The assumption of 15 Kwh/ston for the following 17 years has a cost in net present value of 54 MUS$ as shown in Table 3. Very few planners would predict that inclusion of the actual Wi in their plan would reduce the NPV by 5%. This demonstrates the problem with building production plans using non-evaluated assumptions or, worst yet, to choose to ignore such an important factor in their evaluations.
The 33Kton/day production plan, after smoothing the caving sequence and now including Wi and Cu grade has a net present value of 1149.6 MUS$. This recovers 38 MUS$ of 54 MUS$ initially "lost" because it assumes a more realistic flotation plant. Table 4 shows the production plan and evaluation for this new sequence. Table 4 Production Plan and Evaluation Wi-CU Sequence
Table 3 Production Expected Work Index NPV = 1111.5 MUS$
The important next step is to re-analyze, the caving sequence and column heights now considering Wi and concentrator relations. The resulting caving sequence, as shown in Figure 5, is strongly central and north oriented, because of the value added or subtracted by work index in potential throughput increases, and variable recoveries.
Figure 5 Caving Sequence with Wi Inclusion Massmin 2004
It is necessary to define the optimal production rate based on information from the block model, projecting grades and Wi as function of changes in throughput. This analysis must connect the reserve behaviour to define the optimal throughput with a NPV maximization target. Figure 6 shows the profit and recoveries for a simplified case considering a fixed grade. Is important to highlight that this kind of "static" analysis is only for the academic purposes, as it is impossible to assume that with increased throughput the grades will remain constant. Connecting every block each containing the true differing grades, Wi and cost to the optimal throughput decision, and associating the mining variables such as sequences and column height while always respecting the constraint of maximum column height and extraction speed, is the key to real mine to mill optimization. Building the optimal production plan under the concentrator’s constraints, increases the NPV for the 17 years of mine life by 8,7 % as shown in Table 5, and increases the profit by 14 and 32% during the first two years. (Figure 7). It is necessary to question if all the sequences and column height analysis is necessary. Why not accept the actual sequence and column heights and just go to the maximum capacity of 38 Kton/day. This is similar to Carlisle (1954) who states "Except for gross maladjustment of mill capacity to ore reserves, the optimum rate of recovery in the short run is likely to be close to designed mill capacity. In many mines it is profitable to work the mill at greater than designed capacity".
Figure 6 Profit and Recovery as Function of Throughput static analysis. Santiago Chile, 22-25 August 2004
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Table 5 Expanded Production Plan Optimal Cu-Wi Sequence NPV 1206
especially those based on selective mining, there is no room for optimization, grades are extremely confined to small spaces and there are no mining options. However, even without sequence optimization, it is healthy to estimate the behavior of a flotation plant based on something better than an average value. The next industrial Case analyzes the correlation between Wi and Grades for an Open Pit Mine. 3 ANALYSIS AND CONCLUSIONS
Accepting the original sequence and building a production plan for 38 Kton/day (Table 6) without including the Wi increases the NPV by just 1.3%, while increasing throughput by 15% and forcing the P80 over the constraint of 260-300 um, getting values around 310 um the year 1 and 322.8 um year 9. This is certainly not an acceptable route.
After reviewing the underground case, the following conclusions can be drawn. • Work Index is a variable that in any case justifies its inclusion in mine planning at an early stage, and has an importance equivalent product grades. • Inclusion of Work Index, will imply changes in mining sequences and mining variables as important as column height definition. • In concentrators where recovery is strongly related to product size or P80, work index selection or modification by better mining sequences is a valid way to increase throughput without additional investment. • A flat recovery curve, smooth work index-P80 relationship, plus a flat grade profile, is the manner to increase revenues without additional investment. • Metallurgical recovery improvements by work index reduction have the potential to significantly increase cash flow and NPV. • The improvement in mining sequences by the inclusion of work index data has the potential for significant increases in profitability, as demonstrated the case study presented. • Relating this analysis to initial project design and linking the mine sequence and Wi, with the plant design is an unexplored area with an enormous potential. ACKNOWLEGEMENT Mr. Carlos Valenzuela Vega Metalica Consultores’s Engineering Chief. Metalica Consultores for the support provided to the author. The mines that provided the information used in this work and congratulate them for risking the resources required to generate the Work Index information and including it in the block model.
Figure 7 Profit Comparison Underground Case Table 6 Plan 38 Kton/day Original Sequence
REFERENCES • Bond F.C. (1961) "Crushing and grinding calculations" pp 1-12, British Chemical Engineering, 6, 1960 (Revised 1961 by Allis Chalmers Publication 07R923B). • Camus, J., (2002) "Management of Mineral Resources: Creating Value in the Mining Business", An SME publication ISBN 0-87335-216-5. • Carlisle, Donald (1954) "The economic of a fund resource with particular reference to mining". The American Economic Review. September issue, pp. 595-619. • Lane, K. 1988 "The economic definition of ore" • S. Kelebek, (2000) "Analysis of Andina data on the effect of primary grind size on the recovery of Cu with implications on processing tonnage", An internal Queen’s University report for Metalica, Santiago Chile. • Steane H.A. (1978) "Coarser grind may mean lower metal recovery but higher profit" Canadian Mining Journal, June issue, pp 2-6.
A second and unavoidable question is whether it is possible to get a lower Wi without decreasing Cu grade, is there a correlation between them?. For some mines, 160
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An evolutionary model for underground mining planning José Saavedra, Facultad de Ingeniería, Universidad de los Andes Marco Alfaro, Facultad de Ingeniería, Universidad de los Andes and Metálica Consultores S.A. Jorge Amaya, Centro de Modelamiento Matemático, Universidad de Chile
Abstract In this work we present a methodological approach for finding near optimal solutions to the problem of defining plans for underground mines. This problem of obvious combinatorial nature is intractable by means of traditional techniques (Mixed Integer Programming for example). The approach proposed here is based in the mechanism of natural selection, we construct a Genetic Algorithm to conduct the search of a approximate solution to the problem. In order to acquire this objective, we first need a model which characterizes the gravitational flow of material from the drawpoints, the chosen model for this objective is a cellular automata specifically designed with very simple rules of local evolution. The model was implemented and tested, the time needed for a solution in real type cases is much less than the time human planner needs actually for the same task. Possible extensions to this model are presented.
I. INTRODUCTION Underground mining operations are very complex in nature. The main factors that determine this complexity (between others) are the unsuitable knowledge of ore resources contained in the mine and the parameters that characterize them. From this point of view, we can obviously see that human planners are unable to handling this complexity and as a consequence very poor solutions are obtained. The problem of underground mining planning can be defined as choosing the best production plan in order to maximize the benefits derived from the operations. This planning could be made in three distinct scenarios: short term (operational), mid term (tactical) and long term (strategical). In this paper we focus our attention to the operational case. Many resources are wasted in preparing mining plans. Usually this task takes two or three workers dedicated only to this work. Moreover, they don’t have any tools available and this imposes on the final solutions the human planner bias. Some efforts have been made to solve this situation. From the point of view of classical optimization we can consider a model that has the following features: • The mine is subdivided into blocks of homogeneous dimensions that conform a domain without holes (edge connected). • When we extract a block from a given drawpoint, the blocks that are in the same column descend in one position (precedence restrictions) . • Ore extraction is supposed to be realized in a soft manner, in the sense that adjacent columns can’t show high differences in height, this restriction prevents dilution entry. • Usually the objective function of this kind of models is the Net Present Value (NPV) of the economic result of the business derived from the operations.
Another important source of problems in this deterministic perspective is the combinatorial explosion of the problem. We have the following result: Proposition I.1 If we call Sk,l,m the number of feasible sequences in a sector of k blocks width by l blocks length by
m
blocks height, with m ≤ k,l, the we have (k • l)m < Sk,l,m < (k • l • m)! the proof of this proposition can be found in the work [7]. As an application of this result, if we consider a sector with dimensions then we are in presence of at least possible sequences, if we take 1 second in evaluating each one of this sequences then we need at least 34048129883307965499746321664130 years in order to resolve this problem. So an obvious conclusion is that exhaustive search is a very bad strategy for this class of problems. Of course not every sequence is feasible. One possible choice, in order to reduce this high number of sequences, is to make an algorithm that can generate feasible sequences. It’s not hard to see that this approach faces other problems that are not easy to resolve, for example to decide if a given sequence is feasible. In this paper we propose a model that breaks this classical approach to the problem of underground mining planning. We choose as an alternative an Evolutionary Model because of the flexibility and good empirical results in problems of higher complexity (like the one presented here). II. EVOLUTIONARY ALGORITHMS
this point of view is the dominant one presented in the work [6]. One of the problems that isn’t resolved by this approach is the incorporation of the stochastic behavior of Gravitational Flow. Because of this, we call to this kind of models deterministic. Massmin 2004
Genetic Algorithms (GA from now) were introduced by John Holland in 1975. They are inspired in Darwin’s mechanisms of natural selection. Such mechanisms establish that an individual is generated as a mixture of the genes from his parents by means of crossover, added to this
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process of mixture there is a process called mutation (change in some segments of genetic material). This last mechanism implies in some way evolution because add novel elements not present in the genetic information from parents. Finally, the adaptation to the medium makes that some individuals survive and inherit their genes to his sons. The general form of an Evolutionary Algorithm is the following:
Definition IV.1 Given a set of drawpoints {ηi}ni=l we define a Extraction Chart as a matrix M ∈ Mmxn with η the number of drawpoints and m the number of turns. The coefficient mij of this matrix is defined as:
t := 0; initialize (P(0)); evaluate(P(0));
mij : = tons (expressed in shovels) to extract in drawpoint j in turn i
While Not has been done Do P (t) : = select parents (P(t)); P (t) : = recombinate (P’(t)); P (t) : = mutate (P’’(t)); evaluate (P’’’(t)); P (t + 1) := natural selection(P(t); P’’’(t)); t := t + 1;
Observation IV.1 Extraction Chart M don’t have to be a square matrix. If we call to demand (in shovels) in turn i then we have n
∑m
ik
Next In this Algorithm t is the counter of generations and evaluate (P) implies to evaluate fitness function to every member of population P. This algorithm finish when the fitness value of actual population P(t) in time t don’t innovate or after a fixed number of iterations. The considerations in the moment of implementing GA strategies are: chromosomic representation, population size, fitness function, crossover and mutation operators, crossover and mutation probabilities. For further reading on this technique go to references [3], [5], [7].
" ≤ ·" in the last restriction gives the choice of not satisfy the hole demand, if not we could force our algorithm to extract blocks that gives worse solutions. B. Chromosome Representation The natural chromosome representation for a extraction chart is a matrix like the one previously defined. This will be the formal structure in which we will define the crossover and mutation operators. C. Crossover Operator Given two individuals (matrix) from a given population, for example M;N, we define the associated crossover operator as:
III. MODELS FOR GRAVITATIONAL FLOW If we want to incorporate the stochastic behavior of granular flow in our model we need first consider models for this phenomenon. In the last decade, some efforts have been made. The main results were obtained by chilean well known scientists, Eric Goles [1] and Servet Martínez [4]. These two approaches are similar in the kind of technique used, both are cellular automata models. Another interesting model is the one proposed by Gregorio González [2]. In this work the ideas presented in [1] are refined. Applications to underground mining are presented. The most recent development in this area is the model presented by Marco Alfaro [8]. This model is a Cellular Automata that has the benefit of simple rules of evolution, and as a consequence, it’s possible to obtain a efficient implementation [9]. Independent of the chosen model, it’s fundamental for the proposed methodology to have some gravitational flow model. As we will see in the next section, our model consider a gravitational flow model in the kernel of the evaluation function of the proposed genetic algorithm. IV. PROPOSED EVOLUTIONARY MODEL A. Extraction Charts Operations in underground mining are realized in turns. Such turns are usually of 8 hours each one and each day is divided in three turns. Given a set of drawpoints {ηi }ni=1 a Extraction Chart is an assignment of tons to extract for each drawpoint en each turn. We can if we wish consider another kinds of periods like days, weeks, months, etc. Formally: 162
≤ D1
k =1
• We select randomly a number in the set {1,...,m} (i.e. we select randomly a turn). Let i* such number. • We consider the submatrices of M and N given by M1 = = (mij) i*i=1, j ∈ {1,...,n} and, M2 = (mij)mi=i* +1, j ∈ {1,..., n} and analog for N. • We define the new matrices
N M M = 1 and— = 1 M2 N2 This crossover operator guarantees feasibility from turn to turn of extraction charts, this because we maintain demand inequality in each turn. D. Mutation Operator To define a mutation operator we have to randomly select a turn (row in the extraction matrix). We proceed then to rebalance the selected row with a number randomly chosen between 0 and the demand of the turn, then we distribute randomly in to the selected turn. This operator guarantees feasibility of the selected row (the number chosen is less than demand). E. Fitness Function In order to evaluate the fitness of a extraction chart we have to run a simulation of the chart and then obtain a list of blocks with laws of ore grades. With this information we can obtain the benefit given by that extraction chart incorporating NPV in this calculation.
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F. The Algorithm The search algorithm is described with the following pseudo-code: Algorithm. Inputs: Block Model Location of Drawpoints Population Size n Parameters: Crossover and Mutation Probabilities Iterations Probabilities for Cellular Automata Outputs: Optimal Extraction Chart Algorithm: Initialize Extraction Charts Population (n Extraction Charts): P(0); Fitness Evaluation(P(0)); For t = 0 To V Begin Crossover (P(t)): P(t + 1); Mutation (P(t + 1)); Fitness Evaluation (P(t + 1)); Next Generation Selection (P(t),P(t + 1)); End For
FIG. 1: Best Individual Evolution. As we can see, the algorithm generates a sequence of solutions in which each one is at least equal or better than the previous. B. Calibration of Model Parameters In order to operate this GA, we need calibrate the functional parameters. The used example was the previous one. We will vary crossover and mutation probabilities. To denote the instances of the problem we will use the notation An1/n2M, where n1 is crossover probability, n2 mutation probability, selection method A (Parents and Sons: PH; Only Sons: SH) and parent selection M (Roulette: R; Drawing: S).
Function Fitness Evaluation(P(t)); For i = 1 To n Begin Simulate(Extraction Chart i(t)); ↵ Cellular Automata(Extraction Chart i(t)); ↵ Economic Evaluation(Extraction Chart i(t)); End For End Function
Each instance was runned 30 times and we determine in which generation we reach the optimum. The following table resumes the results.
V. NUMERICAL RESULTS A. Trivial Case This case is a sector of 4 by 4 by 4 blocks, all of them with grade 0, i.e. sterile. This simple example has only two extraction points, one in coordinates (1; 1; 0) and the other in coordinates (3; 2; 0). The obvious solution to this problem is to extract nothing from drawpoints. This example was tested with the following parameters: Application Parameters Turn Number
5
Max. Demand in each Turn
10
GA Parameters Iterations Number
100
Population Size
10
Crossover Probability
0.8
Mutation Probability
0.2
Selection Policy
Between Parents and Sons
Selection Method
Drawing
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0.10 0.51 0.10
Instance
Mean
Standard Deviation
PH80/20S PH90/30S PH80/30S PH80/40S PH80/50S PH90/20S PH100/20S PH100/0S PH100/100S SH100/100S SH80/20S PH100/100R
64.93 50.17 45.70 38.33 35.40 61.13 55.23 100.00 31.70 100.00 100.00 28.97
22.89 21.55 20.36 20.33 15.70 27.64 27.17 0.00 (*) 9.50 0.00 (*) 0.00 (*) 15.01
(*) means that instance doesn’t converge never in 100 iterations.
The solution to this problem was obtained in 63 iterations. 0.02 0.10 0.02
The results of the iterations are summarized in the following picture. This picture illustrate the behavior of the best solution in each iteration:
0.02 0.10 0.02
C. Another Factors We prove many others effects: Maximum Demand, Number of Blocks, Population Size, Drawpoints Number. Almost all of them gives a linear dependence between the number of iterations needed to reach the optimum and the increase of the values. The only factor that shows exponential behavior is Maximum Demand, i.e., if we vary the Maximum Demand parameter then the number of iterations needed to converge grows exponentially. We summarize this effects in the following table:
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Effect
Variation
Result
Max. Demand Number of Blocks Population Size Drawpoints Number
Grow Grow Grow Grow
Exponential Grow Linear Grow Linear Decrease Linear Grow
D. Real Scale Example This example was runned with the following parameters: Application Parameters
VII. CONCLUSIONS
Turn Number
12
Max. Demand in each Turn
100
Number of Drawpoints
171
Number of Blocks
400000 GA Parameters
Iterations Number
34
Population Size
8
Crossover Probability
1
constructed with the same philosophy so we can expect some kind of integration and scale economies between + both models. In the future we hope to construct a model that incorporates gravitational flow simulations and downstream operations. Another source of extensions to the model is to consider operational events as stochastic processes. In this way simulations of extraction charts would be more realistic. In order to accomplish this objective is needed historical data to calibrate the parameters of the involved stochastic processes.
Mutation Probability
1
Selection Policy
Between Parents and Sons
Selection Method
Drawing
The results are summarized in the following table:
As an obvious first conclusion we have that this problem has a very large number of involved variables. In this moment it’s really very difficult consider the hole complexity of this problem. For example at the moment there aren’t appropriate models to handle the breaking of the solid rock. Another important conclusion is that the stochastic nature of the phenomenon is hard to include in the modeling process. Some attempts have been made but at the moment this efforts are in initial development. The proposed methodology could be applied in real type situations. The response times are good compared with human planners. The most important advantages of the methodology is that we can test in an efficient way many choices and that the search procedure is well conducted. With this methodology we are in presence of a flexible model that can be customized to satisfy the planner needs. VIII. ACKNOWLEDGMENTS
Execution Time
3:11
J.S. wish to acknowledge the support of FAI (Fondo de Ayuda a la Investigación), Proyecto Ingeniería 2004, Universidad de Los Andes.
Iterations Number
34
REFERENCES
Results
We can extrapolate this result and see that in real situations it would take about 8 hours to finish the optimization process. VI. EXTENSIONS TO THE MODEL The proposed model don’t consider downstream operations. These operations are in general the most restrictive operations. For example in some downstream operations smooth ore grade is required, the fine mid term promise has to be accomplished, etc. In all of these cases the proposed model don’t give an answer. Recently, an extension of the model proposed in this paper have been implemented [10]. This model uses Genetic Algorithms too and the main characteristics are: 1. Genetic Algorithms mechanisms in the search of solutions. 2. Mixture Models for Ore Unload. 3. Operations are considered as transport problem. 4. Restrictions on the quantity of ore to be extracted from drawpoints are imposed. 5. Capacity constraints in downstream operations are considered. 6. NPV evaluation. 7. Drawpoint Grade behavior is assumed. In real type situations, the response time of this implementation are in the order of 3:00 hrs [10]. The next challenge is to integrate the model proposed in [7] whit the one proposed in [10]. Both models were 164
[1] Eric Goles y Sebastián Peña: Modelamiento y Simulación del Flujo Gravitacional. Informe Final PB510004, Apéndice N°2, 1996. [2] Gregorio González: Estudio del Comportamiento de un Material Granular Mediante Modelos Computacionales. Memoria para optar al título de Ingeniero Civil Matemático, 1999. [3] John R. Koza: Genetic Programming: on the programming of computers by means of natural selection, The MIT Press, Cambridge, Massachussets, 1992. [4] Servet Martínez: Consideraciones Acerca del Modelamiento de Flujo Gravitacional. Informe Final PB5- 10004, Apéndice N°3, 1996. [5] Zbigniew Michalewicz: Genetic Algorithms + Data Structures = Evolution Programs, Springer, 1996. [6] Nelson Morales: Modelos Matemáticos para Planificación Minera. Memoria para optar al título de Ingeniero Civil Matemático, Universidad de Chile, 2002. [7] José Saavedra: Secuencias, Flujo Gravitacional y Evolución en Planificación Minera. Memoria para optar al título de Ingeniero Civil Matemático, Universidad de Chile, 2002. [8] Marco Alfaro, José Saavedra: Predictive Models for Gravitational Flow. To be presented in MassMin 2004. [9] Carlo Calderón, Marco Alfaro, José Saavedra: Computational Model for Simulation and Visualization of Gravitational Flow. To be presented in MassMin 2004. [10] Andrés Donoso: Modelos para la Planificación Operacional de Producción en Minería Subterránea. Memoria para optar al título de Ingeniero Civil Industrial, Universidad de Los Andes, 2004.
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Chapter 5
Gravitational Flow
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Theory of gravity flow (Part 1) Andrés Susaeta, Mining Consultant, Professor Mining Department, University of Chile, Chile
Abstract The results of 4 years of a research project over gravity flow based on sand model experiments, conceptual analysis and back analysis from the Codelco mines is presented. The emphasis of the paper is on a "knew concept" of interactive gravity flow, that has been developed from sand experiments results and conceptual analysis, and confirmed with empirical observations from the 3 underground Caving operations of Codelco. The proposed model for gravity flow is a function of extraction, material properties and layout geometry. This model integrates the experimental and conceptual results to generate a general model of gravity flow or Mass Flow Behavior. Several indexes and concepts to assess and control relevant issues of the gravity phenomena are proposed to characterize and evaluate the Mass Flow Behavior.
1. INTRODUCTION The understanding of the physical phenomena of gravity flow is of utmost importance for the optimization of ore recovery in caving methods, through a proper design of the extraction level and of the draw practice and control. With the results of several experiments with a specially designed sand scale model, conceptual analysis of caving material behavior, specific scaling relations and evaluation of data from the three underground operating mines of Codelco, a model of gravity flow propagation in a panel caving is proposed. The proposed model considers the basic variables that define the stages of a cave before the flow is established: initial state and caving propagation, but does not deal with this two processes. Initial state: The initial state of a caving column corresponds to the "in situ" Rock Mass, that can be characterized through several classification systems. All of them considering IRS, FF and FC and the original stress field. The potential for caving, the prediction of primary fragmentation and the design of the caving layout has to be assessed from this initial state. Caving Propagation: The propagation of the caving, determined by the caving sequence and draw policy, generates a dynamic stress environment. The extraction of 15 to 20 % of the in situ draw column is required to propagate the cave up to the surface or "waste" limit. The proposed gravity flow model does not include the caving propagation phenomena. It mainly deals with the gravity flow generated by extraction in a caved panel / block. The generally accepted knowledge and fundamentals regarding gravity flow during the last century, acquired both from models and mining practice can be summarized in the following points : • A draw point extracted in isolation of its neighbors, generates an isolated draw column. • The diameter of the isolated draw column is a function of basically the fragmentation of the material (and of its moisture). • In order to minimize dilution entry it is necessary to have even draw of neighbor draw points. • The isolated draw diameter is independent of the draw point width (common acceptance except Laubscher). Massmin 2004
• The nearer the draw points are the better the recovery of the ore. • Fragmentation increases with the extraction of the ore (The higher the extraction finer the material). • The higher the draw column, less the total dilution to be extracted (one extraction level better than two). • Interaction between two drawpoints exists if the distance between them is less than 1,5 times the isolated draw diameter. If all of these concepts were possible to achieve in the design and normal operation of a cave mine, the following model wouldn’t be of great need to understand and manage where and how to minimize the loss of ore due to dilution. Fm = f (E (estraction), PM (material properties), Gm (Layout Geometry)) The problem is that with primary rock cave mining, that requires handling of big fragmentation, it has been more and more difficult to achieve even draw (due to secondary blasting, hang ups, loss of draw points due to collapses, etc.) and to design fully interactive layouts that have a proper geometry and ensure good interaction. The proposed model can contribute to the understanding of the gravity flow phenomena, to the control and draw practice and for interactive design of primary rock layouts. 2. PROPOSED GRAVITY FLOW MODEL The model integrates experimental and conceptual results generated in two research programs that totalize more than four years of investigation (CIMM – Fondef 1037). It proposes relations between the control parameters of the variables that define the phenomena of gravity flow, based on analysis over empirical results. The behavior of gravity flow of particulate matter seen under the perspective of mass movement, can be defined as a function of the following variables: where: • Fm :Mass Flow behavior. • E: Extraction of material can be characterized after the parameters defined in the time series of a draw
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point defined in Figure N°1 and through the Uniformity Index. • Pm° : The principle properties of the caved material from the perspective of gravity flow, are the internal friction angle (ø) and moisture. The relation is that the greater ø minor the Isolated Draw Diameter (Dta), and greater the moisture of the material (up to certain limit) minor the Dta, where : ø = f ( δ (Density of caved material), R (Rock characteristics), F (Fragmentation)) with: δ : Density of caved material, corresponds to the apparent density of the caved column expressed in ton/m3. The density depends mainly of the hight of the column of caved material (sv), and the alteration generated by the extraction in the production level. The higher the density, the higher ø. R : Characteristics associated to the rock, corresponding to rugosity, IRS (Intact Rock Strength), angularity of fragments, etc. F : Fragmentation, corresponds to size of particles and gradation. The greater the particle size (D80),greater ø. Additionally the better the size graduation (all sizes present) greater ø. Moisture changes the mechanical properties of the caved material giving it cohesion, and by it diminishing the Dta, up to certain range of moisture, over which liquefaction risk exists, generating "mud flows" that constitute one of the great risks of the method. • Gm: Geometry of the layout considers distance between extraction points and it´s distribution as well as the caving geometry (topography, stress, pillars, etc.)
Fm (caving mass flow behavior) function has at least three different models that characterize its behavior:
"Isolated Flow", "Isolated – Interactive Flow", and "Interactive Flow", that are described as follows. ISOLATED FLOW The main characteristics of the first model that corresponds to "isolated draw" are: • Dta (Isolated draw diameter): A draw point that is extracted in isolation generates a "vertical chimney movement" of a diameter approximately constant in all over its hight. This diameter is only affected by waste due to attrition as any ore shaft. This diameter defined as "isolated draw diameter" (Dta) is defined mainly by the internal friction angle (ø) of the caved material and its moisture. The Dta is independent of the width of the extraction point and is reached at a height of 2 x Dta from the extraction point. • Uniform Settlement: Even if there is isolated draw when draw points are closed and time is given the caved material "settles uniformly". This phenomena is generated by a slow propagation movement within the cave due to equalization of stresses within the cave, towards the isolated draw columns. Lateral movement is generated towards the draw columns producing a uniform settlement in surface. This has been observed both in models and in caved areas. The dynamics (time propagation) of the settlement or "slow density equalization" has not been studied in detail, but is evidently much lower than the flow in active draw. INTERACTIVE FLOW AND ISOLATED – INTERACTIVE FLOW It is proposed that the phenomena of gravity flow is affected by the superposition of two states of the system that generates the movement: (a) Extraction point open and (b) extraction point closed: a) Extraction Point Open Behavior of gravity flow , corresponds to the actual movement of the caved material when the draw point is open (or when the equipment loads the material). According
Figure N°1 – Flow Behavior 168
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to the results of the research team1 and other publications the propagation of the flow in granular materials would happen through subsequent failure of "pressure arcs" formed by caved material. These process generates a movement that is outspread up to the surface as a descent. The propagation is of such nature that the zone near to the isolated draw zone in the low part of the column are affected generating a lateral movement towards the draw zone when distance between the draw points are less than 1,5 times its diameter. b) Extraction Point Closed Continuous media behavior : This corresponds to a physical behavior similar to a continuous media (fluid). When the draw point is closed (open draw) the density over the draw point is much lower than in the immediate surroundings, thus a slow equalization of densities occurs, where sectors of grater density tend to move towards sectors of lower density. This generates a slow arrangement of the material towards "litho static equilibrium" with vertical and lateral movement towards sectors of lower density (zones of open draw movement). In the left the movement is preferentially vertical over the draw point when it is open, also lateral movement happens when draw points are located less than 1,5 times the Dta away. Then with closed draw point (continuous behavior) lateral movement happens due to density equalization. The limit between an isolated and interactive behavior exists for open flow. When draw points are located at a distance of 1,5 times Dta or less the "threshold" of interaction is achieved during open flow, generating a superposition of the two behaviors. These is called "isolated – interactive draw". The graphic representation of "isolated" and "isolated – Interactive" draw have been taken from results from the CIMM physical model tests. The following Figures N°2 and N°3 show sections of the sand model with isolated flow and Isolate – Interactive behavior.
Figure N°3 Isolated – interactive Behavior , Layout "Tipo Teniente" Where : vta is the isolated flow rate vti is the interactive flow rate The difference between the "isolated – interactive" flow and "interactive" flow is that in the latter the material of the upper part of the caving descends uniformly, thus in the isolated-interactive, there is difference between the flow rate over the draw point and over the pillar (nti and nta) forming small craters in the surface of the cave. For distances between draw points of less than 1,2 Dta, the rates over the draw point and over the pillar are equal below the cave surface. The differentiation and management of these two models of behavior is very relevant due to the effect over dilution control and the recovery of a panel cave, normally designed to operate in the limit of interaction (D = 1,5 Dta). Each behavior can be defined as a function of the flow rates Vti (interactive flow rate) and Vta (isolated flow rate): Model of Behavior a) Isolated Flow b) Isolated - Interactive c) Interactive Flow
Relation of Flow Rate (Velocity) vti=0 vta > Vti > 0 vta = Vti
The transition from one model of behavior to the other for a given design and cave material will depend basically upon the parameters that define the Extraction (E). According to results of a research with sand models, stress simulations of the beginning of the gravity flow movement and extensive mine back analysis results gravity flow on a caved ground behaves according to the idealization of the following figure N°4, defined as "Interactive – isolated flow :
Figure N°2 - Isolated Draw Behavior Massmin 2004
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Figure N°4 – Interactive – Isolated Flow Behavior With the above geometry the following Gi index is proposed to characterize the degree of interaction that represents the different behaviors of Fm :
In order to have interaction between the draw points the proposed 1.5 *Dta was confirmed right through results of several sand model tests and conceptual analysis3. One draw point, can have only one of the three behaviors at one instant in time, thus the function Fm (Mass flow Model) should have the following time series:
Figure N°5 – Interaction Degree
Then with Gi definition : Gi = 0 : with isolated flow 0 < Gi < 1 : with isolat–interactive flow Gi = 1 : with completely interactive flow This implies that the "interactive theory" proposed by DH Laubscher is correct when Gi=1, that should be attained with mine designs where Dpe is less than 1.2 * Dta, and there is complete uniformity of draw.
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Figure N°6 – Two draw points Mass Flow Behavior as a function of Draw
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Figure N°7 – Flow Behavior Representation of Fm The change from one kind of behavior to the other will depend on the extraction practice (E), that will mainly vary upon the uniformity with which the different draw points are extracted. Following Figure N°7 shows the different behaviors that conform the proposed gravity flow model: With the proposed behavior the prediction of the dilution entry point, and the way of mixing of the column is much more complex than the model proposed by D.Laubscher, even if total uniformity of draw is obtained, that for primary rock caving´s, would be rare. The best approximation to the effective gravity flow of an LHD operation would be an "isolated – interactive" behavior. It is important to bear in mind that even if there is interaction the flow rate of the area outside the isolated draw column is significantly slower than the isolated draw velocity. According to the previous model the dilution entry as a function of extraction percentage should look as the graph of Figure N°8: 3. CONCLUSIONS The proposed Fm model integrates part of the previous gravity flow concepts (Laubscher) broadening its interpretation. The model is based in the results of physical Massmin 2004
tests with sand and observations of caving operations. The validation of the presented model has been done with empirical results as presented in following paper. The intensive experimental stage with a specially designed sand model is an invaluable tool to observe mass flow behavior. Efforts to keep with these experiments should not be abandoned, specially to measure stress dynamic behavior. 4. AKNOWLEDGMENTS The author acknowledges the important contribution to the presented results, of the CIMM Gravity Flow Research Program and the FONDEF 1037 Project, and that of all the professionals that participated in the experimental stage, back analysis of the Codelco data and conceptual analysis developed in these Projects. Both projects were managed by Hugo Diaz, to a successful end. Special mention to Dennis Laubscher for setting the basement, without his acid critic and our long discussions that led to permanent review of his interpretation of gravity flow and caving design, I wouldn’t have been compelled to give an additional step.
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Figure N°8 – Assessed dilution entry as per Fm Proposed Model 5. REFERENCES • Diaz, H., Susaeta,A, (2000), "Modelamiento del Flujo Gravitacional", Revista Minerales, in Spanish. • Kvapil, R. (1965): "Gravity flow of granular materials in hoppers and bins," Int. J. Rock Mechanics and Mining Science, Parts I and II. Vol. 2. • Kvapil, R. (1982): "The mechanics and design of sublevel caving system," Underground Mining Methods Handbook, W. Hustrulid ed., SME, New York. • Laubscher, D. (1994): "Cave mining-the state of the art," The Journal of The South African Institute of Mining and Metallurgy. • Mansson, A. (1995): "Development of body of motion under controlled gravity flow of bulk solids," Licentiate thesis, Lulea University of Technology, Sweden.
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• Susaeta, et all , "Modelamiento del Flujo Gravitacional", Informe Final, CIMM, 1999. Internal Document in Spanish. • Susaeta, A., Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 2. Internal Document in spanish. • Tamburino, A , Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 4. Internal Document in Spanish. • Verdugo, R., Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 1. Internal Document in Spanish.
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Theory of gravity flow (Part 2) Andrés Susaeta, Mining Consultant, Professor Mining Department, University of Chile, Chile
Abstract The objective of this paper is to introduce some control parameters for Fm (Model of Gravity Flow), and using them prove, through empirical relations and findings from back analysis from Codelco mines, the validity of the proposed model of gravity flow. The uniformity Index is introduced and results of back analysis are presented to show relation between draw and interaction, as well as a methodology to asses and evaluate interactive layout designs. Some relevant results of the research team are applied to propose a methodology to assess Dta (Isolated draw diameter) and evaluate layout interaction in the design stage.
1. INTRODUCTION The proposed model for gravity flow depends on three principal variables. One of them is related with extraction, that can be characterized with the Uniformity Index, that is a practical tool to plan and evaluate the effective uniformity of draw. The relation of this variable with the proposed gravity flow model is very important for design and operation of caving layouts, specially in primary ore, where maintaining uniform draw is not easy. The layout design and its relation to the cave material properties is one of the other control variables of the proposed model and a very relevant issue for a successful caving operation. A methodology to determine the isolated draw diameter for different caving materials at different extraction percentages is proposed, as well as a criteria to ensure LHD layout interaction for all the ore column. 2. UNIFORMITY INDEX The measure of the uniformity of draw is a key issue for draw control practice. It allows both control and proper planning of the call order, as well as very easy graphical representation of the entire mine draw practice. The proposed index is (as per Susaeta, A, Saavedra, J. unpublished):
I .U . = ∆ + Γ ⋅
(t
p
− t min )
2 ⋅n tmax
⋅ ∑ ( t max − ti )
(1)
where: • ∆: Number of inactive draw points in the draw point vicinity. • r : Factor of normalization, equal to 99/89. • tp: Tonnage extracted from draw point p under analysis, in a specific period of time. • ti: Tonnage extracted from draw point i belonging to draw point p vicinity in the same period of time. • tmax: Maximum tonnage extracted in the vicinity of draw point p, in the same period of time. • tmin: Minimum tonnage extracted in the vicinity of draw point p, in the same period of time. • n: Number of draw points belonging to the vicinity of draw point p.
Analysis of the equation As it can be seen the equation is constituted by two parts: • A first integer part (∆), that indicates the number of inactive draw points in the vicinity, in other words draw points without movement during the considered period of time. Then as ∆ grows draw is less uniform. • The second part of equation (1) will be called Specific index of uniformity (I.E.U.). With the normalization factor r, the I.E.U. is a decimal number between 0 and 1, and indicates the uniformity of draw among the active draw points for the selected period of time. So with the above definition: Draw characteristic Completely uniform draw Completely isolated draw
Specific index of uniformity 0 1
For example, a draw point with a configuration of 6 neighbors and a uniformity index of 1,009, indicates that for the draw event analyzed, only one of its neighbors was inactive and the rest of the draw points were extracted in a relatively good uniformity. A key issue regarding uniformity is to define which is the maximum period of time that can be assumed, to consider that there is still "movement" in the draw columns so as to generate interaction between them. A period of three shifts (one day) was determined as the farthest time where there is reasonable correlation between the U.I and interaction for back analysis studies. This result is on line with practical experience and stress measurements of the mayor apex. INTERPRETATION OF THE UNIFORMITY INDEX Due to the vectorial nature of the index where two information’s are integrated in one number it is not possible to categorize its results linearly. The following matrix (Figure N°1) is proposed to characterize the results, that defines every draw event either as uniform, semi uniform or isolated.
By definition the vicinity of draw point p includes that draw point. Then as an example for a layout of a draw point with 6 neighbors cardinal n of the vicinity is 7. Massmin 2004
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Uniformity Index Characterization
Teniente Mine, where vertical up holes were drilled through the mayor apex into the caved material after it was closed. The graph (Figure N°3) where the U.I, represented as percentage of time in isolation against Gi (determined from lithological mapping of the ore over the pillars) shows the relation between these two variables. These results were predictable from the Fm proposed model. They confirm that draw must be uniform ideally 100% of the time. With this the maximum potential of interaction degree (Gi) for the design can be attained. The Gi for any LHD layout, as will be showed will be less than 1, thus the potential of leaving part of the ore reserves over the mayor apex is certain.
Figure N°1 – Uniformity Index example matrix This matrix can be modified to adjust different design patterns and uniformity criteria’s. For programming purposes special matrix have to be defined for the cave boundary, where the total possible number of draw points is less than within the layout. Each call event (one shift) will have an associated category of the U.I. Its graphic representation (Figure N°2) for a control period (whole life of mine, one month, etc.) can be expresed for every draw point in percentage of tonage, or time, within a specific category. The following figure shows an example of the representation of a sector of a case study, where the scale shows percentage of uniformity. The red then is isolated draw. Figure N°3 – Interaction Degree versus Uniformity Index The following table shows the results of the Dilution Entry Point for several representative sectors. They confirm that there is a direct dependency between U.I and Dilution entry point, for points that have had over 70% extraction with uniformity. The higher the percentage of time (or tonnage) extracted with uniformity, the higher the dilution entry point.
Sector
Uniformity Index
% Dilution Entry
%t = % t = Unif. Isolation + semiuni
Figure N°2 - Uniformity Index Graphic Representation in Plan 3. BACK ANALYSIS RESULTS Using the described parameters, and the Fm concepts, the information of all the Codelco Mines was analyzed to determine relations between draw practice and interaction. - Degree of Interaction (Gi) A relation between the Gi (Gi = vti / vta) and the U.I (Uniformity Index) was derived from a sector of El 174
Inca Norte
34
66
57
Inca Central
57
43
41
Quebrada Teniente
19
81
62
Teniente 4
34
66
54
The following graph (Figure N°4) summarizes the behavior of all LHD draw points of Andina III Panel. This graph looks like the proposed behavior of Isolated – interactive flow. The Pedza (Dilution entry point of isolated flow) was determined through a very accurate tracer of dilution (construction material of previous level) and the Pedzi (Dilution entry point from interactive flow) was determined from the change of curve slope. The real dilution percentage is not accurate because not all dilution is ryolite. The improve in the draw practice during the last year has improved the Pedza several points.
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Figure N°4 – Percentage dilution versus Extraction for LHD Area A
Figure N°5 – Relation of Dilution entry (isolated flow) versus draw practice for grizzly 9 x 9 m layout (secondary ore) The analysis of the Pedza (Dilution entry point of isolated flow) for Sector A (Andina III Panel) versus percentage of uniform and semi uniform extracted tonnage (Figure N° 5) show a clear tendency (considering side dilution and other practical effects) that confirms the previous results. The generation of a data base that relates a layout with fragmentation (f) and dilution entry (for isolated and Massmin 2004
interactive flow) is not and easy task because it is very difficult to encounter representative sectors with consistent draw uniformity. The above plus several observed results of the analyzed caving operations confirm the proposed Fm model at a mining scale.
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4. DESIGN PARAMETERS
DESIGN INTERACTION POTENTIAL
CAVE MATERIAL AND ISOLATED DRAW DIAMETER ESTIMATION (DTA) The proposed function Fm (Model of Gravity Flow) is controlled by the variables: Extraction, Material properties and Layout geometry. The material properties of the caved material (Pm) depend of moisture and the internal friction angle (ø). Through conceptual analysis the following expression was derived for ø :
Considering that the limit for interaction between two neighbor draw points at a distance Dpe is :
ø = 100 Cu / (1 + 2.332 Cu)
Dpe < or = 1.5 * Dta
(4)
Then if the following "ideal draw point spacing layout" is considered (equilateral arrangement), the critical point for interaction can be expressed as a relation between areas:
(1)
and
Cu = D60 / D10
(material gradation or uniformity coefficient)
(2)
Where D10, D60 and D80 correspond to fragmentation size (m) passing accumulated 10, 60 and 80 % respectively. As defined through similarity analysis the internal friction angle of the material (f) determines the Dta, so these is the basic parameter to be considered during the design stage. For design purposes the following equations was derived to assess the isolated draw diameter (Dta ): (3) Using standard size distribution for coarse, medium and fine material for different percentages of extraction the following table and curve (Figure N°6) was computed with the above equations:
Figure N° 6 – Isolated draw diameter for different material types as a function of extraction percentage
For design purposes, the layout has to be interactive at an extraction percentage of at least 70% of the ore in the column to obtain a minimum recovery. Ideally interaction should be permanent, that means that the Dta for design is near to the one calculated for a fragmentation of 90% extraction. No detail evaluation of moisture effect was introduced to the above calculations.
Dta as a function of material type (fragmentation) and percentage extraction % Extraction
D60 (m)
D10 (m)
D80 (m)
Cu
ø (°)
Dta = 5*D80 * tg (ø)-7 (m)
Coarse Material
-
-
-
10.00
1.44
0.22
2.15
6.69
40.50
32.53
50.00
0.90
0.15
1.30
6.00
40.21
21.04
90.00
0.70
0.10
1.00
7.00
40.60
14.70
10.00
0.67
0.10
1.00
6.69
40.50
15.10
50.00
0.46
0.08
0.79
5.50
39.97
13.64
90.00
0.43
0.05
0.67
9.28
41.19
8.51
10
0,19
0,05
0,45
4,22
39,11
9,58
50
0,17
0,04
0,35
4,25
39,13
7,41
90
0,15
0,04
0,30
4,29
39,16
6,30
-
Medium Material
Fine Material
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The experimental results for a "Teniente Layout" shows interaction over the mayor apex for distances of Dpe = 2,5 * Dta, as per Figure N°8. The only way to understand this is to assume that the two draw columns of the trench act "as one wider draw point". Interaction over the mayor apex will exist only if all the following conditions are met: a) Dpz (Distance between trench draw points) is less or equal to 1,5 Dta => There exists a Interactive Trench Diameter (Dzi). The interactive trench diameter (DZI) is:
Dzi = Dpz + Dta
Figure N°7 – equilateral layout arrengement
According to following sketch (Figure N°9): The value λ is defined as the distance between the drawpoints and the Dta (isolated draw width). With the interaction index (Ii) defined as active area (isolated draw area) divided by the total area of influence:
Ii =
Aa At
c) (Dpm + Dpz) is less or equal to 1,5 * Dzi
5. CONCLUSIONS
With l = 1.5 (critical limit of interaction) the Ii critic can be calculated for the equilateral layout:
Then the critical interaction limit, expressed as isolated draw width to total area is (Ii critic) 0,403, or at least 40% of the total area to be extracted needs to be under isolated flow behavior in order to have total interaction. Different geometries with the same critical limit (l = 1.5) will have different Ii, as expected. Defining a "Dispersion Index (Id)" for a design as the interaction index divided by the critical interaction index, different layouts can be compared.
id =
b) Operational uniform draw is strictly observed (Interactive draw)
The proposed gravity flow model can be presented as:
Where UI is the uniformity index, Dta the isolated draw diameter, and Id the dispersion index. Back analysis results show that in effective caving operations, Fm varies its
li licritico
The interaction phenomena can be expressed in words, considering the results of models and Fm conceptualization and the critical interaction described above as follows: The isolated draw cylinders form an "open area" (Aa).The caved material surrounding them behaves as a solid due to cohesion, forming "pillars" of these material (Ap – Pillar Area). These pillars will have the highest load at a height where the isolated draw cylinders reach their maximum width (HDta = 2 * Dta). If the stress of the static material, plus the friction with the isolated draw column is less than the ø of the caved material there will be isolated draw only (Gi = 0). If the stress is grater than the internal friction angle of the material the "pillar" will collapse at that location (Interection height = 2 * Dta) contributing material towards the isolated draw cylinder.
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Figure N°8 – Experimental result for a "Teniente" Layout with interactive Draw
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1037 Project, and of all the professionals that participated in the experimental stage, back analysis of the Codelco data and conceptual analysis. Both projects were managed by Hugo Diaz, to a successful end. Special mention to Dennis Laubscher for setting the basement, without his acid critic and discussions that led to permanent review of his interpretation of gravity flow and caving design, I wouldn’t have been compelled to give an additional step.
7. REFERENCES Figure N°9 - Sketch behavior through time in three modes: : "isolated", "interactive-isolated" and "interactive", determined by the Gi (degree of interaction). The proposed control variables of the model have been applied successfully, in planning and draw control optimization in various mines in operation. The Uniformity Index has proven a very handy tool for practical draw control and back analysis of historical draw practice. According to the proposed model, and results of stress monitoring in the sand model, the interaction potential of a layout and its relation to draw practice can be effectively observed from the stress changes over the mayor apex (central pillar). This methodology to evaluate interaction potential of different layouts should be complemented with a devised experiment to measure effective f of the caved material over the mayor apex. Further on, it is recommended that the back analysis effort done by Codelco mines is extended to a broader number of operations to integrate additional panel caving experience into the model. Moisture as a control variable of f remains to be investigated in detail, because it has a very important control over Fm that has not been evaluated in depth.
6. AKNOWLEDGMENTS
• Diaz, H., Susaeta,A, (2000), "Modelamiento del Flujo Gravitacional", Revista Minerales, in Spanish. • Kvapil, R. (1965): "Gravity flow of granular materials in hoppers and bins," Int. J. Rock Mechanics and Mining Science, Parts I and II. Vol. 2. • Kvapil, R. (1982): "The mechanics and design of sublevel caving system," Underground Mining Methods Handbook, W. Hustrulid ed., SME, New York. • Laubscher, D. (1994): "Cave mining-the state of the art," The Journal of The South African Institute of Mining and Metallurgy. • Mansson, A. (1995): "Development of body of motion under controlled gravity flow of bulk solids," Licentiate thesis, Lulea University of Technology, Sweden. • Susaeta, et all , "Modelamiento del Flujo Gravitacional", Informe Final, CIMM, 1999. Internal Document. • Susaeta, A., Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 2. Internal Document. • Tamburino, A , Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 4. Internal Document. • Verdugo, R., Proyecto Fondef 1037, "Desarrollo de un Modelo de Flujo Gravitacional para Minería por Hundimiento de Bloques", Mayo 2002, Sub Proyecto 1. Internal Document.
The author acknowledges the important contribution to the results of the CIMM research program and the FONDEF
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Predictive models for gravitational flow Marco Alfaro, Facultad de Ingeniería, Universidad de los Andes and Metálica Consultores S.A. José Saavedra, Facultad de Ingeniería, Universidad de los Andes
Abstract In underground operations the planner is faced with a very complex decision problem. The complex nature of this decision depends on the complex behavior of ore extraction from drawpoints. Up to now it have been impossible to have good models to predict the grade of ore extracted from the drawpoints. Some attempts have been made up to now to model this phenomena. In this work we present a Cellular Automata based model for Gravitational Flow. The model proposed here exploits the simplicity of local rules of evolution to acquire the objective. The model was implemented and tested with satisfactory results at El Salvador Mine. The proposed model is analyzed from a mathematical point of view in order to characterize its behavior. Some extensions to this basic model are presented. Some observed phenomena is explained with the aid of this model, for example interaction between drawpoints. The mathematical analysis of the model show that the implicit stochastic process involved is diffusive and in some way equivalent to Brownian Motion phenomena in Two-Dimensional space. Finally, the way this model behaves has motivated us to call him the bubbling process.
I. INTRODUCTION
and fast. In Martínez’s model the construction tries to model avalanche phenomenon.
In this paper we present a Cellular Automata which models Gravitational Flow in Underground Mining. The proposed model exhibits very simple rules of evolutions. This simplicity allow us to get simple implementations. The model was initially constructed for two-dimensional case, the extension to three dimensional situations is straightforward. This paper focuses in the mathematical aspects of the model. An important stochastic process, called by us bubbling process is derived from the model. The behavior of this process is equivalent to random walk type models.
A. Goles’s Model This model was presented by Eric Goles and Sebastián Peña [2] in the context of the research developed in CIMM (Mining and Metallurgic Research Center) in 1996. This model presents a first approximation to this kind of models, taking into account physical characteristics of material, for example density and grade of each cell. The model is defined as a Cellular Automaton, i.e., is defined on a two dimensional space, which is subdivided into cells. The cells are defined as entities which certain grade and certain density:
II. CELLULAR AUTOMATA THEORY To give a light introduction to Cellular Automata theory let’s give first a definition adapted from [1]:
npm (1) grade : =
• 100 density = npe + npm npe + npm
Definition II.1 A Cellular Automaton is an array of cells each colored either black or white. At every step there is the a definite rule that determines the color of a given cell from the color of that cell and its neighbors on the step before. The main characteristics of a Cellular Automaton are: 1. His state, which is variable for each cell. 2. His neighborhood, the set of cells which interacts with the cell. 3. The set of rules or program: gives us the changes in state with respect the neighbors.
Where: • npe is the number of sterile particles which are contained in the cell. • npm is the number of ore particles which are contained in the cell. Two kinds of neighborhoods are defined for cell (k,m), depending on transition rules:
Usually the set of cells is IN2 or some subset of it, it’s possible to use other kinds of reticules (Hexagonal for example) to represent cells. III. SOME PRELIMINAIRES MODELS In the last decade, some efforts have been made. The main results were obtained by chilean well known scientists, Eric Goles [2] and Servet Martínez [4]. In Goles’s model the main idea is the existence of two distinct dynamics: slow Massmin 2004
Transition rules are divided into two cases. When (D and , are model parameters) the cell is considered as "not fill
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enough", in this case a fast evolution dynamic is used. The neighborhood used in this situation is the partial one, the cell gets contributions from the three superior cells. Material contribution (sterile or ore), is determined by a parameter called upper neighbor probability (Pv2). The total contribution is then calculated as:
(2) (D - dij ) * Pv2 = Upper Neighbor Contribution (D - dij ) * (1 – Pv2) (3)
= Another Neighbors Contribution 2
Otherwise, i.e., D-dij, < Ø the cell is considered as "almost complete filled", in this case a slow evolution dynamic is used. The neighborhood used is the complete one and the contribution of the cells is distributed proportionally between the five neighbor cells:
dij
(5)
∑
5
d i=1 veci
IV. PROPOSED MODEL This Cellular Automata uses as cells a reticulate of squares (could be rectangles too), in which cells has two possible states: full or empty. At the beginning all cells are full. Then we extract a block (represented by a cell). The generated void must be filled by another block, we see that the upper blocks are in a privileged position to accomplish this objective. Inspired by this idea, we define the transition rule for the evolution of the Cellular Automata. We assign a probability distribution for the neighbors (the upper cells), and then we select according this distribution the block which replaces the void. The void now is in the position of the selected block. This process is repeated until the void reaches the surface (the last level of blocks). In the two dimensional case we can see trough the next figure how look one block with his neighborhood:
= Neighbor Cells CONTRIBUTION
In both cases material contribution comprehend sterile or ore, the contribution in each case is proportional to his content (npe y npm). B. Martínez’s Model This model was proposed by Servet Martínez [4] in the context of the research developed in CIMM (Mining and Metallurgic Research Center) in 1996. Is based on the stability of the particles and how they flow when this stability is loose.
FIG. 2: Block Neighborhood. Two Dimensional Case. After this first void reaches surface, we generate a new void extracting a new block from the same position used before (the positions here represents drawpoints). The ascending void generated by the extraction of a block is called by us as Bubble. The next figure show us a two dimensional example of how one bubble evolve:
This Cellular Automaton considers and array that guarantees strong stability for each cell. This model is based in the following principles: • Three blocks are extracted from the base which originates a void. • In front of any void space, the particles flow fall in vertical way. • In other cases there is an instable equilibrium, fundamental characteristic of the model, in these cases the particles fall according to two priorities. First, the more instable particles fall. Second, if we have particles with equal instability, the one which has the older instability fall, here is assumed that instability is increasing in time. As an example let’s look at one iteration of this model: FIG. 3: Bubble Evolution. V. THE BUBBLING PROCESS The proposed Cellular Automata will be analyzed in two dimensions. This approach simplifies calculations, the extension to three dimensions is straightforward.
FIG. 1 Instability, makes that cells 7, 8 y 9 fall, in that order, because cell 7 has an older instability. Then the same will occur with cells 10,11, 12 y 13, which originates a surface crater.
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A. Some Useful Calculations Given a block model of certain dimensions, it’s not hard to see that the probability of a given bubble to reach position of block (i, j) is: (6)
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Where is a function defined as:
particular bubble, and this probability decreases as we move away from the central column. This result is summarized in the next proposition.
(7)
This calculation is trivial using Total Probabilities Rule. We will call Bubble Probability to the quantity Pr (n (i,j) = 1). If we consider as transition probabilities for the neighbors (left, center, right) then we have:
Proposition VI.1 Denoting central column blocks as Bck (here k denotes height with respect to drawpoint) and denoting bubble probability for this column as then we have the following valid relation when : (9)
lim p
c ,k
=0
k →∞
(8)
It’s very important to characterize the final cavity generated by extraction of ore. We will see that empirical tests require an additional definition in order to study this feature of the model. As an example consider that transition probabilities are . If we consider the blocks that obtain a positive probability of being visited by a bubble then we have as a resultant cavity the one given by Fig. 4 (a). If we consider only the blocks that have bubble probability greater than 0.001 we have situation described if Fig. 4 (b). Finally cases 4 (c) and 4 (d) are for bubble probabilities greater than 0.05 and 0.07 respectively.
This relation guarantees that in absence of height restrictions the final cavity will have ellipsoidal form. This asymptotic behavior is not found in real cases, in all real cases always there is a limitation on sector height. In these cases the final form of the cavity have parabolic form. This kind of behavior can be explained with the aid of a new concept introduced in the next definition: Definition VI.1 Given a sector of height H, which translated to blocks means that sector has height (in blocks). We define Cut Probability, which is denoted by , as the quantity which satisfies:
(10)
Where is understood as the closest value near is interpreted as the value such if we consider only those blocks with bubble probability greater or equal than this number then in the last level of the given sector (at height ) the sum of probabilities is near (Usually takes value 0.95). Mathematical interpretation of this quantity is clear in the sense that blocks with bubble probability greater or equal has an 95% of probabilities to be removed by extractive process. With this tool we can deduce that the final cavity has paraboloidal form. This theoretical fact has very large support from empirical experience. All simulations made using this Cellular Automata gives final cavities whit paraboloidal form. Almost most important, we can deduce from definition that cut probability is a height-dependent quantity. The distribution of bubble probabilities is more likely uniform distribution when he take values greater. VII. CONTINUOUS APPROACH TO SIGNIFICANCE LEVEL FIG. 4: Final Cavities considering distinct bubble probabilities : (a) > 0, (b) > 0:001,(c) > 0:05 , (d) > 0:07 (Labeled from top left to down right)
As we can see, different probabilities give different final cavities. The main question is then: Which probability we need to consider to obtain a good representation of the final cavity? The answer to this question is given in the next section. VI. DEFINING SIGNIFICANCE LEVEL It’s not hard to see that equation 8 give us a probability distribution that in each level approximates a Gaussian one, i.e., the central block (the one in the same column as the drawpoint) has a greater probability of being visited by a Massmin 2004
Our model can be seen as a random walk in one dimension (in the two dimensional case) and a random walk in two dimensions (in the three dimensional case). The way to understand this fundamental fact is to see that if we put together in one level all levels of a given sector then the system evolves in the same way as a random walk process would do. To fix ideas we will analyze our model in the context of two dimensional sectors. The general form of this random walk process could be described mathematically denoting by X(t) to the random variable that represents the walker position in time t. Consider too a one dimensional reticulate (for example Z) and discretization of time T . The rules of evolution of one dimensional random walk are given by:
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(11) X (τ) =
{
X (τ) – 1 X (τ) X (τ) + 1
con probabilidad q con probabilidad p con probabilidad q
A. Real Space Derivation Following B. Hughes [6], let us consider that the sites in our reticulate will be indexed, as usual, with integer coordinates. We denote by the probability that the walker will be in site l after n steps. If random walk steps are i.i.d (independents and identically distributed), with representing the probability that walker actually in position will reach position l in the next step, we have the following fundamental relation: l =∞
(12)
Pn +1(l ) = ∑ p ( l − l ′ ) Pn(l ′)
FIG. 5: function with fixed .
l =∞
this relation comes from Total Probability Theorem. Following calculations and taking limits [5] we finally get a Partial Differential Equation (PDE) for the function p(x; t) (probability density starting from position x in time t):
When we resolve the equation
(17)
(13)
∂ ∂ ∂2 p ( x, t ) = v p ( x, t ) + D 2 p ( x, t ) ∂t ∂x ∂x
Equation 13 is the well-known Diffusion Equation in presence of drag velocity v. D is called diffusion constant. B. Diffusion Equation Solution, Application The solution of equation (13) is very well known in the case with border conditions the solution is:
with as a funtion of t we get that
i.e. the final cavities are of paraboloidal form. The extension to three dimensional case is straightforward.
VIII. APPLICATION TO A REAL CASE: CODELCO- EL SALVADOR MINE First we can see in Fig. 6 how the simulation tool looks.
(14)
which corresponds to Gaussian density functions with average position vt and variance 2Dt in that position. In our case we are in absence of drag velocity, so in this particular case solution of equation (13) has the closed form:
(15)
In order to find significance level in this approach we need to resolve the following equation:
a
(16)
r (a) =
∫ p ( x, t )dx = 0.95
−a
FIG. 6: Simulation Tool
given some value of t. When we resolve this equation in some particular case, for example when t = 200 and D = 1 we find that . The function for this value of is plotted in the next figure: 182
In Fig. 7 and Fig. 8 we present examples obtained from simulations from a real sector of Salvador Mine. This simulations were made in Drawpoint 918N.
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which we call Potential Energy. For each block we can calculate this energy by using the usual formula , where m represents block mass, g is gravity acceleration constant and h is height with respect to sector floor [5]. In second place we have another kind of energy which we call radial Energy, this energy is related to velocity field (which uniforms at greater heights). This energy can be measured as where m is the block mass and v is block movement velocity (which is height dependent) [5]. For each block we define a function that assign energy:
(19) E ( Bijk ) : = Ep ( Bijk ) + Er ( Bijk ) FIG. 7: Accumulated Variation Drawpoint 918N we note that this expression is simple and we only need recalculate if the block is moved [5]. The rule of transition for our Cellular Automaton consists then in make a Stochastic Drawing based in energy values of the neighbors blocks [5]. As the interactions between granular materials is dissipative in nature, we can discount to the energy value of a block certain quantity each time the block is moved, one possible function that incorporate this factor could be: (20) E ( Bijk ) : = Ep ( Bijk ) + Er ( Bijk ) – lε
where ε represents a little energy diminution and l the number of times the block has been moved [5].
FIG. 8: Grade Evolution from Drawpoint 918N
Finally in Fig. 9 we present an Error Map generated with the aid of the tool.
B. Fractional Simulation One limitation of the presented model is that needs a complete extraction of blocks from drawpoints. This limitation is a very big one because of the actual dimensions of blocks [5]. One possible approach is to discretize one given sector in more units. This led the problem of have a model that needs very much memory to run [5]. Another option is to allow the model to extract fractions of blocks. The only modification that this concept imposes on the original model is that we have to maintain for each drawpoint a state number that tells how much ore remain in the last extracted block. If the actual extraction needs more ore that the remaining one the we have to generate another bubble and refresh the state of the drwapoint [5]. X. CONCLUSIONS As we can see, we have a very powerful tool for Gravitational Flow Simulation. His theoretical characteristics and flexibility make them a very good tool that can aid to understand mine operations. This model could serve not only as a prediction tool, we can use it in planning labors or to decide when to close some drawpoint for example. Preliminary computational experiment show us that this model could be better implemented. In that case we hope that simulations could take fewer completion time. This objective really makes sense when one thinks that a model of this nature could be runned almost everyday.
FIG. 9: Grade Evolution from Drawpoint 918N
IX. SOME POSSIBLE EXTENSIONS A. Energetic Model The main idea behind this extension is to modify Cellular Automata rules of evolution. For this purpose we first need to assign to each cell of the model an energy value [5]. First, we can consider each block having one kind of energy
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XI. ACKNOWLEDGMENTS J.S. wish to acknowledge the support of FAI (Fondo de Ayuda a la Investigación), Proyecto Ingeniería 2004, Universidad de Los Andes.
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REFERENCES
[4]
[1] Stephen Wolfram: A New Kind of Science. Wolfram Media, Inc. 2002. [2] Eric Goles y Sebastián Peña: Modelamiento y Simulación del Flujo Gravitacional. Informe Final PB510004, Apéndice N°2, 1996. [3] Gregorio González: Estudio del Comportamiento de un Material Granular Mediante Modelos Computacionales. Memoria para optar al título de Ingeniero Civil Matemático, 1999.
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Servet Martínez: Consideraciones Acerca del Modelamiento de Flujo Gravitacional. Informe Final PB5- 10004, Apéndice N°3, 1996. [5] José Saavedra: Secuencias, Flujo Gravitacional y Evolución en Planificación Minera. Memoria para optar al título de Ingeniero Civil Matemático, Universidad de Chile, 2002. [6] Barry D. Hughes: Random Walks and Random Environments, Volume 1: Random Walks, Clarendon Press, Oxford, 1995.
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Computational model for simulation and Visualization of gravitational flow Carlos Calderón, Facultad de Ciencias Físicas y Matemáticas, Departamento de Ciencias de la Computación, Universidad de Chile Marco Alfaro, Facultad de Ingeniería, Universidad de los Andes and Metalica Consultores S.A José Saavedra, Facultad de Ingeniería, Universidad de los Andes
Abstract Some models have appeared in he last time to modeling the Gravitational Flow from drawpoints. Having a theoretical model for a phenomena is only a part of the task. Another important task is the implementation and validation of the model by means of software techniques. In this paper we explore the considerations, techniques and complexity of algorithms needed for implementing one of such models. We present some results obtained with the implementation and show how this computational model can be integrated with other tools. The graphical aspect have received special attention, powerful libraries such as OPENGL are in the kernel of our implementation. Some examples of the utilization of this model in real type applications are provided. The main advantage of this approach is that simulations are executed in real-time. Conclusions and extensions are presented.
I. INTRODUCTION In underground mining operations there are many factors that influence decision making process. One of the most important factor is the inevitable stochastic behavior of Granular Flow from drawpoints. As underground operations aren’t selective (in contrast with open pit operations), the planner must be careful about quality of extracted ore from drawpoints. It’s fundamental to have models for gravitational flow. This is the first task. In the past some attempts have been made to modeling this phenomenon. Remarkable works are the references [2], [4], [3]. All of these works have in common the technique used: Cellular Automata. The most recent work is the one by Marco Alfaro and José Saavedra [1]. This last model was implemented and validated and some practical issues related to implementation were noted in the development process. In this paper we present the main characteristics of such a system and some visualizations of the system in action. The implemented software allows the analysis and visualization of gravitational flow simulations, this can be done trough a suitable discretization of ore material, this model is called block model. It has some mathematical properties which are reviewed in the work [5]. The main idea behind the algorithm is to generate a bubble (by means of extracting a block from the floor of a given sector), then with the aid of a probabilities matrix (rules of evolution of Cellular Automaton) we evolve this bubble until it reaches the surface [1]. The simplicity of the model and the efficiency of the implementation allow us to make real time simulations, which is extremely useful in terms of visualization and as a consequence a better understanding of the underlying physical phenomenon. II. IMPLEMENTATION METHODOLOGY C++ was the chosen language for implementing the model. The reasons for this decision are: • This language has many desirable characteristics that facilitate the development process, especially, C++ has native support of vector classes which are very useful Massmin 2004
when we work with block models. • Related with the last point is the native support of specialized algorithms that act on these kinds of data structures. • C++ is Object Oriented. This allow us an easy implementation and a desirable extensibility for future versions. This is mainly related to modularity characteristics of Object Oriented languages (Another OO languages are: Java, Object Pascal, etc). • Another important feature of C++ is the ability to work with pointers. This feature makes faster algorithms, because thanks to pointer techniques we don’t need copy and manage the whole object, instead of that we only give the memory direction and this feature can be used to simulate the spatial movement of blocks. C++ is a standard programming language. One of the main features is that OPENGL could be used in conjunction with C++. As we need to develop a powerful graphics software this consideration is essential. III. DATA STRUCTURES Any implementation of an algorithm, is intimately related with data structures used by the algorithm, thus the description of that structure is very important to understand how the algorithm works. The restriction that our simulation must work in real time carry us to design a threedimensional matrix of pointers (or memory directions), where the real data of each block were resident in other distinct memory place without need of moving these data. The location of a block in certain spatial position is made up with the aid of pointers. In simple terms we can say that the block in position (i, j, k) is referenced by Pi,j,k. For example, if we want to describe the vertical movement of a certain block, i.e., move the block from position (i, j, k) to position (i, j, k–1), then we have to do the following pointer operation: Pi,j,k = Pi,j,k–1 Bubble movement is made by using permutation of two pointers, in particular the one who has the initial position of the bubble and the one ho has the final position.
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A void in space is represented by a special kind of pointer: the null pointer. When we make a permutation in this situation there isn’t lose in information. The transition probabilities matrix is a cube of size 3. The bubble is in the central position in the first level of the cube. This matrix is designed to manage several physical phenomena, for example relative movement or avalanche phenomenon, nevertheless to incorporate successfully these phenomena we need the physical information to fill the matrix. IV. TECHNICAL CHARACTERISTICS We call processing capacity to the quantity of material that can be simulated per second, i.e., the quantity of blocks that the simulator extracts per second. Using an example where we fill the block model at the top (i.e. the bubble must travel the whole height of the block model), the processing capacity is about 4.000 blocks per second when the sector has 240 blocks height. The memory space that uses one block is about 20 Bytes taking into account his pointer, this means that in a sector with dimensions 100 length by 100 width by 240 height needs 48 Mb in memory. If we double the number of blocks in each dimension then we need about 400 Mb of memory (note that we will be moving about 20 millions of blocks in each simulation). Some characteristics of the implementation in terms of flexibility are the following: 1. We can fill the block model at the top in the block model. 2. We can orientate the blocks according each possible pair of linear independent vectors. 3. We can locate the drawpoints everywhere in the base of the blocks model. 4. We can establish sample frequency according to user needs. 5. We can give specific seeds to make simulations, this is important when we want to make comparisons. V. BASIC ANALISYS OF ALGORITHMS This little analysis consist in determine the cost of bubbling function, this function carry a bubble from the drawpoint to surface. Bubbling function it’s a recursive function that executes while the bubble doesn’t reach surface, if we call c the cost of bubbling function without recursion. We can see that c depends on the cost of the instructions related with up-to-date bringing of probabilities matrix. To calculate the average cost that takes the algorithmin extract all the possible material from a given drawpoint, assuming that the final cavity has paraboloidal form [1], we sum the cost of extract every block inside this paraboloid. It’s not hard to see that the extraction of block has a cost , because is analogous to get a bubble into position (i, j ,k). If we have a paraboloid, then at height z we will have N(z) blocks to be extracted. From elementary calculus we know that N (z) = , where r2z is paraboloid radius at height z. Let C(z) the computational cost of extracting all blocks at height z then we have:
The total cost Ct is then
If we use asymptotic notation then the cost of extracting all the possible material from a given drawpoint would be O (h3). VI. CALIBRATION OF CELLULAR AUTOMATON MODEL As we saw in a past section we have a probability cube that keeps the transition probabilities. We need calibrate those probabilities if we want to use properly the model. One way to calibrate our model is trough a quantity called Isolated Extraction Diameter(IED), which is characteristic in each kind of material. This physical parameter is related with the paraboloid parameter p (of geometrical nature) in the following way: IEH = 2p EID Where IEH : Isolated Extraction Heigth and IED: Isolated Extraction Diameter. When we have the information needed to calibrate our probabilities, we use an algorithm designed for this purpose. The algorithm is as follow: Initialize-Prob-Matrix Find(IED) If Height(IED) > IEH then Give More Weight to the borders Else If Height(IED) < IEH then Give More Weight to the center Else Return Calibrated Matrix End If This procedure is justified by the fact that if Height(IED) > IEH then paraboloid is very closed so we need to open it. The another case is analogous. This procedure exhibits good convergence and the maximum error is bounded by 5%. In figures 1 and 2 we can see two snapshots of the software in action in the calibration procedure.
If our paraboloid have focus = (0; 0; p) and directrix equation z = -p then we have: FIG. 1: Results obtained from simulation to calibrate model when Height(IED) > IEH.
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Te choice of graphics support libraries in three dimensions was OPENGL, because of his efficiency and the ability to work in several platforms (Windows or Linux/Unix). In graphics terms, blocks grades can be differentiated by they colors, it’s possible to visualize a specific zone of the model that we want to analyze, in particular we can see cuts in the three canonic directions. In figures 4,5 and 6 we can see visualizations while Cellular Automaton is running.
FIG. 2: Results obtained from simulation to calibrate model when Height(IED) < IEH.
VII. MODEL VALIDATION It’s important that the implemented model behaves as theory predicts. In order to acquire this objective, one test was designed to show blocks movement. If one block is moved his color is darker than the ones with none or little movement. We can see that high probabilities zones are darker than low probabilities zones. A snapshot of this test could be appreciated in figure 3.
FIG. 4: Model Visualization 1.
FIG. 5: Model Visualization 2.
FIG. 3: Validation of the Model.
VIII. VISUALIZATION Cellular Automaton visualization if essential for studying his behavior, for this reason visualization windows were implemented, these windows can be accessed at any time. Massmin 2004
FIG. 6: Model Visualization 3.
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IX. NUMERICAL EXPERIENCES In order to test the model we carry out a numerical experiment. We dispose blocks in a box in such a way that 5 layers were formed. Each layer has his own characteristics constant grade. Simulation was made in a box of dimensions: • Width = 1.00 mts. • Length = 1.00 mts. • Height = 2.40 mts. The box was filled by blocks of 1.00 cm. by 1.00 cm. by 1.00 cm. And the layers were defined as: • Material 1 : 3167 [particles per million] • Material 2 : 1213 [particles per million] • Material 3 : 2690 [particles per million] • Material 4 : 7610 [particles per million] • Material 5 : 2570 [particles per million] If we divide the number of particles per million by 1000000 then we obtain the grade of the layer. Some drawpoint were placed in the base of the sector. As an example we will show the graph from results obtained in our simulations (see figures 7 and 8).
FIG. 7: Numerical Experience 1.
As we saw, the number of blocks needed to run the model grow a lot when we double the number of blocks in each dimension, to be more accurate we have to multiply the number of blocks by a constant factor of . Data structure is designed to be small in terms of size. If we want to consider a block model with many blocks we will need more RAM memory if we want to run simulations quickly. But as we saw, when we move about 20 millions blocks in one simulation the memory needs are about 400 Mb. This imposes a natural limitation in our implementation. As the times goes on, this memory limitation would disappear and then the model could be runned very quickly. The benefits of our implementation are clear: real time simulation with good visualization is a very powerful aid for mining industry. For the future, there is a clear line of development. This model could be extended to consider distinct block sizes. At the moment the software only allow the use of a dimensional homogeneous kind of block. If we can conduct research that allow us to implement this feature we will be able to have more realistic simulations. At the other hand, more efficient classes could be developed. The actual model could be refined and the algorithms could be optimized. As an example of this last asseveration we can for example modify bubbling algorithm. Now this algorithm allows one bubble to travel from floor to surface. We can modify this implementation allowing several bubbles to be generated (for the same drawpoint). This can be done because if one bubble is in level k then another bubble in level k-2 will don’t touch this bubble. The calibration algorithm could be useful when research about Isolated Extraction Diameter will be complete. At the moment there is only a little knowledge of this phenomenon and in the future we expect more work in this area. Finally our computational model could be useful for many reasons: • Planning tasks: If we have good simulations then we can reduce uncertainty about grades values. This tool is the one that gives good results with little time effort. • Visualization: The model can be stopped in every step. This allow us to see how blocks movements have developed. • Good execution times. • Real time interface. XII. ACKNOWLEDGMENTS J.S. wish to acknowledge the support of FAI (Fondo de Ayuda a la Investigación), Proyecto Ingeniería 2004, Universidad de Los Andes. REFERENCES
FIG. 8: Numerical Experience 2. X. CONCLUSIONS, DETECTED PROBLEMS AND EXTENSIONS The most biggest detected problem in our implementation is that needs very much resources if we want to carry out good simulations.
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[1] Marco Alfaro, José Saavedra: Predictive Models for Gravitational Flow. To be presented in MassMin 2004. [2] Eric Goles y Sebastián Peña: Modelamiento y Simulación del Flujo Gravitacional. Informe Final PB510004, Apéndice N°2, 1996. [3] Gregorio González: Estudio del Comportamiento de un Material Granular Mediante Modelos Computacionales. Memoria para optar al título de Ingeniero Civil Matemático, 1999. [4] Servet Martínez: Consideraciones Acerca del Modelamiento de Flujo Gravitacional. Informe Final PB5- 10004, Apéndice N°3, 1996. [5] José Saavedra: Secuencias, Flujo Gravitacional y Evolución en Planificación Minera. Memoria para optar al título de Ingeniero Civil Matemático, Universidad de Chile, 2002.
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Simulating gravity flow in sub-level caving with cellular automata Glenn Sharrock, Senior Geotechnical Engineer - AMC Consultants David Beck, Mining Engineer - Beck Mining Engineering Geoff Booth, Senior Mine Geologist - WMC - Leinster Nickel Operations Mike Sandy, Principal Geotechnical Engineer - AMC Consultants
Abstract Block caves and sub-level caving mines are now operating at greater depths and in stronger rocks than ever before. The rules of thumb and tools for design and layout of these mines were not developed for these environments and the resulting cave designs are often in conflict with the operational and geotechnical requirements for large caving operations in moderately to highly stressed environments. In this paper, a new particle flow code - CAVE-SIM (© Sharrock 2003) is described for assessing the economic impact of the changes to cave flow caused by different SLC layouts designed to ensure uninterrupted production. The modelling package enables full integration of economic parameters and geological models into a three dimensional flow model of an operating cave. In CAVE-SIM the user has full control over draw rates from individual drawpoints and there is full tracking of the instantaneous grade and particle size distribution through each drawpoint.
1 INTRODUCTION In recent times, a growing interest has emerged in the simulation of granular flows in block caves (BC) and sublevel caves (SLC). In particular, efforts are presently focused on the development of computer simulation techniques for predicting rock breakage or progressive growth of a caving front and the associated granular flows in the cave void. One promising technique is the recently-developed "Caving Simulator" known as CAVE-SIM. The key motivation for the development of CAVE-SIM is to properly simulate the effects that alternative extraction geometries and draw strategies would have on the economic performance of BC and SLC operations. The analytical engine of CAVE-SIM is based on the cellular automaton (CA) concept first developed by John Von Neumann in 1947. CA can efficently simulate granular flows in an interacting, three-dimensional system of drawpoints. However, before CAVE-SIM is described, a review of existing simulation methods is required.
2 EXISTING METHODS Prior to the development of CAVE-SIM, there were two simulation methods used to study granular flows in SLC, namely stochastic methods (SM) and discrete element methods (DEM). The main difficulty with these methods is the incapability to properly represent realistic geometry and problem size within computational constraints. In contrast, owing to the efficency of the CA concept, CAVE-SIM can currently be applied to problems involving in excess of 30 million particles.
Figure 1: 3D Mine Scale simulations of SLC in CAVE-SIM coupled with MAP3D: n = 30e6 Particles. Massmin 2004
2.1. Stochastic Methods Stochastic Methods assume that gravity flow is a stochastic process; that is, "a process that can be described by a random variable that depends on some stochastic parameter which may be discrete or continuous" (Borowski and Borwein 1991). The stochastic parameter in SLC problems is typically: "the probability of downwards propagation of a particle" or, "upwards propagation of voids". Examples include Nilsson 1988, Bergmark 1975, Heden 1976 and Power 2003. It is important to understand that the physics of granular flow are almost completely ignored in SM (Nedderman 1992). For example, fundamental physical laws such as
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Newton's Laws and cohesive-frictional behavior (Coulomb 1776, Rankine 1857) are not necessarily a hard constraint in SM simulations and particle contacts do not exist. This means that the effects of stress and frictional behavior are not considered. For example, lateral or sideways movement of particles in an SM model are often constrained to a fixed direction and governed by simple kinematic rules, when in reality or in a CA model the material behaviour and the forces acting on the particle can result in complex flow. Complex flow phenomena such as arching, slow-infilling of voids, inrushes and piping cannot be forecast or re-created in an SM model. For these reasons SM cannot offer the best possible solution to SLC type problems; indeed SM have been all but discarded in the field for which they were originally created - bulk solids handling. 2.2. Discrete Element Method The Discrete Element Method is by far the most promising and widely used simulation method in granular science. While SM sacrifices physics in favour of computational efficiency, the DEM provides a detailed treatment of the micro-mechanics of granular media. The DEM, as first proposed by Cundall and Strack (1979) involves computing the contact forces and resulting Newtonian dynamics of individual particles in an assembly. As a result, the distribution of shear and normal force, rotation, velocity and displacement are determined for each particle. In recent times the original two-dimensional scheme introduced by Cundall and Strack (1979) has been extended to three dimensions for six sided solids (Ghaboussi 1990), ellipsoids (Lin and Ng 1997) and super-quadratics (Williams and Pentland 1989). DEM is well suited to modelling the effects of particle shape on granular flow (Sharrock 2003a; Figure 2). While DEM is without doubt the most popular discrete method in granular science, currently it cannot be applied to solving mine scale SLC problems, largely because of computational limitations. This is due in part to a requirement for numerical stability in the central difference approximation to Newton’s second law. In addition, this method also requires computationally intensive near neighbour searches to determine particle contacts within the assembly. With existing computing hardware, DEM is most commonly used for assemblies of 104 or 105 particles for short model time durations. Well known DEM codes include PFC3D, REBOP, FASTDISC and FLOW3D. Because of the accuracy of DEM, it is well suited for detailed gravity flow studies on factors affecting complex flow phenomena in SLC and BC mines. Some excellent work has been undertaken in the International Caving Study on this topic (Brown 2003). For this reason, CAVE-SIM results have been tested against DEM simulations from FLOW3D, a three dimensional particle flow code for sphere clusters (Sharrock 2003a).
Figure 2: Calibration curves for inter-particle contact friction angle vs material internal friction angle; axi-symmetric biaxial test; undertaken with the 3D DEM code, FLOW3D; using cluster particles (Sharrock 2003a). 2.3. Conclusion It is proposed that neither SM nor DEM are ideally suited to "mine scale" SLC simulations. SM does not include the physics of granular flow, while DEM will continue to be limited by the requirements for numerical stability for the forseeable future. In contrast, cellular automata are well suited to extremely large and complex problems in discrete systems. CA permits very large systems to be studied while including conventional physics. For this reason, CA has gained wide acceptance in science and engineering fields concerned with discrete mechanics. 3 CELLULAR AUTOMATA The Cellular Automata (CA) concept was first developed by John Von Neumann in 1947. Von Neumann reasoned that a conceptual machine, or automaton, could be constructed by dividing the physical space of a problem into a large number of "cells" that interact according to partial differential equations describing the physics of a system (Von Neumann 1966). In this approach, cells contain discrete objects that are categorised by individual "state" parameters that evolve dynamically according to the partial differential equations. CA has been used in a number of contexts; in quantitative inquiries such as the investigation of physical materials, engineering problems, mathematics and the traditionally "qualitative" fields such as the social sciences. Examples from engineering include computational fluid dynamics (Cattaneo and Jocher 2002), particle physics (Fox et.al. 1994), gas dynamics (Frisch 1986), stress in granular media (Hemmingson et.al 1997, Sharrock 2003a) and gravity flow of granular media (Baxter and Behringer 1990). CAVE-SIM is designed using CA techniques and models particle friction, particle size distributions and stress for a three-dimensional SLC model, governed by a series of sequencing and draw-rule options .
Figure 3: 3D stress and segregation in a binary mixture: CAVE-SIM (Sharrock 2003). 190
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4 CAVESIM CAVE-SIM can efficiently simulate the physics of very large numbers of particles (ie >30e6 particles), for long simulation times, over short computational periods (i.e. 2% Ni) ore located at the northern end of this block of the disseminated orebody. Development in the upper levels (1145, 1130, and 1115 Levels) was thus concentrated at the southern end of the deposit. Transverse production drifts were used for these sub-levels. The mine plan required 4.5m wide by 4.5m high cross-cuts to be developed from the hangingwall drive to the orebody footwall position. Footwall drives would then be developed to connect the cross-cuts and enable future establishment of production slots between sublevels. As cross-cuts were advanced by electric hydraulic jumbos, weld mesh (5mm diameter wires at 100mm spacings) and friction stabiliser bolts (2.4m long) were installed across the backs and half way down the sidewalls. Sprayed concrete (shotcrete) support which had been used during previous mining in the adjacent development was considered to be too costly. Ground control issues that arose early during the second phase of stoping consisted of: • Extensive unravelling of the rock mass within the hangingwall contact zone The hangingwall contact between felsic volcanic hangingwall rocks and the dunite ultramafics comprised a major shear zone up to ~15m wide. The felsic volcanic rock mass was intensely fractured and contained thick seams of clay gouge. The ultramafic contained numerous mylonitic shears up to ~1m wide. Unravelling would commence immediately after each development cut. During installation of the friction bolts and mesh, rock mass unravelling was exacerbated by the effects of drilling and hammering of the bolts into the rock mass. Due to the inability to tightly pin the mesh against the rock mass, unravelling would continue until the mesh was filled with loose rocks. The very poor rock mass conditions encountered at the southern end of the orebody prompted a change in drift layout from transverse to longitudinal in an effort to avoid the hangingwall shear.
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Table 1: Case Study Orebody & Mine characteristics Orebody Characteristics
Mine A
Mine B
Mine C
Host Rock
Ultramafic
Mafic Sediment
Schist
Strike Length
200m – 300m (~ north-south)
300m (~ north-south)
500m – 700m (~ north-south)
Width
100m
30m
35m
Dip
75°
75°
75°
Production Layout
Transverse
Longitudinal (some transverse)
Longitudinal
Sub-level Interval
15m – 22m
25m
25m
Cross-cut/Drive Spacing
14m - 17.5m
11m
12.5m
Cross-cut/Drive Sizes
4.8m x 4.8m
5.0m x 5.0m
4.5m x 4.5m
Table 2: Rock stress regimes at Case Study Mines Mine Mine A (770m depth)
Mine B(600m depth)
Mine C (600m depth)
Principal Stresses
Magnitude (MPa)
Plunge (°)
Azimuth (°)
Major
111
49
166
Intermediate
83
19
278
Minor
65
35
022
Major
44
0
050
Intermediate
27
0
140
Minor
18
90
–
Major
86
10
266
Intermediate
38
29
170
Minor
31
59
014
• Large wedge failures from the backs of the development Away from the hangingwall contact, large block/wedge failures occurred from the backs of production drifts. The blocks were defined by the intersections of persistent geological structures which had been exposed by formation of the openings. Large spans formed at the intersection of the cross-cuts and footwall drives were particularly prone to block failure during development. • Time dependent rock mass deterioration Relatively shortly after developing the three upper sublevels it was clearly apparent that the rock mass was deteriorating with time. Friction bolts were being ruptured close behind the rock surface. SLC development at this stage was characterised by excessive overbreak, particularly from the backs of the cross-cuts. Loosening within the rock mass was deemed to be excessive. Approximately four years following resumption of operations, with mining still taking place in the upper levels, there was a fundamental change in the approach to the mining and support of SLC development. It was concluded that the previous approach would not provide sufficiently stable cross-cuts, in particular production brows, to allow a sustainable high SLC production rate. The level and intensity of support installed in the SLC development was significantly increased. Later development openings were systematically supported floor to floor with mesh-reinforced fibrecrete (shotcrete reinforced with steel fibres), and reinforced with 5m long 250kN cablebolts installed in rings containing 13 to 15 cablebolts, spaced either 1.2m or 2.5m apart, depending on local rock mass conditions. This ground control scheme significantly reduced overbreak and improved SLC brow conditions for continuous production. Massmin 2004
All earlier development had to be rehabilitated as best as practicably possible to the new ground support standard. The majority of this development had fibrecrete sprayed directly over the installed mesh and down to floor level. The lower sidewalls had fibrecrete applied directly onto the rock surface. Cablebolts were also installed through the fibrecrete using the above described scheme. Due to the time lag between initial development and the upgrade in the level of ground support and reinforcement, ground control issues still occurred in this rehabilitated development. These included: • Large scale failure in cross-cuts Some of the cross-cuts experienced large scale back and/or wall failures (up to several metres deep) even when rehabilitated. These crosscuts were deemed to be unsuitable for SLC drill and blast and/or draw control and were abandoned and backfilled. • Poor SLC brow conditions Due to the high yield capacity of the friction bolts and the limited confinement provided by the weld mesh, the rock mass around the drift opening deformed readily. Rehabilitation was thus often installed after large deformations had occurred. The main problems experienced in these situations were dislocation and blockage of blastholes and loss of brows. The required rehabilitation works were extensive, and were undertaken simultaneously with an increase in production rate; hence the program was very difficult to manage. During 1997 a decision was made to establish a new SLC mining area at a lower level, commencing at ~600m (920 Level) below surface. SLC mining of the upper block would be completed at ~500m (1000 Level) below surface. The
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920 Level required development of 19 cross-cuts up to 100m long. A hangingwall drive had been established several years before, thereby allowing multiple face crosscut development. Stoping was planned to retreat from the southernmost cross-cuts. The major geotechnical issue with establishment of the lower caving area was the adverse influence of (unexpected) very high rock stresses on the new development. Even with the new ground support regime, significant damage was experienced, and the rate and extent of damage to the cross-cuts increased with time. Cross-cuts experienced rapid sidewall closure due to the highly fractured rock mass dilating (bulking up) behind the mesh reinforced fibrecrete layer. The competence of the cross-cut backs also decreased with degradation of the drift pillars. It became clear that the intensive ground support could not prevent the cross-cuts being damaged extensively. A number of factors must be recognised and accounted for in such a situation: • Prioritisation of sub-level development and reducing sublevel service life The initial stages of development on the 920 Level focused more on developing the entire level rather than only in the southern area that was to come into production first. The complexity of the development and ground support was such that it was easier to manage if spread out, rather than concentrated in a particular area. As a consequence, many of the cross-cuts had to remain open for extended periods of time while adjacent cross-cuts were completed, and the initial cross-cuts had to be rehabilitated several times. One cross-cut had to be abandoned due to extreme wall closure. As a result a ‘just in time’ approach was adopted when developing cross-cuts. • Importance of immediate remediation During mining of the 920 Level several of the cross-cuts experienced major back failures. Rather than immediately rehabilitating, or attempting to recover a cross-cut, there was a tendency to delay action for several weeks or months. At the same time, the adjacent cross-cuts were being advanced, leaving the affected cross-cut lagging by up to tens of metres. Production could not commence until all cross-cuts were completed; hence the cross-cuts completed earlier would have to remain open for an extended period of time, giving rise to the possibility of disturbance and damage. It was therefore imperative that high priority was given to rehabilitation of cross-cuts which had experienced major failure. • Timing of the installation of rock reinforcement and support The objective of the ground support scheme was to control the rate and extent of rock mass deformation within the cross-cut development sufficiently to allow production to be completed (i.e. consumption of the drive by the SLC). Delaying installation of ground support and reinforcement frequently led to ground support and reinforcement not being installed to the minimum standard, simply due to sections of development being overlooked and/or poor allocation of resources. Generally, as the quantity of development that had not been completed to standard increased, so did the difficulty in managing the work force and equipment to undertake the work and "catch up" to the schedule. At some stage all new development had to be stopped and resources re-allocated in order that a "catch up" phase of ground support and reinforcement could be undertaken on the older development. 260
Greater difficulties were experienced in installing the rock reinforcement (rock bolts and cablebolts) due to holes becoming blocked/dislocated following movement of the rock mass. Installation problems almost invariably increased if rock reinforcement was not installed progressively with drilling (i.e. rock and cablebolts had to be installed immediately after completion of drilling individual holes). 2.2 Case Study B Underground mining at Mine B commenced in late 1989 after completion of an ~80m deep open pit. Gold mineralisation was contained in a mafic conglomerate between an ultramafic hangingwall and a meta-sediment footwall. Underground mining at Mine B was completed in 2002 at a final depth of ~700m below surface. SLC was chosen to extract the Mine B orebody predominantly due to its width, relatively low grade (800m) and relatively narrow width (>30m) of the orebody, a longitudinal layout with 1 or 2 ore drives was used. On each level the ore drives were generally accessed via several cross-cuts developed from a footwall haulage drive. Even at relatively shallow depths below surface (1.4 mm/day – hazard of uncontrolled ground movement; work to be suspended till further notice. With time, it was found that with convergence varying from 0.8 to 1.4 mm/day there was no visible damage to the rock mass. The criteria were founded to be too stringent often resulting in withdrawal of working personnel where there was no hazard of uncontrolled ground movement. As a result the criteria were modified: 266
• < 1 mm/day – no hazard of uncontrolled ground movement; work can continue. • between 1.0 to 2.0 mm/day – a limited hazard of uncontrolled ground movement, short time activities, under supervision, • >2 mm/day – hazard of uncontrolled ground movement; work to be suspended till further notice Convergence in the North Crusher. Two convergence stations were installed across the crusher in the east-west direction. A distinct change in daily convergence across the crusher were observed in three periods: • For the period of three months before the closure the daily changes in convergence varied moderately from –0.22 to 1.08 mm/day with an average daily increase in convergence of 0.50 mm. • During the period for six weeks from the time of closure to suspension of salvage, changes highly varied between 0 to 2.62 mm/day with an average being 0.9 mm/day. The trend line showed gradual increase in average convergence growing from 0.8 to 1 mm/day. • After suspension of salvage daily change in convergence varied from -0.3 to 1.4 mm with a trend line decreasing from 0.8 to 0.5 mm/day. Changes in daily convergence in the crusher are shown in Fig. 7. Convergence in the drainage drift. Convergence stations in the drainage drift were installed in the same direction (parallel) to the convergence stations in the crusher and monitoring results showed similar behaviour. • For the period of three months before the closure, the daily convergence varied between –0.13 to 0.74 mm with an average daily increase in convergence of 0.45 mm. The trend of velocity averaged 0.4 - 0.5 mm/day. • During the period from the time of closure to suspension of salvage, varied between 0.0 – 0.96 mm/day with an average being 0.5 mm/day. The trend line showed a gradual increase in convergence from 0.45 to 0.55 mm/day. • After suspension of salvage activities the daily convergence varied from –0.21 to 1.68 mm with a trend line decreasing from 0.6 to 0.15 mm/day. Changes in daily convergence in the drainage drift are shown in Figure. 8. Convergence in the conveyer drift. In general the velocity of convergence in this drift did not show significant changes and that could be due to the orientation of the convergence station with measuring pins parallel to the crusher (installed in north-south direction). 4. ROCK MASS BEHAVIOUR LEADING TO SUSPENSION OF SALVAGE OPERATIONS The sequence of geotechnical events indicated that the rock mass around the conveyer level was undergoing high mining induced stress and was progressively failing, (Szwedzicki 2004). Deterioration in ground conditions was accelerating in time and severity, and could result in rock mass instability around the crusher and other places on the conveyor level. Analysis of the sequence of the geotechnical events and subsequently conducted risk analysis resulted in a decision to suspended of salvage operation and to withdraw crews from the crusher. Six weeks after the decision to close crushing operations, the rate of convergence in the crusher increased to 4.91 mm/day and seismic activities were recorded. At that time, a decision to stop salvage operations was made.
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Figure 7: Daily change in convergence in the North Crusher
Figure 8: Daily change in convergence in an access drainage drift
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During the salvage operations the following geotechnical events were noticed: Rock falls in an exhaust drift. The exhaust drift was situated 56 m below the crusher and 200 m to the west of the DOZ caving zone. The first signs of deterioration were noticed about six months before closure. About a month before closure, more than 30 m of the drift subsided for about 0.2 – 0.3 m. Acceleration in deterioration was visible after closure. Loose rocks detaching from the back were found in many places with two major rock falls occurring. A long fracture in the floor of an incline. A fracture was found in the floor of the incline from the conveyor level to the undercut level. The cracks become visible after the closure and continued to extend and open for the next few weeks (the total length of the fracture was about 80 m and the width of the crack was up to 2.5 mm). A few more cracks developed parallel to that main fracture. Rock fall at the intersection of a feeder and a conveyor drift. A fall of ground (10 m3) took place at a brow between the conveyor drift and a feeder about 4 m from a chute. In a few days water started to seep out through cracks at the back. Bolts and mesh were highly corroded and provided very little (if any) support. The water seepage increased in time and spread to adjacent ribs and shoulder of the feeder. Change in water inflow pattern. Rock mass fracturing caused by mining extraction and caving resulted in change in the water inflow pattern. In drifts on undercut and extraction levels, where water used to accumulate due to uneven floor, water disappeared during the period of about six to three months before closure. Conversely in a number of areas water started to appear. The west wall of the North Crusher started to be moist about three months before closure. After the closure, water was seeping and at the time of suspension of salvage the whole west wall was wet and some seepage started to appear at the back of the crusher and along conveyor drift and feeders. At the same time it was noticed that water inflow to draw points increased. Seismic activities. A decision to suspend salvage operations was made after recording three large, and a number of small, seismic events in six hours. All events that took place, except for the last one, were felt for a short period of time – about 10 –15 seconds. The last event lasted for about 15 minutes with "progressive banging" heard in the roof of the crusher. The last event was accompanied by liberation of large quantities of rock dust in the crusher. An immediate geotechnical inspection and results of geotechnical instrumentation did not show any substantial changes in damage in the vicinity of the crusher. The events appeared to be caused by merging of the IOZ and DOZ zones of caving influence. There were also a number of reports that earth tremors were felt at distance varying from 5 to 10 km. Within the next few days it was noticed that in the vicinity of the crusher there was acceleration in damage such as small detachment of rocks, spalling, slabbing, crack propagation and also floor heave. 5. CONTINUATION OF SALVAGE OPERATIONS After seismic activities subsided, daily changes in convergence were reduced to 0.8 mm or less. It was also noticed that ground deterioration ceased. After eight days of monitoring, it was concluded that mining induced stress had re-distributed and that it was safe to continue salvage operations. Salvage was carried out under close geotechnical supervision. With time the number of geotechnical events subsided, convergence reduced and there was very little progressive damage. In retrospection, looking at the cumulative convergence curve, Fig. 2, it can be seen that the decision to close the
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crusher operations was made when the curve displayed a transgressive trend (Szwedzicki, 2003) i.e. daily changes in convergence were increasing. The decision to recommence salvage operations was made when the curve started to display a regressive trend i.e. daily changes in convergence were decreasing. Operational decisions based on geotechnical monitoring and risk management allowed for timely closure of the crusher chamber and for safe and successful salvage of equipment. 6. CONCLUSIONS Propagation of the zone of caving influence of the Deep Ore body Zone Mine (DOZ) resulted in changes in ground conditions at the conveyor level at the Intermediate Ore body Zone (IOZ) Mine. Early warning signs and the rock mass precursory behavior indicated that rock mass was under high mining induced stress. Analysis of geotechnical events proved that: • The first geotechnical warning signs leading to major ground deterioration in the crusher and the whole conveyor level were detected three years prior to the decision of closure. • Geotechnical events such as ground deterioration, rock falls, support damage, convergence and seismic events increased in time and severity. Large scale ground instability was preceded by small scale events. • The recorded geotechnical events were noticed where the rock mass had the lowest mechanical properties, there were geological structures, mining excavations had large open span, and ground support was corroded. • Damage to the structure of the rock mass was followed by a change in water inflow pattern. Warning signs and the sequence of geotechnical events with their acceleration in time and severity formed base for risk management analysis. As a result the following subsequent decisions were made to: • Stop the crushing operations, close crusher chamber and start salvage of the crushing equipment • Suspend salvage of the equipment • Re-commence salvage. Geotechnical inspections, monitoring, analysis and supervision allowed for timely decisions to close the crusher operations and to salvage crusher equipment safely and successfully. REFERENCES • Barber J, Thomas L, Casten T, 2000, Freeport Indonesia’s Deep Ore Zone Mine. Proceedings Mass Min 2000, pp 289-300, Brisbane. • Hubert G, Dirdjosuwondo S, Plaisance R, Thomas L, 2000, Tele-Operations at Freeport to Reduce Wet Muck Hazard. Proceedings Mass Min 2000, pp 173 - 180, Brisbane. • Rachmad L, Sahupala A H, 2003, Update on North Crusher Condition. Internal Report, PT. Freeport Indonesia. • Rachmad L, Widijanto E, 2003, Application of Convergence Monitoring at PT Freeport Indonesia Deep Ore Zone Mine. Proc NARMS-TAC, (editors Hammah at al.) pp. 181 – 189. • Szwedzicki T, 2004, Warning Signs to Geotechnical Failure of Mining Structures. Int J. Surface Mining, June, Vol 18.2. • Szwedzicki T, 2003, Behaviour of Rock Mass Prior to Failure, Int. J. Rock Mechanics, Vol 40. pp 565 – 572.
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Application of convergence monitoring to manage induced stress by mining activities at PT Freeport Indonesia deep ore zone mine Indra Febrian, Chief Engineer of Geotechnical Engineering, Widyo Yudanto, Engineer of Geotechnical Engineering, PT Freeport Indonesia, Tembagapura, Indonesia Enrique Rubio, Consultant Mining Engineer, Gemcom Software International, Vancouver, Canada
Abstract Freeport Indonesia’s Deep Ore Zone (DOZ) mine is located in the East Ertsberg Skarn System deposit. DOZ uses mechanized advanced panel caving method to produce 38,000 tpd with a grade of 0.60% copper equivalent. Currently the mine operates 15 production drifts which run across the economic layout. It is planned that the number of production panels will increase up to 27 by the end of the West extension, reaching productions of 50,000 tpd. One of the main concepts learned from the passed experiences at IOZ and GBT mines was that block cave needs to be fairly instrumented in order to assess rock mass behavior due to mining activities. One of the instrumentation used at DOZ is the convergence monitoring system which consists of measuring the deformation of the production tunnels and undercut tunnels. The convergence system is considered fundamental to understand the rock mass behavior as a response of mining activities such as undercutting and production stage, providing information to mine planning and mine operation sections to maintain ground stability. The average spacing between convergence stations on the undercut level is between 5.0 – 10m apart and on the extraction level is between 15 to 18 m depending on draw point spacing. The results of convergence monitoring have been used as guideline for the undercutting and mucking strategies to manage induced stress. This paper describes the usage of convergence monitoring in understanding rock mass behavior induced by mining activities and its application in DOZ mine.
1 INTRODUCTION
2. CONVERGENCE MONITORING
The Deep Ore Zone (DOZ) is the underground block cave mine at Freeport Indonesia. It is located about 300 meters below the last active mine IOZ (Intermediate Ore Zone) and about 1,200 meters below surface. The IOZ mine, started production in 1983 and finished production by 2001. IOZ operation experienced complex problems such as considerable displacements leading to the collapse of the production drifts. Several lessons were learned in the operation of IOZ , probably one of the most important ones is related to geomechanic instrumentation and its usage to manage caving activity. The intensive instrumentation program consisted of convergence monitoring system, multipoint borehole extensometer, and three different types of relative stress monitoring devices. This paper concentrates the discussion about the convergence monitoring system. In general, monitoring is carried out for two main reasons [1]: • Ensure safety during construction and operation by assessing on ground deformation, ground water pressure, load in support elements. • Assess abutment stress zones at the undercut level and production level ensuring construction is carried on outside the abutment stress zones Calibrate initial assumptions regarding rock mass properties used in planning and design of the DOZ mine
Freeport started implementing intensive ground monitoring by the middle of 90’s following large displacements occurred at IOZ production drifts. Convergence Monitoring was chosen because of its simplicity, repeatability, and easy operation. The main purpose of convergence monitoring at the DOZ Mine was to ensure safety during production, and provide an early warning on the event of excessive movement at the undercut and production level. Another benefit related to the convergence monitoring system is to provide information to assess the status of the induced stresses and rock behavior during undercutting and production stage, providing proper information to mine planning and the mine operation section to maintain ground stability. Tape Extensometer is used for a convergence measuring system developed by Kovari et al (1974). The displacement gauge has a least count of 0.01 mm and a range of 100 mm. The overall accuracy of the convergence measurement is 0.02 mm. Three pins are installed in the wall at each station using 1.0m threadbar and grouting cement, as shown in Figure 1. Measurements are taken every week for each station and almost everyday in areas that may have reported high displacements until the station stopped showing continuous movement then the rate is gradually decreased.
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o The main areas in which monitoring is performed regularly are as follows: 1. Undercut level: Convergence Monitoring stations are 5.0 - 10m apart. 2. Extraction level: Convergence Monitoring station are located every two draw points. The main purpose of the stations located on the undercut level stations is to register the deformation of the undercut drifts occurred due to the abutment stress zones produced by the mining method. The deformation data is analyzed providing the extension and magnitude of the induced stress zones. Ones the abutment stress zone is identified guidelines can be design for safety and blasting design. The stations located on the production level register the deformation of the drifts due to static and dynamic loads transferred throughout the broken muck pile to the production level. The origin of these loads has been fully studied however it is widely agreed that these loads are the result of different draw patterns,
Fig.2. Historical Case Convergence Monitoring • Stress distribution due to draw pattern. In Figure 3 peak no. 4 shows an abnormal displacement in the active production area. Normally, there should not be much movement on the production area since the area is within the stress shadow. However high displacements have been observed due to differential draw and its effect on the overlying broken muck pile. It is believe that when a draw point is not continuously drawn the broken muck pile begins to compact and transfers stresses to the major and minor apex pillars. Using the contour map, the engineer recognizes the anomaly and makes a plan to approach the problem modifying the draw pattern on the daily draw order.
Figure 1 Convergence Installation at DOZ
3. DATA VISUALIZATION For analysis and visualization, the deformation from convergence monitoring is converted into velocity (mm/day). The velocity of each station is then plotted on a displacement velocity contour map using Surfer program. The contour map gives easy-to-read data presentation and valuable information to the mine operation. Below are historical examples of the usage of convergence monitoring data in DOZ mine. • High stress location at the perimeter of undercut boundary. As shown in Figure 2, there are four contour peaks (no.1 to no.4) of high displacement velocity located at the undercut boundary (grey line). This fact proves the theoretical assumption that the concentration of stress will be the most at the undercut boundary due to the abutment stress. Moreover, from the convergence readings and the damage observation in the field, it has been found that the velocity of 0.5 mm/day being a cutoff value of ‘need attention’ displacement. In this case, when the peak shows a 0.5 mm/day or above, the caving operation will be warned to move the cave face immediately with the blasting progress by geotech recommendations.
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Fig.3. Historical Case Convergence Monitoring
4. BLASTING ACTIVITIES As undercutting advances, the stress will be transferred and distributed around the perimeter of the undercutting area. The time of area influenced by the abutment stress is important for planning and scheduling purpose. The area of the abutment zone has been estimated to be 20 to 30 meters from cave front, based on the convergence measurements. In Figure 4, the convergence stations show decreasing of horizontal velocity due to continuous undercut blasting and
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increasing of horizontal velocity due to suspension of blasting. It has been found that horizontal displacement increases if the cave front is not blasted for more than one week.
convergence or displacement. A few conclusions can be drawn studding the behavior of the displacement versus different draw patterns, the following points summarize the observations • As the mucking rate increases, the displacement decreases • At a consistent mucking rate a consistent convergence is observed • Differential draw induces, high convergence
Fig 4. Decreasing of horizontal velocity due to progressive blasting
Based on this fact, it is concluded that the magnitude and shape of the abutment stress zone is influenced by blasting activities. Therefore a few guidelines have been designed in order to avoid high stresses on the undercut level that could eventually induce damage and collapse of the undercut level. Examples of these guidelines are shown as follows: Convergence (mm/day) 0.5-1.0 > 1.0
Blasting 1 ring blasted More than 1 ring blasted
Fig.5. Mucking effect on Convergence Rate Based on the above points can be concluded that there are stresses acting on the production drifts induced by the draw patterns. Therefore it would be worthwhile to find relationships that better explain this phenomenon and eventually could be embedded in a production planning system to measure the impact of stresses given different production strategies. 6. CONCLUSIONS
5. DISPLACEMENT VERSUS DRAW RATE DUE TO MUCKING PLAN From previous mine experience, it has been observed that there is a relationship between displacement rate of production drifts and draw rate. The previous observations have been confirmed at DOZ and it is shown in the data collected in the past 1 year, Figure 5, that there is an inverse relationship between draw rate and Massmin 2004
• The results of Convergence monitoring are used as guidelines to manage the mining induced stress. • Geotechnical Recommendation are given as soon as any one of condition are met: – The convergence rate is in the range of 0.5 mm/day – 1.00mm/day – Cumulative displacement of convergence is more than 30mm.
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Even though the system is simple, the data produced is very useful. The application of convergence is not limited to what is presented in this paper and ongoing investigation is required, providing further challenges to the underground geotechnical engineer. This simple device has proven to be useful to ensure safety during development and operation, and to check the validity of the assumptions relating to rock behavior in the block caving mining method. ACKNOWLEDGEMENTS The author would like to thank the management of PT. Freeport Indonesia for permission to publish this paper. The contribution made by underground personnel involved in block caving mining at DOZ mine, especially Szwedzicki, Tadeusz and M. Stawski, to this paper are gratefully acknowledged.
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REFERENCES • Brady, B.H.G. and Brown, E.T. 1993. Monitoring rock mass performance. In Rock Mechanics for Underground Mining. 2nd ed, 491-496. • Butcher, R.J. 2000. Block Cave Undercutting – Aims, Strategies, Methods, and Management. In Proceedings of Mass Min 2000 Conference, Brisbane, 29 Oct-2 Nov 2000, 406-410. • Butcher, R.J. 2000. The Role of Mass Concrete in Soft Rock Block Cave Mines. In Proceedings of Mass Min 2000 Conference, Brisbane, 29 Oct-2 Nov 2000, 423. • Laubscher, D. H. 1994. Cave mining: state of the art, J. Sth Afr Inst Min Met. 94:279-293.
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Optimization of the mining business through geomechanics: Two Case Histories Carlos Soto, Head, Rock Mechanics Division, Juan C. Cereceda, Specialist Enginee; Francisco Orcaistegui, Geomechanics Engineer, Arcadis Geotécnica, Santiago, Chile
Abstract The economic optimisation of the mining business is illustrated through two case histories, for surface and underground mining: the Mantoverde copper mine (Anglo American Chile); and Navío limestone mine (Lafarge Cement Chile), respectively. In the Mantoverde case, extensive characterisation and analytical studies were conducted to formulate rock slope design guidelines for a main pit wall. Final recommendations reflected an increased final pit slope angle, with the favourable consequences of greater ore recovery, reduced stripping ratio, and obvious production cost reduction. These were not obtained at the expense of overall pit safety, but through a more realistic, reliable assessment of events which actually control pit slope behaviour, as opposed to the adoption of excessively theoretical formulations. For the Navío mine case, two crucial objectives were achieved: technical optimisation of its underground mining method, and maximisation of economic ore recovery for the remainder of its operating life, in an adverse geomechanical environment. A cost-efficient numerical modelling methodology was adopted for all further mining method design, combining two-dimensional and three-dimensional formulations. Modelling results were validated by a displacement monitoring program. In addition to a safe refinement of stope and pillar dimensions, this allowed the formulation of a mining sequence capable of yielding the complex, precise ore-blending requirements of the nearby cementmanufacturing plant. Central conclusions are that mine design optimisation, with moderately increased ore recovery, is feasible and already underway, and that the extraction ratio is now said to be close to a geomechanical optimum. The central aim in these two applications has been one of demonstrating the potential of geomechanics not just as a mere instrument in sophisticated mine design, but as an effective aid in the pursuit of long term mine planning objectives; and ultimately, in the economic optimisation of the mining business.
1 INTRODUCTION
2 DESIGN REFINEMENT OF EAST WALL OF MANTOVERDE OPEN PIT
Some fundamentals of economic optimization of the mining business through geomechanics are presented in this paper through two case histories: the Mantoverde open pit copper mine, and the Navío underground caliza mine. In the first case, Empresa Minera de Mantos Blancos (Anglo American Chile) operates its Mantoverde open pit copper mine in Northern Chile, some 40 km South of Chañaral city, or 1,020 km North of the capital city of Santiago (see Figure 1), whereas Cemento Melón (Lafarge Cement Chile) operates its Navío underground limestone mine at approximately 140 km north of Santiago (Figure 2).
Figure 1: Mantoverde Open Pit. Massmin 2004
Economic optimization of this operation called for slope design refinement for the east wall. This has been the subject of detailed geotechnical studies carried out in recent years, conclusions from which are summarized below. Geological Setting Regional conditions include intense tectonic activity associated to the Atacama fault, the principal structural feature. Extensive outcrops of andesitic rocks, intruded by
Figure 2: Navío underground, Sub-Level Stoping mine. Santiago Chile, 22-25 August 2004
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dykes, are present in the area, as well as smaller tabular bodies of granitic composition. Locally, the Mantoverde fault is the predominant feature, also related to mineralization, and along which various hydrothermal and tectonic breccia formations are found. As shown in Figure 3, the geologic/structural sequence in the East wall comprises, East to West: chloritic andesites, transition zone (parallel to the Mantoverde fault), hydrothermal breccias, gouge zone, "green" breccias, and intrusive rocks.
Figura 4: Dips’ output for joints in Andesites and "green" Breccia combined.
Figure 3: Geological cross-section.
Design Methodologies Extensive characterisation and analytical studies were conducted to arrive at the recommended slope design. This included the adoption of broadly accepted methodologies of geotechnical characterization for various rock types (including laboratory testing of representative specimens); statistical analyses of geological structure, by means of Dips software (see Figure 4); kinematic feasibility analyses for discernible failure modes of pit slopes (mainly wedge and circular failures); and stability assessments by limit equilibrium analytical techniques (Bishop’s simplified method; Figure 5).
Figura 5: Bishop’s simplified analysis for full slope height. Slope Design Recommendations As a result of the above analytical work, two versions of a complete "Slope Design Curve" were formulated, for two different values of the main design criterion: the Safety Factor (SF). Each of these displays, graphically, the complete slope angle / slope height relationship for the full range of slope heights to be encountered in the pit’s operational life. Figure 6 shows these two slope design curves, designated as recommended design (SF= 1.30), and optimistic design (SF= 1.15). Obviously, the main difference between both curves is that, in the optimistic design, a discernible slope steepening is achieved for similar slope heights, thus generating the significant economic benefits which normally arise from a lower stripping ratio.
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Figure 6: Recommended and Optimistic slope design curves for East wall.
Adoption of (at least) the recommended design guidelines is adequately supported by the available information, and reflects a so-called "standard" degree of certainty or reliability of pit slope design. Furthermore, adoption of the optimistic design guidelines is greatly encouraged provided that an efficient, reliable slope monitoring system is rigorously implemented throughout the entire operational life of the pit.
Figure 7: Iso-contours for horizontal and vertical displacements in 350m-high East slope.
In this particular project, the latter recommendation is based on the positive outcome of broad sensitivity analyses performed between the "extremes" represented by both design curves. Such outcome indicated that: • significant operational savings are realized as we move from recommended to optimistic design (impact on stripping ratio); and • such economic benefit far outweighs the combined cost of: cleaning up occasional rock failures which may thus take place in individual benches or steeper working slopes, and implementing the required slope monitoring system which would feed back hypothetical warnings into the slope design process. Finally, for added assurance, the design of the highest (350 m), final pit slope obtained above was verified through numerical modelling, by means of the Flac2d software. Typical results are shown in Figure 7, in terms of isocontours of horizontal and vertical displacements (as FS evolves from 1.1 to 1.2), in which an overall, approximately circular failure surface clearly emerges. Furthermore, as illustrated in Figure 8, the overall failure surface dictating final pit design by numerical modelling very closely approaches that previously identified by analytical, limit equilibrium techniques.
Figure 8: Overall failure surface comparison, analytical and numerical methods.
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3 OPTIMISATION OF UNDERGROUND MINING METHOD AT NAVIO MINE Navío mine is the prime limestone supplier for Lafarge Cement’s manufacturing plant in Central Chile. Overall economic optimization calls for more massive operations, in order to remain competitive with surface operations in the same region. Thus, the mining method at Navío has evolved from earlier forms of Room & Pillar decades ago, to the current version of a more massive Sublevel Open Stoping operation. This mining method optimisation is more acutely needed as mining progresses into deeper geological horizons, with their ensuing problems of high in situ and induced stresses, complex tectonics resulting in severe dislocation of the two main sub-parallel seams which comprise the orebody, and accompanying intense fracturing. A series of geomechanics studies have been carried out in support of mine planning’s two main objectives: technical optimisation of its mining method, and maximisation of economic ore recovery for the remainder of its operating life. Both would be achieved by the latest modification to the mining method, i.e. Sublevel Stoping with a vertical distance between drill drifts increased from 20m to 40m. Geological Setting. The Navio orebody occurs within a marine sedimentary sequence, the La Calera formation, of Jurasic age. Over its 430m thickness, this sequence comprises alternating sandstones, lutites, limestones, and fine conglomerates, flanked at floor and roof by lavas, tuffs, and volcanic breccias. The limestones of economic interest are disposed in two sub-parallel limestone seams, designated as Upper and Lower Seams or Mantos, both with an average thickness of about 12 to 15m, and a dip increasing to about 45º at the greater depths to be mined in the future. A 65m thick inter-seam stratum separates both seams, and plays a key role in the geomechanics interaction between stopes being mined simultaneously. This general arrangement is illustrated in Figure 2 (already introduced), which also depicts a "typical" stope; i.e., one which mines up to full manto thickness, and extends about 45m down dip, and 60 to 80m along strike. The required refinement of the mining method is to be pursued through the optimisation of dimensions for the various pillars involved.
Figure 9: Mining sequences analysed. Subsequently, any particular sequence from the above figure would be further analysed for deeper sectors designated for future mining, as early geomechanics support to mine planning. Typical results from theses preliminary analyses are displayed in Figures 10a and 10b for the radial and longitudinal "sampling" directions, respectively. A figure 10a show that, as a convention, the radial sampling of results is only displayed for the hanging wall of the latest stope, in a given excavation sequence. Additionally, it obviously shows the expected decrease in the safety factor as extraction ratio increases. The minimum value of FS = 1.0 is reached in the last stage, and remains below an acceptable design criterion of say, FS = 1.25, for about 5m into the hanging wall. However, irregularities in the actual excavation may modify this trend. It is important to point out that modelling results were validated by a displacement monitoring program.
Mining sequence optimisation by two-dimensional numerical modelling Design optimisation of the current Sublevel Open Stoping mining method has proceeded in two stages, which are only briefly outlined here. Firstly, a broad series of "generic" designs has been developed, via two-dimensional modelling (Phase2 finite element code), in an attempt to explore the geomechanical impact of simultaneously mining stopes from both seams, for a large number of scenarios or configurations compatible with ore blending requirements. The essential result from this stage, dealt with in the present section, is a limited number of "promising" mining sequences warranting further consideration. Modelled mining sequences Through close interaction with Mine Planning, an extensive search of promising mining sequences was conducted, so as to select a limited number of configurations, satisfying economic and ore-blending requirements, for initial geomechanics feasibility assessment. Figure 9 summarises the best configurations finally analysed by numerical modelling, where the correlation numbers indicate the actual mining sequence adopted in each case. 276
Figure 10: Safety Factor sampled results.
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Mining Method Design Refinement by ThreeDimensional Modelling This second stage undertakes the optimisation or refinement of the previously selected configurations. Through a limited number of more detailed, threedimensional modelling applications, it seeks the satisfactory balance of two central requirements: assurance of local and global stability, and maximisation of ore extraction ratio via minimisation of pillar dimensions Examine 3D, v.4.0 (Rocscience, Toronto) was adopted for this design refinement stage. This is an engineering analysis program to essentially perform stress and displacement analysis for underground excavations in rock, by means of the direct boundary element method. However, its powerful data visualisation tools make it amenable for application to a much wider range of mining and civil engineering problems. The first mining sequence configuration included in Figure 9 has been selected to illustrate this design refinement process, because of its broad representativeness of the Sublevel Open Stoping mining method, and its more complex geomechanics interaction. To generate this three-dimensional model, the original (2-d) model has been expanded along the seams’ strike to incorporate up to three stopes and two "rib pillars" in that direction. As illustrated in the general arrangement of Figure 11, this results in a total of 15 stopes and 10 pillars distributed in both seams. Additional modelling dimensions are spans of 45m down-dip and 60m along-strike, and a 10m pillar thickness.
Figure 12: Central Block visualisation.
Figure 13: Safety factor distribution at central stope walls.
Figure 11: Three-dimensional sub-level stoping model geometry. Various visualization modes are available with this software, as follows. Making use of the "on plane" and "within volumetric grid" visualisation tools, modelling results are briefly outlined below according to three sampling modes, which may termed: "Central Block", "Rib Pillar", and "Pillar & Adjacent Stopes". This geometry for the first mode is illustrated in Figure 12, as a central sampling block or plate that intersects the global 15-stope, 10-pillar model. For this configuration, the actual Factor of Safety ("strength factor") results as sampled within that block volume are shown in Figure 13.
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This last representation allows the analysis of the geomechanical behaviour of stope walls down-dip and along-strike. Specific results captured within this volume are displayed in terms of Safety Factor iso-surfaces inside this volume, but having "put out" all stopes in front of the central block (as well as the block itself), for ease of observation. Total and "zoom" view are included in the above figure. The depth of safety factors below limit equilibrium (FS < 1.0) extends to a maximum of 1.4m into the walls, and laterally for some 4m down-dip and 6.5m along-strike. The most unfavourable safety factor distribution regarding our design criterion (minimum Safety Factor of 1.25) corresponds to an iso-surface which extends some 2.7m into the walls, and laterally for 4.5m down-dip and 13m along-strike. If these design dimensions are to be adopted, appropriate support would be designed and installed to stabilise these volumes. Rib pillar visualisation is also available. This makes use of the "on plane" result display, which captures all results on a cutting plane passing through the middle of a given rib pillar, thus permitting an assessment of conditions in the pillar core, to either confirm or modify its dimensions. The total distribution of Safety Factor iso-contours for this case is shown in Figure 14a, whereas Figure 14b affords better
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visualisation of the same distribution, where the portion (half) of the model in front of the cutting plane has been removed.
based on pillar/stope interaction, are best conducted in this mode.
Figure 16: Safety Factor iso-surfaces inside pillar/stope sampling volume.
RESULTS
Figure 14: Safety Factor iso-contours through rib pillars. Detailed analysis of above results yields a more than adequate range of Safety Factors from 1.8 to 2.0, with the marginally lower values distributed in the central pillar. Rib pillar and adjacent stopes. This visualisation mode captures all results within a vertical plate volume incorporating the sampled rib pillars and about one third of the volume of stopes adjacent to either side, as depicted in Figure 15.
• Typically, rib pillars exhibit acceptable iso-surfaces of Safety Factor values in excess of 1.4 at very shallow depths into them. In turn, main pillars (orthogonal to rib pillars) show marginal instability in some sectors. • These findings raise the interesting potential for an optimised dimensioning of pillars, whereby the thickness of rib pillars may be moderately reduced, and that of main pillars modestly increased. • The above implies a discernible increase in the extraction ratio or overall recovery for this orebody, through a generalised reduction of rib pillar thickness from 10m to 7m. (In some sectors, this may require the "compensation" of increasing main pillar thickness from 15m to 17m to avoid the onset of limit equilibrium conditions). • Alternatively, rib pillar thickness may be left unchanged, with this "surplus" invested in a moderate increase of the along-strike dimension of the stope. • Sensitivity analyses for a broad range of mining sequences show a clear trend towards marginal stability with a further increase in extraction ratio, which suggests that the current mining method is economically efficient. Thus, extraction ratio is said to be close to a "geomechanical optimum". • Two optimised mining sequences emerge from this study. In sectors where the presence and effects of geological structures are more pronounced, the optimum sequence calls for more advanced mining of the lower seam, where the scatter of safety factor distribution is reduced (lower dilution is an added bonus). In sectors without such marked structural influence, the optimum sequence is one of similar and simultaneous advance in both seams, in which mining of the next stope alternates between them. • The results from these studies have important implications for the global stability conditions at Navío mine, as well as for its eventual closure plans.
Figure 15: Sampling volume for pillar/stope interaction. 4 CONCLUSIONS The full potential of this visualisation capability is displayed in Figure 16, with a three-dimensional view of all Safety Factor iso-surfaces contained within this sampling volume, but having removed that portion of the model located immediately in front of such volume, in order to "see through". Sensitivity analyses of pillar dimensioning, 278
The two case histories presented in this paper demonstrate the significant potential of geomechanics as an effective tool in the pursuit of long term mine planning objectives, and ultimately, in the economic optimisation of the mining business.
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In the open pit application, final recommendations reflected increased slope angles for both, working slopes and final pit, with the favourable consequences of: greater ore recovery, reduced stripping ratio, and the obvious reduction in production unit costs. These were not obtained at the expense of overall pit safety, but through a more realistic, reliable assessment of events which actually control pit slope behaviour (as opposed to the adoption of excessively theoretical formulations), and supported by practically oriented sensitivity analyses and the implementation of a monitoring system. For the underground mining case, geomechanics allowed the formulation of: an overall mine design optimisation, through a safe refinement of stope and pillar dimensions; a mining sequence capable of yielding the complex oreblending requirements of the plant; and a moderately increased ore recovery, such that the extraction ratio is now said to be close to a geomechanical optimum. ACKNOWLEDGEMENTS The authors gratefully acknowledge ANGLO AMERICAN (Chile), through Mr. Manuel Schellman, Sr. Geotechnical Engineer of the Mining Vice-Presidency, as well as LAFARGE CEMENT (Chile), through Mr. Sergio Navarrete, Manager of Mineral Resources Planning, for their kind permission to use information from their various geotechnical studies as required to prepare this paper.
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REFERENCES • Brady, B.H.G. & E.T. Brown 1993. Rock Mechanics for Underground Mining (Second Edition), 141-158, Chapman & Hall:London. • Brown, E.T. and Hoek, E. 1978. Trends in Relationships between Measured Rock In-Situ Stresses and Depth. Int. J. Rock Mech. Min. Sci. & Geomech. Abstr. 15, 211-215. • Hoek, E, 1994. Strength of rock and rock masses. ISRM News J, 2(2): 4-16. • Hoek, E, 2002. Rock engineering. Course notes available on line at http://www.rocscience.com. • Hoek, E., and Bray, J.W. 1981. Rock Slope Engineering (Revised Third Edition), The Institution of Mining and Metallurgy: London. • Navarrete, S.A. 1995. Evaluación Técnico- Económica del Método de Explotación Modificado "Sublevel Stoping" para Mina Navío (Internal report). • Soto, C.A., J.C. Cereceda & F.J. Orcaistegui 1999. Estudio Geotécnico de Refinamiento del Método de Explotación de Mina Navío. Report to Empresas Melon S.A.(Blue Circle Cement). • Soto, C.A., Cereceda, J.C., Orcaistegui, F.J., and Hartley, S. 2000. Estudio de Estabilidad de Taludes Pared Este Rajo Mantoverde. Report to Empresa Minera de Mantos Blancos S.A. (Anglo American Chile). • Soto, C.A. & F.J. Orcaistegui 1993. Estudio Diseño Geomecánico Mina Navío, Etapa I. Report to Cemento Melon S.A.
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"Extension" in large open pit slopes and possible consequences T. R. Stacey, School of Mining Engineering, University of the Witwatersrand, South Africa Yu Xianbin, Kunming University of Science and Technology, China and School of Mining Engineering, University of the Witwatersrand, South Africa
Abstract Research has recently been carried out into the occurrence of zones of extension strain in open pit slopes. This showed that very large zones of extension can occur, and this finding represents a significant new aspect in slope stability that has not been considered before. The greatest magnitudes of extension strain occur near the toe of the slope, either in the slope itself, or in the floor of the pit. Results of the research will be presented for different slope heights, slope angles and horizontal to vertical in situ stress ratios. The magnitudes of the strains are considered to be large enough to result in fracturing of intact rock, and the fracture orientations predicted are adverse for slope stability. Fracturing that is extension in nature is common in competent, brittle rocks and often develops with some violence and little or no warning, producing easily measurable seismicity. In the slope situation, the expected physical manifestation of this behaviour would be popping off of rock slabs and plates of rock from slope surfaces and popping up of the pit floor, as well as the formation of new fractures within the rock mass. Such behaviour may cause overall slope failure, or may initiate failure, which may then be driven to overall slope failure by other influencing factors or combinations of factors. In addition to instability resulting from the fracture surfaces themselves, all induced fracture surfaces could interact with natural geological structures to facilitate formation of a significant failure surface. 1. INTRODUCTION Many open pit mines are now being mined at depths in excess of 500m and many are being planned for much greater depths than these. Many existing pits were not originally planned to be mined to such depths, but this mining has been facilitated by developments that have taken place in the capabilities of mining equipment. In addition to the depth factor, open pit mining is taking place in areas in which the horizontal to vertical stress ratio is significantly in excess of unity. An example of this is Western Australia, where the horizontal to vertical stress ratio is commonly quoted to be about 3. Consequently, a 300m deep pit in this environment might well be considered equivalent to a 1000m deep pit in a "normal" stress environment. The result is that many current pits, and the future very deep pits, are being and will be mined in a high stress environment, which means that they might be in "new territory" as far as slope design is concerned. In such situations, it is important to consider the influence of the high stresses, and to assess whether new mechanisms of slope behaviour need to be considered. In the 1960’s and 1970’s considerable research was carried out into the stability of rock slopes in open pit mines. It was also during this period that early development of numerical analysis methods took place, but there was relatively little application of them to the evaluation of slope stability. Most of the work carried out concentrated on the use of limit equilibrium techniques, and there has been surprisingly little development in the technology of stability evaluation of open pit mine slopes in the 30 years since this period. There has been substantial development in the capabilities of numerical modelling packages, making them suitable for analysis of the stability of slopes, but there has been disappointingly little development in the understanding of the behaviour of rock slopes. The result is that, although numerical techniques are now commonly used in the evaluation of the stability of rock slopes, there are no recognized robust, standardized 280
approaches using these techniques. The lack of development in slope stability evaluation is further surprising since, as pointed out by Stacey (1993), many of the pits designed in the 1960’s and 1970’s were reaching their full depths after about 20 to 30 years. In such cases the stress levels around the pit could induce failure of the rock, and it is therefore appropriate that attention should be paid to the effects of stresses on the stability of the slopes of these mines. In this paper, a brief review of the application of stress and numerical analysis techniques to slope stability is presented. This is followed by research results from a programme of stress analyses of two-dimensional and axisymmetric open pit slopes, that are relevant to a new mechanism of behaviour of open pit slopes in hard and brittle rock. Such behaviour is potentially associated with the development of stress induced failure surfaces in which significant seismic activity might be involved. 2. REVIEW OF THE APPLICATION OF NUMERICAL STRESS ANALYSES IN EVALUATING SLOPE STABILITY After the development of the finite element method of stress analysis in the 1960’s, there were early applications of the method to slopes, (for example Duncan and Goodman, 1968; Gates, 1968; Mahtab and Goodman, 1970). Extensive investigations into the elastic stress distributions in two and three dimensional slopes were carried out by Stacey (1970, 1972, 1973). Factors taken into account in these investigations were the angle of inclination of the slopes, the floor width of the open pit, the horizontal in situ stress field, and the value of Poisson’s ratio. One of the important findings of this work was that a large horizontal in situ stress field has a major influence on the stress distributions in slopes. A further important observation from early stress analyses of slopes was the correspondence between failure zones and the presence of tensile stresses (Kalkani, 1976; Kalkani
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and Piteau, 1976; Lee, 1978; Valliappan and Evans, 1980; and Brown et al, 1980). Common application of numerical methods in the evaluation of the stability of rock slopes in recent times is apparent from numerous papers in a recent specialized publication (Hustrulid et al, 2000). A paper by Board et al (1996) represents a significant development in the application of stress analysis techniques to the evaluation of slope stability. In this paper, use is made of measured deformations to calibrate models prior to their use for prediction of slope behaviour into the future. Use was made of both continuum and discontinuum numerical analyses. It was found that the continuum approach could model the discontinuous rock mass successfully, and it was the preferred approach for computational reasons. The steps involved in the approach were: a) development of a conceptual model of the slope taking into account the geological structure and structural regions, the observed deformation behaviour, the measured deformations, failures that have occurred, etc; b) determination of the appropriate constitutive model of rock mass behaviour; c) determination of rock mass properties using rock mass classification; calibration of models – confirmation of the boundary and initial conditions, sensitivity studies to establish appropriate material properties, comparison of model results with observations and measurements, and modifications if necessary; d) application of the "final" model to determine behaviour as a result of future mining. The use of a calibration approach as summarized in the points above is considered to be a good engineering approach provided that extrapolation is not significantly into "new territory". However, satisfactory absolute prediction of behaviour is unlikely to be successful if the following are not used: the correct mechanisms, or combinations of different mechanisms, of behaviour; the correct strength parameters; the correct deformation parameters; and the correct failure criteria. Steps b) and c) above are unlikely to lead to successful prediction except in particular cases. As indicated by Stacey (2000), the use of a rock mass classification approach, which, as indicated above, is commonly used to estimate such parameters, is unlikely to be satisfactory for absolute prediction since all correlations are based on empirical data. Such correlations are "smear" correlations and cannot hope to provide data for successful absolute prediction. If successful predictions are obtained, they are likely to be by luck rather than by confident engineering. Absolute prediction of behaviour, which is very significantly involved in the analysis and design of new open pit mine slopes, requires a very thorough understanding of the mechanisms of deformation and behaviour of the slopes. In more recent application of numerical stress analysis methods to slope stability, Lorig (1999) and Zettler et al (1999) investigated numerically the effect of three dimensions on slope stability. Eberhardt et al (2004) modelled the progressive development of failure in a rock slope using a hybrid finite/discrete element code. From the above it may be concluded that development in the understanding of the behaviour of rock slopes has been very disappointing. Recognising the importance of deformations in rock slopes, it is perhaps surprising that no attention appears to have been given to strains in slopes, as opposed to stresses.
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3. RECENT MODELLING OF OPEN PIT MINE SLOPES A programme of two dimensional (plane strain) and axisymmetric stress analyses of various slope geometries has recently been carried out using finite element analysis. The interpretation of the results concentrated on the strain in the slopes, and this has identified a new and potentially very significant factor in the consideration of rock slope stability in open pit mines – the occurrence of zones of extension strain. In this context, the extension strain is defined as the minimum principal strain _3 (in a compression positive convention) and can be calculated from the principal stresses using the three dimensional elastic equation:
ε3 = [σ3 – _(σ1 + σ2)]/E where σ1, σ2, σ3 are the three principal stresses ν is Poisson’s ratio E is the modulus of elasticity
ε3 depends on all three principal stresses and can be extension in nature, even in a triaxially compressive stress field. It is clear therefore that it will depend on the magnitude of the out-of-plane stress used in the two dimensional stress analysis. The results for the two dimensional analyses presented in this paper have all assumed that the out-of-plane stress corresponds with a plane strain condition. The occurrence of a zone of extension strain around an open pit implies that the rock mass has "expanded" in at least one direction in that zone. It is clear from the above equation that, the smaller the value of E, and the larger the value of Poisson’s ratio, the greater will be the magnitude of extension strain. In the programme of analyses, only elastic analyses were carried out, and the following parameters were taken into account: • Pit depths: pit depths of 400m, 800m and 1200m were considered. Pit geometries, at an intermediate stage of mining, were modelled to determine their influence on the stress and strain distributions. The pit geometries and the excavation steps are illustrated in Figure 1 for one of the slope angles modelled. • Pit slope angles: slope angles of 30o, 40o, 45o, 50o, 55o, 60o, 70o and 75o were analysed in most cases. • Horizontal to vertical in situ stress ratios (k ratios): values of 0.5, 1.0, 2.0, 3.0 and 4.0 were used in most cases. • A value of 0.17 for Poisson’s ratio, and a modulus of elasticity of 80 GPa, were used for all analyses presented in this paper. For elastic analyses, the stress results are independent of the modulus of elasticity used, but the strain results are directly proportional to the value of the modulus. The results of this programme of analyses have been dealt with in detail by Stacey et al (2003) and therefore only selected results will be presented here. Although, as indicated in the Section above, stresses in slopes have been considered over the years, it does not appear that any consideration has previously been given to the occurrence of strains in slopes.
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• significant zones of extension occur in the crest of the slope only for higher k ratios – there is no significant zone of extension strain for k values of 0.5 and 1.0.
Figure 1 Pit geometries and stages of excavation The results reported below, dealing in particular with extension strains, therefore represents a new development in the consideration of open pit mine slope stability. Extension strains around two dimensional slopes The results of the analyses show that substantial zones of extension strain occur around the open pit for all in situ stress conditions and for all slope angles. Figure 2 illustrates the extent of the extension strain zones for a two dimensional slope with an angle of 60o, a height of 800m, under a k ratio of 2.0. From this figure it can be seen that the maximum magnitudes of extension strain occur at the toe of the slope, either in the slope or in the floor of the pit. From the programme of analyses, the following were found: • the magnitudes of extension strain are greater for higher values of k; • the zones of extension strain are larger around the bowl of the pit for flatter slope angles and higher k ratios;
The sizes of the zones of extension can be very substantial, of the order of many hundreds of metres behind the slope face and beneath the pit floor, depending mainly on the k ratio and the pit slope geometry, and have been quantified by their horizontal and vertical extents (in m/100m of slope height). The horizontal extent is represented by the horizontal distance between the toe of the slope and the boundary of the zone. The vertical extent is represented by the vertical distance between the floor of the pit and the boundary of the zone, measured along the vertical axis of symmetry. The results are shown in Figures 3 and 4, illustrating the effects of both the k ratio and the slope angle. In fact the extents of the zones in these figures are measured to where the extension strain exceeds a value of 0.00001, since this is better defined than the point of zero extension strain. The actual zones of extension are therefore slightly greater than indicated. It can be seen from these figures that the zones of extension can be very large, of the order of many hundreds of metres behind the slope face and beneath the pit floor, depending mainly on the k ratio and the pit slope geometry. Extension strains around axisymmetric slopes For the axisymmetric analyses, the result corresponding with Figure 2 for the plane slope is shown in Figure 5. It can be seen that the strains are not as concentrated at the toe as in the two dimensional plane case. The extents of the zones are slightly smaller than those for the plane slopes, but are still very substantial. As for the plane slopes, the sizes of the zones of extension strain can be very substantial. The quantified extents of the extension zones for axisymmetric slopes, corresponding with Figures 3 and 4 above for two dimensional slopes, illustrating the effects of k ratio and slope angle, are given in Figures 6 & 7.
Figure 2 Extension strain distribution for a two dimensional slope, k = 2.0 282
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Figure 3 Horizontal extents of extension strain zones for two dimensional slopes
Figure 4 Vertical extents of extension strain zones for two dimensional slopes
Implications of the occurrence of extension strains on the behaviour of open pit rock slopes These results of the analyses have shown that very large extension strain zones will develop around the slopes and floors of open pits. It is to be noted that the geometry of the pit in plan (such as the occurrence of convex slopes, resulting in "noses" in the pit) will have a significant influence on the occurrence of extension strains.
correspondingly, the magnitudes of the extension strains will also increase. For a 400m deep pit, the maximum extension strain magnitudes calculated vary from about 0.00001 for k of 0.5 to about 0.0001 for a k of 2.0. For a 1200m deep pit the corresponding strain values are about 0.0001 and about 0.0003. In terms of the extension strain criterion for fracture of brittle rock (Stacey, 1981), critical extension strain levels at which fracturing can occur are in the 0.0001 to 0.0003 range, in particular for hard, brittle rock. The stress conditions are therefore conducive for the development of extension fracturing around and adjacent to the toes of slopes (in the pit floor and in the slope face area). From the analyses carried out, these conditions could apply for 400m high slopes when the k ratio is about 2 or greater. They would apply for 1200m high slopes under any k ratio. It can therefore be concluded that failure of the rock material and rock mass will occur, and that this might have a significant negative effect on stability of the slope.
Magnitudes of extension strains The magnitude of the extension strain calculated is directly proportional to the modulus of elasticity used in the analysis. The results presented above are for a modulus of elasticity of 80 GPa. If the modulus of elasticity was 40 GPa instead of 80 GPa, for example, these strains would double. Note, however, that the extents of the zones of extension are not influenced by the modulus. As the height of the slope increases, and as the k ratio increases, the stress levels will increase and,
Figure 5 Extension strain distribution for an axisymmetric slope, k = 2.0 Massmin 2004
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Figure 6 Horizontal extents of extension strain zones for axisymmetric slopes
Figure 8 Probable fracture orientations
loosening effects of blasting. The further implications are that there may be limits to the depths to which open pit operations can progress without the occurrence of stress induced instability. It is to be noted that the type of fracturing likely to develop will be in a brittle rock environment in which "stiff" rock conditions exist. The analyses described above have made use of an elastic modulus of 80 GPa and a Poisson’s ratio of 0.17. These might be considered as high and low respectively for a rock mass. Lower modulus values and higher Poisson’s ratio values will result in greater extension strain magnitudes. Hence the interpretation of the results above probably represents a conservative approach.
Figure 7 Vertical extents of extension strain zones for axisymmetric slopes Orientations of potential fracture surfaces The orientations of extension fractures will be normal to the minimum principal stress. In the proximity of excavation surfaces, the direction of the minimum principal stress is normal to the surface and hence extension fractures will be subparallel to the excavation surface. In Figure 8, an interpretation is given of the probable orientations of fractures in the toe and slope face regions. Therefore inclined fractures are likely to develop in any toe region and slope face parallel fractures are likely to develop behind the face, away from the toe. These orientations are likely to be adverse for stability of the slope – the inclined surfaces will facilitate sliding out from the face and the face parallel surfaces will promote the formation of "columns" or plates. These could be involved in slope failure due to sliding out, or buckling out. From the analyses carried out, the extent of the zone of potential fracturing is very substantial. For example, for an 800m high, axisymmetric 60o slope (k ratio of 2.0), fracturing could occur to a horizontal depth of about 50m into the rock mass behind the toe region and over a height of between 50m and 100m up the slope face. In comparison, for a 1200m pit, fracturing could occur to a horizontal depth of more than 100m behind the toe of the slope, and over a height of about 400m up the slope face. Similar extents apply, for example, for a 45o slope, 1200m high. This illustrates the influence of the depth of the open pit. The implication from this is that adversely orientated fracture planes can develop in the tightly confined rock mass beyond the surficial zone of blast damage, and the potential for fracturing damage is unlikely to be inhibited by the 284
Mode of fracture and failure It is likely that, at greater mining depths, rocks will be more competent and less weathered. Fracturing that is extension in nature is more common in more competent, brittle rocks and often develops with some violence and little or no warning. The mechanism involved would be a form of "strain bursting" which produces easily measurable seismicity, events often being audible as well. In the slope situation, the expected physical manifestation of this behaviour would be popping off of rock slabs and plates of rock from slope surfaces and popping up of the pit floor, as well as the formation of new fractures within the rock mass. There is likely to be some time dependency in the development of fractures. Since the size of the open pit increases slowly, there is an abundance of time for the stresses around the pit to readjust and to interact with the rock material and rock mass, promoting the development of fractures. Such behaviour may cause overall slope failure, or may initiate failure, which may then be driven to overall slope failure by other influencing factors or combinations of factors. In addition to instability resulting from the fracture surfaces themselves, all induced fracture surfaces could interact with natural geological structures (joints, bedding planes, faults, etc) to facilitate formation of a significant failure surface and to reduce the stability of the slope. Extension strains are likely to manifest themselves in the opening up of adversely orientated natural structures and hence in the loosening of the rock mass in a preferential orientation. Influence on groundwater conditions The presence of zones of extension will lead to expansion of the rock mass in a preferred direction. This, the formation of new extension fracture surfaces, the opening up of adversely orientated natural structures, and the interaction of these effects with other natural geological structures, will express itself as a zone of relaxation, in a preferential direction, around the pit. This is likely to have a significant influence on the permeability of the rock mass. This may facilitate the entry of groundwater
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into the rock mass, may allow channelling of water flow in particular directions, and may inhibit flow in other directions. Since groundwater pressures are a very important contributor to the instability of slopes, some of these effects may be detrimental to stability of the slopes. 4. CONCLUSIONS From the research programme of stress analyses and the observations from the case studies, the following conclusions can be drawn: • relatively limited zones of tensile stress occur behind the crest of slopes for in situ stress conditions with k ratios greater than 1.0, and in the pit floor for low k values; • in contrast with the limited occurrence of tensile stresses, very large zones of extension strain can develop around the slopes. The occurrence of extension strains in open pit slopes represents a new, and potentially very important, consideration with regard to slope stability; • the probable orientations of fracture surfaces, adversely inclined out of the slope near the toe and parallel to the slope face away from the toe, are such that they could have a significantly negative influence on the stability of slopes. The geometry of the fracturing could lead to the formation of slabs parallel to the slope face; • the fracture surfaces will probably also provide surfaces that can interact with, or combine with, natural geological structures to form potential failure surfaces within the slope; • fracturing that is extension in nature usually occurs with some violence and suddenness, with little or no warning. There is usually easily measurable seismicity associated with these "strain burst" type of events; • the zones of extension in the slopes represent possible locations of changed rock mass permeability and preferential groundwater flow. REFERENCES • Board, M, Chacon, E, Varona, P and Lorig, L, 1996. Comparative analysis of toppling behaviour at Chuquicamata open-pit mine, Chile, Trans. Instn Min. Metall. (Section A: Min. Industry), 105:A1-A21. • Brown, I, Hittinger, M and Goodman, R, 1980. Finite element analysis of the Nevis Bluff (New Zealand) rock slope failure, Rock Mech., 12:231-245. • Duncan, J M and Goodman, R E, 1968. Finite element analysis of slopes in jointed rock, Contract Report, U S Army Engineer Waterways Experiment Station, Corps of Engineers, No S-68-3:274p. • Eberhardt, E, Stead, D and Coggan, J S, 2004, Numerical analysis of initiation and progressive failure in natural rock slopes – the 1991 Randa rockslide, Int. J. Rock Mech. Min. Sci., 41, pp 69-87. • Gates, R H, 1968. Inelastic analysis of slopes by the finite element method, PhD Thesis (unpublished), University of Illinois, 123p. • Hustrulid, W A, McCarter, M K and Van Zyl, D J A (eds), 2000. Slope Stability in Surface Mining, SME, Colorado, 442p.
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• Kalkani, E C, 1976. Stability analysis of rock slopes, Water Power & Dam Construction, September:47-49. • Kalkani, E C and Piteau, D R, 1976. Finite element analysis of toppling failure at Hell’s Gate Bluffs, British Columbia, Bull. Ass. Engng Geol., 13(4):315-327. • Lee, C F, 1978. Stress relief and cliff stability at a power station near Niagara Falls, Engng Geol., 12:193-204. • Lorig, L 1999. Lessons learned from slope stability studies, in FLAC and Numerical Modeling in Geomechanics, Detournay & Hart (eds), Balkema, Rotterdam, pp 17-21. • Mahtab, M A and Goodman, R E, 1970. Threedimensional finite element analysis of jointed rock slopes, Proc. 2nd Int. Cong. Int. Soc. Rock Mech., Belgrade, 3(Theme 7). Paper No 12. • Stacey, P F, 1993. Pit slope designs for the 21st century, in Innovative Mine Design for the 21st Century, (Ed Bawden and Archibald), Balkema, Rotterdam, pp 3-11. • Stacey, T R, 1970. The stresses surrounding open pit mine slopes, in Planning Open Pit Mines, (Ed P W J van Rensburg), A A Balkema, pp 199-207. • Stacey, T R, 1972. Three-dimensional finite element stress analysis applied to two problems in rock mechanics, Jl S. Afr. Inst. Min. Metall., 72(10):251-256. • Stacey T R, 1973. A three-dimensional consideration of the stresses surrounding open pit mine slopes, Int. J. Rock Mech. Min. Sci., 10:523-533. • Stacey, T R, 1981. A simple extension strain criterion for fracture of brittle rock, Int. J. Rock Mech. Min. Sci., 18:469-474. • Stacey, T R, 2000. Reservations regarding the use of rock mass classifications in rock engineering, Proc. Bergmekanikdag 2000, SveBeFo, Swedish Rock Engineering Research and National Group ISRM, pp 118. • Stacey, T R, Yu Xianbin, Armstrong, R and Keyter, G J, 2003. New slope stability considerations for deep open pit mines, Jl S. Afr. Inst. Min. Metall., 103(6):373-389. • Valliappan, S and Evans, R S, 1980. Finite element analysis of a slope at Illawarra Escarpment, Proc. 3rd Australia-New Zealand Conf. On Geomechanics, Wellington, 2: 241-246. • Zettler, A H, Poisel, R, Roth, W and Preh, A 1999. Slope stability analysis based on the shear reduction technique in 3D, in FLAC and Numerical Modeling in Geomechanics, Detournay & Hart (eds), Balkema, Rotterdam, pp 11-16. • Amadei, B and Stephansson, O, 1997. Rock Stress and its Measurement, 515 p. Kluwer: Dordrecht. • Bandis, S, 1990. Mechanical properties of rock joints. Proceedings International Symposium on Rock Joints, Löen, Norway, (Ed: N Barton and O Stephansson), 125140. Balkema: Rotterdam. • Bandis, S, 1993. Engineering properties and characterization of rock discontinuities. Comprehensive Rock Engineering, (Ed: J A Hudson, E T Brown, C Fairhurst and E Hoek), 1: 155-183. Pergamon Press: Oxford.
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Visual estimation of fragment size distributions in the DOZ block cave A. Srikant, Rock Mechanics Engineer, David E. Nicholas, President, Call & Nicholas, Inc., Tucson, Arizona, USA Lufi Rachmad, Staff Geotechnical Engineer, P.T. Freeport Indonesia, New Orleans, Louisiana, USA
Abstract Predicting fragmentation distributions before the start of block caving is an important planning exercise that helps in the production scheduling and in the selection of equipment for the ore handling system. Though several methods have been developed for the assessment of fragment size distribution in block cave operations, limited efforts have been made for developing correlations between the predicted and observed fragmentation. At P.T. Freeport Indonesia’s DOZ block cave operation, a procedure was developed for assessing the fragmentation at the drawpoints. The observed fragmentation was tabulated along with the estimated height of draw, and the fragment size distributions were then compared with the predicted fragmentation at different draw heights. The results show that while the median fragment size for the predicted and observed fragmentation were similar, the fragmentation measured at the drawpoint included more fines and fewer large blocks. This paper presents the methodology used for assessing the observed fragmentation and shows the results of the correlation studies.
1 INTRODUCTION The mines of the Ertsberg District, in the province of West Papua in Indonesia, are operated by P.T. Freeport Indonesia (PTFI) under contract to the Republic of Indonesia. The Deep Ore Zone (DOZ) mine is a part of the Ertsberg District of Papua, Indonesia.
benefits from the previous experience gained while mining the GBT and the IOZ in the same skarn system. Four major rock types are mined in the DOZ: 1. Ertsberg Diorite 2. Forsterite Skarn 3. Magnetite Skarn 4. DOZ Breccia Fragmentation characteristics for these rock types vary from fine (in the DOZ Breccia) to coarse (in the Diorite). In order to help select the most appropriate production and ore-flow equipment and to assess the productivity from the mine, fragmentation estimates were developed using different available methods. A drawpoint fragmentation mapping scheme was developed to assess the current fragmentation and to correlate between the observed and predicted fragmentation. This paper discusses the development of the fragmentation estimates in the different rock types, the process and results of drawpoint fragmentation mapping, and the correlations between the observed and predicted fragmentation distributions. 2 CHALLENGES IN ESTIMATING FRAGMENTATION IN THE DOZ
Figure 1. Location of PTFI’s Mining Operations The DOZ is a copper-gold deposit found on the northeast flank of the Ertsberg diorite. It comprises the lower elevations of the East Ertsberg Skarn System (EESS). The DOZ mine is an LHD block cave mine using eightcubic-yard loaders at the extraction level from where the broken rock is transferred through ore passes feeding 55ton trucks, which then dump directly into a 54-inch gyratory crusher (Barber et al, 2000). The DOZ is essentially the third lift of the block cave mine that has exploited the East Ertsberg Skarn complex, and 286
At the time of the feasibility study and preparation of the report, very few excavations were available in the DOZ undercut and extraction levels for the collection of the geotechnical information required for development of fragmentation estimates. Three panel drifts at the extraction level and three drill drifts at the undercut level were mapped for estimating the rock-mass and joint-set characteristics in the four major rock types expected in the DOZ. The Block Cave Fragmentation (BCF) program was used to develop primary and secondary fragmentation estimates for the DOZ mine.
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The different rock types in the DOZ exhibit vastly different fragmentation characteristics, from the fine fragmentation in the DOZ Breccia and Magnetite Skarn to the coarse fragmentation in the Diorite. Based on an examination of the drill-core piece lengths from different areas of the DOZ, the fragment size distributions were expected to show a significant variation along strike and by elevation, even within the same rock type. In the absence of joint-set characteristics from all areas of the DOZ, the variation of fragmentation could not be simulated using the available estimation tools like BCF. Since the drill-core data from the DOZ showed significant variation in piece lengths in different parts of the deposit, Call & Nicholas, Inc. (CNI) and PTFI concluded that the drillcore data should be used to help differentiate between the fragmentation characteristics in the different areas of the DOZ. CNI developed a program (Core2Frag) for estimating the primary fragment size distributions based on the core piece length data collected by PTFI from the drill holes in the DOZ. The BCF program was used for the development of secondary fragmentation estimates. 3 FRAGMENTATION ESTIMATES IN THE DOZ The Core2Frag program developed by CNI uses the core piece length information and the joint-set characteristics for estimating the primary fragmentation (Srikant et al, 2004). The drill-hole data, sorted into domains of three panels each along strike, and the joint-set characteristics of four rock types (Forsterite Skarn, Forsterite-Magnetite Skarn, Magnetite Skarn, Diorite) were used to estimate the primary fragmentation. Table 1 shows the structural data used for the Diorite and the Mixed Skarns. (There was little difference in the joint-set information for the different skarns, so the data for all the skarns was combined). Secondary fragmentation estimates and hang-up frequencies were determined using the BCF program (Esterhuizen, 1999). The BCF program uses the primary blocks generated by the Core2Frag program and the caving geometry to estimate the breakage of the primary blocks as they travel through the draw column. The primary and secondary fragmentation estimates for the Forsterite Skarn for Panels 12-15 in the DOZ are shown on Figure 2 (Pratt et al, 2002). The estimates show that about 57 percent of the tons in the first 60 meters of the draw in Panels 12-15 would be larger than two cubic meters. As the draw height increases, the percentage of tons that require secondary breakage reduces, with 20 percent of the tons being larger than two cubic meters at a draw height of 240 meters. In the Diorite, the fragmentation estimates show that about 80 percent of the tons in the first 60 meters of the draw in Panels 12-15 would be larger than two cubic meters, and about 68 percent of the tons would be larger than two cubic meters at a draw height of 240 meters (Figure 3).
Figure 2. Primary and Secondary Fragmentation Estimates for Forsterite Skarn in Panels 12-15 Massmin 2004
Figure 3. Primary and Secondary Fragmentation Estimates for Diorite in Panels 12-15 The fragmentation estimates developed for the DOZ showed that some areas of the DOZ would have significantly coarse fragmentation. Since the fragment size distribution would have an impact on the operations in the DOZ mine, PTFI initiated a program for monitoring the fragmentation at the drawpoint. This program helped to calibrate the fragmentation estimates developed by CNI and develop more accurate estimates for areas that will be undercut at a later stage in the mining of the DOZ block cave. 4 DRAWPOINT MAPPING The development of digital imaging methods has introduced the possibility of correlating some of the fragmentation estimates with direct observations in the field (Girdner et al, 1996). However, the measurement of fragmentation at the drawpoint gives information regarding the secondary fragmentation only. It is also very difficult to take photographs in the wet and dusty areas in the underground and the photographs thus taken are often not suitable for fragmentation analyses using digital imaging methods. Efforts are underway to develop systems for the field assessment of fragmentation at the drawpoints using photographs as well as visual size distribution assessments. A procedure for the DOZ fragmentation visual data collection, called drawpoint mapping, was developed to assess the actual fragmentation size distribution and to provide feedback for the fragmentation prediction tools. This procedure satisfied the following requirements: it is simple, repeatable, and creates no significant interruption to the production. The qualitative information was collected by developing a rating system while the quantitative information was collected by estimating the percentages of particular material sizes in the drawpoint. Mapping of fragmentation at the drawpoints in the DOZ was undertaken from mid-2002. The available data were provided to CNI for analysis up to November 2002. CNI developed correlations to help assess the fragmentation in the unmined areas of the DOZ and the ESZ block caves. The rating system (Figure 4) was based on three categories: (1) drawpoint productivity, (2) hang-up condition, and (3) general block sizes. This qualitative information would provide a general overview of fragmentation condition. In addition, this information could be used for supporting the quantitative information. The quantitative information regarding the material size distribution was collected using the categories shown on Figure 5. The material size distribution was divided into five categories: fines, small block, intermediate block, large block, and oversize.
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appropriate. The oversize and fines were estimated, and the small blocks category was calculated to make up the rest. The draw height information for each drawpoint, which was important for the analysis, was collected through the Cave Management System (CMS). Rock type information was collected through geological mapping at the drawpoints. Currently, there are ten production panels and about 120 active drawpoints in the DOZ. Drawpoint mapping is conducted at least once every two weeks. Table 1 shows the number of observations for each fragmentation domain by rock type and height of draw. The table shows that there is limited data for the Diorite and the Forsterite. Since the fragmentation characteristics of the Forsterite and the Forsterite-Magnetite are similar, these data were combined to provide a greater database for comparison and correlation. Also, though PTFI geologists had classified the DOZ breccia as HALO-1, HALO-2, or HALO-4, during the mapping of the drawpoint muck piles, CNI estimated the fragmentation characteristics for all types of DOZ Breccia as one unit. Table 1. Number of Drawpoint Mapping Observations in the DOZ Rock type
Forsterite
Forsterite-Magnetite
Magnetite
Diorite
Height of Draw 0-60 60-120 120-180 0-60 60-120 120-180 0-60 60-120 120-180 Panels 12-15 Panels 15-18 Panels 18-21 All Panels
Figure 4. Freeport’s DOZ Drawpoint Rating System
The first three size categories represent the material that could pass through the ore pass grizzly. The "large blocks" category represents the material that could be handled by LHDs without any material size reduction required. The "oversize" category represents the material that requires either secondary blasting at the draw point or hang-up blasting. In addition, the length and width of the largest block was recorded.
Figure 5. DOZ Drawpoint Fragmentation Log The main concern for the mine was the large or coarse blocks; however, fines information was also important for understanding the compaction issue inside the transfer raise and for analyzing the effectiveness of the secondary ore handling system. In the original version of the drawpoint log, the observer was required to estimate the percentage of each category. In the later version, counting the numbers of intermediate and large blocks and multiplying those by proportional areabased percentages was found to be easier and more 288
16 12 1 29
21 4 0 25
7 0 0 7
125 64 26 215
87 27 2 116
5 0 0 5
38 18 8 64
29 12 1 42
0 0 0 0
DOZ Breccia
0-60 60-120 120-180 0-60 60-120 120-180 7 0 0 7
0 0 0 0
0 0 0 0
9 4 5 18
101 19 3 123
22 2 0 24
The observations showed variations due to rock type, and are discussed herein. Forsterite: The fragmentation distributions observed in the Forsterite do not show any definite pattern. In panels 12-15, the fragmentation was similar for heights of draw of 0-60 meters and 60-120 meters, with 50 percent of the total volume having a fragment size greater than eight cubic meters. However, the fragmentation for heights of draw of 120-180 meters showed some larger blocks, which made the fragmentation coarser, with 50 percent of the volume having fragment size greater than 100 cubic meters. In panels 15-18, the fragmentation for 0-60 meters of draw is slightly coarser than that for 60-120 meter draw. In panels 18-21, there were no data for draw heights greater than 60 meters. The distribution showed a maximum fragment size of 0.1 cubic meters. The percentage of volume that has blocks greater than two cubic meters for different heights of draw is shown in Table 2. Forsterite-Magnetite: The fragmentation distributions for the ForsteriteMagnetite show that the blocks undergo size reduction as the height of draw increases. The maximum size reduction takes place after 120 meters of draw. Again, the percentage of volume that has blocks greater than two cubic meters for different heights of draw is shown in Table 2. Magnetite: The observed fragmentation in the Magnetite shows some size reduction for the 0-60 and 60-120 meters of height-of-draw (HOD). However, the extent of size reduction is significantly greater in panels 18-21. In panels 12-15, the magnetite was observed to be coarser than in the other panels. Table 2 shows the percentage of volume that has blocks greater than two cubic meters. DOZ Breccia: Observations in the DOZ Breccia show that the material undergoes major size reduction after 60 meters of draw and
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does not break down significantly beyond 120 meters of draw. The percentage of volume that has blocks greater than two cubic meters is shown in Table 2. Diorite: There was very little fragmentation mapping data available from the drawpoints in the Diorite because few drawpoints were opened in the Diorite at the time of this analysis. However, the data collected by PTFI since November 2002 will help better calibrate the fragmentation estimates in this rock type. Table 2. Percentage Volume Greater Than 2 Cubic Meters by Rock Type and Draw Height Rock type
Forsterite
Height of Draw
0-60
Panels 12-15 Panels 15-18 Panels 18-21
28 72 100
Forsterite-Magnetite
60-120 120-180 27 76 -
26 -
Magnetite
0-60 60-120 120-180
0-60
78 74 65
41 87 80
84 77 71
99 99 -
DOZ Breccia
60-120 120-180 46 85 100
-
0-60 72 87 90
60-120 120-180 94 93 88
95 95 -
5 CALIBRATING THE RESULTS OF CORE2FRAG USING DRAWPOINT MAPPING The observations at the drawpoints at the DOZ provided a unique opportunity to compare the predicted and observed fragmentation for the different rock types for several heights of draw. Drawpoint fragmentation information for a particular "fragmentation domain" was compared to the predicted fragmentation for the same fragmentation domain, for the same rock type and height of draw. Observed and predicted fragmentation distributions were compared only for those drawpoints that had at least 80 percent of one rock type. The observed and predicted fragmentation distributions in Panels 12-15, 15-18, and 18-21 were compared for the Forsterite/Forsterite-Magnetite, Magnetite, and DOZ Breccia for HOD of 0-60 meters, 60-120 meters, and 120180 meters. The data were plotted in several ways to understand the relationships between the two distributions. The details of the comparison and the inferences therefrom are discussed in this section. The data collection procedure for drawpoint mapping was designed so that the PTFI geotech engineers could make a quick assessment of the block sizes and estimate the percentage in each of the categories. The procedure was based on block side lengths, because it is easier for the engineers to estimate lengths rather than areas or volumes. However, the fragmentation distributions from Core2Frag are based on block volumes. In order to compare the two sets of information, CNI converted the linear fragment size distributions to volumetric distributions assuming that the blocks had broken down to a stable aspect ratio of 1:1.2:1.2. The dimensions of the largest block observed in the draw point were used to estimate its volume. The fragment size distribution was then adjusted accordingly. The height of draw for each draw point was recorded based on the month-end HOD computed from the tons drawn from the drawpoint as reported by the PTFI Cave Management System. These data were then sorted into the HOD Sectors used by CNI for comparison to the predicted fragmentation. The fragmentation distributions observed at the various drawpoints were sorted by rock type and domain. The predicted and observed fragmentation distributions in the Forsterite/Forsterite-Magnetite Skarn are plotted on Figure 6. The observed fragmentation distribution represents the fragmentation at the surface of the muck pile in the draw points (Srikant and Nicholas, 2003).
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Figure 6. Predicted and Observed Fragment Size Distributions for Forsterite/Forsterite-Magnetite Skarn in Panels 12-15. Calibration of the fragmentation estimates in the Diorite could not be performed in a similar way because of the limited number of observations. Furthermore, the predicted fragmentation within the Diorite was observed to be about an order of magnitude greater than that in the Forsterite and Forsterite-Magnetite Skarns. A methodology was therefore developed for calibrating the estimates in the Diorite based on the observations in the Forsterite and ForsteriteMagnetite Skarns by correlating the observed percentages and the percentages predicted by Core2Frag for the different draw heights. This methodology will be refined as more observations are collected in the drawpoints in the Diorite. 6 CONCLUSIONS Based on this study, the following conclusions are made regarding the correlations between the predicted and observed fragment size distributions: 1. In general, the percent of the total volume greater than 12 cubic meters is similar in both the observed and predicted fragmentation. The 80th percentile block volumes in the predicted and observed fragment size distributions are similar. 2. The procedures for estimating fragmentation underestimate the size fragments below 0.1 cubic meters in volume. 3. The observed and predicted fragment size distributions show similar relationships to the height of draw. As the draw height increases, more of the smaller blocks are created. 4. The correlation between the two distributions appears to be related to rock type and may be dependent on rock properties. 5. Additional data is required for generating the correlation for the Diorite. The process of calibration carried out in the DOZ is limited in its extent and applicability because of the small number of observations in the database. Ideally, it would have been more effective to have several observations in different rock types, in different draw height categories, so that a rationale for calibration could be developed. It is therefore important that the process of data collection be continued so that more reliable correlations can be developed between the predicted and observed fragment size distributions. The following problems were noted in the data collection and correlation process:
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1. The observations are biased towards the fines as the boulders may be covered with fine gained material and may not be visible during mapping of the muck pile surface. 2. The mapping of the drawpoints does not include volumes of the broken material, thus necessitating the assumptions regarding the aspect ratio and shape of the blocks. 3. Adequate data were not available in all rock types for all draw heights. The applicability of the correlation is thus limited to the draw heights and rock types in which sufficient data are available. However, the method-ology for the breaking of the blocks in the draw column can be developed based on the observations. 4. The drawpoint fragmentation mapping was done on a daily basis. However, it is a well documented fact that unless the material is known to be well graded, the distribution of fragment sizes within the muck pile at the drawpoint is best assessed by taking several measurements while the muck is being drawn out of the drawpoint. As a result, the observed fragmentation is an instantaneous measurement and may not be representative of the fragmentation at that height of draw. In order to improve the results of the Core2Frag program and provide better estimates of the expected fragmentation, the authors suggest that the following future work be undertaken: 1. Continue data collection in the different rock types so that an adequate database of observations is created for correlation to the predicted fragment size distributions. 2. Improve the data collection process through digital photography of the muck pile at the drawpoints. 3. Compare observed fragmentation at the drawpoint with fragmentation measured by sieving to establish correlations between observed 2D and measured 3D volumetric fragment size distributions.
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4. Assess other operational parameters that are affected by the changing fragmentation so that correlations between Core2Frag results and parameters such as secondary blasting requirements, loader productivity, etc. can be developed. 5. Perform similar analyses in other block caves and other mining operations to evaluate the Core2Frag program and the process of calibration. 7 ACKNOWLEDGEMENTS The authors thank PTFI for permission to publish this paper. The assistance of geotech engineers at PTFI in the collection of the fragmentation data from the DOZ mine is also gratefully acknowledged. 8 REFERENCES • Barber, J., Thomas, L., and Casten, T., "Freeport Indonesia’s Deep Ore Zone Mine", Proc. MassMin 2000, AusIMM, Brisbane, November 2000. • Esterhuizen, G.S., BCF Version 3.0 – A Program to Predict Block Cave Fragmentation - Technical Reference and User’s Guide, Littleton, Colorado, 1999. • Girdner, K.K., Kemeny, J.M., Srikant, A., and McGill, R., "The Split system for analyzing the size distribution of fragmented rock", Measurement of Blast Fragmentation (eds. John Franklin et al), A.A. Balkema, Rotterdam, 1996. • Pratt, R.W., Srikant, A., Nicholas, D.E. and Flint, D.C., Analysis of DOZ Fragmentation, CNI Report, May 2002. • Srikant, A. and Nicholas, D.E., Calibration of Fragmentation in the DOZ Block Cave, CNI Report, September 2003. • Srikant, A. and Nicholas, D.E., "Assessment of Primary Fragmentation from Drill Core Data", Proc. MassMin 2004, Santiago, August 2004.
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Chapter 7
Ore Pass Design
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Experiments regarding grain pressure in silos H.A. Janssen, Engineer in Bremen (Translated from German by William Hustrulid, University of Utah, Salt Lake City, Utah and Norbert Krauland, Consultant, Boliden, Sweden)
Abstract Translators Note: One of the "classic" papers referred to in many subsequent papers dealing with the flow of granular materials is the paper by H.A. Janssen published in 1895 (Janssen, 1895). The paper was one of the first, if not the first, to recognize that the full weight of the contents of a bin/ore pass are not carried on the bottom of the bin but much of the weight is actually transferred to the walls. The paper included the development of the governing equations as well as the experiments performed. Use of the hydraulic radius concept in bin/ore pass design most certainly originated based on some of the contained ideas. In spite of the importance of the paper, the translators were unable to find a published or even an unpublished version in English. This would suggest that those citing the reference were either fluent in German or, more likely, simply included it for completeness sake. It is felt that this paper might be of interest to others trying to understand and explain granular flow associated with mass mining systems.The division of the paper into sections was done by the authors to conform to the present publication style
1 INTRODUCTION In the past decades, the grain production (export) from the grain-producing countries of the earth to Europe’s civilized countries has made sensational advances. Considering the lengths of the required transport paths, this development could only be achieved by keeping production costs low. These are due to low wages, the extensive use of machines both for the preparation of the soil and for gathering the harvest as well as the means for grain storage and transportation. With respect to low costs, in the United States of America the development of large-scale grain elevators (storage bins) has been essential (Arndt, 1892). The purpose of these grain storage facilities is to store the grain arriving in rail cars until it is reloaded onto ships for river or sea transport. The extensive handling facilities for weighing, storing and reloading the grain onto ships have very high capacities. In some instances, the handling capacity exceeds several hundreds of tons per hour. The interior of such silos (in North America called "elevators") consists of a number of vertical tube-like compartments that extend from ground level to the roof. Earlier, the basic shape of these compartments (in plan) was exclusively rectangular. The common walls were constructed in a wooden framework covered by thick boards. Recently, the walls of the bins have been constructed of masonry with steel reinforcing as well. In plan section the compartments have hexagonal shapes like honeycombs (Arndt, 1892). The grain is dumped into the upper end of the bin through a small opening. The grain is removed from the bin through closable openings in the floor. The sizes of the bins are very diverse. The largest have a capacity of up to about 250t. In this case, the height of the grain in the elevator is about 25m. It is clear that the large bin content must exert considerable pressure against the side walls and the floor. In the different construction textbooks, mention is actually made about the necessity of providing enough support to the bin walls. However, no guidance is provided regarding the determination of the grain pressure against the enclosing bin walls. The formulas for the pressure exerted by liquids are not applicable for grain because, in the latter case, the pressure transfer is greatly influenced by the Massmin 2004
friction of the individual grains against each other. In addition, the formulas for the calculation of earth pressure against retaining walls are also not applicable because these formulas, in general, have been developed for straight walls. For a bin, the content is completely surrounded by vertical walls. 2 EXPERIMENTAL SETUP To my knowledge, the only publication about tests regarding the measurement of grain pressure in bins/silos (and then only about the floor pressure) is found in Engineering (Roberts, 1882). These tests demonstrate that the pressure of the grain in a silo against the foundation increase with increasing height until a certain height is reached. Beyond that critical height there is no further increase of the bottom pressure. The largest pressure developed on the bottom is dependent upon the crosssection of the bin. It should be approximately proportional to the diameter of the largest circle inscribed in the crosssection. As it appeared valuable to obtain information on the increase of pressure in the silo from the surface down to any given depth, I performed a series of thorough (systematic) tests, the results of which are presented here. Four wooden test cells of square cross section with side lengths of 20, 30, 40 and 60 cm were constructed. For all of the tests, these cells were mounted on four adjustable screw supports denoted by S in Figures 1 and 2. The lower end of each cell consists of a well-fitting sliding floor. This cell floor rests on a decimal balance. At the start of the tests, the system is brought into balance by adjusting the counter balance weight G so that the weight of the bin bottom is cancelled out. In this balance position, a support is placed below the counter weight scale G and an additional weight is added. Before the start of the test, the grain to be poured into the apparatus and its containers are weighed. Grain is now poured into the cell until the counter weight scale G starts moving upward. By weighing the pouring container once again, the amount of grain remaining in the container can be obtained. Hence, the amount of grain in the bin generating
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Figure 3. System of adding weights to the board placed on top of the grain to simulate greater column heights.
3 RESULTS The test results summarized in Table 1 are graphically presented in Figures 4 through 7. The smooth pressure curves obtained in tests 2 through 6 suggest that significant observational errors were not present. Observational errors for Test No. 1 are proportionally higher since it was the first one performed. With continued use of the apparatus, the procedures were improved and hence observational errors reduced. If the content of the bin consists of a fluid, the bottom pressure (Translator: more correctly the measured force on the bottom) is exactly equal to the weight of the bin contents. The tests with grain, however, reveal a substantial reduction in the bottom pressure due to the frictional resistance between the bin walls and the grain. This frictional resistance becomes so large with increasing depth that the pressure increase is no longer noticeable. Therefore the friction between the grain and the bin wall must be equal to the weight of the enclosed grain layers.
Figure 1. Front view of the model bin.
4 BIN WALL PRESSURE The magnitude of the largest grain pressure against the surrounding walls in this case, and obviously the highest value achieved, is calculated in the following manner: Psmax = maximum pressure of the grain against the bin wall f = friction coefficient between the grain and the bin wall s = side length of the square bin cross-section dh = height of the grain layer γ = density of the grain
Figure 2. Side view of the model bin. the floor pressure can be known exactly. After placing addition weights on the counter weight scale G, the bin is lifted by means of the support screws S until the scale is brought back into balance. Now the apparatus is once again ready for use and the test is continued. In this way, it was possible with a single filling of the bin to determine the floor pressures corresponding to various filling heights.
From Figure 8 one finds that: Psmax • f • 4s • dh = γ • s2 • dh or Psmax = γ•s/4f
(1)
Under the assumption that the friction coefficient f remains constant even for varying grain pressures against
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Figure 4. The measured bin floor force as a function of the amount of grain for Test 1. the bin walls, the maximum pressure for different bins of square cross-section is proportional to the side length. 5 DETERMINATION OF WALL FRICTION The value of f can be determined by the following experiment. The flat vessel shown in Figure 9 was filled with wheat and then a wooden board was placed on top of the wheat. By a string attached to the middle of the board and connected to a spring balance F, the board could be pulled over the grain. By placing weights on the upper surface, the pressure (Translators Note: A dimension problem enters again here since "pressure" between the board and the grain is in reality a force.) between the wooden board and the grain could be precisely controlled. The results of these tests are presented in Tables 2a-2f. Tests 10 and 11 were performed in exactly the same way on different days. The difference in the ascertained friction coefficients is about 8% which may be due to differences in the humidity of the air. The maximum value of f, f = 0.346, was obtained in Test 8. The minimum value, f = 0.302, was obtained in Tests 10 and 12
Figure 7. The measured bin floor force as a function of the amount of grain for Tests 5 and 6. From Figure 8 one finds that:
Figure 5. The measured bin floor force as a function of the amount of grain for Tests 2 and 3.
Figure 8. Diagrammatic cut through the bin
Figure 6. The measured bin floor force as a function of the amount of grain for Test 4. Massmin 2004
Figure 9. Diagrammatic view of the shear test.
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These investigations in which the surface pressure has been gradually increased by a factor of seven from the initial ones show that the maximum friction coefficient has not been significantly affected. Thus the friction coefficient can be assumed, with sufficient precision, to remain constant for different grain heights.
But it is known that when x=0 then p=0
6 MATHEMATICAL DERIVATIONS Hence The following mathematical derivations are presented under the assumption that the side pressures in the grain are proportional to the respective vertical pressures.
x0 = 0 Thus
Given: P = the total force of the grain on the bin floor p = vertical pressure of the grain on a surface ps = lateral pressure of the grain f = friction coefficient between the grain and the bin wall K = psF and p K = psf p s = side length of the square cross-section u = perimeter of the square cross-section = 4 s F = bin cross-sectional plan area = s2 x = height of the superincumbent pile of grain in the bin γ = density of the bin contents e = base of the natural logarithms From Figure 10, the force equilibrium on the elemental slice may be expressed by
loge (1-Kup) = -Kux Fγ F This equation may then be rewritten as 1 – Kup = e-Kux/f Fγ Knowing that F = s2 and u = 4s, then 1 – e-4Kx/s = 4Kp sγ Solving for the bottom vertical pressure p gives
F(p+dp-p)= (γ F dx)-(f ps u dx)
p = s γ (1- e-4Kx/s) 4K
Simplifying and rearranging terms yields
Finally, solving for the bottom force P one finds that
dp = γ dx – K p (µ/F) dx
P = s3γ (1-e-4Kx/s) 4K
(2)
(3)
By additional rearranging one can write 7 COMPARISON OF THEORY AND EXPERIMENT dp = dx γ (1-Kup) Fγ
In equations (2) and (3), only the value for K is unknown. This can be easily determined from experiments (tests) 1 through 6. In Test 4 (wheat with _ = 0.8 g/cm3, s = 40 cm = 4 dm), a maximum floor force of 63 kg corresponding to a pressure of 3.94 kg/dm2 was obtained. In this case, the circumferential friction force for a pile of height of 1 dm is obtained using 16 γ = 16 (0.8) = 12.8 kg = ps f u Thus psf = 12.8 = 12.8 = 0.8 u 16 Since K = ps f p Then K = 0.8 = 0.203 3.94
Figure 10. The forces acting on the slice. As a simplification, I set
This can be integrated to yield -F loge (1-Kup) = Ku(x-x0) Fγ 296
K = 0.2 γ = 0.80 Equations (2) and (3) then take, respectively, the following simplified forms Santiago Chile, 22-25 August 2004
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p = s ( 1 γ e-0.8x/s)
(2a)
P = s3 ( 1 γ e-0.8x/s)
(2b)
A comparison of the experimental results from test No. 4 and the predicted grain pressures using equation (2b) is presented in Figure 11. Better agreement is obtained if one selects a larger value of K. This larger K value would occur if it is determined from pressure measurements at a smaller height. It cannot be excluded that an observational error enters when determining the maximum pressure. In the same way as described above, the K values for the other tests can be determined. Thus: Test No. 1 No. 2 and 3 No. 4 No. 6
K 0.211 0.235 0.203 0.227
1.15(0.7p) = about 0.8p As a starting point for the calculation of bin wall strength, a uniform loading of magnitude ps = 0.75 p = s ( 0.75)( 1 – e-0.8x/s)
(4a)
will provide enough precision. Even if the above tests and mathematical results are considered enough for the evaluation of the required strengths of the silo floor and of the silo walls, a simpler calculation method may be desired. We are on the safe side for the common grain types if we assume wheat with a density of 0.8 kg/dm3 as the bin content and we use a value of K = 0.2. The grain pressure against the bin floor is calculated with formulas (2a) and (3a). From the graphical presentation in Figure 13 one can read the value of Z = ( 1 – e-0.8x/s)
The observed deviation can be explained by the small variation of f. From Tests No. 2 and 3 the value is about 15.5% greater then that of Test No. 4. In the experiments conducted to determine the magnitude of the friction coefficients (experiments 7 through 12), the values of f were found to vary from f = 0.302 to 0.348 which is a range of 14.5 %. For the mathematical determination of the lateral pressure against the bin walls, one could choose as the largest value that obtained in Test No. 2 and No. 3, f = 0.346, and as the smallest value that obtained in Test No. 4, f = 0.302. The value of
included in formulas (2a) and (3a) for any value of x/s.
ps = Kp f Figure 11. Comparison of the theoretical curve and the experimental results for Test 4.
calculated from Tests No. 2 and 3 is ps = 0.235p = 0.68 p 0.346 From Test No. 4 it becomes ps = 0.203p = 0.675p 0.302 The average is taken as ps = 0.7p This is the average lateral pressure acting against the side walls of a bin with a square cross-sectional area. However, it is to be assumed that in the proximity of the corner of the bin cross-section, the surface pressure is less then this average value. On the other hand, in the center part of the wall a larger value would be obtained. Figure 12 shows the approximate values of pressure transfer to the sidewalls. It is taken, as a starting point, that the grain pressure radiates outward from the center of the bin. At the bin wall a pressure of p sin _ is exerted. Here _ is the angle at which the radial pressure meets the wall. The highest value of the pressure against the wall occurs in the middle of the bin wall and is about 1.15 times the average pressure or about
Figure 12. Variation of the side force as a function of the position along the perimeter of a square bin.
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8 EXAMPLES Example 1 Calculation of the floor pressure for a bin with a plan section area 4m x 4m (s=4m). Pile Height(x) (m)
x/s
Z (from Fig 13)
p=sZ (t/m2)
P = s3z (t)
1 2 3 4 5 6 8 10 12 14
0.25 0.50 0.75 1.00 1.25 1.50 2.00 2.50 3.00 3.50
0.18 0.33 0.45 0.55 0.63 0.70 0.80 0.86 0.91 0.94
0.72 1.32 1.80 2.20 2.52 2.80 3.20 3.44 3.64 3.76
11.5 21.1 28.8 35.2 40.3 44.7 51.2 55.0 58.2 60.1
The corresponding lateral pressure applied to the bin walls can now be determined by multiplying the appropriate vertical pressure by the factor 0.75. It is easiest to read the magnitude of
obtained. The same thing occurred when it was attempted to determine the floor pressure at various positions on the floor by special small "openings". In my opinion this result is due to an arch forming in the grain column because of the decrease in cross-section. This strongly influences the pressure transfer to the smaller cross section. It was decided not to carry out a test with a movable bin wall extending over the whole height of the bin (from top to bottom) because this would have required considerable effort and because the above described calculation achieved the same goal. 10 TESTS WITH OTHER GRAINS In addition to the wheat experiments, experiments were also carried out with rye and corn (maize). The results of these tests agree, in general, very well with the results of the tests described above. For rye with a density of 0.75 kg/dm3, the pressures were about 20% lower then those of wheat. Corn which has the same density as wheat (_ = 0.80 kg/dm3) produced, because of its smooth surface, a floor pressure that was 22% higher. For bins intended for corn storage, the strength of the bin walls and the bin floor has to be increased by 22%. 11 CONCLUSIONS
w = 0.75( 1 – e-0.8x/s) from Fig. 14 for the desired x/s values. Example 2 Calculation of the lateral pressure for a silo with a plan cross-sectional area of 3m x 3m (s = 3m) Pile Height(x) (m)
x/s
w (from Fig 14)
p=sw (t/m2)
1.5 3.0 4.5 6.0 8.0 10.0 12.0
0.50 1.00 1.50 2.00 2.67 3.33 4.00
0.245 0.410 0.530 0.595 0.655 0.690 0.720
0.735 1.230 1.590 1.790 1.970 2.070 2.160
The previously described experiments from which the above described calculations where derived were carried out with small scale models as the production (manufacture) of larger scale test bins would have been associated with considerable cost. However, they should be clarifying in a number of respects and it is hoped that these results can be confirmed by tests in larger scale; such tests are anticipated to be conducted by a company well known in mill construction. I retain the right to provide additional information on this subject (topic) in the future. ACKNOWLEDGEMENTS:
9 DIRECT MEASUREMENT OF PRESSURE Originally it was the intention of the author to determine the lateral pressure of grain directly by experiments. For this purpose one of the experimental cells was provided with a moveable sidewall door. The load on the door could be applied by means of an angular lever and a suitable gravity load. The pressure required for opening the sidewall door could now be determined by calculation. However, during the testing, the opening of the door by the pressure of the cell contents was so slow that accurate results could not be
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The authors would like to express their thanks to Mr. Stephan Mahnke for help in procuring the articles by Janssen and Arndt. Ms. Alyson Boye assisted in the preparation of the manuscript. REFERENCES • Arndt, C. 1892. Die Silospeicher von Galatz und Braila (The Storage Silos at Galatz and Braila). Zeitschrift Des Vereines Deutscher Ingenieure. Volume 36, No. 34, August 20, pp 973-978. • Janssen, H.A. 1895, Versuche Über Getreidedruck in Silozellen (Experiments Regarding Grain Pressure in Silos). Zeitschrift Des Vereines Deutscher Ingenieure. Volume 39, No. 35 August 31, pp 1045-1049. • Roberts, I. 1892. On The Pressure of Wheat Stored in Elongated Cells or Bins. Engineering. Volume 34. Oct 27. p 399.
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Figure 13. Calculation of the bottom force for a 4m x 4m bin. Table 1a. Floor pressure measurement results from Test No. 1 in Cell 1. Plan cross-sectional area = 20 cm x 20 cm, wheat with density of 0.8 kg/dm3. Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
2 2.5 2.7 4 1.2 4.4 4.6 5 5.5 6 6.2 6.4 6.6 6.8 7 7.2 7.4 7.5 7.6
2.53 3.5 3.9 6 7.1 8 9.2 10 12 16 17.9 20 21.8 24 26 28 36 50 62
Table 1b. Floor pressure measurement results from Test No. 2 in Cell 2. Plan cross-sectional area = 30 cm x 30 cm, wheat with density of 0.8 kg/dm3. Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
12.5 14 16.5 18 19 21 22 22.5 23
18.5 22.25 31.8 238.1 44.4 65 78.5 90 1802
2The
test apparatus has a capacity of only 90 kg wheat which can be handled without too much difficulty. After Massmin 2004
establishing the floor pressure for the 90 kg pile, the top surface of the grain was smoothed out and a suitable board placed upon it. Weights of 22.5 kg are then placed on the board as shown in Figure 3. This is equivalent to the force exerted on the bin floor by a 90 kg grain pile. The measured 23 kg force on the bin floor corresponds to a pile of 2 x 90 kg = 180 kg wheat. Table 1c. Floor pressure measurement results from Test No. 3 in Cell 2. Plan cross-sectional area = 30 cm x 30 cm, wheat with density of 0.8 kg/dm3. Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
12.5 15 17.5 19.5 21 22 23
19.5 27 36.5 49 69.7 90 1803
3This value is obtained from Test No. 2 by adding another 22 kg weight on the column of grain.
Table 1d. Floor pressure measurement results from Test No. 4 in Cell 3. Plan cross-sectional area = 40 cm x 40 cm, wheat with density of 0.8 kg/dm3. Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
30 35 40 45 50 55 58 60 63
40 52.5 66 85 106.8 138.2 165 192 3844
4This value is obtained similar to Test No. 2 by laying a weight of 60 kg on the column of wheat.
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Table 1e. Floor pressure measurement results from Test No. 5 in Cell 4. Plan cross-sectional area = 60 cm x 60 cm, wheat with density of 0.8 kg/dm3.
Coefficient of friction = f = 3.77/10.9 = 0.346 Table 2c. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 9.
Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
Normal Force (kg)
Pulling Force (kg)
‘80 120 140 155 165 170 175 180 185
105 185 245 300 350 380 440 510 540
20.9
6.98 6.73 6.78 7.23 7.28 7.23 7.18 6.98 6.98 6.73
Table 1f. Floor pressure measurement results from Test No. 6 in Cell 4. Plan cross-sectional area = 60 cm x 60 cm, wheat with density of 0.8 kg/dm3. Measured Floor Pressure (kg)
Amount of Wheat in Cell (kg)
100 150 170 180 185 190
140 290 430 500 540 8905
Average pulling force = 7.01 kg Coefficient of friction = f = 7.01/20.9 = 0.335 Table 2d. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 10. Normal Force (kg)
Pulling Force (kg)
30.9
9.23 9.03 9.58 9.23 9.63 9.28 9.43 9.18
5This value is obtained by laying a weight of 165 kg on the grain column of 540 kg. Table 2a. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 7. Normal Force (kg)
Pulling Force (kg)
5.9
1.93 1.93 1.93 1.93 1.98 1.88 2.03 2.08 1.93 1.98
Average pulling force = 1.96 kg Coefficient of friction = f = 1.96/5.9 = 0.332
Table 2e. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 11. Normal Force (kg)
Pulling Force (kg)
30.9
10.23 10.18 10.08 9.93 9.93 10.03 10.38 10.03
Average pulling force = 10.10 kg Coefficient of friction = f = 10.10/30.9 = 0.327
Table 2b. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 8.
Table 2f. Results of the sliding friction tests conducted with the wooden board and grain. Test No. 12.
Normal Force (kg)
Pulling Force (kg)
Normal Force (kg)
Pulling Force (kg)
10.9
3.73 3.73 3.78 3.78 3.83 3.83 3.78 3.78 3.73 3.73
40.9
12.53 12.48 12.38 11.93 12.98 12.48 12.28 12.53
Average pulling force = 3.77 kg 300
Average pulling force = 9.32 kg Coefficient of friction = f = 9.32/30.9 = 0.302
Average pulling force = 12.36 kg Coefficient of friction = f = 12.36/40.9 = 0.302
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Some remarks on ore pass design guidelines William Hustrulid and Changshou Sun, Department of Mining Engineering, University of Utah, SLC, UT, USA Graham Mustoe, Division of Engineering, Colorado School of Mines, Golden, CO, USA
Abstract Mass mining systems rely heavily on the use of ore passes to transfer material from the production to the haulage levels. Hang-ups can severely impair production capacity, present safety problems in their removal, and can significantly reduce ore pass life. Although there are some rules of thumb relating ore pass geometry and fragment size, the database is weak. The paper begins by presenting a critical review of the literature concerning ore pass design rules. The results of an extensive set of laboratory experiments involving different ore pass geometries and material characteristics are then presented. Some numerical simulations using OREPASS 3D, the 3D discrete element program developed at the Colorado School of Mines are then presented. The paper concludes with some recommendations.
1 INTRODUCTION The trend in underground mining systems has been to increased scale. This is reflected in the size of the openings, the block/sublevel heights, and the extraction pattern geometries. As the size of the extraction equipment and drawpoints is increased, so too is the maximum size of the pieces to be handled. Some of the boulders are too large to be moved from the drawpoint and must be treated in place. Others can be moved to special drifts where they are drilled and blasted or otherwise broken. Finally there are the majority of the boulders which are transported along with the smaller muck to an ore pass. The special case of muck being transported directly to crushers is an interesting one but will not be dealt with here. There are a number of philosophies concerning orepass management. In some cases, the loaders will dump to an open hole such as shown in Figure 1A. Guidelines are given to the loader operator regarding the limiting boulder size so that hangups should not occur. The proper judging of boulder size and following the instructions is left up to the loader operator. In other cases, positive control is exerted over the maximum size of the boulders being transferred to the pass in the form of either a scalper grizzly as shown in Figure 1B or the full grid grizzly shown in Figure 1C. The disadvantage of these two systems as compared to the open hole is that some means must be provided to break the boulders caught by the grizzly. In older times, this function was performed by men with double jacks. Today, this function is done by heavy hydraulic hammers often operated remotely. The placement of a grizzly between the loader and the open hole introduces a dumping delay in the loading operation and, as such, acts as a production rate limiter. As a compromise, some operations dump down a short open hole to a grizzly level where boulder breaking can occur in parallel rather than in series with the loader. There are limits applied to the orepass size based on economics, excavation techniques, and layout considerations. With regard to the latter, as the diameter of the orepasses is increased in a panel caving sector, the pillars become smaller and the Massmin 2004
Figure 1. Different Boulder Control Alternatives spans larger presenting the possibility for areal stability problems. A limiting orepass diameter given such practical considerations is of the order of 3 to 3.5m. With the increase in mining scale and the emphasis on minimizing specific development, the number of orepasses serving a panel is minimized. However with the increased mining scale, the tonnage which must be passed through each ore pass is greatly increased from earlier times. To pass the extra tonnage, linings of various types are sometimes installed. Their presence can reduce the useable diameter from the 3.5m diameter, as excavated, to a 3m finished diameter, for example. Although this might not seem to be a major change, this must be evaluated with regard to changing the hangup potential. Over the years, some guidelines have been developed regarding the relationship between the size of pieces to be passed, the size of the orepass and the hangup potential. This paper will review those guidelines and the data supporting them, describe some of the results of a recent laboratory ore pass study, present some ore pass simulations using OREPASS3D, and conclude with some suggested guideline revisions. 2 REVIEW OF CURRENT DESIGN GUIDELINES The most complete set of laboratory studies relating particle size, ore pass size and the formation of hang-ups was published by Aytaman (1960). In these studies, the source material was screened into the ten size fractions given in Table 1.
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Table 1. Screen Fractions Used in the Hang-up Study by Aytaman. Aytaman (1960). -
0.742 0.525 0.371 0.263 0.185 0.131 0.093 0.065 0.046 0.033
+ + + + + + + + + +
0.525 0.371 0.263 0.185 0.131 0.093 0.065 0.046 0.033 0.023
ins ins ins ins ins ins ins ins ins ins
Flow tests were then performed using each of the fractions in cylindrical steel tubes having nominal opening diameters of 3", 2", 1-1/2", 1-1/4", 1", _" and _". The tubes were filled and then the bottom closure was opened. It was noted whether the material (1) flowed out, (2) flowed intermittently, or (3) did not flow at all. The (x,y) points corresponding to the maximum screen opening (d) for each pair of screens and the actual tube diameter (D) were then plotted with an indication as to which of the above flow classes they belonged. Limiting lines representing the noflow and free-flow conditions were then superimposed. As seen in Figure 2, Aytaman (1960) found that the regions were defined by: D/d ≥ 4.21 flow always occurs, free flow 2.24 < D/d < 4.21 flow may or may not occur, probable flow D/d ≤ 2.24 flow never occurs, no-flow Aytaman (1960) indicates that d is the "largest" particle size. Although performed in the
bored to a diameter of 4 m, particles with a diameter of less than 1.0 m will flow freely through the channel, particles greater than 2.0 m will not flow and sizes in between will produce intermittent hang-ups." In this case, Just (1980) has modified the mesh-size definition of the "largest particle size" used by Aytaman (1960) to a "largest particle size" defined by the diameter. With regard to designing within the intermittent flow regime, Just (1980) suggests that one might make an analysis of the type made by Driver (1972) in which the number of interruptions to the free flow of material is expressed in terms of the extracted material. Although his predicted blockages per ton of material passed seem to be, in general, very high, it may be worthwhile reviewing the approach. On the basis of such curves, one might make an ore pass size – material size decision based upon an acceptability of blockages criterion. In 1983, Engineers International, Inc. submitted their final report "Guidelines for Open-Pit Ore Pass Design" to the U.S. Bureau of Mines (Engineers International, Inc., 1983). In the report, they include Table 2. By way of explanation, they indicate that "Table 2 summarizes the empirical evidence for design against interlocking arch formation in terms of the ratio of ore pass dimension (D) or outlet dimension (Do) and the frequency of arch formation. Table 2 provides design guidance with respect to the prevention of interlocking arches. In order to use the information in this table, one must decide which ore pass or outlet dimension and which particle dimension should be used to form the ratio D/d or Do/d. In the case of circular ore passes and outlets handling roughly equi-dimensional rock fragments, the situation is unambiguous. In the case of rectangular or square ore passes and outlets, arching is apt to occur across the smallest dimension. However, such an arch may not stop flow over the entire area of the ore pass or slot. In fact, slot outlets are considered to be more active than circular outlets because arching can only occur in one direction with slot outlets while two mutually perpendicular directions are possible over the circular outlets. As a practical matter, the dimensions of rectangular ore passes are unlikely to differ so much as to be considered slots, so that the least dimension of a rectangular ore pass or outlet should be used as D or Do in the table." Table 2. Ratio of Ore Pass or Outlet Dimension to Particle Dimension versus Relative Interlocking Frequency. After Engineers International, Inc. (1983).
Figure 2. Ore Pass Design Data and Guidelines By Aytaman (1960). laboratory with certain fractions rather than a full spectrum of size distributions, this is considered by the present authors to be the most comprehensive data set available. In 1980 Just (1980) presented his version of the Aytaman (1960) results in which he simplified the limits to be D/d = 4 and D/d = 2. He indicates "If material containing a com-plete size range of particles is involved, the flow regime is determined on the basis of the largest particle size. For exam-ple, if the ore pass is raise302
Dimension Ratio
Relative Interlocking Frequency
D/d or Do/d > 5 3< D/d or Do/d < 5 D/d or Do/d < 3
very low, almost certain flow often, flow uncertain very high, almost certain no-flow
Where: D = ore pass diameter Do = draw point (outlet) diameter d = maximum size of muck
They continue to say "In the case of rock fragments that are slab-shaped rather than equi-dimensional, a very conservative approach would be to use the largest dimension of the fragments as d." This, apparently, is the dimension they have suggested for use with their table. As was discussed in the introduction, in a panel caving layout the space available for locating an ore pass is very limited. It is important that the opening dimension be "big enough" but not "too big" since this
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translates into major expense. Given the high multiplication factor, there cannot be any confusion in the appropriate fragment dimension to be used. Engineers International used a number of sources to arrive at their table including the earlier cited work of Aytaman (1960). New information of their own was collected during eight mine visits. The data set included (1) both circular and rectangular openings, (2) open holes and grizzlies, and (3) various levels of hang-up formation. In a few cases two different ore pass sizes were used at the same mine. Using their data, the present authors have constructed Figure 3 with the axes being minimum ore pass dimension and grizzly opening. Superimposed on the figure are their recommended limits of D/d = 5 and D/d = 3. In this case, the grizzly opening was used to define d. Although some points lie on the wrong side of the appropriate bounding line, the general fit is not too bad. The Aytaman (1960) limits have also been superimposed on Figure 4 and these seem to describe the data even better in the view of the present authors.
Figure 4. Ore Pass Design Data and Guidelines From Lessard and Hadjigeorgiou (2003).
Figure 3. Ore Pass Design Data and Guidelines From Engineers International, Inc. (1983). Lessard and Hadjigeorgiou (2003 a,b) have reported the results of an extensive research program aimed at the development of design tools to minimize the occurrence of ore pass interlocking hang-ups in metal mines. The program involved the collection of field data from mines in Quebec coupled with numerical simulations involving the use of PFC2D and PFC3D. Figure 4 is a summary of their field results. In their plot, the ore pass dimension D was based upon mine plans while the fragment size d was dictated by grizzly dimensions or, in the absence of a grizzly, by image a anylysis of the blasted ore. The x axis was labelled "Largest Rock Size" which has been re-labelled "Grizzly Opening" by the present authors. They superimposed the lines corresponding to D/d = 2, that indicated as the flow limit by Jenike(1961), and D/d = 5 suggested by International Engineers, Inc (1983) as the free flow limit. They found D/d = 4 to be the empirical limit for ensuring free-flow and suggest this to be a more suitable design guideline for use in Quebec mines than earlier ones from the literature. As reference, the present authors have superimposed the Aytaman (1960) limits of D/d = 2.24 and D/d = 4.21.
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It is evident from this short review that guidelines do exist for either (1) selecting ore pass dimensions based on a given fragment size or (2) indicating the sizes of fragments that can be passed through an ore pass of a given dimension. The broad width of the "intermittent flow zone" is a problem for those trying to "optimize" mine layouts. This topic will be addressed again in a later section. There is, however, a consistency problem with regard to the appropriate fragment dimension to be used. Different authors mean different things when they refer to fragment dimension ‘d’. No standard definition of ‘d’ currently exists. This seems to be a critical point which needs addressing. It appears logical to the present authors to select the intermediate particle dimension as that which should be selected as ‘d’ when applying the design rules. Figure 1, introduced earlier, shows an open raise, a raise equipped with a scalper grizzly and one equipped with a square grid grizzly. It will be assumed that a typical particle is ellipsoidal in shape with dimensions L x B x T (length x breadth x thickness). For the open raise, the intermediate dimension is limiting presuming the long axis rotates to point downward during flow (as suggested by Jenike (1961)). For the square grid grizzly, it is also the intermediate dimension that controls. For the scalper grizzly, it is the minimum dimension that controls. In this case, the use of the intermediate dimension for design would be conservative. 3 LABORATORY ORE PASS FLOW STUDIES The laboratory ore pass flow studies of Aytaman (1960) involved testing certain size fractions while holding all other factors constant. Engineers International, Inc (1983) and Lessard et al (2003b) reported actual ore pass data for a variety of mine conditions. Because actual flow in ore passes involves a rather wide range of conditions, it was decided to try and evaluate the applicability of the ore pass design guidelines under a few of them in the laboratory. Based on the literature review and mine experience a number of different factors which influence the operation of an ore pass system might be taken into account: • Ore pass size • Ore pass shape
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• • • • • •
Wall condition: smooth or rough. Orientation: vertical or inclined Operational mode: dump while empty or partially full Boulder size and shape Particle distribution Moisture condition
All of these were studied to some degree in the current laboratory study. Due to space limitations, only a few of the results can be reported here. The interested reader is referred to the report by Sun and Hustrulid (2004). A local quarry supplied quantities of their "road base" mix and "concrete" mix for use in the tests. These mixes could then be combined in different ratios to achieve other mixtures. The one reported in some detail here is the socalled "mine" mix which is patterned after the fragmentation distribution of the primary ore from the Esmeralda section, El Teniente mine. The laboratory size distribution is shown in Table 3. The fraction –23mm + 19.5mm was examined in some detail with measurements made of a number of the individual pieces. The results are given in Table 4. As might be expected, - the minimum dimension > 19.5 mm - the intermediate dimension = 23mm As shown, the average aspect ratios are: L/B = 1.49 B/T = 1.47 At the Esmeralda sector, when using 6 yd3 LHDs it is considered that the largest boulders have dimensions of the order of 1.5m:1.2m:0.85m (Diaz, 2003). The corresponding aspect ratios would be, respectively, 1.25 and 1.41. The Esmeralda ore passes are 3.5m in diameter and are equipped with a 40" x 40"grizzly (Barraza et al, 2000). The D/d ratio is 3.45 and hangups are not indicated to be a problem. Figure 5 shows a picture of the mine mix. A cube with a side length of 14mm has been used for scale. The properties of the mix were determined using a variety of tests. The results are summarized in Table 5. A series of ore pass tests were conducted in much the same way as done by Aytaman (1960) but not as extensive. ABS tubes of 1-1/2", 2", 2-1/2", 3", and 4" internal diameter were selected for the primary tests. Each tube was 60" in length. Although for most of the tests smooth-walled tubes were used, some tests were performed using ABS tubes with a ring structure to simulate the rough wall conditions in ore passes after some time of use. The primary testing orientation was with the tubes mounted vertical. Some tests were also performed with the tubes mounted at 50o and 60o from the horizontal. One test series was performed comparing tubes of circular and square cross-section. Table 3. Size Distribution of the Mine Mix. Fraction - 23.0 mm + 19.5 mm - 19.5 mm + 16.0 mm - 16.0 mm + 12.5 mm -12.5 mm + 9.5 mm - 9.5 mm + 6.5 mm - 6.5 mm + 2.38 mm - 2.38 mm
Weight Percent 9 10 11 15 20 30 5
The tests were conducted using two extreme case LHD dumping modes: (1) dumping of the mixture into an empty tube; and (2) completely filling the tube and then releasing 304
the bottom to simulate full ore pass draw. For Case 1, no hang-ups were observed in any of the tests. Since the intermediate dimension of the largest particle in the different mixtures was less than the tube size, this was to be expected. For Case 2, the tubes were completely filled in a number of steps giving time for compaction and inter-locks to form. This was exactly the way that Aytaman (1960) conducted his tests. Upon opening the bottom of the tube, the flow conditions are noted. For certain geometries and material conditions, the material flows directly out. For others, however, only a portion flows out before there is a blockage. Upon removing the blockage, the
Table 4. Dimensions of the –23mm+19.5mm Fraction Particles Average Aspect Aspect Max Inter Min Dimension Ratio 1 Ratio 2 Sample (L) mm (B) mm (T) mm mm L/B B/T 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23
27 31 38 33 35 30 32 29 31 42 29 28 25 26 32 30 27 28 35 32 27 27 26
Average 30.43
21 20 20 23 22 20 21 17 20 17 21 22 23 22 21 22 23 22 15 23 21 23 21
16 13 12 18 21 12 15 15 16 11 9 13 13 18 13 15 17 16 12 14 15 17 14
21.33 21.33 23.33 24.67 26 20.67 22.67 20.33 22.33 23.33 19.67 21 20.33 22 22 22.33 22.33 22 20.67 23 21 22.33 20.33
1.29 1.55 1.9 1.43 1.59 1.5 1.52 1.71 1.55 2.47 1.38 1.27 1.09 1.18 1.52 1.36 1.17 1.27 2.33 1.39 1.29 1.17 1.24
1.31 1.54 1.67 1.28 1.05 1.67 1.4 1.13 1.25 1.55 2.33 1.69 1.77 1.22 1.62 1.47 1.35 1.38 1.25 1.64 1.4 1.35 1.5
20.87
14.57
21.96
1.49
1.47
Table 5. Properties of the Mine Mix Property
Value
Solid density Bulk density Void ratio Friction angle (wall and material)
2.5 g/cm3 1.62 g/cm3 0.54
Internal friction
24° (smooth surface sliding test) 29° (rough surface sliding test) 37° (angle of repose) 35.9° (direct shear)
flow may start again but then stop. On the other hand, it may continue to flow. Each test is repeated four times with the average number of hang-ups reported. The results using the mine mix in vertical smooth tubes are shown in Table 6. The results are quite surprising based upon conventional ore pass design guidance. Free-flow
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conditions were obtained for D/d ratios only somewhat greater than 2.2 rather than the 4.21 value suggested by Aytaman (1960), for
consider the results obtained with the "concrete mix" and the "road base mix" under the same test conditions. Pictures of the two mixes are shown in figures 7 and 8, respectively. The particle size distributions are given in Table 7.
Figure 5. Picture of the Mine Mix
Figure 7. The Concrete Mix
Figure 8. The Road Base Mix
Figure 6. Test Setup Using Vertical Tubes Table 6. Results of Vertical, Filled, Smooth Wall Ore Pass Tests Using the Mine Mix Tube Diameter (mm)
Number of Hang-ups
D/d
102 76.5 63.8 51 38.3
0 0 0 1 3.5
4.4 3.3 2.8 2.2 1.7
example. The one clear difference between the tests is that here a wide distribution of particle sizes was used whereas his tests involved a very narrow range. To examine the importance of particle size distribution it is of interest to Massmin 2004
As can be seen, there is a major difference in the distributions. About 87% of the concrete mix lies in the range –23mm+12.5mm. Only about 20% of the road-base mix falls in this range. For the road-base mix, about 80% falls below 12.5mm whereas for the concrete mix the corresponding value is about 13%. The results of the filled ore pass flow studies for these mixtures are presented in Table 8. A number of interesting observations can be made. First, the concrete mix with a high percentage of relatively large particles behaves similarly to the Aytaman (1960) results. For a D/d ratio of 1.7, there is essentially no flow. For a ratio of 3.3, on the other hand, there is free-flow. The fact that a large number of inter-locking hang-ups are observed is because about 23% of the mix is made up of particles represented by ‘d’. For the road base mix with a very high proportion of fines, there was still free-flow at a ratio of 2.2 and there were relatively few stoppages even for a ratio of 1.7. This is to be expected since the percentage of the mix upon which the ‘d’ in the D/d ratio is based represents only about 3% of the total. The important conclusion is that the particle distribution plays a major role in determining the frequency of occurrence of inter-locking hang-ups and must be considered when formulating ore pass design guidelines.
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4 NUMERICAL SIMULATIONS OREPASS3D, developed at the Colorado School of Mines under the auspices of the Western Mining Resource Center, . is a user-friendly program for evaluating different ore pass/chute designs. Here, it has been used to examine the Table 7. Size Distribution for the Concrete and Road Base Mixes Weight Percent Concrete Mix Road Base Mix
Fraction - 23.0 mm + 19.5 mm - 19.5 mm + 16.0 mm - 16.0 mm + 12.5 mm -12.5 mm + 9.5 mm - 9.5 mm + 4.75 mm - 4.75 mm + 2.38 mm - 2.38 mm
22.8 26.1 38.3 11.0 0.9 0.5 0.4
Two modified cases were then generated. In the first modified case (Case 1M), the dimensions of the particles were proportionally increased from the standard size by 25%. The D/d ratio drops to 2.0. In the second modified case, the dimensions of the particles were proportionally decreased from the standard size by 25% and D/d = 3.33. The result of the Case 1M (Figure 10) simulation shows that the majority of particles fall as a large chunk of material shortly after the gate is opened. However, a small chunk of particles hang-up at the top of the ore pass for at least 5 seconds. This is a surprising result especially since the standard case hung up.
3.3 10.0 6.4 24.5 10.0 22.0 23.8
Table 8. Results of Vertical, Filled, Smooth Wall Ore Pass Tests Using the Concrete and Road Base Mixes
Tube Diameter 102 76.5 63.8 51 38.3
Number of Hang-ups Concrete Road Base Mix Mix 0 0 2.5 3.5 10
0 0 0 0 1.5
D/d 4.4 3.3 2.8 2.2 1.7
flow characteristics of three different size particles in a 3m diameter, 32 m high, vertical ore pass. For the standard case, the ore pass was filled with ellipsoidal particles of dimension 1.8m x 1.2m x 0.85m. Using the intermediate dimension of 1.2m and the ore pass diameter of 3m gives D/d = 2.5. The coefficient of friction for ore/ore and ore/wall interactions was specified as 0.577 which corresponds to an angle of friction of 300. After the particles settled the lower gate was opened for 10 seconds. By modeling the 32 m. section as a 2m and a 30m , the option of opening the gates at different times was offered. The result of the standard case shows that after the gate is opened, a few particles fall. However, the remainder of the particles hang-up (see Figure 9). This hang-up is maintained for at least 10 seconds.
Figure 10. The Results for Case 1M with 25% Bigger Particles. The Initial and Final Configurations of the Particles Show Particle Chunks The results of Case 2M (Figure 11) involving the 25 percent smaller particles show a majority of the particles falling freely immediately after the gate is opened. In this case the ore particles loosen up as they flow and do not show the chunky motion seen in Case 1M. Overall these results are as expected – the smaller particles flow more freely. One further analysis (called ripple) was performed using the standard size particles. Here, the diameter of the lower part (3m height) of the vertical section of the ore pass was varied from 3m to 3.1m and then back to 3m. This was done to assess the likelihood of particle arching being generated by a small variation in the diameter of the ore pass. The results of this case (Figure 12) showed that ore particles flowed relatively freely and did not show any noticeable arching. As illustrated by these few examples, a great many ore pass design possibilities can be easily studied, with this powerful DEM software. Unfortunately, it is very limited with respect to the range of particle sizes that can be included in any one problem. This is a major limitation given the size distribution flow dependence observed in the laboratory studies.
Figure 9. Hang-up in the Standard Case 306
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Figure 11. The Results of Case 2M with 25% Smaller Particles. The Initial and Final Configuration of Particles Showing Free Flow of Particles
Figure 12. Standard Case Particles with a Lower Ore Pass Geometry "Ripple.": Initial and Final Configuration of Particles Showing the Free Flow of Particles 5 CONCLUDING REMARKS The proper design of ore pass systems to avoid hang-ups is important for successful mass mining. A number of different control techniques are available. For a given ore pass dimension, the size of the boulders must be controlled. One way of accomplishing this is through instructions to the loader operator, another more certain technique is through the placement of a grizzly. The traditional approach to design is through design ratios involving the ore pass dimension and the maximum size fragment. Different authors have selected different particle dimensions to be used with the ratios. It is strongly recommended that the intermediate fragment dimension be selected as the standard. This corresponds to the opening dimension of a square grid grizzly. The current laboratory test series has shown that the occurrence of inter-locking hang-ups is dependent on the size distribution as well as on the absolute size of the pieces present. Three different size distributions
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were studied while maintaining the same maximum particle size. Free-flow conditions were achieved under three different D/d ratios, namely 3.3, 2.8 and 2.2 when the percentage of the largest size particles present assumed values of 23%, 9%, and 3.3%, respectively. For the mine mix, an ore pass diameter to d (intermediate boulder dimension) ratio of 2.8 should be enough to assure freeflow. Applying this rule at the Esmeralda mine, assuming boulders having an intermediate dimension of 1m (the size of the grizzly opening), one would expect that an ore pass 2.8m in diameter should be sufficient. As indicated earlier, the actual ore pass diameter is 3.5m and hang-up problems have not resulted. In reviewing the field data collected by Lessard and Hadjigeorgiou (2003b) it is seen that there are many instances where free-flow conditions are observed for D/d ratios in the range of 2 - 4. From the lab results these can be explained by the broken muck having a special distribution. The present authors suggest trying to correlate the measured size distributions with their hang-up observations. Perhaps this will be done in the future. It was initially thought that numerical techniques such as OREPASS3D, PCF2D, or PFC3D might be of use for developing design guidelines. Unfortunately, at the present time the range of particle sizes that can be studied is extremely limited. The models can easily show hang-ups when the particle size distribution is narrow and the particles are relatively large compared to the ore pass dimension. However, it is clear that single particles only slightly smaller than the ore pass will fall through without arching. Although it appears to be going back in time, the authors suggest conducting some simple laboratory tests when trying to decide on ore pass designs. For new sections in existing mines, fragmentation distributions should be available. For new mines, it may be possible to predict fragmentation curves with sufficient accuracy based upon core measurements. Almost 40 years ago, Kvapil (1965b) published a nomogram to be used in conjunction with some basic relationships for ore pass dimensioning. Through the nomogram, the effect of the component distribution (coarse, medium, fines) and the characteristics of the individual components could be included in the ore pass size decision. It may be time to revisit this approach making modifications, as necessary, to fit modern practice. In summary, it would appear that, except in very special fragmentation cases, the required D/d ratios for free-flow are certainly much less than 5. As shown by the experiments and indirectly by the field data from the Quebec mines, the ratio can even sometimes be less than 3. Further work on the influence of size distribution on ore pass hangup formation is certainly required. The work reported in this paper will hopefully lead to a new way of approaching the problem and new, practical design rules. ACKNOWLEDGEMENTS This publication was supported by Cooperative Agreement number U60/CCU816929-02 from the Department of Health and Human Services, Center for Disease Control and Prevention (CDC). Its contents are solely the responsibility of the authors and do not necessarily represent the views of the Department of Health and Human Services, CDC. Support provided by the Department of Health and Human Services, CDC, is greatly acknowledged. The work presented is part of the Health and Safety research activities currently being carried out at Western Mining Resource Center (WMRC) at the Colorado School of Mines. Ms. Alyson Boye of the Department of
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Mining Engineering, University of Utah assisted in the preparation of the manuscript. Mr. Rob Byrnes assisted in performing the model tests. REFERENCES • Aytaman, V. 1960. Causes of "Hanging" in Ore Chutes. Canadian Mining Journal. November (pp 77-81), December (pp 71-75) and January 1961 (pp 41-45). • Barraza, M., and P. Crorkan. 2000. Esmeralda Mine Exploitation Project. MassMin 2000 (G. Chitombo, editor). Australian Institute of Mining and Metallurgy. Pp 267-278. • Diaz, G. 2003. Personal Communication • Driver, D.J. 1972. Sublevel Stope Extraction Control Design. Unpublished Thesis, University of Queensland. • Engineers International, Inc. 1983. Guidelines for OpenPit Ore Pass Design (Volumes I and II). Final Report on Contract J0205041 Submitted to the U.S. Bureau of Mines. September. • Hadjigeorgiou, J., and J.-F. Lessard. 2003. The Case for Liners in Ore Pass Systems. 3rd International Seminar on Surface Support Liners. Montreal. August. 14pp.
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• Jenike, A.W. 1961. Gravity Flow of Bulk Solids. Bulletin No. 108. The University of Utah. • Just, G.D. 1980. Rock Fragmentation and the design of Underground Materials Handling Systems. CIM Bulletin. February. Pp 45 – 51. • Kvapil, R. 1965a. Gravity Flow of Granular Materials in Hoppers and Bins – Part 1. Int. J. Rock Mech. Mining Sci. Vol 2. pp 35-41. • Kvapil, R. 1965b. Gravity Flow of Granular Materials in Hoppers and Bins – Part 2. Int. J. Rock Mech. Mining Sci. Vol 2. pp 277-304. • Lessard, J.F., and J. Hadjigeorgiou. 2003a. Ore Pass Systems in Quebec Underground Mines. Mine Planning and Equipment Selection. Kalgoorlie, WA. April 23-25. pp 509-521. • Lessard, J.F., and J. Hadjigeorgiou. 2003b. Design Tools to Minimize the Occurrence of Ore Pass Interlocking Hang-ups in Metal Mines. Technology Roadmap for Rock Mechanics. South African Institute of Mining and Metallurgy. Pp 757-762. • Sun, C., and W. Hustrulid. 2004. Ore Pass Study – Final Report. Submitted to the Western Mining Resource Center (WMRC), Colorado School of Mines.
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Geomechanical criteria for orepass design – El Teniente Mine, Codelco Chile Eduardo Rojas, Antonio Bonani, Eugenio Santander, El Teniente Division, Codelco Chile
Abstract El Teniente mine is a large coper deposit. The current daily production rate is over 100,000 tons per day of ore. Using orepasses this ore is loaded into trains or trucks. Orepass overbreak is an important problem that affects the mining proces. In some cases, an ore pass can reach 2 or 3 times its original diameter. This effect shortens the orepass life and may produce instabilities affecting the galleries an upper levels. The damage is principally associated to the: Where (and When) an ore pass is excavated and when orepass production begins in relation to the caving front position (initial stress environment and stress changes). Favorable rockmass conditions for orepass construction and operation are associated with the specific ratio between the major principal stress and UCS. According to previous concepts, design criteria were defined for the construction and operation of orepasses in Tte. 4 South sector.
INTRODUCTION At El Teniente Mine, caving considers a process that begins from extracting ore, and finishes in the concentration plant. Ore extraction includes some sequential activities which are: ore extraction, vertical flow (through ore passes) and haulage. In this system, orepasses are the most vulnerable structures because design does not take into account the latest expertise of rockmass behaviour affected by induced stresses. To date, 30 Mt of primary rock have been handled through 169 orepasses. In many cases, orepass overbreak determines its working life, so affecting mining structure. In relation to the last point, a methodology to evaluate orepass stability in primary rock based upon field stresses and rock mass conditions is introduced in this paper. Typical orepass problems at Tte. 4 Sur Sector are presented to validate this method. GENERAL OVERVIEW El Teniente Mine is a Codelco-Chile underground copper mine. It is located in the foothills of the Andes in the central zone of Chile (South America), about 70 km SSE from the capital city, Santiago. The El Teniente porphyry copper orebody is one of the largest known copper deposits in the world. It includes andesite, diorite and hydrothermal breccias of the Miocene era as the main lithologies. A chimney of subvolcanic breccias known as the "Braden Pipe" post dates the copper-molybdenum mineralisation. It has an inverted cone shape and the hydrothermal mineralisation is distributed round this pipe over a variable radial extension of 400 m to 800 m, with mineralogical associations of variable strength. The mineralisation has two very different forms, secondary ore is located near the surface and primary mineralisation occurs at greater depth. The primary ore can be described as a high cohesion and impermeable rock mass. The stockwork veins, containing the original mineralogy, are sealed. According to Massmin 2004
geomechanical behavior, the primary rockmass can exhibit brittle, often violent failure under a high stress condition. El Teniente Mine began operations in 1906. Since then, various exploitation methods have been used in productive sectors located in secondary mineral. The methods ranged from "raised work over mineral" combined with shrinkage stoping and pillar recovery to block caving and panel caving. Knowledge gained over the years concerning primary ore exploitation with conventional panel caving (320 million tons extracted to date) has indicated that the advance of the caving face is the main cause of gallery damage in levels below the UCL. Experience has also shown that a variation of conventional panel caving sequence, knows as the "preundercut," method reduces the degree of gallery damage in the levels below the UCL, as well as the possibility of rockbursts associated with the advance of the undercut face. Pre-undercutting basically consists of advancing the undercut ahead of all development in the lower levels. All production level development is made behind the cave front and under the caved area. CAVING METHOD DESCRIPTIONS Caving requires the control of the breakage of a great volume of rock, induced by undercutting and draw point extraction, and hauling this broken ore from production level to benefit plant. At El Teniente Mine, gravity ore flow is possible due the favourable location of the concentration plant. In this way, ore passes are a key factor in ore flow process. Physical and mechanical rock mass properties have changed as the mine has deepned. Primary rock has minor grade, greater hardness and coarse fragmentation than secondary ore. INDUCED STRESS IN ROCK MASS A great quantity of information related to panel caving variants and their use in El Teniente Mine can be found in technical books and papers.
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NIVEL HUNDIMIENTO ZANJA
NIVEL DE PRODUCCION
NIVEL MARTILLOS PICADORES
PIQUES DE TRASPASO
Figura 3. Induced stress state in rock mass by the Conventional Panel Caving
NIVEL TRANSPORTE
Figura 1. Panel Caving schematic view condition
Stress conditions induced by the caving front are analyzed to identify the parameters that control orepass stability. Three zones of rock mass conditions induced by caving front stress are recognized (Pre-mining, Transition and 0Relaxation (Rojas E., Molina R., Cavieres P. 2001). Therefore, several distances are incorporated into mining design to implement safe working conditions in the mine. The Induced stress state generated by pre-undercutting caving is showed in figure 2. In this case all production drifts are developing in the de-stressed zones, which is more favourable than the transition zone, excepting the galleries in the ventilation and haulage levels which are constructed in the pre-mining zone.
Some induced stress states in the rockmass are generated by conventional panel caving, but in this case all excavations are constructed during the premining state of stress so they are affected by stress changed (direction and magnitude) due to the caving front advance. Geomechanical Evaluation of orepasses at the Tte.4 Sur sector Rock mass properties: Rock mass deformability and strength were calculated for each lithology by scaling intact rock properties to the mining level (1e4 m3 of rock (table 1)).
Table 1. Geotechnical properties of rock mass at the 1e4 m3 scaled according to Hoek and Brown Criteria Rock Mass (Hoek & Brown) Parameter Em [GPa]
Andesita
Diorita
Bx hidro Anh Turmalina
43
36
14
0.19
0.26
0.16
Φm [°]
35
37
41
σm [MPa]
56
42
37
σtm [MPa]
6
2
1.5
Cm [MPa]
11
8
6
75-85
70-90
85-90
νm
GSI
Em: Young Modulus, νm: Poisson ratio, σcm: uniaxial compresion strength, σtm: tensile strength. Cm: Cohesión. Figura 2. Induced stress state in rock mass by Preundercutting Panel Caving
Φm: internal friction angle, , GSI: Geological Strength Index Geology Primary Andesite is the principal rock within the Tte. 4 Sur sector (46%), coarse phorphyry (42%), hydrothermal breccias (10%), and also some bodies of lamprophyre and latite. Geology is presented in figure 4.
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Table 2. Minor joints at 27L Op 44 and 25L Op 44 Orepasses
Joint
27L Op 44 Orepass 25L Op 44 Orepass DIP DIP/Direction DIP DIP/Direction
1
40
113/293
70
305/125
2
70
135/315
72
330/150
3
41
36/221
60
65/245
Tabla 3. In situ field of stress at Tte. 4 Sur Sector
Figure 4. Tte. 4 Sur Geology and Structural general plan view
The Primary rock mass at El Teniente Mine is characterized by stockwork including faults and veins with different length and trace, width, filling and frequency (figure 5).
Stress
Magnitude [MPa]
Bearing [°]
Inclination [°]
σ1
-33
130
35
σ2
-25
10
10
σ3
-12
330
53
Ore pass geotechnical problems at The Tte. 4 Sur Materials handling encompasses ore extraction from draw points using LHDs which dump into ore passes that conduct the mineral to the haulage level (Teniente 5 level) where it is loaded into trains and then transported to the principal ore passes of the mine.
Figura 5. Andesite Primary lithology (stockworks) According by principal, medium and minor size joints can be defined. JOINTS IN THE ROCK MASS Principal joints Figure 4 shows major joints at Tte. 4 Sur Sector, which are principally faults (>95%), the most relevant include: P, S, Sur – Sur 1 faults systems and also the Lamprophyre Dyke. Minor joints These correspond to geological structures with trace lengths less than excavation size (generally 4 meters). The structural geology of two orepasses is presented in Table 2. Rock mass field Stress The state of insitu stress within The Tte. 4 Sur sector has been obtained from measurements and monitoring using the CSIRO Hollow Inclusion cells of stresses. The results are presented at Table 3.
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Figura 6. Isometric view of a typical orepass at Tte. 4 Sur Size reduction of ore is performed by pick hammers located on the Ten Sub-4 level, roughly 33 meters below the production level. Figure 6 shows a scheme of Tte. 4 Sur sector. The ore pass system of Tte. 4 Sur consists of a pair of branches from two drifts located in the vicinity of the orepass. These branches coverage into one orepass at the reduction level. Construction and reinforcement Different practices have been used for orepass construction at El Teniente Mine: Drop Raising (long blast hole) and Raiseborer technique where a pilot guide of 1.83m is realized previous to manual slyping manual. Orepass construction includes the reinforcement of the orepass collar, ("brocal") (3m in length), with steel rings and concrete.
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Ore pass construction requires manual slyping to obtain the final cross section of the orepass, installing minicage cables, full length grouted, from the beginning of the ore pass at the production level, to the end of the excavation at the rock hammer chambers, and also the segment of the orepass between rock hammer level and Haulage level.
Ore pass behaviour Overbreak has been a major orepass problem, shortening – in some cases – the life time of these systems and affecting their availability which does not 50%. Figures 9 and 10 indicates ore pass damage (overbreak) conditions, with only low quantities of rock handled, such that they must be closed.
Figura 9. Damage at 9 Op 54 Orepass (Production 350.000 ton)
Figure 7. General View of the support of the orepass
Figure 10. Damage (Overbreak) at 21-23 Op 49 orepass (Production: 21 Op=250.000 ton, 23 Op=900.000 ton)
Figura 8. Cable bolting system installed through ore pass section. Cable used (minicage), Cable length varies between 3 and 3.5 meters. Ore Production Until the present about 220 Mton have been from produced at Tte. 4 Sur sector, using 120 ore passes. Information about haulage drifts and their distance to the caving front at the moment of starting production is encountered at Table 3. Tabla 4 Orepasses production at Tte 4 Sur sector Haulage galleries (XC)
312
Production (TON)
Caving front to the start production orepasses distance (mts)
8AN
1.000.000-1.600.000
(-13) - (25)
2 AN
1.200.000-2.400.,,,
(6) - (31)
6 AS
1.400.000-2.600.000
(-20) - (0)
14 AS
500.000-1.700.000
(42) - (73)
21 AS
600.000-1.600.000
(20) - (50)
29 AS
250.000-1.600.000
(10) - (32)
39 AS
300.000-1.400.000
(21) - (82)
44 AS
50.000-900.000
(30) - (90)
49 AS
20.000-2.000.000
(22) - (92)
Geotechnical analysis of orepass design Many studies has been carried out to understand the damage mechanisms that affects orepasses, which in general terms have been associated to stress and geological conditions, overbreak has inititated due to the impact of rock boulders into the ore pass walls. Methodology of analysis (Hypothesis) Two factors are relevant if we consider the rock mass state (premining, transition and relaxation) due to the panel caving method: 1.Where (and When) the ore pass is excavated in relation to caving front position? (initial stress environment) 2. When ore pass production begins, related to caving front position? (stress changes) Support of this hypothesis is based upon the level of field stresses (80 to 100 MPa) induced by the extensive caving front (>350m). See figure 11. Taking into account the uniaxial compressive strength (UCS) of intact primary rock between 100 to 130 MPa and considering the proximity of the caving front, fracturing of excavation walls and roof is generated.
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Figura 11. Principal Stress at the excavation related to the distance from the caving front Figure 13. Graph represents the relationships between s1/sci>0.4 and the location of caving front. Proposal for Design´s Methodology The methodology is based upon stress estimation considering the σmax/σci ratio, where smax is the maximum shear stress round the excavation and σci is the uniaxial compressive strength (UCS). The maximum stress (σmax) was determined based upon a back-analysis of excavations located in different sectors, such as: Ten 4 Sur, Esmeralda and Sub-6. Accordingly, it is possible to establish that σmax is coincident with the maximum principal stress σ1, in the abutment stress (transition) zone. Furthermore, when σmax(=σ1)/σci, >0.4 a progressive damage of rockmass round the excavation begins, which is reflected by increased fracturing (previous and new fractures), key-block formation and overbreaking. Figure 12 shows the plotted values that represent σ1/σci > 0.4 ratio. Back-analysis at The Tte. 4 Sur Information about orepass location respect to the caving front and overbreak initiation at Tte. 4 Sur was compiled, establishing that the greatest damage of productive orepasses is generated at the stress induced zone (abutment zone). See Figure 14.
Figure 14. Orepass location from caving front and overbreal initiation at Tte. 4 Sur Greatest ore pass damage (overbreak) is generated when production starts before or during abutment stress induction. On the other hand, stress fracturing at the abutment zone begins when s1/sci is greatest than 0.4.
Figura 12. Plotted values of s1/sci design and operation
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establish orepass
Results and applicability of proposed methodology During the year 2000 the proposed methodology was established (in the Teniente 4 Sur sector mine). Over the last four years, 20 orepasses have been incorporated into production and these have not experienced orepass overbreak. The good orepass geomechanical conditions have increased their availability to about 80%. The results of this study indicate that ore pass behaviour (excavated in primary rock at El Teniente Mine), could be predicted based upon the induced stresses during their operation. Therefore, the key factor define the moment of orepass excavation and production. Several examples of orepasses which have been excavated in the vicinity of a collapsed orepasss in the relaxation zone, reflect a normal operation of them, without an adverse overbreak, which indicates the small effect of induced stresses at the relaxation zone.
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Finally, it is important to establish other factors which affect orepass overbreak, when production starts. a) Rock mass strength (lithology and structures) b) Design of orepass geometry c) Support/reinforcement system d) Orepass operation (e.g., operation without orepass overflow)
ACKNOWLEDGEMENTS
When orepasses are excavated in zones where σ1/σci>0.4 it is best to install an appropriate support system, capable of absorbing rock mass deformation and overbreak.
[1] Brummer R. (1998). Design of Orepasses, CAMIRO Mining Division Limited. [2] Brzovic A. (2001). Sistema de Clasificación geotécnico – Roca Primaria Mina El Teniente. API T01M101. Informe: SGL-I-050/2003. [3] Cavieres P., Rojas E., (1993).Hundimiento Avanzado: una variante, al método de Explotación de Hundimiento por Paneles en Mina El Teniente. CODELCO CHILE. Publicación 44ª convención IIMCH. [4] Constanzo H., Rojas E. (2000). Problemática Geomecánica de los Piques en la Mina T-4 Sur, Informe Interno Planificación. División El Teniente – Codelco Chile. [5] IM2 (2001). Estudios de Traspaso de Mineral, Proyecto IM2 43/99. [6] Rojas E. (2000); Evaluación Geomecánica de la estabilidad de los Piques de Traspaso en Roca Primaria - Mina El Teniente. Trabajo presentado en Programa de Especialización en Innovación Tecnológica de Geomecánica y Geotécnica aplicada a la Minería (Diplomado, Universidad de Chile). [7] Rojas E., Molina R., Cavieres P (2001); Preundercut caving in El Teniente Mine, Chile; SME 2001; Undergournd Mining Method. [8] Santander E., Bonani A., Rojas E. (2004). Evaluación Geomecánica del estado de Sistemas de Traspaso. Sector Reservas Norte – Area Invariante. SPL-I-0102004.
CONCLUSIONS AND RECOMENDATIONS 1. Orepass damage generated during the panel caving method is controlled by Where (and When) the ore pass is excavated in relation to the caving front position (initial stress environment) and when ore pass production begins, related to caving front position (stress changes). 2. A progressive orepass damage (overbreak) round the excavation periphery during orepass life is a result of rock mass stress and geological condition, after overbreak has been initiated due to the impact of rock boulders on the orepass walls. 3. A support system capable of to absorbing rock mass deformation and keeping the rock fractures in place must be designed for overbreak control, Steel rings (1 inch thick) and concrete have been demonstrated to be a good reinforcement system. 4. Analyzing the formation of keyblocks surrounding an orepass is an important analysis for overbreak evaluation. RECOMMENDATIONS 1. Position orepass construction and operation within a relaxation state of stress (behind the caving front), where σ1/σci ratio is lower than 0.4 (σ1/σci 250 ~22 ~50 ~45 ~50 ~45 ~55
0 ~10 ~13 ~20 ~25 ~20 ~30
Rupture of one or two wires Wedge/strand interface slip Wedge/strand interface slip Wedge/strand interface slip Wedge/strand interface slip Wedge/strand interface slip Wedge/strand interface slip
Note: Residual force was that recorded during sliding. Actual residual force will be 0kN when strand completely pulled through wedge.
Table 4 provides a summary of the peak and residual forces measured for all the specimens tested and the failure modes. Apart from the control test specimens 1 and 2 used to confirm that anchors can mobilise the strength of the strand (but not necessarily its full elongation potential of at least ~3.5%), all the anchors failed by slipping of the strand within the anchors. This is attributed to the inability of the wedge to slide sufficiently relative to the barrel. Figure 9 shows the extent of wedge movement relative to the barrel that can be expected when anchors are loaded to cause rupture of the strand at ~250kN. Also note that the wedges protrude from the base of the barrel for this anchor.
# Barrel/wedge interface glued before load applied. * Strand force ~36kN
Figure 9: Appearance of new anchors (Specimens 1 & 2) after failure of one and two strand wires.
Figure 8:Comparison of new anchor with anchor subjected to 6 months exposure in a mildly corrosive artificial environment.
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The total wedge movement relative to the barrel was measured to be ~10mm. The movement is associated initially with the teeth embedding in the outer wires of the strand and then secondly the barrel expanding radially outwards. The radial stresses in barrels and the associated radial expansion have been studied both experimentally and
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theoretically by Marceau et al. (2001, 2003). It is possible for barrels to expand unacceptably if the barrels are too thin or made from steel with too low yield strength. The radial forces are higher when the barrel/wedge interface is lubricated. Figure 10 shows the appearance of anchor specimens 3 and 4 after failure by strand slip. The test for specimen 3 was stopped to show the extent of material shaved from the outer strand wires by the wedge. All tests in which wedge/strand interface slip occurred were similar. The results for test specimens 6 and 8, shown in Figure 11 and Figure 12, are typical of the results obtained for the other specimens. Figure 13 shows the effects of sliding on the strand wires. A summary of wedge positions before and after testing and the total wedge movement are summarised in Table 5. It is of significance that, other than for test specimens 3 and 4 (shown in Figure 10) in which the barrel/wedge interface was deliberately prevented from sliding, a small amount of wedge movement did initially occur. This suggests that the wedge is prevented from further sliding by the corrosion products built up on the exposed wedge surface near the wedge taper.
Figure 11: Typical force-wedge displacement (Test 6 and Test 8).
Figure 12: Typical force-strand displacement (Test 6 and Test 8).
Figure 10: Appearance of anchor specimens 3 and 4 after failure by wedge/strand slip.
Table 5: Summary of average wedge positions before and after testing. No.
Initial Wedge Outstand (mm)
Final Wedge Outstand (mm)
Wedge Movement (mm)
1/2 3 4 5 6 7 8
6.5 1.3 0.4 2.7 2.1 1.6 0.4
-4.0 1.3 0.4 1.5 1.7 1.5 -0.1
10 2 (1 hr) 18.5
12-30 >100 125
1.5 indicates undamaged rock and mb < 0.5 indicates completely damaged rock.
• Cundall, P, C. Carranza-Torres & R. Hart, 2003. A new constitutive model based on the Hoek-Brown criterion. • Guest, A and Van Hout, G, 2000. An application of Linear Programming for Block Cave Draw Control, in proceedings MASSMIN 2000, Brisbane, pp 461-468. • Hannweg, L and Van Hout, G, 2001. Draw Control at Koffiefontein Mine, in proceedings Mine Mechanistaion and Automation, 2001, Johannesburg, pp • Preece, M, 1996. Front Caving – a solution to waste dilution at Koffiefontein Mine for the extraction of the 370m to 490m ore reserve. Massive Mining Methods, Randburg. The South African Institute of Mining and Metallurgy.
From the discussions detailed, it can be seen that a number of factors contributed to the sudden massive failure experienced at Koffiefontein Mine. No single action can be held accountable for this. The Front Cave mining method has many inherent rules to make it work – if these are all strictly adhered to, there would be no problem with the method. In the case of Koffiefontein Mine, production constraints and requirements may have contributed to the failure of the Front Cave. ACKNOWLEDGEMENTS The authors are grateful to all their colleagues that helped them during the development of this work. Also, the authors wish to acknowledge the permission of the Director, Operations to publish this paper. REFERENCES
Figure 3: Tunnel deformation and rock damage around extraction tunnels
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Henderson’s new 7210 production level K. Keskimaki, B. Nelson, M. Callahan, R. Golden, S. Teuscher, C. deWolfe, A. Hansen, Climax Molybdenum Company, Henderson Mine Empire, CO USA
Abstract Henderson’s new 7210 panel cave production level is currently under development, with plans to initiate the undercut in October 2004. This level will provide nearly all the production from Henderson for the next 20 years. The 7210 level design is similar to 7700 level with the following improvements: high lift cave, wider bell spacing, enhanced wire meshing and shotcrete for drift support, a redesigned drawpoint brow, alternative roadway construction methods and the addition of dewatering drifts. These improvements will reduce ore development costs by over 50% from historic costs. The initial undercut area will be monitored with TDR to ensure that voids that potentially could cause an air blast are not allowed to form. LHDs with 7.4 m3 buckets will feed bins that transfer the ore from 7210 production level to 7065 truck level, where remotely controlled loading chutes are utilized to load 72 tonne side dump haul trucks. From there the ore will flow across the existing 24 km conveyor system to the concentrator located on the opposite side of the continental divide.
1 INTRODUCTION 1.1 General Description The new 7210 level is as deep as 1550 m below the peak of the overlying mountain, with the crusher reclaim located at a depth of 1643 m. This makes Henderson one of the deepest caving operations in the world. (Figure 1). The largest improvement in development costs comes from mining a portion of the orebody in a 240 m to 340 m high cave. Another major change to the way the cave is mined is an increase in the drift spacing from 24.4 m to 30.5 m. This will reduce the number of bells which will need to be mined to exploit this orebody. Other improvements included in the new lower level include an increase in the size of the bucket on the existing and new LHDs from 6.7 m3 to 7.4 m3 and improved support in the drawpoints. Finally, roads have been improved on the 7065 haulage level by changing construction methods to allow for less expensive new installation and easier repairs. The dewatering drifts ease the maintenance of the roads by positively draining away the water as well. These improvements have and will increase productivity and will also make the operation much safer. These improvements combined with a behavioral based safety program and a commitment to "Zero" (to not get hurt at work or at home), have driven the incident rate four times lower than it was 15 years ago, and three times lower than five years ago. Although larger, the modern equipment greatly reduces the risk to the operator due to enhanced designs, better guarding, and more ergonomic cabs with sound suppression and dust filtration.
2 GEOTECHNICAL ISSUES ON THE LOWER LEVEL CAVE The Henderson molybdenum deposit is comrised of two partially overlapping orebodies that lie 1080 m to 1600 m beneath the summit of Red Mountain. The orebodies are entirely within a Tertiary rhyolite porphyry intrusive complex ranging in age from 24 to 30 million years that has intruded Precambrian granite. The deposit is elliptical in plan, with overall dimensions of 670 m x 910 m. In section, it is arcuate with an overall height of 550 m. The mineralization Massmin 2004
Figure 1: Henderson Mine Cross Section is relatively continuous in the orebody and consists of molybdenite and quartz in random, intersecting, closely spaced veinlets. (Rech et al, 2000) The ore of the Henderson deposit has exhibited the strength characteristics of a competent granite or rhyolite, with uniaxial compressive strengths typically ranging between 100 Mpa and 275 Mpa. Past experience on the 8100 and 7700 production levels has been that the rock caves fairly readily, more so than might be expected of a rock of this compressive strength. This is believed to be due to the lubricating influence of molybdenite coatings and fillings on geologic structures. The low friction angle of molybdenite leads to easy shear and tensile failure along mineralized structures. Ore grade has historically been a good indicator of both rock competency, and cavability at Henderson (Rech et al, 2000). Low grade areas, particularly in the high-silica zones on the 8100 production level, correlated closely with areas where weight problems were experienced (Rech et al, 1992). These relationships are expected to continue to hold true for the 7210 production level. Models showing ore
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classification by molybdenum grade will be created to help predict geomechanical conditions. It is also expected that models of other geomechanical parameters such as RQD and a rock mass classification such as RMR or Q will be developed. Historically, Henderson’s RQD has ranged from 0 to 100, averaging 49, and RMR has been found to range from 27 to approximately 60 (Rech et al, 2000). Orderly progression of the growth of the new cave on the 7210 production level is vital to ensure safe working conditions and achievement of production goals. A plan to monitor growth of the cave using Time Domain Reflectometry (TDR) cables has been developed. The current plan assumes installation of TDR cables in 13 vertical holes approximately 120 m long. Those holes will be drilled from exhausted areas of the overlying 7700 production level. The layout of these holes on approximately 50 m centers and their near-vertical orientation should provide good data of the geometry of the new lower level cave as it grows up towards the existing cave. A similar system of vertical and sub-vertical TDR cables was successfully used at Henderson to monitor the growth of the 7755 level cave towards the 8100 production level (Rech and Watson, 1994). It will also be important to monitor the eastward growth of the 7210 cave as it crosses over the critical boundary from low-lift to high-lift conditions. For this purpose, two additional fans of inclined holes are planned to be drilled from the former 7500 rail haulage level. TDR cables installed in inclined holes on the northeast corner of the 7755 undercut level are currently showing the eastward progression of the cave in response to the shooting of the undercut. Where safely possible, cave growth will also be monitored by routine drift inspections on the existing upper levels which are located in the path of the cave growth from the 7270 undercut level. 3 LAYOUT & DEVELOPMENT The 8100 and 7700 production levels were oriented with production drifts at a bearing of due north. The 7210 production level is rotated to an azimuth of 64 degrees. (Figure 2). This orientation better follows the orientation of the orebody and reduces the amount of boundary drifting. It
Figure 2 – 7210 Level General Layout also reduces the amount of drifting required on the 7065 truck haulage level. The 7210 production level is located 18.3 m below the undercut level. Production drift spacing is 30.5 m with drawpoints spaced at 17.1 m intervals. The entry angle into the drawpoints is 56 degrees. (Figure 3). 3.1 Bell Development The increase in the production drift spacing from 24.4 m to 30.5 m necessitated a redesign of the V-cut and bell. The location of the brow relative to the production drift was not changed when the drift spacing was increased. This resulted in a doubling in the length of the V-cut from 7.4 m to 14.8 m. The V-cut has 3 drill set-up locations from each end of the drawpoint crosscut rather than the single set-up in the past (Figure 4). The new drawpoint spacing and brow lcation results in a much more uniform area of influence for the drawpoints than with the previous layout. The bell development drilling was also redesigned with the increase in the production drift spacing. The previous
Figure 3 – 7210 Drawpoint Layout 398
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method of bell development required holes to cross each other from the two adjacent undercut drifts. This required the offsetting of bell development rings from drift to drift. Although the offset would theoretically eliminate the intersection of bell development holes from the adjacent drifts, it was still a common occurrence. The alternating offsetting of the rings from drift to drift also resulted in bell shapes that were inconsistent from drift to drift. The new bell and undercut ring spacing is 2.13 m with 8 rings per bell. This new bell development ring design has eliminated the issue of intersecting holes and has made the bell shape consistent from drift to drift. In addition, the slope from the major apex of the bell down to the brow is now defined with holes from the near drift resulting in a consistent and well defined surface. The previous design defined this slope with the ends of holes from the opposite drift which resulted in a ragged and inconsistent slope, especially when the ends of these holes were lost and couldn’t be loaded (Figure 4). Figure 4 – 7210 Ring and V-cut Layout 3.2 Production Drift and Drawpoint Support As part of lean mining at Henderson, an evaluation of the support for the production drifts and draw points was conducted. The poured concrete linings required significant manpower to install and maintain and in most cases were over-designed for the draw life of the area. The current design is an installation of 100 mm by 100 mm by 12 gauge wire mesh with 1.5 m split bolts covered with 100 mm of shotcrete as soon as the excavation of the draw point is completed. This flexible support is more forgiving under abutment loads than the poured concrete linings. This process requires minimal re-shotcreting to maintain safe production drifts for the draw life of the area. The steel brow section is the only the portion that is in poured concrete. (Figure 5). 3.3 Bins and Orepasses The ore bins for the new production level will be configured similarly to the bins used to feed ore from the 7700 production level to the 7500 rail haulage level. The major difference is that the new bins will be developed directly over the truck haulage level for the center loading chutes rather than the offset configuration used with the side-loading chutes for with the trains.
Spacing of the ore bins varies from 102 m to 137 m depending on factors such as the overall length of the production drift and the ore column height and tonnage in the particular area. Ore storage bins are oriented with the long axis parallel to the 7210 level production drifts, and half way between the two production drifts that each bin serves. (Figure 6). The bins are developed off drifts driven perpendicular to the production drifts on the 7150 ventilation level. Accessing the bins from this level allows the exhausting of some ventilation air down the orepasses, reducing the pistoning effect and resulting dusting on the production level. The two 60 degree, bored 2.1 m diameter production orepasses feed the bin from opposite sides and opposite ends of the storage bin. The bins are developed by mining a 5.5 m wide by 25 m long bin cutout on the 7150 ventilation level. A 2.1 m diameter by 18.5 m long raise is bored at a declination of 65 degrees from the 7150 level down to the truck loading chute cutout on the production level. The bin is then mined by drilling from the 7150 level and slashing to the bored raise. The resulting bin capacity is approximately 675 tonnes.
Figure 5 – 7210 Drawpoint Support Massmin 2004
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3.5 Shops Currently, the haul truck fleet, consisting of four 72 tonne haulage trucks, is maintained in the 7065 haulage level shop. This shop consists of a primary crane bay with four adjoining secondary work bays. An 18 tonne overhead bridge crane having 32 meters of travel and a maximum working height of 6 meters serves the crane bay. In addition to the haul trucks, mine support and development equipment is also maintained in the 7065 shop. A shop will be built on the 7210 production level to support the seven production LHDs. This shop will consist of a crane bay with two adjoining secondary work bays. An 18 tonne bridge crane having 10 meters of travel and a maximum working height of 4 meters is currently planned. Like the 7065 shop, the 7210 shop will also be used to maintain mine support and development equipment. The fuel system for the 7210 and 7065 levels will consist of a 26,000 liter tank and an automated fuel transfer system fed from the existing 8035 fuel storage system that is fed via a fuel line from the surface. The 8035 storage has a total capacity of 80,000 liters. Oil and lube is provided to both the 7210 and 7065 levels using totes, refilled from surface storage facilities. 4 HIGH LIFT CAVING
Figure 6 – 7210 Ore Bin Layout
The rock above the front of the truck loading chute is reinforced by installing cable bolts from the 7150 level. These bolts are installed after the orepass is bored from 7150 to the 7065 chute but before the bin is slashed. Additional ground support such as grouting and additional bolting will be added as geologic conditions warrant. 3.4 Ventilation The ventilation system at Henderson is an exhaust system that uses a single 930 kW fan on the surface of the 7 m diameter No. 1 exhaust shaft and six 225 kW underground fans that feed the 9.8 m diameter No. 5 exhaust shaft. Intake air is fed to the mine through the 8.5 m diameter No. 2 shaft, the 7 m diameter No. 3 shaft and the 16 km conveyor tunnel. The No. 2 shaft also serves as primary access into the mine. The new production level will not require deepening of the ventilation shafts nor new ventilation drifts to be driven to these shafts. Ventilation raises were bored from the perimeter of the previous intake and exhaust ventilation levels down to the new ventilation levels. The main access ramp to the new production level also serves as a primary source of intake air. Production area ventilation consists of intake air that is fed up 2.1 m diameter raises that are northeast of the initial production area. A temporary boundary drift was driven at the southwest end of the initial cave area. Ventilation doors will be installed in all production drifts along this southwest boundary. A 2.1 m diameter exhaust raise will be bored from each of these drifts to the 7150 exhaust level. All intake and exhaust raises will have pneumatically controlled doors so that ventilation can be shut off from inactive production drifts. The 7025 drainage level doubles as a secondary exhaust ventilation level. Each of the truck loading chutes has a 1.2 m diameter exhaust raise at the back of the chute that draws air down to the drainage level. This design ensures that the dust generated during truck loading operations is pulled away from the truck operator. 400
The word "lift" refers to the height of the ore column extracted by a production level of draw points in any caving operation. Traditionally, at Henderson, production levels were developed to extract a nominal 122 m lift. This means that every 122 m vertically in the ore body, a new production level with drawpoints and the associated levels to support the production level must be mined. Therefore, it stands to reason, that if the lifts can be made higher, the amount of development mining will be decreased. As production progressed on the 7700 level, engineers and operators began to study and design the next lower production level. The optimum elevation for the new level was determined to be 7210. This would result in a slightly higher lift of 150 m to mine the ore beneath 7700 level. At this time, only the western portion of the 7700 level had been mined and the idea emerged that the eastern portion of the ore above the 7700 level could be mined from the 7210 level. This would result in a nominal 240 m to 340 m lift for the eastern portion of 7210 while a 150 m lift would remain for the western portion of 7210 beneath 7700 level. If the 270 m lift could be done, it would cut the development mining costs per ton of ore in half. Loss of ore, however, could quickly negate these savings if there were to be major problems. To address these concerns, Henderson personnel benchmarked several other caving mines around the world that utilized high lift caves. Codelco’s El Teniente and Andina mines in Chile and the DOZ mine at Freeport, Indonesia were visited and found to be using high lifts successfully. The Philex Mine in the Philippines, although not visited, had been using high lifts for years. The information gathered from this benchmarking, along with the experience at Henderson of observing very vertical caving from cave monitoring systems and observation of the cave propagation to the surface, led to the decision to design and develop a high lift cave for the eastern portion of the 7210 level. In addition to cutting development costs per ton in half, a high lift cave has other advantages. Henderson’s experience has been that dilution from surrounding waste rock in a cave is less when there is not an existing cave above. Dilution occurs when fine rock from an older cave above sifts down through the coarser ore from the new cave. This phenomenon is avoided with the high lift. There will also be no steel from the drawpoints of an exhausted production level above reporting to the lower level with a
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high lift. Secondary blasting will be minimized as the percentage of very coarse ore from the bottom of ore columns will be low. There will also be several challenges with a high lift. While fine ore will minimize secondary blasting, it can also create dust problems and problems with grinding at sag mills at the concentrator. Fine ore will be present for the upper two thirds of a high lift column as compared to the upper one third of a 122 m column. Since twice the tonnage is undercut with each cave shot, the undercutting process will be much slower. As a result the abutment loading in front of the cave will sit in one spot for a longer time, potentially resulting in more damage to the pillars. This will create the need for additional ground support on the undercut level as well as additional support for the drawpoints. Each drawpoint will be used to extract twice the tonnage creating a need for a longer lasting brow. At Henderson the benefits outweigh the challenges for a high lift cave. These challenges will be addressed with the design of the production and undercut levels. 5 PRODUCTIVITY AND LEAN MINING 5.1 Productivity for LHDs The Henderson mine LHD production rate is expected to increase from 270 tonnes per hour per LHD to 300 tonnes per hour per LHD on the 7210 production level even though the average tram distance will increase from 50 meters to 60 meters. This improvement is a direct result of recently increasing the size of LHD buckets from 6.7 m3 to 7.4 m3. 5.2 Lean Mining Historically a large portion of mining cost at the Henderson Mine has been associated with substantial development mining leads to ensure cave capacity that would support any production demand. Recent market conditions have forced Henderson to reevaluate this process and develop a lean mining approach to long development leads. Development drifting, undercut drifting, undercut drilling, bell development, draw points and undercutting blasting are now in the same year they are scheduled for production. Mining activities that are significantly above our current capabilities are addressed by using a contract mining company. 5.3 Contracted mine development In lieu of adding staff and mining equipment to complete the development of the 7210 level with company forces, Henderson chose to hire a mining contractor to complete a portion of this work (6.5 km). The advantages, with over 50 per cent of the work completed, include less disruption to normal mine production as the existing employees and supervisors can focus on their normal activities. Henderson supplies four contractor support personnel, a project manager, and consumables like rock bolts and wire, and shotcrete as required. 6 7065 HAULAGE LEVEL In 1999, Henderson replaced its 23-year-old rail haulage system with a 24 km conveyor system fed from an underground crusher. To feed this crusher, Henderson purchased four side dumping 72 tonne haul trucks. These are backed up by two 36 tonne, rear dumping haul trucks used to initially develop the level and access ramps. Truck loading is accomplished by remote controlled pneumatic/mechanical center loading chutes. Truck traffic is controlled by a timed light system, similar to a city traffic control system. Production is currently achieved through the use of two or three of the 72 tonne trucks. The trucks are rigid frame Massmin 2004
and utilize a 5-axle system. The third and fourth axles are the drive axles while all but the center axle provide steering. Average cycle time is seven minutes, with an average productivity of 560 tonnes per operating hour. Although the trucks are rated at 72 tonnes, they are typically loaded to only 68 tonnes due to bed liners. The truck beds are lined with seven tonnes of 500 Brinell hardness liner plate to minimize wear on bed surfaces. Choice of the rigid-framed, side-dump trucks came from extensive benchmarking trips to other underground and surface mines. The rigid frame design has a longer life than the articulated trucks typically found in underground mines. And, when the 72 tonne trucks were ordered, the largest underground articulated truck available had a capacity of only 60 tonnes. Although tractor/trailer combinations were available with larger capacities, benchmarking indicated that rigid frame trucks were capable of faster dump cycle times since no stabilizing jacks are required for dumping. In addition, rigid frame trucks traditionally require less maintenance (Tyler, Keskimaki and Stewart, 2000). Issues with the trucks have been the under-designed axles which have been replaced with heavy duty axles and steering linkages which have also been upgraded. Potential improvements for the future include replacing the heavy liner plate with lighter plate, and modifications to the loading process to more evenly distribute the weight and increase capacity. Currently, 7 center-loading chutes are used to load the haul trucks. A truck backs under a chute, and then loads as the truck pulls forward. The chute is operated via an infrared controlled remote, with video monitoring in the cab and the chute and video is also available at a central dispatch control room. Chutes are designed for between two million and six million tones of ore and have liner plates ranging from 37 mm to 75 mm in thicknesses. Chute supporting steel is embedded in concrete and a concrete floor is poured in the chute and entrance. The haulage trucks are currently operated via two-way traffic with an average tram distance of 1000 m (Figure 7). The trucks are operated in 6.1 m wide by 5.5 m high drifts, which are wide enough for a single truck. The traffic control system is a municipal off-the-shelf system programmed for our truck route. This system is flexible and allows for changing traffic patterns. Installation and maintenance of the system is performed by the instrumentation staff at the mine (Tyler et al 2000). The haulage system is currently running at 20,000 tonnes per day utilizing two trucks. Development is nearly complete to allow the haulage system to operate as oneway traffic for increased productivity. 6.1 Haulage road construction Haul roads on the 7065 haulage level are vastly improved from when the haulage level started in late 1999. Improvements in drainage, road construction and maintenance have been instituted and truck productivity has increased. While ramping down to the haulage level, large water inflows were encountered. Original estimates were 150 l/m to 200 l/m based on core holes drilled through the proposed crusher centerline. Actual flows in the declines to the haulage level peaked at 1,500 l/m to 2,000 l/m of 50 degree C water when the 7065 level was reached (Callahan et al 2000). The discrepancy was due to the nature of the geology of the haulage level. The crusher itself is designed to be in the barren, but stronger Ute granite. Most of the haul route is located outside the core in surrounding stocks (Carten et al 1988). Water courses abound in the contacts
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Figure 7 – 7065 Haulage Level Layout
between the stocks and are intersected throughout the haulage route. As haulage was first initiated on the level, a series of pumps and drainage pipes were installed. Although this system was adequate for the first year, as the route expanded, road inundation and damage occurred with resulting loss in production. A secondary system of sump pumps and concrete ditches was installed and a dedicated drainage maintenance worker was assigned to keep the water from damaging road integrity. Finally in the fall of 2002, production had to be halted for one week in order to install concreted roadway in 245 m of the north end of the haulage route due to inability to maintain the roadway. It was decided to mine small (2.1 m wide by 2.7 m high) drainage drifts five meters below the roadway and drill drain holes down to these drifts to handle the water inflow. This was highly successful, and eventually was expanded to all of the wet areas of the haul route. All future haul route drifts are designed with this accompanying drain drift in order to facilitate water removal. When the haul roads were originally designed, several methods of construction were analyzed. These included concrete roads with imbedded drainage, muck haul roads with a poured concrete lean mat and several engineered gravel roads using geotextiles and sub-drains. The first haulage roads were built from mine muck from the drift development process. Development and haulage were scheduled concurrently; therefore these first primitive development roads became main haul roads. A maintenance 402
program using motor graders and a vibratory compactor was instituted to keep these roads passable for the 72 tonne trucks. Studies and benchmarking were undertaken to develop a longer lasting, easier to maintain road. Concrete roads were eliminated due to installation cost and previous experience that had shown a rough ride and a short life (less than 6 months). Three types of an engineered road system were chosen and tried. The first roadway system (Type 1) was a sub-base of ballast, a layer of geotextile, and a top layer of mixed mine muck and raisebore cuttings. The second roadway system (Type 2) investigated was a sub-base of ballast, a honey-combed grid system 15 mm thick filled with ballast, a layer of geotextile and a ballast and mine muck mix top layer. The third system (Type 3) examined was a ballast layer, a layer of geotextile, a ballast layer, a 15 mm thick geogrid layer filled with ballast and a top layer of a class 5 road base. The purpose of the ballast in all cases was to allow sub-drainage below the road. All three systems appeared to work at first, but in wet areas all failed within three to six months. In fact, the Type 1 system installed in the north end of the haul route (TNHW) failed completely and had to be replaced. After drainage drifting was installed, testing was begun for new types of road building materials and methods. A road improvement team developed a cement-modified soil road using run-of-mine muck with a lean concrete fill underneath where needed, and raisebore cuttings as a top-layer. This was not only less expensive than any other options (other than muck only) it was workable for a good riding surface and lasted much longer.
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Table 1 - Summary Haul Roads - Cost Per Running Meter (average 425 mm depth, 6.4 m width) Ballast, geotextile, mine muck / raisebore cuttings
$48
Ballast, geocell, ballast, geotextile, ballast / mine muck
$53
Ballast, geotextile, ballast, geogrid, class 5 road base
$49
Mine Muck with lean concrete fill
$29
Mine Muck without lean concrete fill
$11
Cement Modified Soil with lean concrete fill
$41
Cement Modified Soil without lean concrete fill
$24
High strength concrete road
$52
Road maintenance is performed with two motor graders and a vibratory compactor. Water dripping from the back onto the road is controlled with small sections of brattice cloth nailed to the shotcrete, or for large sections, a "tent" of brattice cloth suspended from the center of the drift and draped out to the ribs. 7 MINE DEWATERING Mine dewatering is done via a series of pumping and settling stages. Water is collected at the 6920 level via drainholes and sump pumps. The collected water is pumped up 35 meters via 56 kW vertical turbine pumps to a slurry tank. Water is pumped from there up 80 meters via two in-line 75 kW horizontal pumps to a solids settling area. The cleared water is transferred to a holding area where it is pumped up 200 meters with 300 kw vertical turbine pumps. Finally, water is pumped completely out of the mine 730 meters vertically with 1,300 kW six-stage horizontal pumps. All mine water is processed through a surface water treatment plant. The average amount pumped from the entire mine is 70 l/s. 8 COMMUNICATIONS FOR THE 7210 PRODUCTION LEVEL. The communications for 7210 will consist of three major categories: Voice communications, the business Local Area Network (LAN) and the Process Control Network. Voice communications will take place in two forms, the telephone system and the leaky feeder radio system. The telephone system will be a standard copper wire phone system. This system will be present in the offices, shops and certain other key spots on the level. Touch tone telephones will also be used to clear the blasting area in conjunction with the mine’s Automated Brass System. The second mode of voice communications for the level will be the leaky feeder radio system. This system will be the
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same as the one used on other levels of the mine and involves the stringing of leaky feeder antenna throughout the working areas. As a result, this will provide radio communications to everyone working on the level. The business LAN will be used for data and ordinary business needs. Fiber optic cable will extend the Henderson LAN to the new 7210 level. This LAN will be extended to the offices, lunchrooms and the shops. This network will be used for e-mail, entering time for payroll, internet access, the warehousing and accounting systems, etc. The process control system is a PLC based control system. This system communicates using a combination of fiber optic and copper cabling to monitor and control various equipment needed to operate the level’s pumping and ventilation systems. Fans, pumps, air doors, CO monitors and other equipment can be monitored and controlled from central computer stations with this system. A radar system will also be used in conjunction with the process control system to monitor the level of ore in the ore passes. This information will be available to the truck haulage dispatcher’s computer for dispatching decisions. REFERENCES • Callahan, M F, Keskimaki, K W, and Rech, W D, 2000. A Case History of Crusher Level Development at Henderson, Proceedings MassMin 2000, Brisbane, pp. 307- 323. • Carten, R B; Geraghty E P; Walker, B M and Shannon, J R; 1988. "Cyclic Development of Igneous Features and their Relationship to High-Temperature Hydrothermal Features in the Henderson Porphyry Molybdenum Deposit, Colorado"; Economic Geology, Vol 83, pp 266296. • Lorig, L, and Brandshaug, T, 1996. Rock Mechanics Analyses of the 7175-Level Mining at the Henderson Mine, Minneapolis, MN: Itasca Consulting Group. • Nelson, B V and Fronapfel, L C, 2002. "Recent Ventilation Improvements at the Henderson Mine"; SME annual meeting, Phoenix. • Rech, W D, Jensen, E B, Hauk, G, and Stewart, D R, 1992. The Application of Geostatistical Software to the Management of Panel Caving Operations, in Proceedings MassMin 92, pp 275-281. • Rech, W D, and Watson, D K, 1994. Cave Initiation and Growth Monitoring at the Henderson Mine, SME Annual Meeting, Albuquerque, NM. • Rech, W D, Keskimaki, K W, and Stewart, D R, 2000. An Update on Cave Development and Draw Control at the Henderson Mine, in Proceedings MassMin 2000, Brisbane, pp. 495-505. • Tyler, W D; Keskimaki, K W and Stewart, D S; 2000. "The New Henderson Mine Truck Haulage System – The Last Step to a Totally Trackless Mine", MassMin 2000, Brisbane, pp317-323, SAIMM.
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The DOZ Mine – A Case History of a mine startup Timothy Casten, Senior Manager of Underground Planning, Brian Clark, Manager Maintenance, Banu Ganesia, Manager Production, John Barber, Leon Thomas, Vice President, Underground Mines Division, P.T. Freeport Indonesia
Abstract The Deep Ore Zone (DOZ) Mine is the latest underground block caving mine to be developed at the P.T. Freeport Indonesia mine site in West Papua, Indonesia. The original feasibility design was for a 25,000 tpd production rate. This was achieved 17 months ahead of schedule and under budget. During this period the mine reserve base was increased and the mine underwent an expansion to 35,000tpd with a 50,000tpd option currently being considered. The mine has been a very successful project for P.T. Freeport Indonesia and this paper attempts to capture some the valuable lessons learned during the start-up phase and on through the ramp-up to full production. High capacity block caving mines are being planned to supply P.T. Freeport’s future production ore, these lessons learned are the keys to successful implementation of new projects. The mine initiated undercutting in November 2000 and has ramped up to a sustainable rate of over 40,000tpd. The mine is an advanced undercut, mechanized block cave, utilizing a truck haulage level and a gyratory crusher. This paper will describe some of the issues encountered during mine development, pre-production and the production ramp-up periods and how they were dealt with. The paper will discuss the key lessons learned in bringing the mine up to full production and describes the methodology employed to reduce activity scheduling interference, improve work quality, how bottlenecks in production were overcome and some of the critical challenges dealt with along the way.
1 INTRODUCTION The DOZ Mine is part of the P.T. Freeport Indonesia (PTFI) mining complex. It is the third vertical lift in the East Ertsberg Skarn System (EESS) deposit. The mining complex began block caving operations in 1980 with the Gunung Bijih Timur – East Ertsberg (GBT) block caving mine. This mine achieved a maximum production rate of 28,000 tpd and was depleted in 1994. The mine started with a gravity slushing system on the extraction level but quickly converted to a more efficient LHD system. The mine used a trench slot post undercutting method. The Intermediate Ore Zone (IOZ) block caving mine began production in 1994 and ramped up to a maximum production rate of 32,000 tpd. This was designed as a LHD extraction system with rockbreakers and grizzlies located at the north and south fringes of the extraction level. The sized ore was loaded onto coarse ore conveyors using vibrating feeders and conveyed to one of two Jaw crushers. The mine began with an El Teniente style layout on the Extraction level but had to convert to a herringbone layout due to wet muck problems. A conventional post undercutting system was used. The IOZ mine was depleted in 2003. The DOZ mine was designed as a fully mechanized block caving mine using LHD’s and a Truck Haulage level. The ore handling system was designed to handle the largest particle that the 54" x 77" crusher would accept. This was accomplished by the use of a 1 meter grizzly on the extraction level, 4.1m diameter orepasses from the Extraction to the Truck Haulage level, chain gate style chutes and 55 tonne capacity trucks direct dumping into the crusher. 404
The DOZ Mine commenced pre-production development in 1997 and initiated caving in November 2000. The feasibility study called for a production ramp up to 25,000 tpd by January 2004, a little more than a three year production ramp up period before achieving steady state. During the development and pre-production periods exploration drilling continued and the orebody was expanded.
Figure 1: DOZ and ESZ orebody reserve growth Operational ore demands from the rest of the PTFI mining area challenged the production ramp-up plan and required that the DOZ Mine achieve 25,000tpd by September of 2002, nearly a year and half ahead of the original schedule.
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The additional ore reserves and production demands also made an expansion to 35,000tpd economically viable and this expansion occurred between October 2002 and March 2003. The DOZ mine produced at an average of 40,500tpd in 2003 to meet additional ore demands for the operation. As of writing this paper the DOZ has produced at a peak of 58,800 tpd with an average YTD production rate of 45,000 tpd. Table 1: DOZ/ESZ Reserves as of 2003 Mine
Tonnes (Millions)
Cu (%)
Au (g/t)
Ag (g/t)
DOZ
184
0.96
0.65
5.11
ESZ
121
0.54
0.90
1.64
In 2000 an adjacent orebody named the Ertsberg Stockwork Zone (ESZ) was added to the reserve base. The ESZ is located on the south west side of the DOZ orebody. The two orebodies are geologically distinct but physically contiguous. Figure 1 shows the DOZ and ESZ reserve growth since 1995. Table 1 shows the current ore reserves for the DOZ and ESZ mines. A study is currently underway to include the ESZ into the DOZ mining sequence and increase the average production output to 50,000tpd. This paper examines some of the key lessons learned in the DOZ mine startup period and focuses on some of the critical bottlenecks encountered and the methodology applied to successfully overcome them. The intent of this paper is to focus on the key issues in some detail while giving mention to the other project drivers. 2 GEOLOGY AND GEOTECHNICAL The geology of the EESS and the DOZ Mine has been described in detail by others (Coutts, 1998). The key issue to note in the DOZ is the extreme variability of rock types encountered in the ore zone across the strike of the orebody from the footwall to the hangingwall. Starting at the north (hangingwall) and proceeding to the south, ground conditions change from very poor to very good. Rock strengths (UCS) vary from a high of 219 MPa in some
Figure 2: DOZ Simplified Rock Quality – Undercut level massive magnetite to less than 10 MPa in the DOZ breccia. RMR varies from a low of 25 in the poorest ground to a high of 65 in the most competent ground. In order to plan and design for this variability in rock quality a series of simplified plans were produced. These plans formed the basis for the design and scheduling of activities for the development, pre-production ground support and caving teams. Figure 2 shows a typical plan view of the east side of the DOZ. The mine has been divided into three rock quality categories; good, medium and bad. This simplification allowed planners and supervision to look ahead and prepare for a change in ground support, caving rates etc. 3 KEY ISSUES AND LESSONS LEARNED The IOZ mine was in active production during the feasibility, design and development phase of the DOZ mine. Several lessons learned from the IOZ mine were applied to the DOZ mine design. In addition, new technologies to the operation were applied in the DOZ such as advanced
Figure 3: Section Looking South – Undercut Transition Massmin 2004
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undercutting and improved QA/QC systems. This section discusses some of the key changes applied to the DOZ during the ramp-up process. 3.1. Caving Methodology The DOZ mine initially employed a conventional post undercutting system during cave initiation. The post undercutting method had been used in the IOZ Mine and was familiar to management and crews alike. The post undercutting system had caused stress damage problems in the IOZ resulting in the premature closure of some drawpoints and additional repair activities. In addition to the post undercutting method the DOZ also initially incorporated a 10m vertical stagger to accommodate an existing exploration drift. This was driven at grade along the strike of the orebody to help prove up the reserve by adding drilling stations. The grade of the drift was such that it intersected the planned extraction and the undercut levels as it progressed from West to East. Figure 3 shows this development in section, looking south. The cave was initiated between Panels 13 (P13) and 14 and advanced to the east. Early on in the undercutting sequence at DOZ, management decided to change to the Advanced Undercut method. This change resulted in a significant reduction of stress related damage to the DOZ Extraction levels and has been a major factor in the reduction of panel and drawpoint repair requirements. The layout of the Extraction and Undercut levels to incorporate the existing exploration drift and change in undercutting methodology required significant design and construction work. The extraction and undercut level elevations were set to match the gradient of the exploration drift where possible. This resulted in a series of split level drawbells between P15 and P16. Significant engineering and caving resources were required to open these drawbells. The success rate of the staggered drawbell blasting was fairly poor with numerous remnant crown pillars that required remedial drilling and often incurred significant delays (Sinuhaji). Changing extraction elevations and undercutting methods could have resulted in a significant delay to the mine production ramp-up as the cave moved east. Although the caving rate was slower than planned during this period the DOZ caving team were able to manage the transition safely and successfully. Since November 2000, over 85,000 square meters of undercutting have been accomplished with no accidents. The issues encountered in this transition and the methods applied to overcome them helped formulate the scheduling rules of thumb discussed in section 4 below. 3.2. Quality Control The quality control on ground support installation and concrete placement was not under an independent QA/QC group in the IOZ. The post undercutting system applied considerable abutment stress to the opened panels. This ground weight and poor quality resulted in some drawpoints and panels failing prematurely and required them to be repaired earlier than anticipated. The DOZ mine has a dedicated team of QA/QC engineers and technicians who monitor the development and pre-production ground support process on a daily basis. Having the Field Engineering group in place at the start of the pre-production process was critical in reducing short term re-work issues required due to poor materials or practices and has improved the long term quality of ground support and development significantly. Having field engineers working directly with operations on a daily basis has greatly assisted in the successful completion of several critical projects such as the main crusher installation, chute modifications and undercut and drawbell blasting issues. 406
Figure 4: IOZ/DOZ mine lintel set comparison 3.3. Pre-Production Ground Support A review of the key problem areas was undertaken in the IOZ, and the drawpoint design and construction was identified as an area that could be improved. The lintel sets were redesigned by the onsite central engineering group and were fabricated in the operations Light Industrial Park (LIP) rather than being built as required by the underground shops. Figure 4 shows a comparison of the IOZ and DOZ mine standard lintel set designs. Closing an active drawpoint for repair requires the panel section to be closed for a significant period of time, potentially impacting access up to 8-10 other drawpoints. This can often result in additional damage to the panel as the drawpoints are not being pulled and the area begins to take weight from the cave. The other active drawpoints are typically pulled harder to maintain tonnage calls resulting in further uneven draw. The new lintel sets are fabricated with shear studs to tie them to the formed concrete and with pockets that allowed for bolts to pin the sets to the ribs. The bolts have been very effective in preventing the posts from buckling and kicking out under load. The bolts and shear studs have also protected the lintels from secondary blasting damage and impacts from the loaders. In practice the new lintel sets have performed extremely well. To date three drawpoint lintel sets have required repair out of 270 opened drawpoints. 4 PRODUCTION RAMP UP ACTIVITY SCHEDULING The Extraction Level posed the greatest problem in short term scheduling for the numerous activities planned in the area. Once production was initiated in a panel, production LHD’s and secondary breakage units as well as drawbelling drills, pre-production ground support crews, construction crews and development crews would require access through the same area. As each panel only has two access points, every intervention by a separate crew caused a safety risk and disruption to the other working crews. With the application of an advanced undercut system the required activities by different crews on the Undercut Level increased and posed a similar scheduling problem. The scheduling of activities became a trade off between a Just-In-Time method and an Early Completion method. In
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the JIT method tasks were deferred as late as possible and a series of crew interventions were required to be managed in the panel and undercut areas. This deferred capital expenditure but increased the risk of slipping the schedule. In addition, the advanced undercutting method requires that development and construction of drawpoints and drawbells be delayed as long as possible to improve the stress distribution of the abutment. This is the typical methodology applied to most projects but also tends to have the least flexibility during the start-up period. The Early Completion method reduces crew interactions by bringing forward activities and separating work areas, effectively decoupling sets of activities. This tends to accelerate activities but reduces risk. The JIT method was found to be not flexible enough to deal with delays or issues such as unexpected ground conditions, material delivery delays or re-work issues. At mine start-up there are typically very few working areas available on the Extraction and Undercut Levels. Any delays or loss of access tend to have a domino effect on the crew activities and overall schedule. A small cave front with limited alternative working areas places every undercut ring and drawbell on the critical path. The savings generated by deferring activities were soon consumed by the lower availability of working places, resultant crew inefficiencies and overall schedule slippages. The ability to move crews to alternative working places as required greatly improved overall productivity and allowed the mine to accelerate production. Although this may have accelerated some capital this was more than offset by the accelerated production that resulted. By accelerating development and pre-production activities, the opportunity for faster drawbell opening and undercutting was realized. The ability to keep all of the crews working in an available heading, with minimal interruptions, ultimately drove the accelerated production ramp up. It was recognized that the price of delay far outweighed the cost of an Early Completion methodology. Although the exploration drift discussed previously presented a significant design and construction hurdle it did effectively split the undercut level of the mine into two sections. This presented an opportunity to defer some undercut development by utilizing this existing access. The splitting of the undercut into manageable sections allowed the crews to focus on the critical path areas while also providing a multiple work place environment. Although not originally planned for the West side of the mine the success of the East side split has driven the change in design for the
Figure 5: Undercut level showing existing and planned mid access drifts and deferred activities. Massmin 2004
West. Figure 5 shows the planned undercut level designs with mid access drift. This methodology was carried down to the Extraction level with the production panels being split into two sections typically at the centrally located orepass. The split is accomplished by driving the drawbell drift through to the adjacent panel and creating a mid-fringe drift for access, ventilation and to multiple activities in a single panel. The development of the drawbell drift ahead of the undercut is an out of sequence event for the advanced undercutting process, but the benefits of this mid-access more than covers the cost of additional ground support required to compensate for the open drift. A set of guidelines were produced to assist in the mine scheduling process: Extraction Level • Panels are split into two sections, North and South, typically at the centrally located grizzly tip. • A mid drawbell drift is chosen and driven between panels. This effectively allows each panel section to have two access points and allows for two concurrent activities per panel with minimal interference between crews. • Separate panel sections to contain development, preproduction ground support and construction, caving and production activities. • The panel section schedules are driven by the drawbell opening sequence. • Cave advance is based on the maximum lead/lags allowable in the differing ground types based on Figure 2. • The draw belling must be completed after one month of the undercut completion to prevent packing of the blasted muck in the undercut. Undercut Level • The drill drift level development is split into two sections using a mid-access drift. • This created a multiple heading environment for the development and caving crews compressing schedules by deferring non-critical activities. • The splitting of the undercut also deferred capital by sectioning off longer term development. • Pre-drilling and pre-charging of undercut holes is utilized to reduce activities around the open brow and improve undercutting efficiencies. Figure 6 shows the production ramp up of the DOZ Mine in tonnes per day by month. There is a clear change in production rates after about the first 16 months of production. In this first phase the mine ramped up at approximately 700tpd/month. This slower rate reflected a smaller cave front and limited working areas. Numerous crews working in the same locale caused interference and delay. It also reflects the typical mine start-up issues normally encountered such as the problems with the chutes and the transition from advanced to post undercutting as discussed in this paper. The second phase shows a ramp up rate of 1900 tpd/month up to an average monthly rate of over 45,000 tpd. Three main factors were responsible for this change in production rate; 1) the de-bottlenecking of the caving areas by the application of the guidelines discussed above, 2) the re-design of the ore handling chutes and, 3) the completion of the transition from the post to the advanced undercutting method. The goal of any ramp up is to compress the start-up period as much as possible. Although the different key production drivers encountered in the DOZ are discussed within this paper, the ability of the project team to recognize and react to these issues was the real key in exceeding the planned production rates.
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5 ORE HANDLING SYSTEMS The ore handling system in the IOZ utilized rockbreakers at the north and south fringes of each panel. The LHD’s dumped ore onto 0.4m x 0.4m grizzlies, sizing it for feeding onto coarse ore conveyor belts that transferred the ore to one of two jaw crushers. During the design phase it was anticipated that a coarser ore would be found in the DOZ mine than was evident in other block cave mines within the PT Freeport contract of work. Several options were considered on how to handle the larger fragmentation and ultimately a ‘coarse ore’ handling system was decided upon and designed into the system. The final option selected was to use one meter spaced grizzlies on the extraction level on top of a 4.1meter diameter raise. A single bored orepass was designed for each extraction level panel which services up to 20 draw points. At the bottom of the 35 to 45 meter long raise a 2.4-meter wide center loading chute was installed which loads into Elphinstone AD55 haul trucks. The trucks then dump to a 54" x 77" gyratory crusher. The ore passes, when kept three-quarters full, were designed to give up to 800 tonnes of storage per orepass. With up to 15 ore passes in production during peak production considerable tonnes could be ‘stored’ which would allow the mine to compensate for disruptions in the system. This storage was also envisioned to enable the extraction level to be disconnected from the truck haulage level. The DOZ Mine began production in 1999 with three installed chutes or LP’s. During the initial period of mine startup several issues became apparent with the chute installations that were significantly reducing the flow of ore and acting as the main bottleneck to production. Ramp up production targets were not being achieved and considerable damage was being done to the existing chute installations in order keep the ore flowing through the chutes. The DOZ ore material ranges from blocky, hard diorites to soft, sticky breccias. The ore types behave significantly different from a handling point of view. The mine was initially undercut in the high grade breccia areas which resulted in large amounts of sticky ore being delivered to the raises and chutes without the ability to blend with coarser material. Being identified as the highest priority obstacle to reaching production goals considerable effort went into redesigning the chutes. A campaign of chute re-design was embarked upon by the underground maintenance, construction and engineering groups to overcome this
Figure 6: DOZ Mine production ramp-up 408
obstacle. Many variations of the standard chute design were tried along with some unique solutions. There were three major problems with the original chute design that needed to be overcome; 1) the chute structural design, 2) the chute control systems and, 3) the ore flow through the chute due to sticky ore packing in the top of the chute. 5.1. Material Flow Issues The biggest obstacle to overcome in increasing production was the sticky ore encountered early on in the mine start-up from the high grade DOZ breccia material. Tests revealed that the angle of repose of the sticky ore was at a minimum of 60 degrees and could be as high as 90 degrees (Jenike and Johanson). Although the raises were bored at an inclination of 75 degrees and the chutes were inclined to 45 degrees in the fully open position the sticky material quickly compacted and blocked the raises at this junction. This required a campaign of hang-up blasting in the lower section of the raise that often resulted in significant damage to the chutes. High pressure water hoses were tried at the chutes but in some cases made the problem worse by adding water to the breccias and marble ore types making them stickier and harder to handle. This created a cleanup problem around the chutes and occasionally caused the crusher pocket to plug and hang-up. In the short term the transfer raises were run empty to prevent the sticky muck from compacting and hanging up. This removed the approximately 800 tonnes of storage from the orepasses and trucks had to queue at chutes waiting for a full load. This impacted the achievable production rate significantly. Additional damage was also seen at the chutes due to the ore impacts caused by the empty raises. Many different vibrator configurations were tested to help vibrate the packed ore loose but these met with limited success. The chute vibrators also resulted in increased damage to the entire structure and were not considered to be a long-term solution. The concept of floating the floor was developed so that the rest of the chute structure would be isolated from the vibration. With the concept of turning the chute floor into a vibrating plate a chute was selected for trial. The floor was detached from the chute and suspended with the use of heavy springs placed around the floor. The exciter mechanism was attached to the floor bottom and connected to an electric motor by V-belts. In effect, the chute was converted into a vibrating feeder with the floor at a 45 degree angle. The modification turned out to be an immediate success so all the existing chutes were modified while a longer-term design was finalized.
Figure 7: Cross section of DOZ chute showing design modifications Santiago Chile, 22-25 August 2004
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5.2. Chute Structure RE-Design The chute design also exposed several key header beams to the muck flow. Although these had been armored with liner plate the repeated impacts seen from running the raises empty resulted in accelerated wear and loosening of these header beams. Several of these beams made it to the crusher, jamming it, and requiring extended shut downs to remove them. Chute downtime for repair was also considerable. Widening the chute and elongating the floor helped overcome the ore flow and structural design issues. Different widths were tested, with the optimum being 3.1meters. This width moved the sides of the chute out of the flow sufficiently enough to alleviate damage and stop bridging of the ore. Floor elongation was accomplished by excavating a hitch under the footwall of the raise. The chute floor was extended underneath the footwall into this hitch. A secondary support set was put at the back of the hitch to prevent erosion of the raise. The head beams were then recessed into the hitches, protecting the beams, improving chute down times and reducing transient steel feed to the ore flow system. 5.3. Chute Control Systems A hydraulic system had been designed to move the chain gates and lips and control the flow of muck. In order to save money these were set up so they could be shared between two chutes and would switch across chutes using PLC systems. The hydraulics proved to be difficult to install, impacted two chutes if they were out of operation, and required considerable maintenance. As the hydraulic cylinders tended to hold the chutes quite rigidly they were easily damaged by rock falling onto the chute through the open raise. The hydraulic chutes were replaced with pneumatic units that were quicker and cheaper to install, easier to maintain, and had a more flexible response to impacts. This reduced the control equipment cost by almost 85% and installation time was reduced to three days versus three weeks. The systems have proved to be very dependable with little or no downtime. An air intensifier is used on the system to provide sufficient power for the ‘lip up’ function. This air-toair intensifier has proved reliable and boosts the compressed air by 50% of operating parameters to ensure adequate power is available to raise full chute lips. As new chutes were constructed these changes were made in the field and the results monitored. The campaign was successful and currently transfer raises are run full and the chutes are performing well. Figure 7 shows the key areas modified to improve the chute design. 6 CONCLUSIONS PT Freeport Indonesia has been block caving at the East Ertsberg Mining district since 1980. The DOZ Mine is the third lift in the sequence. The operation has ramped up the DOZ mine faster than initially anticipated and has incorporated lessons learned from the previous mines as
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well as new technologies. The DOZ mine achieved full production 17 months ahead of schedule and was under budget. The mine is still undergoing expansion and is studying a 50,000tpd mining rate. Every mine start-up encounters challenges and unexpected issues. The mine start-up environment also often presents unexpected opportunities. The management, engineering and operations team must be flexible and reactive enough to cope with these setbacks as well as leveraging any opportunities presented. A block cave mine start-up is a series of interlinked activities that often require close coordination and scheduling. The initial nature of a block cave often places individual undercut rings and drawbells on the critical path. The ability to decouple these activities and allow the crews to work with minimal interruptions in a multiple heading environment is a major driver in growing the cave and accelerating the production ramp-up process. At 40,000tpd the DOZ mine is one of the worlds largest block caving mines. The continued success of the DOZ Mine is critical to PTFI’s future operations. As the Grasberg Open Pit finishes mining in 2015 the operation will be supplied by ore from underground mining operations for the next several decades. Future block caving mines at the PTFI mining complex are currently being planned to produce at rates over 100,000tpd. Block cave mines are capital intensive projects requiring high expenditures in the early years of development. A large component of a block caving mines value is driven by the speed at which it can begin producing at full capacity. The ability quickly bring into production large capacity block caving mines is a key driver for the PTFI future operations. Successful projects like the DOZ not only provide ore to the mill but generate confidence for the future of the operation. ACKNOWLEDGEMENTS The authors would like to thank PTFI Management for allowing the publication of this paper, Rod Mainland and the DOZ Field Engineering Group for their work on the QA/QC systems and Rodger Bonney and 68 Engineering for the improved structural designs. REFERENCES • Coutts, et al., "Geology of the Deep Ore Zone, Ertsberg East Skarn System, Irian Jaya", PacRim 99, Bali, Indonesia, October, 1999. • Casten, Sinuhaji, Poedjono, Flumerfelt, "The Application of Advanced Undercutting at P.T. Freeport Indonesia’s Deep Ore Zone Mine", Operators Conference, Townsville, Australia, 2002. • Jenike and Johanson, Internal Report on Ore Material Characteristics, 2002. • Barber, Thomas, Casten, "Freeport Indonesia’s Deep Ore Zone Mine", MassMin 2000.
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Undercutting at E26 lift 2 Northparkes AC (Tony) Silveira, Senior Mining Engineer, Northparkes Mines
Abstract Northparkes Mines completed the undercut on the Lift 2 block cave in January 2004. This paper discusses the designs considered at the time of the feasibility study, and the various changes made to these designs prior to and during the undercutting operation in the Lift2 block cave. This paper also highlights operational issues encountered whilst mining the Lift 2 undercut and ways in which these issues were overcome. A summary of critical success factors is also included.
1.0 INTRODUCTION
million tonnes per annum are expected from Lift 2 over the next six years.
Northparkes Mines (NPM) is located in New South Wales in Australia, approximately 350 km to the northeast of Sydney. Production presently comes from the E27 open pit and the E26 underground mine. The E26 underground mine is the first mine in Australia to employ the block caving method of mining. Extraction of cave ore commenced in 1997 from the Lift 1 extraction horizon, some 480 m below the surface.
1.1 The Lift 2 Undercut The Lift 2 undercut is a narrow inclined advanced undercut or continuous void, extracted for the purpose of inducing caving directly above it. Some of the critical factors related to the extraction of this undercut include: • No remnant pillars are to be left since they could act as transfer points for stresses from the Lift 2 block, directly onto the extraction drives located below. • Creation of the stress shadow. In this shadow region, further development of the extraction level drives can be progressed. • The progress of the Lift 2 project hinged on the advance rated achieved in the undercut, and preparation of the extraction level for production.
Figure 1: Section showing layout of the Lift 1 & Lift 2.
Ore above the Lift 1 extraction horizon has been exhausted and the mine has developed a second, lower production zone, known as Lift 2. The Lift 2 extraction horizon is located some 350 m below that of Lift 1 (Figure 1). The Lift 2 Undercut, which is the initial stage of setting up for production in a block cave, is located some 10 m above the Lift 2 extraction level. Planned production rates of five 410
Figure 2: Plan view showing undercut drives with the access drive on the east and the slot drive on the west.
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The Lift 2 undercut consisted of 14 parallel drill drives running from east to west. All these drives were linked together on their eastern end by an access drive, and on their western end by a slot drive (Figure 2). Areas between the 14 drill drives were designed to be mined out as inclined or flat portions of the undercut, generating an overall crinkle/corrugated shape (Figure 4). The inclined portions of the undercut were designed to provide major apex pillars above the extraction level drives, while the flat portions of the undercut lie directly above the drawbell crowns. This is similar to that used at Palabora mine in South Africa. 2.0 UNDERCUT DESIGN The main design changes to improve operational, safety and cost aspects are covered in this section. The initial Lift 2 undercut design consisted of 14 parallel drill drives only connected at the flat pillar portions. A major change made to this design involved putting in a western slot drive (shaded section in Figure 2) that linked all the drives together on the western perimeter of the undercut. This western drive allowed early access into some of the western ends of the undercut drives, as well as making slot opening easier as the eight slots were now designed as vertical slots (previously inclined). The undercut drill pattern consisted of 89 mm diameter blast holes in a combination of long incline, short incline and dual flat rings. All these rings were drilled at an apparent ring burden of 2.1 m (true burden of 2.0 m). The initial undercut slotting design proposed at feasibility consisted of a series of 14 inclined slots drilled in a meshing manner above the major apexes on the extraction level (Figure 3). Mining of inclined slots is an inherently difficult task. The development of the western slot drive led to a reduction in the number of slot rises (from 14 to 8), as well as the elimination of inclined slots altogether. This reduced the risks associated with mining inclined slots. During the feasibility stage, the long and short inclined rings consisted of fanned inclined holes with five to six holes per ring (Figure 4). The angle of the major apexes (i.e. the angle of the flattest inclined holes) was designed to be 50 degrees from the horizontal. This was to ensure that the fired dirt cleared the apexes and flowed into the drawbells. This angle was increased to 54 degrees. The biggest change made to the undercut inclined ring design involved reducing the number of holes from six per ring to a final design of three per ring. The drill design went
Figure 3: Planned slot designs Massmin 2004
Figure 4: Planned ring designs through several stages of review and the decision to go to three holes per ring was made after feedback from Palabora mine indicated that it was not necessary to drill and fire the brow portion of the inclined pillar holes above the flats. This area fell in on a regular basis without being drilled or blasted. Palabora at the time were using four 64 mm blast holes on incline ring pillar holes, and four 76 mm blast holes on the flat pillar ring holes. NPM opted for an 89 mm sized blast hole pattern leading to: • Cost savings with fewer holes drilled and charged (only three per ring). • Reduced redrills due to ease of loading primers into hole, even in partially dislocated ground. • Time saving with quicker drilling and charging. • Greater face velocity of the fired dirt, throwing it further into the cave, thereby reducing undercut bogging. • Better fragmentation due to higher powder factor. • Reduction of potential hole closures, dislocations and redrills. This last point was a major success and was also a result of the quick turnaround maintained on all the undercut faces, where no face was kept standing for any extended period of time. On average, each undercut face was fired once a week. 2.1 Inclined Ring Design Figures 5 and 6 below highlight some of the changes made to the inclined rings design. Shaded portion "A" in Figure 5 was the region into which the long inclined rings originally extended. All the drill rings were designed and eventually drilled at a forward dump/tilt angle of 20 degrees. The reasons for the 20 degrees forward tilt angle were: • Reduction in the possibility of loss in overall void height during firing. The firing direction of the toes of the holes of forward tilted rings, helps reduce the likelihood of pillar formation at the toes. • Keeping the falling rocks (if any) away from reporting to the open brow, as the ring was tilted forward. • Better packing of the fired dirt onto the previously fired rill in the cave, due to the directional throw due to forward tilting. • This would thereby reduce the amount of swell reporting to the brow, thereby reducing the swell removal required. The design change made involved flattening of the apex of the inclined void, reducing the drill metres by approximately 1.5 m per long inclined ring. This apex region was a "tight corner" in the firing sequence of the ring, at the time. This last metre was also the section where possible deviation in the longest holes would be most prominent. The number of holes per ring was also reduced from four to three per ring.
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• Ease with drilling and blasting i.e. smaller length holes. • Reduction in the risk of pillar formation due to hole deviation. • Ease of bogging, as access is available from both sides. • Minimal tele-remoting required (avoiding delays and cost). • More flexibility with drilling headings i.e. more drill headings to work with. • Potential reduction in the amount of misfires in the flat pillar region. This dual flats pillar comprised of 3 x 5 m long holes drilled from either end and fired within the same shot (Figure 7). Each flat ring had a jumbo stab hole in the floor to enable ease with bogging of the fired dirt. After mining about 20 % of the undercut it was found that these jumbo stab holes were no longer required as the lowest hole in the flats performed adequately by ensuring breakage of the ground. Figure 5: Initial changes made to the planned inclined rings design
Figure 7: Initial dual flats ring design Modelling at the time indicated that as the undercut progressed, hole dislocation could occur at the collar and toe of the drill holes if a half flat pillar was excavated. Modelling also indicated that hole dislocation would only occur at the collar if the full flat pillar was excavated. In reality however, no such problems were encountered with the excavation of the full flat pillar. 2.3 Slot Design A total of eight "6-reamer" up-hole vertical slots, of lengths varying from 11 m to 14 m, were successfully fired to start the initial firing faces within the undercut (Figure 8).
Figure 6: Final inclined rings design 2.2 Flat Ring Design During initial development of the drill drives, it was found that additional wall overbreak had the potential to reduce the pillar width. Following a geotechnical evaluation, the decision was taken to increase the pillar width from 8 to 10 m to ensure stability of the flat portion of the undercut during undercut retreat. Palabora mine at the time were using approximately 13 m flat pillars. This flat pillar was fired in two separate blasts, with a lead/lag between the 8.3 m and 5.5 m sections. NPM however evaluated the geotechnical and production risks of a lead lag within a 10 m flat pillar, and opted for a "dual flats" pillar. This involved drilling the flat pillars from both adjacent drill drive accesses with 5 m long holes (Figure 7), instead of the single flat 10 m holes. All the drill rings were designed and eventually drilled at a forward dump/tilt angle of 20 degrees. The reasons for this 20 degrees forward tilt angle were ease with bogging of the fired dirt around the corner (turning radius of the LHD) and better visual of the fired toes from a safe location (after firing). Bogging is a term commonly used in Australia to describe rock removal by an LHD. It is also referred to as lashing (South Africa) and mucking (USA). The flat portion was initially designed to have three holes per ring, with these flat sections designed as single flats with the holes being 10 m long. Changes were made to the design wherein the risks of single flats with 10 long holes, were evaluated against the use of dual flats with 5 m long holes. NPM adopted the dual flat rings for a variety of reasons some of which include: 412
Figure 8: Pattern for rises/slots
Santiago Chile, 22-25 August 2004
Massmin 2004
The 6-reamer slot pattern is a pattern widely used in many mines for its forgiving nature with respect to drill hole deviation. Drilling of the slots is a critical part of the operation. Here a single experienced driller was given the responsibility of drilling the inner box holes of this rise. Lower feed and rotation pressures were required for drilling of the initial eight box holes of each slot. Lead guide tube and T-51 rods, drilling 89 mm shot holes and 152 mm reamer holes. The drill string combination used drop centre "Retrac" drill bits to achieve better accuracy. "Retrac" bits have a partial cutting action when the drill rods are retracted from a drill hole, due to the serrated edge on the tail end of the bit. After firing the vertical slots and using CMS (Cavity Monitoring System) to confirm the void, a series of inclined holes were fired to open up the void to the required excavation profile. The design for these slot-stripping rings used on the first four slots, consisted of four rings on either side of the slot rises. While firing the initial slot rises it was found that quite often the first slot stripping row was lost due to over break from the rise firing. After four of these rises and slots were successfully excavated, the number of slot stripping rings was dropped from four to three on either side of the slot (Figure 9).
Figure 9: Design for slot stripping rings showing extraction level located below
3.0 OPERATIONAL ISSUES This section describes some of the operational issues encountered and the ways in which they were overcome. 3.1 Drilling The contract between the owner and the contractor specified a tube drill string requirement to ensure drilling accuracy. In order to maintain drilling flexibility with the two Solomatic 720 drill rigs on the undercut and extraction levels, the following were agreed on: • Undercut level (89 mm holes) would use T51 speed rods and drop-centre 89 mm Retrac drill bits. • Extraction level (64 mm holes) would use T38 speed rods and drop-centre 64 mm Retrac drill bits. • Use of 700 Series drifters on both drill rigs. 3.2 Hole Preparation No individual was allowed access past the last flat ring fired so it was not possible to physically be under the collars of the inclined rings at the brow (to clean and prepare them for charging). A 1.5 m bund wall was erected at the brow where the last flat ring had been fired. This bund was put in place to protect personnel from injury from any rocks falling from the brow area or from within the undercut void. Figure 10 below illustrates the access restrictions to personnel at the undercut face. Massmin 2004
Figure 10: Plan view showing Brow Access Limits
3.3 Charging and Firing Charging and firing activities were carried out using a specialist explosives contractor’s charge-up personnel. Two people were used to get this task completed within the time frame of 3 hours required. On average the time spent on charging two rings (two sets of long and short inclines and flats = 126 m2) in two headings was 2.5 to 3.0 hours. As the production rates were ramped up later in 2003, two shots (four headings = 252 m2) were quite regularly blasted at the end of a shift. All the undercut rises were charged with IKON electronic detonators. The precise timing and intrinsic safety benefits of these IKON detonators made them suitable for the rise firing application. Their cost however, made them uneconomical to be used on standard undercut ring firings as the Enduradet detonators are available at approximately 25% the cost of the IKON detonators. The Feasibility Study considered single ring firings for the entire undercut. After analysing the benefits of multiple ring firings, a decision was taken to carry out all undercut ring firings as "2- ring" firings instead. Each firing comprised of two dual flat rings, two long incline rings and two short incline rings. Some of the benefits of 2- ring firings include: • Reduction in the number of times that charging of holes needed to take place in the undercut. • Reduction in the exposure of personnel to poor brow conditions and hazards at the undercut face. • Better throw characteristics of the fired dirt. Feasibility assumed the use of a "square" brow, or no lead/lag between the inclined and flat rings. Geotechnical modelling was then carried out and confirmed that this could be increased to two rings. The two ring lead lag was then adopted and used for a variety of benefits some of which include: • Improved stability at the brow area. • Less dirt required to be bogged from the area adjacent to the collars of the flat holes. • Reduction in exposure of personnel to the brow area when charging the flats, and the undercut in general. Production drilling and blasting in the undercut proceeded with no major setbacks. The use of a robotic arm for charging helped increase productivity and this is reflected in the increased undercutting rates achieved. The introduction of this machine, coupled with completion of development mining in other areas of the mine, resulted in an increase of LHD and other resources available to the undercut. This all contributed to the improved production rates from October 2003 on.
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3.4 Swapping of Incline Rings After commencement of the undercut blasting, it was noticed that the long inclined holes, which lagged behind the short inclined holes and were damaged from adjacent firings (see DD6 & DD7 in Figure 11). DD7 shows a lag firing of the long inclined rings after a lead firing of short inclined rings has gone off in DD6. The right hand side shows a lag firing of the short inclined rings in DD9 after a lead firing of long inclined rings has gone off in DD8. Some of the advantages of swapping the inclined rings included: • Opportunity to confirm breakage at the major apex after firing the long incline holes in DD8, DD10 and DD12 through the short inclined holes and confirming breakage into the void. • Less likelihood of experiencing hole dislocations and damage in the long incline holes. • The swapped long incline holes pulled the ground effectively providing the correct height and void size.
Figure 11: Ring Layout DD7 & 8
4.0 SUMMARY The Lift 2 undercut drill and blast design has proven to be a successful design. NPM has safely and efficiently completed the undercut within a period of 11 months (February to January 2004). During this period a total of 65,457 metres were drilled (planned metres totalled 64,754 m). Less than 1000 metres or 1.5% of the total metres drilled in the undercut were re-drill metres due to hole dislocations/closures or angle errors. A total of 33,768 square metres were blasted in the undercut, with the advance rate peaking at 6,101 square metres per month in December 2003. An average of 45% of the fired dirt was removed from the undercut, which was significantly below the 60% forecast.
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4.1 Critical Success Factors The following aspects are considered to be key areas in the success of the Lift 2 undercut: • The use of the 6-reamer vertical rise/slot (Figure 8) with the benefits listed (slot design section). • The use of three slot stripping rings on either side of this slot, (Figure 9). • Keeping the short inclined rings as the lag firing and the long inclined rings as the lead firing coupled with the use of dual flats in the flat portion of the undercut. Drilling 89 mm holes in these flats making the jumbo lifter holes redundant (Figure 7). • Use of one ring lead-lag in the dual flats regions in order to reduce early failure of the flat belly portion. The "2-ring" lead-lag used in the drives, between the flat and incline allowed face shape to be maintained whilst providing some flexibility in activity. • Choke firing of the undercut rings with opening the brow every four firings limited personnel exposure to an open brow. It also helped reduce the quantities of fired dirt bogged. Opening the brow for void confirmation was essential. • The specialist explosives contractor using a robotic arm charge up machine contributed to increased productivity and improved safety. • Strict adherence to the lead lag rules ensured no geotechnical problems during mining of the Lift 2 undercut. Good ground support systems ensured stability of excavations and the shotcreted undercut drives all stood up well during drilling and blasting operations. • Steady advance rates reduced the possibility of stressinduced damage. By not leaving the undercut standing for any extended period of time NPM were able to achieve significant targets in a safe and efficient manner. • A good Quality Assurance programme and constant supervision contributed a great deal to the success of the Lift 2 undercut. The lack of experienced operators was an issue during initial stages of the undercut. Engineering staff maintained records on swell removal and ensured that continuous void confirmation is carried out as and when planned. • Provision of adequate resources, e.g. service crews for fixing of sprays, ventilation and assisting with hole prep work ensured that drilling, charge-up and swell removal were completed in a timely manner. • All undercut development, as well as drill and blast designs, were done in-house by NPM engineers and technical specialists. ACKNOWLEDGEMENTS The author wishes to thank those personnel who have contributed to this paper especially those involved with the design and mining of the Lift #2 undercut. The author would also like to thank Northparkes mines for permission to publish this paper.
Santiago Chile, 22-25 August 2004
Massmin 2004
LHD versus mechanized grizzly in III panel of Andina Augusto Aguayo, Claudio Campos, Manuel Mansilla, Jorge Sougarret, Andrés Susaeta Andina Division, Codelco Chile
Abstract The secondary ore of the III Panel of Andina Division is being mined concurrently through a mechanized grizzly and a LHD layout. The design criteria as well as layouts and materials handling system are presented. The retro analysis of 7 years of mining results with a comparison between methods regarding recovery, dilution, stability, costs, etc. is presented in a quantitative and qualitative way. Results show that grizzlies have been a much better system than LHD, from the economic and technical standpoint. Weakness of both methods are described. The analysis intends to project the advantages of the full gravity system (grizzlies) over intermediate hauling systems (LHD). The potential projection of this experience to future designs in more competent rock is considered, with concepts of continuous mining and preconditioning techniques.
1. GENERAL CONDITIONS OF ANDINA MINE
mostly in primary rock, mining secondary and grizzlies in secondary rock.
The Rio Blanco Ore deposit is located at an altitude of 3240 m (undercut level) in the Andean range approximately 50 km to the north of the city of Santiago. The Mediterranean climate is characterized by cold and rainy winter seasons and hot and dry summer conditions. The average snow downfall over the crater is over 10 meters. Total underground tonnage is of 45.000 ton/day. The mine has been caved since 1970 in three different panels. The actual III Panel is being mined by grizzlies and LHD concurrently. The LHD layouts are located
2. DESIGN PARAMETERS Table N°1 presents the principal design parameters of the LHD and Grizzly layouts of the Andina III Panel. These designs have been in operation for more than 8 years. Figure N°1a and b shows a typical design plan of the LHD system and in Figure N°2a and b the III Panel Grizzly design is presented.
Table N°1 - Design Parameters Parameters Rock - Fragmentation Assessment - Primary - Secondary ore column - Lithology Design Geometry
Grizzly
LHD
3% > 1 m3 0m 220 - 300 m Granodiorite, Andesite, Magmatic, granodiorite porphyry Riolite
10% > 1 m3 0 - 120 m 220 - 450 m Granodiorite, Andesite, Magmatic, granodiorite porphyry
9x9 9 x 11 (limited sectors)
13 x 13 13 x 15 ( Ore pass singularity)
Grizzly layout
Teniente layout
Gravitational to truck
7 yd3 LHD
Ore pass (2 x 2 m)
Ore Pass (2.5 x 2.5 m)
50 & 80 ton truck
50 & 80 ton truck
1200 US$/m2
750 US$/m2
Total caved area up to 2003
56.159 m2
96.176 m2
Ore Reserves (Up to 2003 caving)
36 Million t
151 Million t
Extraction system Ore Transfer Haulage Development cost (up to caving) - Project
Average in situ ore column height
280 m
256 m
Estimated Dilution Grade (%Cut)
0.5% Cu
0.5% Cu
65%
75%
Project % Dilution entry point Project extraction rate Massmin 2004
0,44
ton/m2*day
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0,56 ton/m2*day 415
Figure 2a Figure 1a
Figure 1b
Figure 2b
Figure N° 3 shows a profile of the primary/secondary rock interface with the extension of the LHD and grizzly layouts. As it is shown the LHD as well as the grizzlies are primarily extracting secondary ore. The dilution of the III Panel is mainly the remnant of the II and I Panel, which had very high grade. The main dilution (sterile rock) is of riolite coming from the north of the subsidence area, but through the years have migrated and diluted most of the subsidence area. The riolite is used as a lithologic tracer of dilution percentage. 3. OPERATIONAL RESULTS The operational results to compare Grizzlies and LHD for the III Panel consider a combination of results from all information up to 2003 and some specific sectors for each method that have been mined up to closure of the draw points. • Caving Sequence • Following figure N°4 shows the effective and future caving sequence of the III Panel by year, starting in 1995 for the grizzly (in blue) and LHD sectors (in red). • Development and operational cost Following costs are effective average results for representative areas and tonnage. 416
Figure 3
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Primary rock presence in the lower part of the column, poor performance of the ore pass system (due to mud), hang ups, etc. The dilution entry point for LHD sector has been lower than planned, as well as the projected productivity. The following Table N°4 shows the planned and effective production by sector up to 2003: Table N°4 Effective tonnage production by year for LHD and Grizzly UN 1995 1996 1997 1998 1999 2000 2001 2002 2003 Total
Figure 4 Development: As a result of a more dense draw pattern (9 x 9 meters), the grizzly design is more intense in development and construction (specially drop boxes and loading chutes), thus generating a higher development cost. Following Table N°2 shows Grizzly and LHD development cost from caving to transport level in US$/m2. Table N°2 Development Cost
Development Cost
UN
LHD
GRIZZLY
US$/m2
770
1.200
Operating: Operating costs are presented up to transport (not included) due to differences in distances, for comparison purposes of both methods. Repairs of the production infrastructure (production, ore passes, chutes, etc.) are included in the cost. The total cost per ton including operation and development has been estimated considering the effective tonnage per square meter produced by each method, accounting the collapses. Table N°3 summarizes these results. Table N°3 Operational Cost Cost Operational Total Cost
Grizzlies
LHD
0,773 US$7t
1,094 US$/t
2,05 US$/t
2,24 US$/t
• Total Extracted Tonnage – Productivity Total Tonnage : Production was initiated in the grizzly sector, effective produced tonnage was under the planned schedule during the first two years. After that the situation was reverted, from 1997 to 2003 production from the grizzlies have been more than 5 million tons higher than the original project. The effective average dilution entry point for the isolated draw zone (Pedza) for this sector is at 75% extraction, that is 10% higher than the projected value (65%). Productivity has been higher than planned also. In the other hand, the LHD sector has systematically been under the planned productivity, due to a series of factors: Massmin 2004
Grizzly Real
kton 615 2,647 5,845 4,840 5,999 7,084 6,649 6,318 6,828 46,824
Program
kton 1,124 4,090 4,208 4,861 5,091 4,765 5,944 5,885 5,476 41,444
LHD Real
kton
624 4,570 6,829 8,581 8,904 8,969 8,353 46,832
Program
hton
2,945 4,679 9,340 11,435 10,256 10,315 10,724 59,694
• Grade & Fine Copper Production Even considering the low Copper prices that as an effect generated a lower caving area in the III Panel, and lower cut off grades, the average effective grade is higher than the planned one, mainly due to higher grades in the grizzly sector, explained by the higher dilution entry point and probably higher dilution grades. In the LHD sector grades have been a little lower than planned, with also a lower tonnage extracted. The dilution entry point was lower than estimated. Table N°5 shows extracted average grades by year for both sectors. Table N°5 Average extracted grade by sector 1995 1996 1997 1998 1999 2000 2001
2002 2003
Total
Grizzly Real
1.089 1.405 1.308 1.259 1,182 1.307 1.308 1.285 1.355 1.293
Program
1.155 1.188 1.187 1.154 1.161 1.324 1.339 1.203 1.039 1.199
LHD Real
1.690 1.584 1.439 1.338 1.179 1.016 1.035 1.236
Program
1.687 1.518 1.257 1.176 1.205 1.203 1.202 1.255
When comparing the global balance of fine copper produced, the result shows approximately 62.000 tons below the project, mainly due to lower production in the LHD sector. Table N°6 shows the results expressed in fine Copper production for both sectors, against the planned tonnage. Table N°6 Fine Copper Production by sector UN 1995 1996 1997 1998 1999 2000 2001 2002 2003 Total Grizzly Real
kton 6,7
37,2 76,4 60,9 70,9 92,6 87,0
81,2
92,5
605,4
Program
kton 13,0 48,6 49,9 56,1 59,1 63,1 79,6
70,8
56,9
497,1
86,5
578,7
LHD Real
kton 0,0
0,0
10,6 72,4 98,3 114,8 105,0 91,1
Program
hton 0,0
0,0
49,7 71,0 117,4 134,4 123,6 124,1 128,9 749,1
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• Draw Practice - Uniformity of Draw: One of the relevant parameters of draw control in a panel caving operation is the uniformity of draw. This parameter is being controlled with the Uniformity Index (IU). Figure N°5 presents the results for the index, indicating the % of total tonnage per year drawn with uniformity (uniform and semi uniform draw).
Figure N°5: Tonnage drawn with Uniformity
4. BACK ANALYSIS OF PRODUCTION RESULTS • Recovery of reserves The Grizzly sector has had a 18% higher recovery of mining reserves than the LHD sector to date. The column height were similar thus the grizzly sector has recovered up to 350 m columns, compared to the 296 m of LHD. The good recovery of the grizzlies can be explained due to the very good interaction of a 81 m2 draw pattern, compared with a 175,5 m2 LHD layout, for the same caved material.
• % Dilution Entry Point Represents the extracted percentage of the in situ column when the dilution is reported in the draw points. In the case of Andina the dilution corresponds to caved material from the II Panel, overlaying the III Panel. The steel arches, concrete and other materials from this level are a very good tracer of the dilution, and have been mapped in every draw point. This dilution corresponds to the "isolated column dilution" or Pedza. The expected dilution from the "interactive zone" should be determined when the permanent increase in dilution percentage starts. The mapping of a specific lithology (ryolite) that has contaminated the subsidence area from the north side of the crater, that can be easily recognized from the andesite and granodiorite As seen in Figures N°6 and N°7 where all the mapping information has been summarized for Area A – LHD and Areas 1 and 2 Grizzlies of the III Panel, there is a very good correlation between the Pedza appearance and the increase of ryolite percentage for both cases (not considering the effect of side dilution that is incorporating early ryolite). There is also a clear change in the dilution curve slope showing the Pedzi (Dilution from the interactive zone). The summary of the total tonnages extracted, caved areas, dilution entry points, etc. for the sectors presented in the figures is presented in Table N°7. • Available Area The strength of each design can be evaluated considering the geomechanic behavior through the III Panel caved area. For the grizzly system a total area of 56.159 m2 have been caved over 9 years, without the loss of a single draw point due to collapses. With the LHD system a total of 96.176 m2 have been caved most of the development in primary rock, with a total loss due to collapses of 26.369 m2, that corresponds to 27,4% of the total caved area. Evidently this is one of the greatest weakness of the designed system.
Figure N°6: LHD Area A % dilution entry curve. 418
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Figure N°7 – Grizzlies Area 1 and 2 % dilution entry curves
Table N°7 Draw Results Item
Grizzlies
LHD
ton/m2
952
795
Total Area
m2
22.607
22.626
Total Tonnage
ton
21.526.627
17.995.873
Total % Extraction
%
145%
118%
Total % Ryolite extraction
%
6,7%
6,7%
Ryolite Tonnage
ton
1.351.803
1.032.012
Pedza
%
73%
64%
Pedzi
%
89%
79%
Unit productivity(without collapses)
• System Productivity Effective productivity of each system has been evaluated considering the average tonnage produced per active square meter, by year. In following Table N°8 productivity for each sector by year is summarized. The grizzly sector had a program of 150 t/m2 and has an average of 177 t/m2 per day. The LHD had a program of 185 t/m2 and the effective productivity (not considering the collapsed area) is 136 t/m2. • Ore pass availability and Mud Problems Moisture coming from snow and water precipitated in the subsidence area generates mud conditions for both the grizzly and LHD systems. The vulnerability of the LHD due to ore pass closure due to mud problems and repairs is high (great area to one box). Low ore pass availability has generated low productivity of the LHD system. • Summary With the above information, Table N°9 was computed summarizing the main results of the two caving extraction methods used in the Andina III Panel. Massmin 2004
Table N°8 – Productivity (t/m2) 1995 1996 1997 1998 1999 2000 2001 2002 2003 Total Grizzlies Real t/m2
91
149
260
215
163
181
188
134
210
177
Program t/m2
71
129
138
128
143
180
204
175
182
150
LHD Real
t/m2
70
168
144
151
153
135
129
136
Program
t/m2
70
157
190
217
212
222
214
195
• Ore Reserves Recovery The ore reserves recovery has been calculated with the total extracted tonnage per caved area (this includes the collapsed areas) the total productivity for the grizzly sector is of 940 t/m 2 against 667 t/m2 of the LHD.
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Table N°9 Summary of back analysis results Index
Observation
Grizzlies
LHD
From under cut to transport level plus caving
1200
770
Design strength
Strength of design is evaluated against collapsed area
Total caved area = 56159 m2 Total collapsed area = 0 m2 % Area loss = 0%
Total caved area = 96176 m2 Total collapsed area = 26369 m2 % Area loss = 27,4%
Unit Productivity
Ton per open productive area per periiod
177 t/m2
136 t/m2
Total tonnage per closed area (with collapsed area percentage)
940 t/m2
667 t/m2
Total Cost
Operation to truck transport (exclusive) with repairs)
0,773 US$/t
1.094 US$/t
Total Cost
Operation plus development
2,05 US$/t
2,24 US$/t
% Pedza
% Extraction
73
64
% Pedzi
% Extraction
89
79
% of mining reserves recovered
136%
118%
Investment (cost/m2)
Ore Reserves Recovery
In situ reserve recovery
5. CONCLUSIONS AND RECOMMENDATIONS With the compiled results the conclusions are: • Total cost including development and operations, considering collapses areas, is 9,2% lower for Grizzlies against LHD. • Operational problems with LHD are higher due to high vulnerability of the ore pass system. • Ore reserve recovery (due to 81 m2 layout against 169 m2) is much higher in grizzly layout (41% higher). • Dilution control even with the lack of uniformity in the grizzly sector has been much better in grizzlies, obtaining same total tonnage of dilution (ryolite) for 20% additional recovery.
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RECOMMENDATIONS: • For secondary ore, or for cave fragmentation that can be handled by a gravitational system, it should be preferred over a LHD extraction system. • If fragmentation of the cave can be managed (ie: through preconditioning of the rock) a gravity based method, should be much efficient than an LHD alternative. • The LHD design should be reviewed to endure stability of the future developments of the III Panel.
Santiago Chile, 22-25 August 2004
Massmin 2004
New mine level project at El Teniente Patricio Yáñez, Mine Engineering Superintendent, Rigoberto Molina, Geomechanics Engineer, El Teniente Division, Codelco Chile
Abstract El Teniente, one of CODELCO’s mining Divisions, is carrying out an expansion plan of the underground panel caving operation and the processing facilities in order to reach a 130,000 tonnes per day, plan known as the El Teniente Development Plan. For the production plan and future development plans, the exploitation of the deposit above elevation 2,120 meter above sea level, has been considered until year 2014. From this year on, it will necessary to incorporate new mining projects, located below that level, in order to maintain the long term production rates. In this context, starting from year 2014 El Teniente will incorporate the New Mine Level project, which will become the most important underground panel caving project and will sustain the production plans in the long term, exploiting only primary ore. The undercut level will be located at 1,880 meters above sea level. The new level will be divided in five mining sectors, including 1,371 million tons as total ore reserves with a 0.96% copper grade, covering an area of 1,606,000 square meters. The initial production rate will be 2,000 tons per day and will reach a maximum production close to 130,000 tons per day in the long term. The New Mine Level Project will deepen the exploitation of the deposit 100 meters below the current deepest level of the mine (the main transport level, railroad Teniente 8) and will incorporate blocks with an average of 300 meters height approximately. The location of such future mining level requires the introduction of relevant changes to the main infrastructure of the mine, such as a new main transport level, new service shafts, primary crushers chambers, new drainage and ventilation levels. Other important topics of the project to be discussed are the elevation of the future mining level, the slope strategy with the mining going down to the new level (connection to the crater), the interaction with the current operation sectors and with the main infrastructure of the upper levels of the mine, the mining plan and its production capacity, the mining method design and the geological and geomechanical studies. All these aspects configure the main focus of the studies currently under development at the pre-feasibility engineering stage and will be emphasised on this paper.
1. INTRODUCTION The El Teniente Codelco’s Division, a state-owned company, is located 80 kilometers South of Santiago at 2,200 meters above sea level (Figure 1). It exploits the largest underground cooper mine in the world. The mine has been exploited since 1905 and it includes 2,400 kilometers of galleries. This division annually produces close to 450,000 tons of cooper fine, bars and cathodes. As a result of the ore treatment, 4,750 tons of molybdenum and more than 800,000 tons of sulphuric acid are also produced each year. The 130,000 tpd production plan is supported by the exploitation of ore reserves located over the 2,120 meters above sea level elevation (Teniente Sub6 level) up to the year 2014. After that time, new mining projects located below that level should be incorporated in order to maintain the long term production in agreement with the plan rates. In that way, the "New Mine Level " is added to the production plan starting on year 2014. Its reserves are situated below the current main transport level (Teniente 8, elevation 1,980 meters above sea level). It will include five production sectors (Norte, Andes Norte, Andes Sur, Andes Central, Sector Sur y Pacifico).
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Figure 1: División El Teniente Location
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Figure 3: New Mine Level Sector Location
shape, close to 1 Km in diameter, with a sub vertical axe and more than 2 Km of vertical extension. It includes ten main lithological units which are presented in Figure 3. A main set of strike slip N-E faults are present in the northern orebody volume, N, N1 and N2 faults and P and P1 faults in southern area. The width range varies from 5 a 120cm with traces in th range of 100m to 500m. These faults are projected to the lower levels, including the New Mine Level.
Figure 2: New Mine Level Sector Location
Table 1: RMR Laubscher’s Classification Properties
2. NEW MINE LEVEL PROJECT Location The New Mine Level project is located in a zone defined by the coordinates -660N y 1,300N and 0E y 1,580E, as shown in Figure 2. The ore column is defined by the level 1,880 meters above sea level and the current production sectors Esmeralda (2,210 meters above sea level) and Teniente Sub6 (2,120 meters above sea level). The ore columns are 240 meters for the sectors under Teniente Sub6 and 340 meters for those under the Esmeralda sector. Almost 97 % of the project reserves are primary ore. Geology The El Teniente orebody is located in the central Chilean Andes and it is classified as a gigantic deposit. It is one of the largest cooper-molybdenum deposit, with more than 75 millions of cooper fine tons. The size of the economical mineralized rock mass is 3 Km. in N-S direction and 1.5 Km in E-W direction. The underground mine has already exploited a 1 Km rockmass column from the Teniente "J" level to the Teniente Sub6 level. The mafic complex is an aggregate of gabbros, diabasas, basalts, basaltic porphyries and basaltic andesites. It I configured as a tridemensional set of fractures, veins and veinlets ("Stockwork"), with pyrite, chalcopyrite, bornite and molybdenite mineralization. A relevant brecchia complex (Braden Brecchia) is located in the central zone of the orebody. It corresponds to a complex of hydrothermal brecchias, with an inverted cone 422
ff/m3
RMR
Andesite HT (Late Hydrothermal ) Hw
5-8
50-60
Andesite HP (Principal hydrothermal) Principal)
8-11
42-53
UNIT
Diorite (primary)
4-9
55-65
Dacite (primary)
4-6
65-70
Hydrothermal Brecchia
3-5
60-70
Primary Igneous Brecchia
3-7
55-65
Table 2: Intact Rock Properties Parameter
Andesite Diorite Dacite Braden Brecchia
Density [ton/m3]
2,8
2,7
2,7
2,60
Deformability Modulus (E) [GPa]
60
45
30
25
Poisson Ratio(n)
0,16
0,21
0,18
0,23
From Hoek & Brown criteria (σci) [MPa]
120
140
110
90
From Hoek & Brown criteria, (Mi)
9,1
9,2
20,2
11,6
Cohesion ( c )
23
23
19
10
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The Rock Mass Rating (RMR) Laubscher’s (1990) classification values are shown in Table 1 for the main lithological orebody units. Table 2 presents the intact rock properties for the andesite, diorite, Braden brecchia and dacitic porphory units. Geomechanics Guidelines for caving starting and propagation The New Mine Level will initiate the caving using a descending mining, avoiding an excessive tensile stress in the surrounding rockmass and a relevant confinement in the top of the caving cavity. If a standard connection method is used, caving progressing upwards, the caving will be more difficult. It will probably increase the recorded induced seismicity. The abutment stress generated in the borders of the cavity and the increase in the seismically active volume above this cavity will create a higher radiated energy events in relation to the size of the ruptures originating those events. The starting point of the induced caving corresponds to a connection from the crater to the undercut level (downwards connection) instead of a connection from the undercut level to the crater as used in the standard connection (upwards connection), Figure 4. Once the undercut level is reached, the undercut zone is expanded by the incorporation of drawbells as in a standard caving method. A favorable condition is created by a gradual increase of the seismically active volume as the mining goes down. It increases from a cero volume to its maximum. In the other side, the active volume decreases from a maximum at the starting of the undercutting as the caving progress upwards in a standard caving method.
Figure 5: Stress Field zones Fragmentation Five zone have been defined around the central Brecchia, included in the exploitation area of the New Mine Level. The fragmentation has been classified in a relative scale including small size material, medium, coarse, very coarse and blocks. The classification zones are presented in Figure 6. Subsidence Table 3 shows the subsidence angles for the New Mine Level sectors. Those values have been derived from empirical models validated against damage zones surveyed in the current productive sectors.
Figure 4: Difference Natural Caving and Induced Caving
Stress Condition The stress field as been estimated resulting a principal major stress of 60 MPa and a minor principal stress of 33 MPa at a 1,800 meters above sea level elevation in the central zone (see Figure 5). The higher stress concentration are induced in the eastern zone (high mountain zone, zone 3 in figure 5) and in the border of the Braden Pipe (zone 2 in the same figure). This estimation has been obtain from the stress measurements existing in the production sectors and in the bottom level of the mine, Teniente 8 at an 1,980 meters above sea level elevation. Currently, efforts are made in order to improve this estimations using numerical models and/or hydraulic fracturing stress measurements. Massmin 2004
Figure 6: Fragmentation Zones Level 1,880 masl
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Table 3: Subsidence Angles of NML Sectors Sector
Andes Norte
Subsidence Angles (Collapse/Fracture)
Tabla 4: Permisible Distances for the NML Sector
North
South
East
West
63º/55º
N/A
63º/55º
68º/55º
Andes Central
63º/55º
N/A
63º/55º
68º/55º
Andes Sur
66º/58º
65º/58º
63º/57º
68º/55º
Sur
74º/65º
71º/64º
68º/60º
70º/62º
Permissible Distances NML (meters) From
To
Undercutting-Extraction
60
100
Undercutting- Developments (Production Level)
20
25
Undercutting – Mining Preparation(Undercut level)
40
70
Note: The range depends on the NML zone Seismicity ining induced seismicity is associated to the breaking of the rockmass. Under same rockmass characteristics, distances from seismic event sources and underground work local conditions the seismic radiated energy can damage galleries creating rockbursts. Empirical evidences show that the most seismically unfavorable stage of the caving process is the initial caving before connecting to the upper caved level. The seismic risk during this stage will be minimized by the use of the deepening connection. Permisible Distances The application of the Panel Caving method define three zones with different permissible distances. • Undercutting front – extraction face distance • Undercutting front – Production level Development Advancing face • Undercutting front – Prepared Area in Undercutting Level Undercutting front – Production level Developme Tabla 4 shows the permissible distances for the New Mine Level.
Reserves The project reserves are close to 1,371 millions tons, with an average cooper grade of 0.96%, a molybdenum grade of 0.024%. From the total reserves 57% are classified as proven, considering a total area of 1,606,000 square, Tabla 5. Mining Plan The principle used for defining the extraction sequence for each sector was to maximize the economical revenues of the reserve exploitation. In addition, other technical criteria were utilized as the interaction between sectors under exploitation (subsidence) and the geomechanical rockmass characteristics which are predominant in relation to economical criteria due to their relation with the technical feasibility of the exploitation. Figure 7 presents the exploitation sequence for the NML sectors. The project expansion strategy for the starting of its production operations is divided in two stages. The first one called "going down" involves the connection of the current crater with the undercutting of the New Mine Level
Figure 7: Exploitation Sequence for NML Sectors 424
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Tabla 5: NML Sector Reserves Total Reserves % % % Area Proven Probable Possible
SECTOR
Mton
% Cu
% Mo
Norte
157.8
0.9
0.015
22
39
38
171
Andes Norte
216.6
1.03
0.022
20
44
36
309
Andes Central
269.1
1.12
0.025
22
41
36
350
Andes Sur
433.2
0.89
0.022
6
43
51
506
Sur
136.0
0.81
0.021
11
40
49
145
Pacífico
158.3
0.90
0.031
5
15
80
124
TOTAL
1,371.0
0.96
0.024
15
42
43
1,606
• The typical production gallery section is 5.0 m width and 4.5 m height. • The distance between drawbells crosscuts is 20 m and between galleries is 34.64 m, defining an influence area of 346.4 m2 for each extraction point. The tonnage to be extracted by each point will be in the range of 224,000 and 317,000 tons of ore, depending on the column heights. • The orepass system is configured by 3.5 m diameter orepasses, with a 200 m spacing in each production gallery. This means an influence area of 6,928 m2 involving 20 extraction points for each orepass. Then, an ore tonnage ranging from 4.4 to 6.3 millions tons would be transferred depending on the column heights of the corresponding sector • The secondary blasting reduction will be mechanized using hydrofracturing equipments • Automatic LHD equipment is considered.
and the second one the expansion of the Panel Caving operations. The 130,000 tpd mining plan integrates the first NML sector to the mine production during year 2014. Its initial production rate will be 2,000 tpd and it will reach a 130,000 tpd production as permanent regime production after the year 2025. Figure 8 presents the 130,000 tpd mining plan and the NML production.
Figure 9: NML Production Module
Figure 8: The 130,000 tpd Mining Plan Exploitation Method A present-day approach for mining involving large production volumes has been adopted. It will be highly technified, with automated production operations and in agreement with high environmental and safety standards The Panel caving method has been selected for the New Mine Level, using the pre-undercut or advance undercut variant with the exception of the initial area where a SubLevel Caving or an Inclined Caving are going to be applied. A "going down" option for the initial mining will generate a more favorable seismic rockmass response. Besides a faster connection to the upper caved level and a shorter time to reach a permanent regime production rate are expected. The main characteristics of the mining method to be applied are the followings: • Four main levels: undercutting, production, haulage and ventilation. • A production module involving a Teniente type extraction mesh. The production module appears in Figure 9. • Production galleries and drawbell crosscuts in agreement with the use of 13 yd3 LHD equipment. • Drawbell crosscuts with a 60º angle in relation to the production galleries. Massmin 2004
Ore Handling The general criteria to be considered for the ore handling system design are: a high production capacity, handling of a primary coarse ore, operational flexibility, low operational cost and high productivity, low operational risk associated to the technology and an easy construction. From this point of view, the project has defined the use of automatic 13 yd3 LHD equipment for the production level. They will load the ore at the extraction points and dumped it at the orepasses. At the transport level, ore will be loaded to 80 tons trucks by means of mechanized plate feeders. At each loading point the production from 2 orepasses will be handled. The typical gallery section will be 5.5 by 5.0 m to allow the automated truck operation. Ore will be transported by the trucks to the centralized primary crusher chambers located out of the extraction zone. These chambers are provided with shooters, 1.5 by 1.5 m grizllies and permanent pickhammers. The ore will be transferred to the rotating crushers, 60" x 89" size, with a 1.5 m feeder capacity. The resulting material should be less than 8". It will be transported by a conveyor belt system to the concentrator plant. The conveyor belts will be 60" wide, 4.1 m/s speed, a maximum capacity of 5,500 tph and different lengths and slopes. Table 6 and Figure 10 provided the main specifications and a diagram of the ore handling system.
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Table 6: NML Ore Handling System Specification Fase Secondary Reduction Extraction (13 yd3 LHD) Ore transfer Feeder Loading Haulage (80 t truck) Pickhammer Grizlly (per crusher)
Specification
Notes
4.9 hours/shift (LHD)
Ratio > 1.8 m
4,000 t/shift
Average distance 50 m.
3.5 m diameter ore pass
Material Size < 1.8 m
4 trucks/minute
2 ore pass/feeder
2,250 t/shift
Haulage distance 1,500 m.
2.7 hours/shift (truck)
Reduced Ratio 1.5 m. – 1.8 m. range
4 grizllies, 1 pickhammer each
Principal Indicators The conceptual engineering studies currently under development show the following figures: • First sector commissioning date year 2014 • Permanent regime production capacity 130,000 tpd • Level lifetime > 30 years • Maximum rate per exploitation front 45,000 tpd • Area added for each sector < 28,000 m2/year • Estimated El Teniente personnel 476 personnel • Estimated contractors 1,599 personnel • Estimated productivity (El Teniente personnel) > 280 tons/man-day • Estimated productivity (including contarctors) > 70 tons/man-day • Mine investments 663 MUS$ • Average operational cost 2.43 US$/ton - Mining preparation 1.06 US$/ton - Extraction and ore transfer 1.01 US$/ton - Primary crushing and main transport 0.36 US$/ton
The estimated low costs are generated by the NML project correspond mainly to the mining preparation low cost due to the increased sized of the extraction mesh and a decrease in the annual new area required due to the higher rockmass columns, automation process, a single exploitation level and preparation and maintenance outsourcing. Key Factors The conceptual engineering studies have detected same key factors regarding the technical and economical success of the project. The main factor are the followings: • Geological and Geometalurgical Exploration Plan. - In-situ geological resources - Geometalurgical model - Hydraulic drainage • Geomechanical Exploration and Studies - Connection to crater method validation or initiate induced caving
Figure 10: NML Ore Handling System 426
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Figure 11: NML Project Global Plan - Stress field estimation - Seismic risk evaluation • Mining and Metalurgical Planning - Exploitation sequence and production capacity validation - Simultaneous productive operation of different sectors (subsidence and interference) • Mine design and Engineering - Project master plan validation (construction analysis) including engineering, construction, supplies and commissioning. Figure 11, shows the project global plan. - Technological developments (process automation) - Exploitation method and its variant post-evaluation.
•
•
•
•
3. CONCLUSIONS • The New Mine Level is one of the most important project among those with a conceptual engineering study at the El Teniente Division. It will incorporate 1,371 millions tons of ore reserves to the production plan. Production from this project will be included in the year 2014 planned production. The first project sector (Andes Norte) will support the 130,000 tpd long term production rate. Exploiting the only the sectors located above the 2,120 meters above sea level elevation, that rate will be sustained only to the year 2014. The development of this project will generate a large impact in the main mine infrastructure due to excavation of a new infrastructure below the current deepest level existing in the mine (Teniente 8 Level, 1,980 meters above sea level elevation).
• •
•
•
•
ACKNOWLEDGEMENTS • The autors thank at the "Gerencia de Recursos Mineros y Desarrollo" and "Gerencia de Minas" of the El Teniente Division, CODELCO CHILE, for the authorization to publish this paper. • REFERENCES • Meléndez L., Villanueva J. and Yánez P., "Resumen Ingeniería de Perfil Panel 8", Report Nº PL-542/1999, November 1999. • Meléndez L., Villanueva J., and Yánez P., "Ingeniería de Perfil Proyecto Panel 8 – Selección de Reservas", anexos, December 1999. • Celhay F., "Programa de Sondajes Proyecto Panel 8", Report GL-042/2000, March 2000. • Yáñez P., "Ingeniería de Perfil Proyecto de Explotación Panel 8 – Informe Final", Report Nº PDT-I-003/2000, December 1999. • Yáñez P., "Ingeniería de perfil Proyecto de Explotación Massmin 2004
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Panel 8 – Anexos", Report Nº PDT-I-003/2000, December 1999. JRI Ingeniería, "Proyecto Ingeniería de Perfil Plan Exploratorio Mina Concentrador – Informe Final Volumen I", September 2001, in contrat for El Teniente División. JRI Ingeniería, "Proyecto Ingeniería de Perfil Plan Exploratorio Mina Concentrador – Volumen II – Area Mina", September 2001, in contract for El Teniente División. JRI Ingeniería, "Proyecto Ingeniería de Perfil Plan Exploratorio Mina Concentrador – Informe Final Area Planta", September 2001, in contract for El Teniente División. Parraguez R., "Topicos a Considerar en la Ingenería Conceptual Plan Exploratorio NNM (Panel 8), I Etapa", Report Nº PL-059/02, February 2002. Yánez P., "API T01M201: Proyecto Nuevo Nivel Mina – Panel 8 Ingeniería Conceptual", Report Nº PDT-I001/2000, July 2000. Yánez P., "API T03M202: Proyecto de Explotación Nuevo Nivel Mina", Report SPD-I-MI-022/02, July 2002. Celhay F., and Zúñiga P., "Proyecto Panel 8 Antecedentes Geológicos", Report GL-I-59/02, July 2002. Celhay F., and Letelier M., "Estimación de Arsénico y Antimonio Sector Sur Yacimiento El Teniente", Report SGL-I-054/03, July 2003. Celhay F., and Letelier M., "Estimación de Recursos CuT y MoT Sector Norte Central", Report SGL-I-131/03, December 2003. Molina R., Parraguez R., Catalano J., and Derk Ingeniería y Geología Ltda., "Análisis Geomecánico Plan Exploratorio Teniente ( P.E.T.)", Report DI-CT-SP-008, July 2002, in contract for El Teniente División. Molina R., Parraguez R., Catalano J., and Derk Ingeniería y Geología Ltda., "Análisis Geomecánico Proyecto Aumento Capacidad Mina, Ingeniería Conceptual - Etapa I", Report DI-CT-SP-004, July 2002, in contract for El Teniente División. JRI Ingeniería, "Informe Ingeniería Conceptual Proyecto Aumento Capacidad Mina – Concentrador 150 ktpd, Vol 1 al 6", Report SDS-03/02, August 2002, in contract for El Teniente División. Molina R., Parraguez R., Catalano J., and Derk Ingeniería y Geología Ltda., "Informe de Hundibilidad en Mina Esmeralda, Fase I", Report DI-CT-SP-024, October 2002, in contract for El Teniente División. Molina R., Parraguez R., Campos J., and Derk Ingeniería y Geología Ltda., "Informe de Hundibilidad en Mina Esmeralda, Fase II", Report DI-CT-SP-029, November 2002, in contract for El Teniente División. Espinosa C., and CIIM T y S, "Sistema Estándar de Control Operacional de Fragmentación, SECOF", Report P-106745, November 2002, in contract for El Teniente División.
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• León H., and TI, "Sistema Estándar de Control Operacional de Fragmentación, SECOF, Levantamiento de Variables", Report TI-432/03, March 2003. • Dunlop R., "Estudio de Caracterización del Macizo Rocoso por Método Geofísico para Reservas bajo Teniente 8", Report SPL-I-027/03, July 2003.
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• Cavieres P., and Vasquez P., "Aplicación del Sistema Standard de Control de Fragmentación SECOF, al sector Teniente 4 Sur", Report SPL-I-032/03, September 2003. • Molina R., Parraguez R., and Others, "Análisis Geomecánico Plan de Negocio y Desarrollo 2004", Report SPL-I-004/04, February 2004.
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Massmin 2004
KGHM – The largest underground mine in the world Andrzej Zablocki, Atlas Copco Chilena S.A.C., Piotr Kijewski, CBPM Cuprum (Poland)
Abstract Intensive exploration during the middle of the 20th century in the south-west region of Poland led to a discovery of one of the world largest copper ore-bearing deposits. The excavation of the deposit is concentrated in three mining sectors: Lubin, Polkowice-Sieroszowice and Rudna. All of them belong to a joint-stock company KGHM Polska Miedz S.A. comprising three dressing plants, two smelters and one copper rolling mill. Annual extraction of approximately 30 Mill tons of ore with an average copper content of 2% makes it the world largest underground mine of fine copper production (530,000 t. in 2003). Geology, resources, geotechnical conditions and the variation of mining methods for a depth of 1000 m and one of the world’s most unusual mineral deposits of such a large excavation volume is described in this paper.
INTRODUCTION Poland belongs to a group of states that disposes of large copper ore reserves. These deposits are concentrated in the southwest of the country, in two geological units: the northern Foresudetic basin and in the monoclinic grounds of Foresudetic (fig. 1). These are the sedimentary deposits of a variety of sulphuric mineralisation. Mining production is currently carried out in the monoclinic grounds of Foresudetic, in the region of Lubin, since in the northern basin of Sudetic exploitation ceased due to exhaustion of reserves (region of Zlotoryja or because of economic causes (region of Boleslawice).
1) Eocretaceous, 2)Triassic, 3)Permian, 4)Paloeozoic formation (cristalic), 5)Fault belts 6)Line of geological section, 7)Exploatation zone.
Fig. 1 Drawing of geological structure of the region showing copper deposits in western Poland Massmin 2004
Deposits currently being exploited were discovered in 1957 by Dr. Engineer Jan Wyzykowski of the National Institute of Geology. The first documentation was elaborated in 1959 and was the base for the first mining investments in1960 and for the following exploitation tasks, indispensable to recognize the geological-mining conditions and to document the new portions of the deposit. The first shaft reached the ore level on March 20th, 1960, and the first mining units ("Lubin", "Polkowice") reached 25% of its productive capacity in January 1969. In the years 1970-1980 mining investments allowed making use of new sectors in the deposit; "Rudna," "Sieroszowice", and also increasing the exploitation to its current annual level of more than 28-29 million tons of ore. Parallel to the mining construction, ore treatment plants have been created as well as copper foundries in Legnica, Glogow and Cedynia. Features of the Deposit The copper mine located in the mountainous region of Foresudetic appears on the contact area between the sandstone sediments called rotliegendes and marine formations of Zechstein. In general it is a one stratum deposit, with an irregular roof and floor surface with a slight 6 degree dip. The characteristic deposit lithostratigraphic level is the layer of copper slate of Zechstein with a high concentration of copper, silver and other accompanying metals. Mineralisation also includes contiguously slate located on the floors of the rotliegendes and carbonated rocks of Zechstein on the roof. The thickness of the deposit is quite variable, fluctuating between 0.5 m and up to more than 20 m. Due to its significant thickness the "Rudna" sector can be particularly distinguished where close to 45% of the resources are located in the group of more than 6 m with a maximum of 26 m. The deposit is exploited at a depth of 600 to 1,000 m and has been recognized to a depth of 1,600 m. The difference in depth of the deposit is related to the rock stratum monoclinic declination. In the limits of the deposit of each mining division the quantitative differentiation of the mining variations stands out. Thus, in the "Lubin" sector the sandstone ore conforms to 71.5%, carbonated ore 17.5% and slate 11%. The predominance of the sandstone ore stands out even more, 84.5% in the "Rudna" sector, with 10.5 carbonated ore and 5% slate.
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In the "Polkowice-Sieroszowice" sector the superiority of the carbonated ore of 64.2% can be noted over the sandstone ore and slate, 23.6% and 14% respectively. The copper content in each sector of the deposit is different. A general trend appears to be a higher quantity of copper in the deposit where it is less thick, which is related to an increased participation of the more mineralized slate. This shows an average content of 10% of copper whereas in the sandstone ore this fluctuates between 1-3%. The characteristic feature of the deposit is the high contents of silver which often exceeds 100 ppm over the average exploited ore with a level of 60 ppm. Thus the total production of over 1,200 tons puts KGHM within the group of the main world producers of silver.
1)Documented deposit, 2)Exploatation zone, 3)GlogoW Gleboki reserve zone, 4)Battery of shafts, 5)Individual shafts
Fig. 2 Copper-bearing Zone
Base of Reserves The mining ownership in whose limits the Mining Divisions of KGHM Polska Miedz S.A. carry out the exploitation of the deposit, reaches about 416 km2 (fig.2), assuring a base of reserves indispensable for maintaining the high production of copper and silver. According to principles adopted in Poland, balance resources comply with that part of these geological resources which makes the exploitation technically possible and
economically justifiable. On this basis, the balance of reserves has been determined in each Mining Division (tab.1). Included in the balance resources possible to exploit in the near future, is part of the deposit "Glogow Gleboki", which is today the reserve sector. Reserves of this part of the deposit can be developed making use of the current mining structure of divisions "Rudna" and "PolkowiceSieroszowice". Balance reserves in the reserve sector comprise 688.4 million tons of ore and 14.9 million tons of copper, that allow increasing the quantity of reserves of the balance to 2,555 million tons of ore and copper to 45 million tons. Mining Operations and geotechnical conditions The mine gradually increased the production (table 2) split in terms of ore hoisting between Lubin (8 Million t/y), Polkowice-Sieroszowice (10 Million t/y) and Rudna (11Million t/y). In addition, KGHM is exploring to the north-east of its existing mines and has already identified a further 700 Mt of mineralization in the Glogow Gleboki area grading 2.2% copper. Right now the new shaft is being developed in this sector. At the moment the shaft is 1050 m deep and its final depth will be 1246 m. Up to 635 m depth the freezing method was used in order to be able to sink it. In total the mines have 26 vertical shafts with diameters ranging from 6 m to 7.5 m and depths that vary from 632 m at Lubin to 1,120 m at Rudna All of the shafts were sunk using the freezing method as they pass through a major aquifer. The most modern shaft belongs to Rudna mine where the depth of the loading level is 1,022 m; it is equipped with two twin-skip Koepe friction hoists. Each unit consists of a fourrope hoist with 5.5 m diameter friction drum powered by a 3,600 kW motor. Each of the skips has a capacity of 300 kN and runs at a speed of 20 m/s. Access to the deposit from the shafts and preparatory workings is by a network of drifts located directly under the strong dolomite roof and above the sandstone, along the dip of the ore zone. The main drifts are equipped with conveyor systems for ore transport. Evaluation of mining methods is shown in fig. 3. Originally the long wall method was used based on coal mining experiences but very soon abandoned because of limited possibilities to mechanise the work and low productivity. The predominant method used is room and pillar, although this is adapted to the prevailing seam thickness and geotechnical conditions. For instance, deposits up to 5 m thick are currently mined by room and pillar with roof caving or roof deflection, considered to be more effective and safer since it enables full mechanisation. The technology behind mining deposits 5 to 7 m thick is based on advanced-fracturing and post-failure capacity of pillars.
TABLE 1: Level of Reserves year 2000 Mining Divisions
Reserves
Average grade
Quantity
(Million tons)
Cu %
Ag (ppm)
Lubin
423.888
1.29
64
5.462
27.109
Polkowice-Sieroszowice
490.693
2.60
53
12.758
25.624
Rudna
652.063
1.83
141
11.911
26.717
1.568.744
1.92
51
30.131
79.660
Total 430
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Cu (Thousand t)
Ag (t)
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TABLE 2: Annual Production of fine copper and silver at KGHM Polska Miedz S.A. Year Fine copper (ton.) Silver (ton.)
1975
1980
1985
1990
1995
2000
2003
235,244
348,849
376,106
325,319
405,739
486,602
529,616
487.1
690.8
801.4
840.0
964.3
1,119.0
1,357.9
Fig 3. Evaluation of mining methods at KGHM
When exploiting a deposit with thickness of more than 7 m, as well as the exploitation of protection pillars at the surface, a hydraulic filling is applied. The roof opening reaches 150 m, and the longest edges of the pillars are located perpendicular to the exploitation front line. Deposits below 3 m thick, such as most of those encountered in Polkowice-Sieroszowice, are excavated with a special selective mining method using low-profile equipment (fig.4), permitting to decrease dilution. This will be even more important in the future as with the increasing depth of the deposit, the seams are getting thinner.
Figure No. 5. General layout of exploitation for the seam thickness at Polkowice-Sieroszowice capacity. For this reason, extensive rock reinforcement is used, compromising standard mechanical and resin grouted 1.6 m and 2.6 m bolts and 5-7 m cable bolting, mainly at drift crossings SUMMARY
Fig 4 Atlas Copco low profile drilling unit The variation of the method for a height of 3 m at Polkowice-Sieroszowice mine is shown in figure 5. Geotechnically, the deposits are intersected by a multitude of faults. An especially dangerous feature of the rock is its ability to accumulate high amounts of energy, which can occasionally explosively release as a rock bursts. The deposit is stressed, and is subject to seismic events and rock burst, so KGHM has developed a sophisticated monitoring system and makes m3 Because of the variety of the thickness of different parts of the deposit different size of to over 8 m3 bucket capacity, which together with trucks take ore to conveyors.equipment is used. For example, mucking is based on a large fleet of LHDs ranging from 1,5 extensive use of rock bolts. Whilst the immediately overlying rock is normally strong, in some places weak layers of shales decrease the roof bearing Massmin 2004
The Zechstein copper ore deposit in the region of Lubin consists of an important base of resources for copper mining in Poland . The mining ground together with reserves foreseen for the near future exploitation, industrial reserves amount to 935 million tons of ore and 47.6 million tons of copper. There are also balance reserves documented in adjacent land that form a potential base of raw material. • Ore mining that reached in 2003 more than 30 million tons, is accomplished at three Mining Divisions, related mutually by the structure of main drifts, horizontal and vertical transportation, ventilation and ore treatment. • The homogeneity of the deposit’s geological structure, mining level, technical and technological conditions of the mining tasks, indicate that in Poland lies the largest underground copper mine in the world and also the second largest producer of silver, owned by consortium KGHM polska Miedz S.A. • The mine’s technology will continue to evolve and this may well include some even more radical mining techniques in the future. REFERENCES • Atlas Copco, 2003. Copper Mining in the Sudetic Mountains, Underground Mining Methods. • Atlas Copco, 2003. Ericsson M. Mining trends, Underground Mining Methods.
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• Casteel K, Jan. / Febr. 1998. Low, long Lena, World Mining Equipment. • Katowice, 1971. Monograph przemyslu miedziowego w Polsce. • Krakow, 2002. Speczik At: Czterdziesci lat polskiej miedzi. • Lubin, 1996. Monograph KGHM Polska Miedz S.A.
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• Potts A, November 2003. Mining the Kupferschiefer, Mining Magazine. • Wroclaw, 1998. Chronicle of Polish Copper. • Zablocki A. Kijewski P: Está en Polonia la mayor mina subterranean de cobre? Minería Chilena, April 2004.
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Open Benching at EKATI diamond mine – Koala North: Case Study Jaroslav Jakubec, SRK Consulting, Canada, Larry Long, BHP Billiton Diamonds Inc
Abstract EKATI Diamond Mine was the first diamond mine to be developed near Lac de Gras in the Northwest Territories of Canada. The first production came from the Panda open pit in 1998. Current operations are based on mining multiple pipes by the open pit method, and Koala North pipe has been developed and mined underground. Koala North underground project was undertaken to test the underground mining method and to provide access to the lower elevations of the Panda and Koala pipes, which will also be mined from underground once the open pit operations are completed.. The Koala North underground mine, North America’s first underground diamond mine, formally opened in November 2002. It is being developed as an open-benching, mechanized, trackless operation. This paper documents the experience from the first two years of open benching mining method applied to kimberlites in Arctic conditions.
1 INTRODUCTION The EKATI Diamond Mine, operated, and 80% owned by BHP Billiton Diamonds Inc. (BHPB), 10% owned by Stewart Blussom, and 10% owned by Chuck Fipke, is Canada’s first diamond mine. The EKATI Diamond Mine is located in the heart of Arctic in Northwest Territories of Canada, approximately 300 km northeast of Yellowknife and 200 km south of Arctic Circle – see Figure 1. The mine lease area is entirely covered by treeless tundra and approximately one third of the surface is covered by lakes.
The upper 40 meters of the Koala North pipe was mined in late 2000 as a small open pit to provide grade and geotechnical information and a prepared surface for the transition to underground mining. The Koala North pipe has been selected as a trial underground mine for the purposes of testing mining methods and to provide access to the lower elevations of the Panda and Koala pipes which will be also developed as underground operations once the open pit mining is completed. The trial mining decision was made primarily because of uncertainty in the several aspects of open benching and mass underground mining at large in the northern Arctic environment. Although this mining method was successfully used on several De Beers diamond operations in South Africa, it has not been tested in this setting. 2 GENERAL GEOLOGY Koala North kimberlite pipe intruded into Archean granitoids within central Slave Structural Province in the Northwest Territories. It belongs to Lac de Gras group of kimberlites and it is located between Koala and Panda kimberlite bodies along the North-East trending structure (Figure 2).
Figure 1: Geographical location of EKATI Diamond Mine. Since its opening in 1998, the EKATI Diamond Mine has produced more than 15 million carats. Its annual output contributes to approximately 6% of world diamond production by value. As of 2002 more then 150 kimberlite occurrences had been found at EKATI claims that cover more than 3,400 km2. Currently six kimberlite pipes are included in the mining plan and nearly all of the diamond production has been from open pit mining of multiple pipes. However, as some pits deepened the decision was made to convert some of them into underground mines.
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Figure 2: Aerial view (looking approximately north) across the central development area at EKATI Diamond Mine.
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3 KOALA NORTH PIPE GEOLOGY The Koala North body forms steep-sided pipe with inverted cone morphology typical of most kimberlites in the Lac de Gras area (Figure 3). It was covered with 15 to 20 m of boulder - and gravel-dominated glacial till overburden and it was located in a depression formerly occupied by the Koala Lake. Immediate country rocks are formed by competent, granodiorite of the Koala Batholith. In general the kimberlite outline is roughly circular with wall rock contacts dipping steeply inward at angle of approximately 86°. The pipe has irregular geometry in the upper part where the eastern margin of the pipe flattens out considerably. It was suggested that the irregular morphology of the upper portion of the pipe is possibly controlled by a fault that predates kimberlite emplacement. The dominant infill lithology at Koala North is crudely bedded to massive, relatively mud-rich volcaniclastic kimberlite. This includes fine- to medium-grained crater sediments, ash/mud-rich to olivine-rich resedimented volcaniclastic kimberlite (RKV) and primary volcaniclastic or pyroclastic kimberlite (PVK). Current drilling at Koala North has intersected kimberlite down to the 155 m elevation (~ 270 m below the top of the pipe) and indicates that it is comprised exclusively of volcaniclastic kimberlite material to this depth.
The rock mass weathering susceptibility in the mining context is the strength deterioration within the life of excavation due to the exposure to moisture. The degree of weathering is influenced mainly by the clay minerals within the kimberlite rock types. Weathering of the kimberlite can adversely impact on ground support performance, production blasting and trafficability. Weathering can also promote the generation of mud as experienced in South African operations and it is an important issue that needs to be addressed in any kimberlite mining. An accelerated weathering tests performed on the core revealed that the majority of the Koala North kimberlite has low weathering susceptibility except above mentioned clay-rich kimberlite and mudstone, which form only 5-10% of the rock mass. It is often found that poor quality rock mass is present adjacent to the kimberlite body contacts. This could create mining difficulties in both open pit and underground operations. From the geotechnical point of view, two types of contact zone are usually recognized: the internal contact zone within the kimberlite body and the external contact zone developed in the country rocks. In case of Koala North the external contact zones are variable both in the geometry and competency. Typically, as the development in the country rock approaches the kimberlite pipe contact, there is a zone of granodiorite approximately 1m to 5m in width with increased joint frequency and in some areas there is also a narrow (approximately 1m to 2 m wide) transition zone at the pipe contact in which kimberlite stringers occupy joints within the granodiorite. Although the contact zones are not equally developed around the entire perimeter of the pipe, their presence negatively impact on the dilution. Based on the physical properties of the individual rock types and geological zones, a geotechnical domain model was developed subdividing the rock mass in the mining area into eight geotechnical domains. The rock mass parameters for individual domains are illustrated in Table 1. 5 MINING METHOD The selected mining method for Koala North is open benching. The decision was made as a result of technical, economical and safety risk assessments. Competent country rocks, favorable geometry, relatively competent kimberlite, and most importantly, the arctic context of the projects played an important role in the mining method decision making process. Table 1: Rock mass rating values for individual geotechnical domains
Figure 3: An isometric view of the Koala North pipe and underground development (after Jakubec et al 2003). 4 GEOTECHNICAL CHARACTERIZATION The initial geotechnical information was obtained from the exploration drillholes using Laubscher’s rock mass rating (RMR) classification system (Laubscher, 1990) for both kimberlite and country rock masses. This was later updated with data obtained from the open pit and it is continuously updated with ongoing geotechnical mapping using updated Laubscher’s classification system (Laubscher and Jakubec, 2001). In comparison with other known pipes in the vicinity of the Koala North it was very clear that majority of the kimberlite rocks are relatively competent. The exceptions are clay rich intervals within the RVK units that are relatively weak and show low rock mass competency. Such clay units form irregular bodies and are also highly susceptible to degradation and weathering when exposed to moisture. 434
Geotechnical Domain
RMR
Overburden
---------
Near-surface Granodiorite
35 - 55
Granodiorite
60 - 75
External contact zone
40 - 55
Internal contact zone
18 - 30
Upper RVK
20 - 40
Lower RVK
45 -65
Clay-rich RVK
15 - 30
The natural caving option was rejected because the size of the pipe in terms of the Hydraulic Radius (plan area/perimeter) varies from HR=18m at the surface to HR=10m at the base of the current study zone. These are small values in terms of caving and would require very weak material for an assured cave.
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The open benching is a top-down retreat mining method, similar to sublevel caving but without the caved waste behind the drawn ore.
When the crosscuts are fully developed, the slot drifts connect individual tunnels and the level is ready for the slot development and production blast. Production stope with apex pillars is shown in Figure 5. Individual production levels have been developed at 15 m spacing. The mining front maintained 3 levels in production, maintaining approximately a 45° slope for stability purposes. In the plan view, the front maintained a concave shape with the boundary drawpoints lagging behind the central drawpoint. 6 ACHIEVEMENTS IN ARCTIC CONTEXT
Figure 4: Schematic vertical section of open benching with individual production level geometries (after Jakubec et al 2003). An access ramp to the underground workings was developed from the surface down to the first production level at 2385. The ramp has an arched profile of 5.5 x 5.5 m and all the development drilling was conducted using brine solution due to the presence of permafrost and cold air temperatures. Accesses to the individual production levels are developed from the main ramp at regular intervals. These drives provide access for stope production, exploratory diamond drilling and installation of the mining infrastructure such as sumps, electrical installations and refuge bays. All the level accesses have flat back square profiles of 5 x 5 m. Production crosscuts were developed into and across the kimberlite pipe for slot access, stope drilling and production mucking. These cross-cuts are designed in an arched profile of 4.0 m wide x 4.0 m high. Due to the susceptibility of kimberlite to weathering, all the development and production blast-hole drilling has been completed dry.
6.1 Production rates and Dilution The production rates 1,500 wtpd achieved as planned. Koala North underground operation experiences more dilution than the open pits during normal mining operations. The current average dilution rate is approximately 17%. Half of the dilution is eliminated underground at the draw points, and half is sorted on surface. It has to be noted that approximately 75 % of the dilution is from upper 2385 level, compared to 25 % from 2370 level. This is due to several factors including pit bottom blasting damage, relaxation of the rock mass at the upper edge of the pipe contacts and also due to the underground production blast design for the upper level. Once the rock mass is damaged and "loosen up" then the effect of thaw and freeze up - ice jacking can create small scale rock falls resulting in dilution. 6.2 Operational Issues Freezing muck pile During the winter months the broken muck pile in the stope after the production blast will freeze up if not removed. The level of freezing depends on water content. This can potentially result in operational problems if measures to mitigate the impact would not be implemented. The key to the successful ore recovery is removal of freshly blasted muckpile as soon as possible and prevention of water access to the stope. Any muck that left behind was primarily on the apexes between the drawpoints was recovered on the next level below when it was blasted. Although sometimes experienced, the truck "carry back" of the ore due to freezing has never been a serious problem. Stabilizing effect of frozen pipe walls and the destabilizing effect of ice jacking When the Koala North open pit walls and the "glory hole" walls are completely frozen from October to May no stability problem as observed on exposed pipe walls. In late May with the walls warming up due to the longer exposure to sun and warmer temperatures followed by freezing during the short night small scale rock falls were experienced. This is mainly due to the "ice jacking" effect on open joints – see figure 6. Trafficability and Ice build up When the Panda Ramp had advanced to the point that it was developed below the permafrost water began to seep through the joints in the rock mass. Ice began to build up on the haulage road and it quickly became a problem with equipment moving on the ramp. After evaluating of the alternatives the decision was made to install heaters on the fresh air intake raise and within a few days of commissioning the system the roads were free of ice. During the winter there is also ice build up on the escape ladder way and cleaning procedures were developed to combat the problem.
Figure 5: Production stope after the mucking is completed. Note that frozen muck on top of apex pillar is standing up in very steep angles. Massmin 2004
Fogging - cold and heated air After the heaters were installed on Panda fresh air intake raise a pocket of fog would develop where the unheated air
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The overall volume of shotcrete used per meter of kimberlite development is 1.4 m3. It was found during the testing that the shotcrete set up poorly at -18°C, and never attained full strength, but from –5°C to 20°C, the compressive strength averages 38 MPa which is acceptable for Koala North application.
Figure 6: View into the open stope from the open pit. Note well developed apex pillars and stable blasting face. The main source of wallrock dilution came from the upper levels, combination of poor blasting, pit bottom relaxation and ice jacking. from Koala North mixed with the heated air. A "fog zone" was created between 2245 Level and 2205 Level of the Koala North Ramp. The fog would begin to dissipate once temperatures rose above –30°C. Procedures were developed that allowed traffic to move safely through this zone and prohibited pedestrians. Impact of cold on productivity, men and equipment Temperatures in the underground workings follow the temperature outside with a significant delay of approximately 1 month. Long development headings were still +4.50°C after 3 weeks of –10°C to -15°C outside. The main reason is heat generated by the working equipment at the headings and as well as low airflow due to leakage in the vent ducting. Before the heaters were installed, some of the mine areas could became very cold in the winter months. The open drawpoints in the kimberlite could easily reach 50°C with the wind chill factor. The underground crews work an 11-hour shift but their effective time at the face is 8.5 hours. Based on the assessment of the performance of the underground crews in the winter months, the rotation of the underground personnel was changed from 6 weeks in and 3 weeks out to 4 weeks in and 2 week out. The main reason for this was the impact of the severely cold temperatures. It was felt that six weeks proved to be too long to be exposed to that work environment. The primary impact the cold had on the equipment was on the hydraulic systems. This was especially true at the beginning of the project when ruptured hydraulic hoses were a common occurrence. This was because the equipment was more frequently exposed to the low temperatures on the surface outside the mine. Indirectly, there was also increased damage to the equipment due to the use of brine for drilling. This issue is discussed below separately. Shotcrete mix for cold climate Current shotcrete mix used underground comprises of 480 kg ready mixed shotcrete (Type 30 include silica fume) and 1120 kg of aggregate. In the winter 27.5 kg of CaCl is added to prevent freezing. In order to achieve good quality product it was important to introduce the QA/QC program that covers entire process from batching to application. Daily inspections by the geology/geotechnical staff are the best tool to inspect shotcrete performance. 436
Development and Production drilling Two issues have to be considered: development drilling in permafrost and production long hole drilling in the kimberlite. Due to the susceptibility of kimberlite to weathering, the production drillholes had to be drilled dry. This proved to be very successful and the only problem that has to be combated in the context of the cold climate is re-drill due to the icing up of the drillholes. Although this problem could have a significant impact, it is experienced only during the spring snow melt and freeze-up period, while in the remainder of the year there were virtually no issues with redrilling. The development in the permafrost granite required drilling with brine. Some increased corrosion problems were experienced on underground equipment. It was also found that high quality two-stage settling of solids out of recirculated brine is essential. Roadways construction The key to the successful roadway maintenance system in any underground mine is to keep water out of the running surface. In diamond mines this rule even more important due to the weathering susceptibility of the kimberlite. Harsh winter condition can add another level of difficulties to the roadway maintenance especially during the spring time of the year. In Koala North all the drawpoints were driven at +4% grade to allow water to drain quickly and for ease of equipment recovery. Three roadway designs were tested on 2355 mining level; 150mm granite crush, graded kimberlite, and "geogrid" beneath 100mm crush. The 100mm granite crush as required on kimberlite floor provided the optimal acceptable running surface. Kimberlite roadways when frozen from November to June - provide excellent pavement surface. 6.3 Stress Release In order to accelerate the development of an access to Koala and Panda pipes it was decided to temporarily suspend the mining of Koala North. The production levels that were already developed were retreated all the way to the granite contacts but no slot was excavated on level 2340 below. While the level 2355 was approximately half way mined out, some shotcrete cracking occurred in the tunnels and on level 2340 below. The intensity of cracking on level 2340 increased towards the slot end of the pipe. Stress measurements in the vicinity of the pipe were conducted by CANMET and numerical modeling using FLAC3D was undertaken by ITASCA. The model was calibrated to the mining sequence and the level of cracking was reproduced. The results from the Koala North modeling are used to evaluate the geometry, support and mining strategy for other underground projects. 7 CONCLUSIONS The successful introduction of open benching at Koala North kimberlite pipe is very encouraging. Although this mining method was successful on several De Beers diamond operations in South Africa, it has not been tested
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in the harsh arctic environment prior to commissioning the Koala North underground operation. The long term data are not available yet and initial experiences are influenced by the learning process. For example the 2003 spring snowmelt in comparison with previous year had only relatively minor production impact and in total only three shifts were lost due to water issues, and approx. one additional week of 25% reduced production. While it is probably too early to properly assess all the aspects of the operation very valuable lessons were learned. The knowledge gained at Koala North will contribute to the planning processes for the other underground mining projects
REFERENCES • Jakubec, J., Long, L., Nowicki, T., Dyck D., 2003. Underground Geotechnical and Geological Investigations at EKATI Mine – Koala North: Case Study. 8th International Kimberlite Conference Abstracts. Victoria, Canada • Laubscher, D.H, Jakubec, J., 2001. The MRMR Rock Mass Classification for Jointed Rock Masses. Underground Mining Methods, Society for Mining, Mettalurgy, and Exploration, Inc. (SME). Littleton, Colorado, pp. 475 – 483. • Laubscher, D.H., 1990. A geomechanics classification system for the rating of rock mass in mine design. J.S. Afr. Inst. Min. Metall. vol. 90. no. 10. South Africa, pp.257 - 273.
ACKNOWLEDGEMENTS The authors would like to thank BHP Billiton Diamonds for permission to publish this paper. Also, the help of Tyla Hay of SRK in making this paper a reality is greatly appreciated.
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438
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Chapter 11
Mine Planning II: Case Histories
440
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The development and implementation of a mixed integer programming model for production scheduling at the Kiruna Mine Mark Kuchta, Department of Mining Engineering, Colorado School of Mines
Abstract Production scheduling is a critical aspect of planning the operation of an underground mine. While mathematical optimization techniques have been widely applied to production scheduling for surface mines, the application of optimization techniques such as Mixed Integer Programming (MIP) to multi-time period underground mine scheduling has been severely limited due to the large number of integer variables and complex sets of constraints required, which results in unrealistically long solution times. A new long-term MIP-based production scheduling model has been developed and integrated into the mine planning system at LKAB’s Kiruna mine. The optimization model uses a new block data format for which production data for mining areas is preprocessed into monthly production quantities. This data structure allows for a significant reduction in the number of integer variables required. Algorithms to determine the earliest and latest possible start dates for production areas further reduce the number of integer variables. In this paper, the model development is outlined and the results are illustrated using practical scheduling examples from the Kiruna mine.
1 INTRODUCTION Production scheduling for large underground mines is often a complex and time-consuming task. Due to the complexity of the problem, heuristic scheduling algorithms are often used. While heuristic algorithms may produce useable schedules, there is no way of knowing how close the schedule is to some desired optimum, for example, a schedule that minimizes total mining costs. Mathematical programming methods provide a means of producing "optimal" schedules. While the application of such techniques to open pit mine scheduling problems has been widely used, their practical application has been limited in underground mining. Mixed Integer Programming (MIP) is a mathematical programming technique that can be applied to the underground scheduling problem. The practical use of MIP has been hindered because models must incorporate a large number of decision variables, many of them restricted to assume integer values. The large number of integer variables required for model formulation results in commensurately long solution times that may be unacceptable for practical planning purposes. By preprocessing the production data and through careful model formulation, it is possible to reduce the number of integer variables and thus greatly reduce solution times. This paper summarizes a new mixed integer programming model developed for the Kiruna Mine. The scheduling system developed has been integrated into the mine’s computerized mine planning system and is used for both monthly and strategic planning. Five-year production schedules with monthly fidelity can usually be produced in about 10 minutes. 2 THE KIRUNA MINE The Kiruna mine is located in Northern Sweden above the Arctic Circle. It is owned and operated by the Swedish mining company LKAB. The ore body is a world-class highgrade magnetite deposit, approximately 4 km long, 80 m Massmin 2004
wide, with a strike roughly in the north-south direction and a dip of about 70 degrees. Two basic ore types are found insitu: a low phosphorous content very high grade magnetite ore known as B ore, and a high phosphorous magnetite ore known as D ore. The ore body is characterized by clear-cut boundaries at the hangingwall and footwall contacts and between the two primary in-situ ore types. From the two in-situ ore types, the mine produces three raw ore types based primarily on their phosphorous contents known as B1, B2 and D3 ore. The three run-ofmine ore types supply four mills. The B1 ore type is used for fines production and feeds a single mill, the B2 ore type is used for pellets production in a mill not equipped with a floatation circuit, and the D3 ore is used to supply two palletizing plants, each equipped with flotation circuits for excess phosphorous removal. There is very little stockpiling available between the mine and the mill, which places enormous demands on the mine to carefully plan the extraction sequence so that the proper quantities of the three ore types can be delivered to the four mills. Kiruna produces about 24 million tons per year using large scale sublevel caving (Fig. 1). Transversal sublevel caving is normally used with mining proceeding from the hangingwall to the footwall. The spacing between sublevels is about 27 m and the spacing between crosscuts is about 25 m. The production drifts are 7 m wide and 5 m high. From the production drifts, rings of holes are drilled upward in a fan-shaped pattern at an inclination of 10 degrees forward (towards the hangingwall), with spacing between rings of 3 to 3.5 m. Each production ring contains around 10,000 tons of ore. The orebody is currently being mined from the 1045 m main transportation level (Fig. 2). The mine is divided vertically into 10 production areas, each with an extent of 400 to 500 m. One or two large capacity Load Haul Dump units (LHDs) operate within each production area on a given sublevel. The area where an LHD operates is known as a machine placement. A series of ore passes known as a shaft group runs from approximately the center of each
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production area down to the main transportation level. The ore passes extend down to the 1045 m transportation level where the ore is transported by large trains to the main crushers and then hoisted to the surface through a series of vertical shafts. Mining begins at the uppermost sublevels and proceeds sequentially downwards.
MP1
MP2
1, 50
1, 10
2, 40
2, 20
3, 30
3, 30
4, 20
4, 40
5, 10
Figure 3. Two production areas MP1 and MP2 with 5 and 4 production blocks, respectively. The block number and reserves in tons are given.
Indices b = production block t = time period Sets B V
Figure 1. Mining by sublevel caving.
Sb
= set of all production blocks = set of blocks for which a sequencing constraint exists = set of blocks by which block b is constrained
Parameters T = number of time periods Rb = total tons available in block b Dt = production demand for time period t Variables Xb,t = tons mined in block b during time period t Yb,t = 1 if block completely mined by time period t, 0 otherwise Ut = under production from demand in time period t Ot = over production from demand in time period t Objective t
∑
minimize
(Ut + Ot)
t=1
Figure 2. The 1045 m main transportation level at the Kiruna Mine.
Constraints
∑ Xbt + Ut + Ot = Dt , ∀t
demand:
b
3 BASIC MODEL FORMULATION t
A number of different types of sequencing constraints are required for MIP scheduling models. The most common is a constraint that simply requires that all of a constraining block be mined before mining can commence in a constrained block. Constraints of this type are always required when the reserves for a block are greater than the tons that can be mined from a block in a single time period. Such constraints require one integer variable per constrained block and time period. The number of integers required increases exponentially as number of time periods increases. Figure 3 shows a series of blocks that must be mined sequentially, i.e., an overlying block must be mined before the block immediately under it can be mined. Assume that a production schedule is to be determined which minimizes the deviation from demand for ore in each time period. The mathematical formulation for producing such a schedule using MIP is given below:
442
sequence 1:
∑ Xb,u > = Yb,t Rb ,
∀bt ≤ T
u=1
t
sequence 2:
∑ Xb,u < = Yb,t Rb , ∀b∈ V, b’ ∈ Sb, t u=1
t
reserves:
∑ Xb,t < = Rb
, ∀b
u=1
Xb,t , Ut, Ot ≥ = 0, Ybt binary Sequencing constraint 1 turns the binary variable Yb,t for block b to 1 in the time period in which all the reserves for that block have been mined. Sequencing constraint 2 insures that nothing can be mined from the constrained block b until all the material from the constraining block b’ has been mined. The reserve constraint insures that no more than the available reserves can be mined from a given
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block, and the demand constraint insures that the demand for ore is met in each time period while allowing for both over and under production. The objective function seeks to minimize the over and under production. Using this type of sequencing constraint, a production schedule consisting of five time periods for the two production areas shown in Figure 3 would require a total of 20+15 = 35 integer variables. Several attempts at multi-period scheduling using mixedinteger programming have been made at the Kiruna Mine using basic model formulations similar in structure to that described above (Almgren, 1994) and (Dagdelen, et al., 2002). In both cases the models could not be solved for multiple time periods due to the excessively large number of integer variables required and the resulting long solution times. The need for a sequencing constraint that insures that all the available tons from a constraining block be mined before anything can be mined from a constrained block can be completely eliminated by preprocessing the production data into production blocks that exactly contain the amount of material that can be mined during one time period. Integer variables are still required for scheduling; however, only one integer variable is required per production area per time period regardless of the number of blocks contained within the production area. With such a formulation, the number of integer variables increases linearly rather than exponentially as the number of time periods increases. The mathematical formulation for a production schedule that minimizes the deviation from demand for ore using preprocessed block data is as follows: Indices a = production area b = production block t = time period Sets A Ba Bt
= set of all production areas = set of blocks within production area A = set of blocks that can be mined in time period t.
Parameters T = number of time periods Rb = total tons available in block b Dt = production demand for time period t Variables Ya,t = 1 if mining starts in production area a in period t, 0 otherwise Ut = under production from demand in time period t Ot = over production from demand in time period t Objective t
minimize:
∑ (Ut + Ot ) t =1
Constraints
Notice that the formulation contains no sequencing constraints. As with the previous model, the demand constraint insures that the demand for ore is met in each time period while allowing for both over and under production, and the objective function seeks to minimize the over and under production. The constraint "limit one" insures that at most one Y variable for each production area can be 1. The key to the formulation is the demand constraint, which for 5 time periods using the data given in Figure 3 would expand as follows: 1) 50*Y1,1 + 10*Y2,1 + U1 – O1 = D1 2) 40*Y1,1 + 50*Y1,2 + 20*Y2,1 + 10*Y2,2 + U1 – O2 =D2 3) 30*Y1,1 + 40*Y1,2 + 50*Y1,3 + 30*Y2,1 + 20*Y2,2 + 10*Y2,3 + U1 – O2 =D3 4) 20*Y1,1 + 30*Y1,2 + 40*Y1,3 + 50*Y1,4 + 40*Y2,1 + 30*Y2,2 + 20*Y2,3 + 10*Y2,4 + U1 – O2 =D4 5) 10*Y1,1 + 20*Y1,2 + 30*Y1,3 + 40*Y1,4 + 50*Y1,5 + 40*Y2,2 + 30*Y2,3 + 20*Y2,4 + 10*Y2,5 + U1 – O2 =D5 Figure 4 shows the overall structure of the formulation in spreadsheet format. The reserves for two production areas are shown in column 1. Columns 3 to 7 give the time periods. The value of the Y variable for each time period for each production area is given along with the corresponding production. The total production, demand, and deviation from demand for each time period are shown in the last three rows. As can be seen with this simple example the demand of 50 units per time period can be met exactly by starting mining in production area 1 in time period 1, and in production area 2 in time period 2. Time period
1
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Reserves MP1 12345-
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Figure 4. Spreadsheet representation of a production schedule using preprocessed production data.
t
demand:
∑ ∑ Rb Ya,t + Ut – Ot = Dt ∀t a ∈ Α b∈ Ba ∩ Bt t
limit one:
∑ Yat ≤ 1 t =1
Ut, Ot ≥ 0, Yat binary
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, ∀a
The formulation for this simple example requires 10 integer variables compared to 35 required with the previous formulation. While in this simple example the reduction is small, with large model containing many production areas and hundreds of production blocks, the reduction in the number of integer variables can be dramatic. Often, solutions for large models can be found quickly using the second model, whereas using the first model solutions Santiago Chile, 22-25 August 2004
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simply cannot be found at all in a reasonable amount of time. 4 MODEL EXTENSIONS The basic model structure using preprocessed production data outlined in the previous section has been applied to the problem of long-term production scheduling at the Kiruna mine. For planning purposes, the ore body is divided into 100 m blocks, extending from the hangingwall to the footwall, and with the block height equal to the sublevel height. Estimates of the amount of B1, B2, and D3 ore that can be extracted from each 100m block are then made. These estimates take into account the blending of the in-situ high phosphorous D-ore with the in-situ low phosphorous B ore during extraction (Kuchta, 2002). Results from mining of the upper sublevels are incorporated into the block estimates. A machine placement, defined as the area on a sublevel within a production area where an LHD operates usually consisting of one to four 100m blocks, is the basic planning unit for long term strategic production scheduling. For each machine placement, the 100m block data is processed into blocks that contain the amount of the three ore types that can be extracted in one month. These block estimates take into account the planed extraction sequence for the block as well as the estimated LHD capacity. This preprocessing is key to the model performance since it eliminates the need for binary variables that sequence the extraction of blocks within a machine placement. The block data for a typical machine placement is shown in Figure 5. A five-year strategic plan typically contains up to 60 machine placements each with an average of 15 to 18 monthly production blocks representing a total of over 125 million tons of ore.
constraint that limits the number of LHDs that can be active simultaneously within the same shaft group, and (iv) limit start, a constraint that limits the number of machine placement s that can be started during a planning period. The horizontal and vertical sequencing constraints make use of the existing binary variables that represent whether to start each machine placement in a given time period. As such, these constraints do not require introducing any additional integer variables. The complete mathematic formulation for the Kiruna MIP scheduling model can be found in (Kuchta, et al., 2003 and Kuchta, et al., 2004). The first year of a typical 5-year schedule produced using this model is shown in Figure 6. Twenty machine placements were scheduled in the first year yielding a production of 22 million tons.
Figure 6. Example of a one-year monthly schedule.
5 VARIABLE REDUCTION
Figure 5. Production plan for a typical machine placement. The objective of the MIP scheduler is to determine the start dates for each machine placement such that the deviations from the planned quantities for the three ore types (B1, B2, and D3 ore) are minimized, subject to various operational constraints that are tied to the mine layout and mining aspects associated with the sublevel caving mining method. The major operational constraints are: (i) vertical sequencing, a constraint that insures that at least 50% of an overlying machine placement has been mined before mining commences in an underlying machine placement, (ii) horizontal sequencing, a constraint that insures that mining commences in the machine placement to the right and left of a given machine placement when 50% of the given machine placement has been mined, (iii) shaft group, a 444
The principle variables in the model are binary variables indicating in which time period the mining of a given machine placement is to start. Various mining sequencing requirements can be used to eliminate binary variables corresponding to time periods when it can be shown that it would not be possible to start the mining of a machine placement. In most cases it is possible to use a simple procedure for establishing the earliest possible start date for each machine placement. Figure 7 shows a small section of a mining layout that is representative of the Kiruna Mine. There are two production areas, each with its own shaft group, SG1 and SG2. There are two sublevels, L1 and L2, and each production area on each sublevel has been divided into two machine placements. For simplicity, each machine placement is assumed to contain 10 monthly production blocks, each with the same tonnage. Referring to Figure 7, if machine placement MP1 is assigned a start date of 1, the vertical sequencing constraint requiring that 50% of the overlying machine placement be mined before mining of an underlying machine placement can begin would require the earliest possible start date for MP5 located immediately under MP1 to be 6. All binary
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with 60 monthly time periods. Using the early satrt and late start variable reduction procedures described above, the model size will typically be reduced to around 900 integer variables. The difference in solution times with such a reduction is huge. By preprocessing the production data into monthly production quantities and applying additional variable reduction techniques, five-year production schedules with monthly fidelity can usually be produced in about 10 minutes. 6 SUMMARY
Figure 7. Small section of a mining layout.
variables used to indicate whether or not to start mining MP5 in time periods 1 to 5 can be removed from the model. Additionally, the number of machine placements that can be active within a given shaft group is limited, i.e., mining of a given machine placement cannot begin until an ore pass becomes available. Assume both MP1 and MP2 are started in time period 1, and that the maximum number of machine placements allowed to be active in shaft group 1 is 2. This restriction would establish an earliest possible start date for both MP5 and MP6 of 11. The binary variables corresponding to this delay can also be eliminated. Finally, the left and right sequencing constraints that require that mining commence in an adjacent machine placement when 50% of the given machine placement has been mined can also be used to establish the earliest possible start date for a machine placement In some cases it is also possible to use a simple procedure for establishing the latest possible start date of a machine placement. If the start date for any one machine placement on a sublevel is known, the left and right sequencing constraints that require that mining commence in an adjacent machine placement when 50% of the given machine placement has been mined can be used to establish a latest possible start date for all other machine placements on the same sublevel. For example, if MP2 is started in time period 1, the latest possible start dates for both MP1 and MP3 would be 6. All binary variables for a machine placement greater than the latest possible start date can be eliminated. A complete description of the early start and late start algorithms described above as well as an implementation in the AMPL programming language is given in Topal (2003). A five-year strategic plan typically contains up to 60 machine placements and with no variable reduction would require 3600 integer variables for a production schedule
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A new mixed integer-programming model has been developed and is currently being used for long-term production scheduling at LKAB’s Kiruna mine. The model makes use of a new data structure based on preprocessing the data for each production area into blocks containing the amount of ore that can be mined in one time period. This data structure significantly reduces the number of integer variables required. Algorithms that determine the earliest and latest possible start dates reduce the number of integers required further. Five-year production schedules using the new system can usually be produced in less than 10 minutes. Research is currently being conducted in extending the model towards shorter term weekly planning. The techniques described in this paper have been shown to be useful for developing production schedules for a large mine using the sublevel caving mining method. It should also be possible to apply this model to operations using other underground bulk mining methods such as sublevel stoping. ACKNOWLEDGEMENTS The authors would like to thank LKAB for the opportunity to work on this challenging project and for permission to publish these results. REFERENCES • Almgren,T., 1994, "An Approach to Long Range Production and Development Planning with Application to the Kiruna Mine, Sweden", Lulea University of Technology, Doctoral Thesis number 1994:143D. • Dagdelen, K., Kuchta, M., Topal, E., 2002, "Linear Programming Model Applied to Scheduling of Iron Ore Production at the Kiruna Mine, Kiruna, Sweden", Transactions of the Society for Mining, Metallurgy, And Exploration, Inc., Vol. 312, 2002, pp 194-198 • Kuchta, M., 2002, "Predicting Run-of-Mine Ore Grades for Large-Scale Sub-Level Caving at LKAB’s Kiruna Mine, Transactions of the Society for Mining, Metallurgy, And Exploration, Inc., Vol. 312, 2002, pp 74-80 • Kuchta, M., Newman, A., and Topal, E., 2003, "Long Term Production Scheduling at LKAB’s Kiruna Mine", Mining Engineering, The Society of Mining, Metallurgy, and Exploration, Inc., Vol. 55, No. 4, April 2003, pp 35-40 • Kuchta, M., Newman, A., and Topal, E., 2004, "Implementing a Production Schedule at LKAB’s Kiruna Mine, Interfaces, Vol. 34, no 2, March-April 2004, pp 124134 • Topal, E., 2003, "Advanced Underground Mine Scheduling Using Mixed Integer Programming", PhD Dissertation T-5733, Colorado School of Mines, Golden, Colorado.
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Complex cutoff grade optimization at the Kiruna Mine Emmanuel Henry, Senior Geostatistician, AMEC, Canada Kjell Klippmark, President, KGS, formerly Mine Manager, LKAB, Sweden
Abstract LKAB operates the Kiruna mine, a large-scale sublevel caving (SLC) operation in Northern Sweden, and provides the market with iron products of different qualities and prices. Defining an optimal draw point closure cutoff grade is a key concern for SLC mines. It has a direct effect on operational costs, mineral resource recovery, market deliveries and corporate financial results in the short and long term. Lane’s (1988) Cutoff Grade Theory gives a framework to address this issue, but must be adapted to account for several particularities at the Kiruna Mine. First, the grade-tonnage curve must account for draw point dilution, a particular feature of SLC, and second, complex process structures must be modeled with several parallel flows. Each flow has its own blending, cost, and capacity constraints, which eventually become non-linear. Finally, fluctuations in the market capacity for the various iron ore products impose different marketing capacity constraints. The paper describes a cutoff grade optimization model that was built specifically for the Kiruna Mine. Based on market conditions, the model determines the necessary cutoff grade to meet a required return rate, optimal production rates, and the location of limiting capacities (bottlenecks in the process). The underlying assumptions, results and potential significant economical benefits are also discussed.
1 INTRODUCTION LKAB operates the Kiruna Mine, a large-scale sublevel caving (SLC) operation in Northern Sweden, and provides the market with iron fines and pellets. Iron ore producers are subject not only to variations in the metal price, negotiated once a year between major market suppliers and buyers, but also to demand, which is very dependent of the global economic conjuncture. The higher the economic conjuncture (e g the more cars sold in the World), the more steel consumed, and thus the more iron ore bought. In order to keep competitive with major low cost open-pit producers, LKAB invested heavily in process optimization and cost reduction. One aspect that was overlooked until recently was the potential impact of an optimization of the cutoff grade on LKABs financial performance. A cutoff grade optimization theory, intrinsically a process optimization tool, was developed by Taylor in 1972. He demonstrated that maximizing the cash flow of a mine by using an optimal cutoff grade is a function of costs, capacities, and a grade-tonnage curve. Lane (1988) integrated financial constraints, such as the rate of return, and long-term resource consumption optimization constraints in the model, and made it a strategic planning tool. An important conclusion of both theories is that the break-even cutoff grade that is applied by many mines is rarely optimal (see Hall, 2003). A model was developed specifically for the Kiruna mine in order to demonstrate the economic potential of adapting the cutoff grade to market capacity (Henry, 2003). Adaptation of Lane’s theory, however, was not straightforward since process flows are very complex, with several ore qualities giving rise to several products processed and sold in parallel. A very particular feature in SLC mining operations is the distortion of the in situ grade-tonnage curve due to waste-rock dilution at draw points. 2 BACKGROUND Iron ore grades in the Kiruna mine run-of-mine (ROM) depend on in situ iron grades in the raw magnetite and 446
waste-rock dilution (expressed as the weight percentage of waste-rock in a tonne of ROM). Internal waste-rock dilution comes from in situ waste-rock intrusions in the SLC production rings and external dilution comes from the blending of caved waste-rock with newly fragmented ore at draw points. External dilution increases with increasing draw, defined as the percentage of the tonnage produced from a ring compared to its theoretical in situ tonnage (usually 10,000 t). Iron grades are relatively constant in the magnetite and external waste-rock dilution is the most influent factor on the ROM iron grade. ROM waste-rock content is monitored on the mine’s main haulage level but it is not dynamically controlled. Annual averages have been relatively constant around 36-37 %, and when market capacity is high, plants generally complain that the ROM contains too much waste-rock and not enough magnetite to fulfill market capacity, and ask waste-rock dilution to be lowered. The mine usually answers that lowering waste-rock dilution by decreasing the draw would increase mine development costs, and more important, would result in "high-grading" of the resource. This controversy motivated the construction of a wasterock dilution optimizer to find the waste-rock dilution that would maximize the value of the mine in function of the market capacity, using Lane’s Theory (Lane, 1988). 3 PROCESS MODELING The first step in the construction of a cutoff grade optimizer consists in modeling the mining process sensu largo (i.e. from mine development to marketing). The process is described by material flows, capacity structure, and cost structure. 3.1 Material Flow Model The Kiruna mine produces at least three qualities of finished products (f p) from three ROM qualities. Lowphosphorus and low-alkali material (B1) is used to produce fines (KBF) in a plant in Kiruna. Low-phosphorus and highalkali material (B2) is used to produce pellets (SPBO) in a
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plant located in Svappavaara, 50 km south of Kiruna. Finally, high-phosphorus material (D3) is processed into pellets (KPBO) in Kiruna. Processing of the in situ iron resource is divided in the following steps (terminology derived from Lane, 1988; see Figure 1): • Mining: Tunneling and production drilling of SLC rings. • Treating, Step 1: Ore fragmentation, loading, and hoisting. • Treating, Step 2: Primary grinding and primary separation. • Treating, Step 3: - Secondary grinding and enrichment into KBF, or, - Transport by train to Svappavaara, secondary grinding, enrichment, and pelletizing into SPBO, or - Secondary grinding, enrichment, and pelletizing into KPBO. • Marketing: Transport by train to harbors and shipment to customers. The model considers a tonne of ROM loaded from a draw point and dumped in an ore pass containing a proportion of grb (gråberg = waste-rock, in Swedish) of external wasterock and a proportion 1-grb of constant iron grade magnetite (internal dilution and iron grade variations are ignored). For each 1-grb tonne of magnetite loaded and dumped in an ore pass: • 1-grb t is hoisted. • 1-grb t goes through the primary grinding and separation. • A part of (1-grb) is treated in the KBF-process. • A part of (1-grb) is treated in the Svappavaara-process. • A part of (1-grb) is treated in the Kiruna KA1/KK-process. • A part of (1-grb) is treated in the Kiruna KA2/KK-process. For each grb tonne of external waste-rock loaded and dumped in an ore pass: • grb t is hoisted. • grb t goes through the primary grinding and separation. • A part of grb is separated away in the primary separation line and is dumped on waste dumps. • A part of grb follows in the KBF-process, where it is separated and stored in tailing dams. • A part of grb follows in the Svappavaara-process, where it is separated in the enrichment plant and stored in tailing dams. • A part of grb follows in the KA1 enrichment process, where it is separated and stored in tailing dams. • A part of grb follows in the KA2 enrichment process, where it is separated and stored in tailing dams.
Mining: Ore Development
Treating, Step 1: Ore production and hoisting
Treating, Step 2: Grinding and primary separation
The proportions of magnetite (1-grb) and waste-rock (grb) entering each sub-process (KBF, Svappavaara, KA1, and KA2) were determined by analyzing historical data. They will most likely vary with the differential demand for KBF, SPBO, and KPBO, and will also depend on grb. Some of the magnetite which is fed into the primary separator and following enrichment plants is lost with the separated waste-rock, resulting in magnetite recovery less than 100 %. Recoveries achieved in each of the KBF, Svappavaara, and KA/KK processes are different and functions of grb. The proportion grb is directly related to the draw of a SLC ring. The greater the draw, the greater the proportion grb. 3.2 Capacity Structure Capacity constraints limit the quantity of material that can be transported and processed at each step of the process. A simplified capacity structure breakdown into major process unit is proposed: • Annual mine development capacity (tunneling and production drilling): M. • Annual ore production capacity (blasting, LHD loading, hoisting, and primary grinding and separation): H1. • Annual enrichment and pelletizing capacity: H2. • Annual market capacity (train transport, harbour stockpiles, demand): K. Capacities are rarely constant and are often multivariabledependent. Hoisting capacity, for example, depends on the proportions of B1, B2, and D3 ore qualities produced since these qualities must be hoisted separately. Plant capacities are dependent not only on the waste-rock proportion, but also the moisture content and the fragmentation quality. A high waste-rock proportion requires the raw material to be re-circulated several times before an adequate quantity of waste is removed, resulting in decreased plant capacities. Whenever possible, non-linear relations should be accounted for in the model. Although non-linear effects between waste-rock dilution and plant capacities are suspected, a valid mathematical model was not available at the time of the model construction and capacities were therefore modeled constant. Mine development capacity must be expressed in tonnes of resource, while for practical reasons mine personnel reports it in meters of tunnel or in meters of blast-holes drilled. A draw point can produce an almost infinite quantity
Treating, Step 3: KBF process
Marketing: KBF fines
Treating, Step 3: Svappavaara process
Treating, Step 3: KA1 process
Marketing: SPBO pellets
Treating, Step 3: KK1/KK2 process
Marketing: KPBO pellets
Treating, Step 3: KA2 process
Figure 1: Ore flow model at the Kiruna mine. Massmin 2004
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of raw material, but in the model, it is fair to assume that after loading 150 % of its nominal in situ 10,000 t, a ring does not produce any magnetite. Thus only 15,000 t useful tonnes can be produced from a draw point. There is a 3 mspacing between production rings, and about 65 % of the total tunnel length developed in Kiruna is for production purpose. This implies that about 4,6 m of tunnels must be driven for each 15,000 t developed in situ resource. For example, a tunneling capacity of 15,000 m/yr would then correspond to 3,261 draw points and about 48,9 Mt of resource. The nominal hoisting capacity is 26 Mt/yr. If the real LHD production capacity was only 22 Mt/yr, this real capacity should be used as parameter H1 in the optimization. The optimization must reflect the actual process state, not theoretical ones. This is also true for all the other capacities: M, H2 and K. Capacities H1, H2, and K must also account for the material produced by tunneling, since it produces a significant amount of ore and waste-rock, around 1 Mt/yr. For example, if the true hoisting capacity is 23 Mt and 1 Mt of material are produced by tunneling, the final capacity available for ore production (from rings) should be reduced to 22 Mt. If the market can take 11 Mt f p only, and the mine development generates 0.5 Mt f p, then the optimization should be run using a total demand of 10.5 Mt f p. Ideally, the model should be dynamic and optimize the cutoff grade as well as the quantity of tunnel needed at the same time. However, this desirable level of sophistication would require integrating long-term mine planning, market planning, and certainly a resource block-model, in order to account for mining infrastructures that must be developed in advanced (e g new ramps). One of the biggest assumptions in the model is that the market capacity K is a summation of several semidependent market capacities for KBF, SPBO, and KPBO. Modeling individual and differential capacities would be
more accurate but would also require a much more sophisticated model which would integrate individual gradetonnage curves for B1, B2, and D3 ROMs, taking into account quality mixes at draw points. 3.3 Cost Structure Each sub-process induces fixed and variable costs. The following variable costs are accounted for in the optimization model: • Mine development. • Ore production, excluding primary grinding and separation. • Primary grinding and separation. • Waste-rock handling at primary separation. • KBF-production. • Waste-rock handling in KBF-process. • Svappavaara enrichment and pelletizing. • Waste-rock handling in Svappavaara. • KA1 enrichment. • Waste-rock handling in KA1. • KA2 enrichment. • Waste-rock handling in KA2. • KK pelletizing. • Train transport and harbours. Fixed costs were all regrouped in a single variable. The separation between fixed and variable costs is essentially a question of time perspective. Most personnel costs, for example, could be considered as fixed costs. Most material costs, like rock bolts in the mine, or contracted maintenance costs were considered variable. The model incorporates the very important opportunity cost defined by Lane (1988). This cost should be evaluated considering the present state of the remaining mining reserves, market capacity forecasts, and the cost of capital. It is indeed the parameter that links short-term process optimization (immediate cash flow maximization) to mine
Figure 2: Waste-rock dilution at draw points. 448
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value maximization on long-term (net present value maximization). 3.4 Marketing Price Although market capacity was not modeled differentially, marketing price was different for KBF, SPBO, and KPBO. In iron marketing, it is not unusual to discount prices over the official price negotiated every year in Germany and Japan. This is used primarily when iron producers wish to enter a new market or in periods of low conjuncture. It is important to account for these discounts in the optimization. 4 GRADE-TONNAGE CURVE The second step in a cutoff grade optimization consists in defining a grade-tonnage curve. The diluted grade-tonnage curve, not the in situ grade-tonnage curve, must be used. Although there are some theoretical curves describing waste-rock dilution as a function of increasing draw, they could not be validated on real curves obtained from test draw points. Rings with early waste-rock dilution are generally stopped early, giving low draws, while rings with late dilution are stopped late, giving high draws. Although waste-rock dilution is monitored for most rings at the Kiruna mine, this makes it impossible to build an unbiased average wasterock dilution curve (the higher the draw, the higher the bias). Figure 2 shows the diluted grade-tonnage curve finally used in the optimization. It relies upon the two following points: • The very first tonnes from a draw point are expected to be waste-rock-free, provided that the crosscut is located in magnetite (this is an approximation since on the footwall side, the lowest part of the ring is composed of wasterock). Average waste-rock dilution at 0 % draw should therefore be close to 0 %.
• The average waste-rock dilution for the whole year in 2002 in Kiruna was 36 %, which corresponded to an average draw of 104 %. This curve is one of the most uncertain parameters in the optimization.
5 OPTIMIZATION PROGRAM The model was programmed in Excel using 37 input flow, capacity, cost, price and grade-tonnage-curve parameters. For a series of draw increments between 0 % and 150 %, it calculates the average cumulative waste-rock dilution in the material produced from a draw point, the increment of present value per resource unit utilized using functions adapted from Lane (1988), the net cash flow before taxes generated by a ring and by the whole Kiruna operation, and the numbers of consumed rings and produced tonnes of finished products. Lane (1988) demonstrates that in order to maximize the value of a mine, it is enough to maximize the increment of present value per resource unit utilized, which is a function of costs (variable, fixed, and opportunity costs), price, grade-tonnage curve, recoveries, and process limiting capacity. The program calculates the increment in present value if M, H1, H2, or K are the limiting capacities, alternatively. Figure 3 illustrates a configuration where the market capacity (K) is low, 10.5 Mt. The optimum waste-rock dilution is then determined by following the lowest curve pass (M limiting between 0 % and 21 % waste-rock dilution, K between 21 % and 44 %, and H1 limiting above 44 %) and finding the maximum on this pass, which occurs at about 36 %. At this waste-rock dilution level, the increment in present value is 13.9 kr.
Figure 3: Increment in present value of resource utilized in function of the average waste-rock dilution, market capacity = 10.5 Mt. Massmin 2004
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Figure 4: Increment in present value of resource utilized in function of the average waste-rock dilution, market capacity = 15 Mt. 6 HOW MUCH IS THIS WORTH? The value maximization curve in Figure 3 is relatively flat around the optimal 36 %, indicating a departure to this optimum has little impact on the mine’s value. In order to understand the impact of market capacity on the optimal waste-rock dilution, the market capacity was increased to 15 Mt. The resulting optimal waste-rock dilution is then 30 %, as illustrated in Figure 4. The optimum waste-rock dilution is then determined by the balancing mine development and hoisting capacities, M and H1. The optimal waste-rock dilution is also lower than for a market capacity of 10.5 Mt: Richer material is sent into the process, so that more finished products are made from each tonne loaded at draw points (less waste-rock circulates in the system). The process is economically more efficient, at the cost of increased resource consumption, however. A decrease in these bottleneck capacities M and H1 would have a direct impact on the net economy of the mine in periods of high market capacity. Similarly, increasing H1, by renting additional LHDs for example, would increase the mine value and the optimal waste-rock dilution. Investing in enrichment or pelletizing capacity (H2) increase, however, would have no effect on the mine value and would be a waste of money. Producing at 36 % waste-rock dilution instead of 30 % when the market capacity is 15 Mt generates a significant loss of value, as detailed in Table 1. A direct cash flow loss of 242 Mkr is incurred as less finished products can be shipped to customers, thus less net earnings can be generated. Rings are "sold" for a net earning of 318,919 kr instead of their potential maximum value of 358,743 kr. Obviously, the value of such an optimal waste-rock policy will vary considerably with the market level. It is during high market capacity periods that it will have the most significant economical impact. 450
Table 1: Cash flow generated for two waste-rock dilution scenarios, 30 % and 36 %, and market capacity = 15 Mt Waste-rock dilution
30 %
36 %
Cash flow1 generated by an in situ tonne of magnetite developed (kr/t)
23,9
21,3
Cash flow1 generated by a ring (kr/ring)
358,743
318,919
897
655
2,500
2,056
-
242
Global operation cash flow1 (Mkr) Number of consumed rings Direct
loss2
(Mkr)
1-Net before tax; 2-Direct loss = 897 – 655 = 242 Mkr
7 THE DRAW POINT DILUTION CURVE, A STRATEGIC TOOL The optimal waste-rock dilution calculated from the model relies heavily on the grade-tonnage curve in Figure 2. The sensitivity of the optimal waste-rock dilution to the gradetonnage curve was tested for a market capacity of 15 Mt. The new grade-tonnage curve is illustrated in Figure 5. Draw points are now assumed to produce only waste-rock from 120 % draw and above. The resulting optimum waste-rock dilution is 37 %, for an increment in present value of about 15.7 kr/t, as showed in Figure 6. It corresponds to a draw of 88 %. If the true curve was the one of Figure 2, but the optimization was performed on the curve of Figure 5, the (incorrect) optimal waste-rock dilution would cost about 280 Mkr. This amount (remember it is for one year only!) gives a perspective on the value of a research and development program aiming at determining the true grade-tonnage curve achieved in operation.
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Figure 5: Waste-rock dilution at draw points, second scenario.
Figure 6: Increment in present value of resource utilized in function of the average waste-rock dilution, market capacity = 15 Mt, and grade-tonnage curve as in Figure 5. Massmin 2004
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8 CONCLUSIONS
REFERENCES
Cutoff grade optimization could have a significant impact on the Kiruna mine short- and long-term economic performances and value. Net cash flow increases in the order of several hundreds of million Swedish Crowns could be achieved in high market capacity years. Implementation of a cutoff grade control would need relatively minor adjustments in the way the Kiruna operation is run; particularly, no major investment would be needed. Further investigations are required, however, first to determine a reliable grade-tonnage curve, then to test the various assumptions on the process model. In particular, the model should account for the three run-of-mine ore qualities and for the differential demands in KBF, SPBO, and KPBO. A similar model was built for the Malmberget mine, LKAB’s second operation, where differential demand for fines and pellets were modeled. This second model demonstrated the importance of modeling differential demands on the mine value and optimal operating strategy.
• Hall, B, 2003. How mining companies improve share price by destroying shareholder value, or how the junior geologist and engineer determine the CEO’s bonus. CIM Conference 2003, Montreal, Canada. • Henry, E, 2003. Värdemaximering av Kiruna gruvan med hjälp av en optimal gråbergsinblandningspolicy (Value maximization of the Kiruna Mine using an optimal wasterock grade policy; in Swedish). LKAB internal report. • Lane, K, 1988. The economic definition of ore. Mining Journal Books Ltd, London, United Kingdom. • Taylor, H K, 1972. General background theory of cutoff grade. Trans Min Metall, Section A, Vol 81, July, p 160179.
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Production scheduling at Finsch diamond mine, South Africa Otto Richter, Mining Engineer, Block 4 Project Planning, Finsch Mine - De Beers, South Africa Tony Diering, Principal Consultant, Gemcom Software International, Canada
Abstract De Beers’ Finsch Diamond Mine in South Africa is in its final stages of developing the new Block 4 project which will introduce block caving at Finsch Mine for the first time. There are several new and innovative technologies which will be employed for Block 4. As a result of this, it has been necessary to pay particular attention to production scheduling for the build up of production in the early years. The combination of geotechnical constraints and the need for special conditions around the use of fully autonomous Load Haul Dumpers and Dump Trucks , whilst satisfying the need to maximize Net Present Value (NPV), has resulted in a unique set of conditions for the production schedules. In particular, the ventilation and vehicle access requirements impose unwanted production constraints on active draw points which makes the generation of a smooth caving front much more challenging.
1 INTRODUCTION Finsch Diamond Mine is situated some 165km North West of Kimberley, the diamond city of South Africa. Finsch Mine started as an open cast mine in 1964, changing over to underground operations in 1990 after reaching a pit bottom depth of 430m. Underground operations were planned to extract ore from the kimberlite pipe through a modified open stoping method. Country rock instabilities and resulting sidewall failures poses major challenges to the current mining method in Blocks 2 and 3, resulting in the development of a block cave in Block 4. The Extraction Level will be situated on 630mL and the Undercut Level ±20m above on 610mL, with an average column height of around 80-110m above the Undercut Level. Undercut and Extraction tunnels run parallel to each other as can be seen in figure 2. Loading in the Extraction tunnels will only be performed from one side of the kimberlite pipe. This paper describes how the various facets of undercut development, draw bell development and production buildup to full production have been modeled. This required extensive modifications to the PC-BC Block Cave system in order to accommodate these complex constraints. The end result is an improved confidence in the ability to achieve the required production rates as well as overall risk reduction for this significant project.
Figure 1: Finsch Mine - De Beers mining blocks.
2 GEOLOGICAL MODEL Due to the variance in grade between the different kimberlite facies within the kimberlite pipe, the direction of Undercut and Extraction level advance has a major effect on the Net Present Value (NPV) of the project. Initial planning, based on the geological information available at the time, was done for the Undercut to start in the North-Western corner as indicated by the arrow in Figure 3 on 1999’s lithology. As exploration drilling and sampling progressed, it was realized however that the highest grade (F8) portion of the ore body was not situated where originally anticipated, but was located against the contact on the opposite side of the kimberlite pipe as shown on 2004 Lithology. Massmin 2004
Figure 2: Extraction / Undercut Tunnel relationship.
Through scheduling the updated ore body in PCBC it was proven that it would make financial sense to start the Undercut on the opposite side of the kimberlite pipe
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Figure 4: Block 4 Undercut Ring Blasting schedule.
to build up to a maximum undercutting rate of 1,350m2 per month over a period of about 7 months. A down side to the straight undercut face and the advance direction is the prolonged period required for the undercutting process to progress across the width of the kimberlite pipe and the associated prolonged duration each Extraction tunnel will be exposed to both production loading at the rear end and draw bell installation in the front. 4 EXTRACTION LEVEL SCHEDULING Draw bell installation and hence production buildup will follow in the shadow of the Undercut face at an angle of 450. Two areas of the Extraction level scheduling will be discussed, namely: • Drawpoint prioritization and • Development and support constraints
Figure 3: Changes in Geological Model extracting the higher grade ore first. More detail is given on this in Section 3. The kimberlite pipe also contains internal "floating" portions of Basaltic low grade ore. In order to understand the effect these grade and rock type variances will have on the overall mined grade and ore composition being sent to the treatment plant over time when mining in different sequences and at various rates, multiple runs were performed in PCBC in order to come up with a schedule that would best utilize the given resource model. 3 UNDERCUT LEVEL SCHEDULING Geotechnical constraints and current infrastructure were the two major role players in the design of the undercut face shape. From a purely financial / NPV point of view, the Undercut would ideally have to be mined in a chevron VShape, starting in the highest grade F8 kimberlite in the South-Eastern corner and advancing across the kimberlite pipe in an North-Westerly direction. The associated challenge of moving past two precursors and the current sizer excavations on 620mL in this manner however resulted in the undercut being designed to move across the kimberlite pipe with a straight face, starting on the Eastern side of the pipe at the one precursor and moving across the pipe in a West-South-Westerly direction towards the other precursor as shown in Figure 4. Undercut rings are designed at a spacing of 2m and leadlags of 7m between adjacent undercut tunnels. Undercut ring blasting commenced early May 2004 and is scheduled 454
4.1. DRAWPOINT PRIORITIZATION In order to optimize the Net Present Value of the project within the given Geotechnical and Mining Constraints, various scenarios were investigated using PCBC in order to test how sensitive the overall grade and dilution are to the prioritization of the drawpoints. Four main options were investigated, namely: • Emphasis on new draw points • Emphasis on old drawpoints • Emphasis on all drawpoints according to remaining tons in drawpoint • NPV optimization The results clearly showed that the low cave height placed a major constraint on the degree to which the emphasis on the drawpoints could be varied. Should the emphasis be varied too much, drawpoint availability was seriously compromised due to earlier drawpoint depletions, premature dilution ingress and resulting grade dilution. It was found that despite this constraint, the grade can be increased by up to 13% in the earlier years by placing more priority on the new, undiluted drawpoints without compromising the total ore resource. It was also found that the period at which the cave would be producing at its designed capacity of 3.8Mta can be extended by up to 12 months. This was done by placing the emphasis on the drawpoints according to its remaining tons. NPV optimization was also performed by scheduling the cave using LINDO, a linear programming option within PCBC.
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Care had to be taken not to compromise the cave angle through this option as priority is placed on the drawpoints that will yield the highest NPV value, hence the full benefit of NPV optimization could not be enjoyed. This option proved however that even within a given tons call, there is usually room for financial optimization. 4.2. DEVELOPMENT AND SUPPORT Poor kimberlite rock conditions (Mining Rock Mass Rating of 19-25) resulted in major support work required to ensure the stability of the Extraction Level tunnels. Table 1 shows a typical support requirement per meter Extraction Tunnel.
Table 1: Typical Extraction level tunnel support Unit
Size
Quantity/m
Threaded bolts
25mm x 3.8m
15 (0.5m spacing)
Vibro mesh
9mm
2.0m (Side wall)
Woven mesh
5mm
Hanging to Vibro
Cable straps
18mm
5 (Side wall)
Cable straps
15mm
9 (Hanging wall)
Wet-crete
30Mpa
200mm thick
Fiber-crete
40Mpa
50mm thick
challenges with regards to premature waste ingress, will have to be contended with. The total number of activities taking place at any one time could not be reduced as this would mean deviation to the current design, nor could the number of activities taking place at any one time be reduced as this would result in an extended production build-up, negatively affecting the NPV. In order to increase the total number of ventilation districts, a double return airway system was implemented. This would allow for development and support as well as production loading to be performed in the same tunnel at the same time. Making use of a double return airway effectively splits the extraction tunnel into four separate ventilation phases (Figure 5) during its operation with the following advantages: • Allows for smaller ventilation zones to be in operation at any point in time, thus less activities per ventilation zone, • Allows for separation between autonomous production LHDs and employees involved in development and support, • Allows for separate ventilation zones in a single tunnel, thus multiple activities can be performed in a single tunnel, • Ensures that production loading can be better controlled across the position of the ventilation barrier, thus ensuring an improved cave back angle and a reduced chance of premature waste ingress as can be seen from Figure 6. Intake air:
Return air:
Added to the poor rock conditions is the decomposing characteristic of Kimberlite in the presence of water due to a high clay content, resulting in dry development with a full cover of sealant and shotcrete to be sprayed after each development round. This had dramatic effects on the development and support cycle times, resulting in slow linear advance rates utilizing mechanized continuous operations on multiple development ends using continuous operations. Draw bell installation was also negatively affected, requiring around 3 months per draw bell with a total of 3 draw bells planned per month. In addition to development and support taking place simultaneously in the same tunnel, time had to be allowed for production loading. Production loading is as important in the daily activity schedule as the development and support activities, as this ensures continuous movement of the production columns, which prevents re-compaction of the broken kimberlite. Re-compaction of the broken kimberlite would result in stress build-up which often leads to infrastructure damage and tunnel collapse. Uninterrupted production loading allows for faster payback of invested capital and an improved Net Present Value (NPV). In order to allow for more time to perform the above activities, multiple ventilation districts had to be created in order to facilitate the scheduling of multiple activities in the same tunnel. 5 MULTIPLE ACTIVITY SCHEDULING Making use of a central return airway in a block cave to create multiple ventilation districts in the same tunnel is fairly common practice when making use of a V-Shape undercut face. Combining a single return airway with the long single sided loading extraction tunnels of Block 4 and the relatively slow advance rates however resulted in a very slow production build-up unless cave angles were allowed to exceed angles of up to 600. The Block 4 portion of the ore body has a height to width ratio in the order of 1:2.5. Should such steep cave angles be allowed in the early stages of the cave, it would soon hole into the current operations and waste failures, posing major mining Massmin 2004
Figure 5: Four ventilation stages of an Extraction Tunnel
6 PCBC SOFTWARE MODIFICATIONS Block 4 currently makes use of a general scheduling package to schedule the horizontal development and support activities. From this, the drawpoint commissioning
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dates are given as input into the PCBC cave scheduling and simulation software. In order to cater for the constraints explained above, a number of alterations had to be incorporated into the PCBC software in order to allow the accurate scheduling of the expected Block 4 tons and grade profiles. The most important changes are explained next.
large columns, the variance between the PCBC generated draw cones and the actual trough and ring designs are negligibly small, but in Block 4’s case these tons contribute 12% of the total column tons and therefore required more accurate representation. This was done through a quick import function where the correct trough and undercut heights and its associated tons and volumes as per the blasting designs are imported for each of the drawpoints, replacing the slices as created in PCBC, as can be seen in the illustration below.
Figure 7: Slice file modification for trough and undercut ring slices
Figure 6: Improved cave back angle
6.1. DATA SECURITY Data security was ensured by converting the Desktop edition of PCBC to support full SQL Server databases. Scheduling is performed directly from the geological block model as created by the Geologists in the GEMS SQL Database, ensuring that all systems make use of the same information from Geology and Geotechnical, to Survey, to Mine Planning and Production. Data security will become more important once production commences and schedules have to be modified for actual tonnage production and draw point development status. 6.2. MULTIPLE GRADE ELEMENTS In order to ensure accurate diamond grade forecasts, PCBC had to be modified to handle multiple grade elements when scheduling. In order to cater for the various kimberlite facies encountered, the slice file had to be expanded to cater for 15 different grade elements. When performing a production schedule with the expanded slice file, information is obtained on both the total production values in terms of tons and grades, as well as on the individual kimberlite facies groups, improving the functionality and accuracy of Diamond Value Management, Treatment Plant flow control and accurate grade predictions. 6.3. SLICE HEIGHT Block 4’s relatively low column height required the bottom two slices of the slice file to be modified to more accurately reflect the trough and undercut ring tons. When scheduling 456
6.4. DRAWPOINT SEQUENCING PCBC caters for scheduling in different time intervals, e.g. monthly, quarterly, annually, etc. In order to make use of the drawpoint dates as scheduled by the development and support scheduling package, PCBC had to be modified to generate a drawpoint commissioning sequence making use of these dates, irrespective of what time interval is used to schedule in PCBC. This was accomplished successfully by stating both the drawpoint commissioning dates and the PCBC periods in number format. PCBC then compares the drawpoint commissioning dates with the PCBC scheduling periods and automatically places the drawpoints in the correct sequence and time period. In earlier versions the time period was based on a text string and required manual modifications to suit the PCBC scheduling periods. 6.5. USER INTERFACE It can be seen from the information above that multiple records of data are associated with each drawpoint. In order to ensure easy access and modifications to this information, PCBC was modified to easily export/import data to/from MS Excel as it is often easier to modify data in MS Excel than in the SQL Database environment. Any information associated with a list of drawpoints (Drawpoint Status, limiting draw heights, descriptions, tunnel associations, production curves, hang-up curves, etc) can now be easily formatted in MS Excel and import into the PCBC Database to update drawpoint records. 6.6. PRODUCTION RATE CONSTRAINTS Production rates are constrained by various methods within PCBC. The first level of constraint is on the individual drawpoint production rates. In order to ensure that the cave matures adequately before higher production rates are called for, each drawpoint’s daily production rate is increased based on the cumulative tons being drawn from
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that particular drawpoint to date. The production rate curve below is defined within PCBC as a normal XY-Curve, stating tons per square meter of drawpoint area per day versus cumulative percentage drawn from that drawpoint. Individual drawpoints can be assigned individual curves or similar drawpoints can be grouped together to make use of a single curve. The initial stage of the drawpoint curve has a very high rate during draw bell construction. There after it follows a slow buildup, starting at about 50t/85mm per drawpoint per day and maturing over a range of about 18,000t to a maximum rate of about 120t/200mm per drawpoint per day for a drawpoint area of influence of 225m2. A drawpoint will reach maturity within about 9 months based on the stated daily tons. In addition to the constraints on individual drawpoints, a constraint can be set per production block. This is typically applied where the sum of the total available tons in a single tunnel is greater than what a single loader can realistically produce from that tunnel in the same time period.
Table 2: Typical hang-up frequency table production time is calculated and hence a more realistic prediction of tons and grades is achieved. 7 CONCLUSIONS
Figure 8: Typical drawpoint production rate curve
The maximum capacity of a single loader is calculated for each production block / tunnel based on the tons weighted average tramming distance for that tunnel in that time period. This function was further enhanced within PCBC by importing the average cycle time per drawpoint and based on the total available loading time per period, PCBC will constrain the tons called for if it exceeds the maximum possible tons which can be loaded from that tunnel based on the individual drawpoint cycle times. 6.7. HANG-UP CONSTRAINT In order to further enhance the accurate prediction of production tons and grades, the available time for production loading had to be enhanced. This was done by incorporating the expected number of hang-up’s or blockages expected due to oversize rock at any point in time in a drawpoint’s life, as can be defined by a XY- Curve per drawpoint of number of occurrences per 1000 tons drawn versus accumulative tons/height drawn to date. By doing this, the total number of blockages expected per drawpoint can be calculate for every period based on the tons loaded in that same period. Converting these blockages to an associated "down time" where the drawpoint will not be available for loading, the total time available for production can be more accurately calculated. Through an iterative process the most realistic available Massmin 2004
The Block 4 project is currently standing on the doorstep of moving from planning to actual implementation with the undercut that commenced early May 2004. A lot of the assumptions used in Block 4 are based on current operations, both inside and outside of De Beers, yet these assumptions had to be adapted and tried and tested where possible to suit the current project. The Planning Phase of any project is the stage where major cost savings or losses are built into the project and should be adequately resourced with regards to time, money and people and initial assumptions should be re-evaluated and re-calibrated in order to ensure that applicable, accurate values are used at all times. Proper planning and scheduling ensures easier implementation which in turn reduces risk to the success of the project to a large degree. Installing a block cave requires intense scheduling of multiple activities. It is therefore important to create an auditable schedule that can be tracked in order to highlight areas lagging behind and to identify areas of opportunity. Initial planning should therefore be realistic and rather on the conservative side than on the optimistic side, especially for a first of its kind project like the Block 4 operation. Software applications are however not always designed to best apply the rules and constraints applicable to each project and therefore software customization is often inevitable. Working closely with the software manufacturer results in a final product that is better suited for the task at hand and that will benefit the mining industry at large. The Block 4 schedule is the product of many iterations. Only through comparison of multiple scenarios could the schedule be optimized to what it is now. Having only a large number of scenarios however do not always proof to be valuable information, but merely a lot of data. It is important that the mine planning personnel, the software manufacturer and the mine production personnel all understand and agree to the problem at hand in order to turn heaps of data into usable information. The best solution to a problem is usually only found once the correct information is available. Production pressure and NPV often play major roles when making decisions. Even though these are good indicating factors, they should not be used as over riding factors to break the basic rules of mining. The cave should dictate the production build-up, not vice versa (Don’t let the tail wag the dog). The value of accurate, upfront information should never be underestimated. Exploration work well ahead of time is crucial in order to ensure proper extension of current production blocks well in time, ensuring fewer complications
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during change over between production blocks. The current plan is only as good as the information it is based on. ACKNOWLEDGEMENTS The authors want to acknowledge the permission given by De Beers – South Africa to publish this technical paper as well as the time and dedication of Gemcom staff in the modification of the PCBC cave simulation and scheduling software.
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REFERENCES • Diering, T, 2000. PC-BC: A block cave design and draw control system. Proceedings MassMin 2000, Brisbane, pp. 469-484. • Diering, T, 2004. Computational considerations for production scheduling of block cave mines. Proceedings MassMin 2004, Santiago.
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Application of simulation to improved planning at Esmeralda, El Teniente Mine, Chile Mauricio Barraza, El Teniente Division, Codelco, Chile Matt Rohrer, Automod, Brooks Automation, SLC UT, USA William Hustrulid, Department of Mining Engineering, University of Utah, SLC, UT, USA
Abstract The mining systems used to mine the ore at El Teniente have evolved over the years. Today the primary mining system, conventional panel caving, provides about 75 percent of the daily production. In 1997, panel caving with pre-undercutting was introduced in the Esmeralda sector. Since that time, sector production has increased from 250 tpd to about 40,000 tpd. This rate is scheduled to increase to 45,000 tpd by year 2005. One of the major advantages of pre-undercut panel caving over conventional panel caving is the much higher physical availability of the production area. This is due to the fact that the development and construction work on this level has been done under largely de-stressed conditions. In conventional panel caving, the development and construction on the production level is done under the existing high in situ stress conditions. These stresses are then increased even higher by the abutment stress associated with the passage of the undercutting front. This means that the production level can be severely damaged and in need of repair even before actual production begins. Repair and rebuild of the openings is an on-going task in conventional panel caving. An important question to be answered is how best to utilize the high physical availability of the production level offered by pre-undercutting. This paper demonstrates through the use of Automod simulation how the utilization of the available production area can be maximized. Because of the geomechanics constraints imposed upon the mine design and execution, the planning and sequencing of the development and construction operations on the production and undercutting levels is much more complicated than with conventional panel caving. Simulation offers the opportunity to easily explore many different planning possibilities in a very short time.
1 INTRODUCTION Simulation is defined as the imitation of a real-world process or system over time. It is an indispensable problem-solving methodology for the solution of many real world problems and can be used both to describe and analyze the behavior of a given system and also to aid in the design of a new system. Both existing and conceptual systems can be modeled with simulation (Banks 1998).
Table 1: Several Types of Simulations (Brunner, 2001). Static Simulation
Static simulations, are used when the effect (appearance, load-bearing capacity, etc) is observed only at a single instant of time.
Discrete Event Simulation
Discrete event simulations (sometimes called dynamic simulations) are typically used to model systems whose states change at discrete points in time. The beginning of drill cycle, a skip reaching the surface, and the contents of an ore pass reaching a critically low level are all examples of discrete events. Material flow systems, where liquid or bulk solids move at constant rates, are candidates for discrete event simulation.
Continuous Simulation
Continuous simulation is used to model nonlinear behavior (chemical reactions, heat transfer) using equations.
Stochastic Simulation
The most common stochastic simulation is the Monte Carlo simulation. For example, if you know the probabilities that the various components of an LHD will independently fail today, and if you want to know the probability that the entire LHD will fail today, you could run a static experiment a thousand times (or million times) to see how likely it is that the LHD will fail. Although this particular problem might also be solved mathematically, a stochastic Monte Carlo simulation might be easier to set up for some types of problem.
2 GENERAL SIMULATION CONCEPTS The goal of any simulation is to mimic something. Typically, there is a stimulus/response interaction (Banks 1998). Some examples of this are: • In a flight simulator, the goal is to mimic the visual and tactile interaction between a person and a physical system (the aircraft). • In a circuit simulator, the stimulus is the input signal and the goals are to mimic the response of a physical circuit component. • In an automobile simulator, the goal is to simulate the response of the suspension to road bumps, the air flow around a vehicle and many other physical factors. All of these examples are primarily physical models. A brief description of different types of simulations is represented in Table 1. 2.1 Simulation Basics Simulations can be thought of as a framework for describing the operation of a system. Describing a system Massmin 2004
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using this framework can sometimes be very beneficial even if the model is never used for experiments (Banks 1998). • Resources and entities In a simulation-oriented system description, entities flow through a series of resources. Entities are units of work or units of traffic. They queue up for and use resources according to logical rules. Resources are generally constrained. In a simple model, an entity might represent some quantity of material (ore or waste), and the resources might be people, an LHD (equipment), or a storage area. In a more complex model, there might be entities that represent the logical controllers of the system (human managers as well as automated controls). • Interactions: Simulation can capture interactions in a way no static technique can. There are many, many dynamic cascading events in an underground mine. A given task may be subject, for example, to waiting for a blast, waiting for repairs, waiting for adjacent work to be completed, and so forth. • Random variables: Most simulations include representations of random variations. Using point estimates (averages) as inputs to any kind of model can cause interactions to be overlooked. This is typically most obvious in the case of failure modeling but many other aspects of mine behavior can be sampled from a probability distribution. 2.2 Simulation Inputs Model inputs fall into these broad categories: logical data, system description data, process data, and demand data (Banks 1998, Brunner 2001, Hollberg et al. 2002, Pegden et al. 1995). • Logical data: Logical data includes all rules for operating the system. What activities must halt when a blast is scheduled? Who decides which scoop should be assigned to a pending mucking task and how is that decision made? What shift schedule do the jumbo drill operators follow? Is there a stope sequence that is to be followed in the model or will the model make sequencing decisions dynamically; and if so, how? • System description data: System description data represent the physical system. This includes mine geometry, material properties, and the equipment list. It is often useful to break unique geometries into individual material blocks that may have unique properties. • Process data: Process data are the rates and speeds that constrain system performance. Operator and equipment performance, equipment failure, hoisting rates, conveyor speeds, and similar data fall into this category. • Demand data: Demand data drive the model. In other contexts, the demand data are how many cars are we trying to make or how many boxes need to be shipped today. Demand data are not typically a major factor in a mine model because the model is generally set to go all out given the other constraints in place. However, there may be cases when the model is set to start certain material blocks at certain times or to stop before the end of any shift once a certain tonnage is hoisted. The goal in such case is to see how the equipment and operators are utilized. Data for a mining model are often difficult to gather and reduce to a usable form. The logical data may be incompletely understood by a single individual and a team of people may have trouble agreeing on what practices are or will be followed underground. 2.3 Simulation Output The output from a simulation can assume various forms. Although simulation outputs are usually statistical, they need not be. There are two reasons for this. First, an 460
important benefit of a well-done simulation project may simply be the insights gained by describing the system in the simulation framework. This includes gathering, analyzing and processing the raw data as well as developing the logical rules for system operation. Second, animation is sometimes used demonstrate to analysts, operators, management, and others that the model is valid (Banks 1998, Brunner 2001, Hollberg et al. 2002, Pegden et al. 1995). • Statistics: Statistical output generated by a model can be derived from a standard output report or customized to suit the model user. Typical outputs include resource statistics, queuing statistics, and other summary information. The model can also produce any other measurable statistic that is consistent with the model’s level of detail. This could include, for example, the duration of various activities, time plots, a trace file for understanding the model logic, and a resource table (schedule). Under the heading of resource statistics, a simple example would be the utilization of a drill. In the calculation of drill utilization one could divide the operational time by; • total time • total time minus maintenance shift and shutdowns • time when crew members are available to run the drill • operating time divided by the sum of operating and repair time • something else It is important that there be agreement as to the correct calculation. Time plots, whether generated by the simulation software or plotted from data produced by the model, can provide many insights. • Animation: A simulation model can be animated in many ways. Animation can, for example, be in 2-D or 3-D space. It can (1) show people and equipment moving or only show material state changes, (2) be pictorial or schematic, (3) be to scale or not to scale and (4) be delivered with or separately from the model. All of these forms of animation are valuable for verifying and validating the model. They are also extremely useful for achieving buy-in from other operating personnel as well as from top management. A model is not going to be considered useful unless everyone understands that it is valid and it is not going to be considered at all if people do not understand what it is. 2.4 Steps in a Simulation Study The flowchart shown in Figure 1 provides a set of steps to guide a model builder in a thorough and sound simulation study. 2.5 Simulation in Inderground Mining There are many ways simulation can be used in the planning and operation of mines (Table 2). The variables include: • Model objectives • Model focus • Model time frame The objectives for modeling a mining operation can vary widely. Typical objectives may include some or all of the following (Banks 1998, Brunner 2001): • Equipment type comparisons • Specific capital purchase decisions • Mine plan analysis (discussion of integrated planning and modeling) • Mining method comparisons • Operating policy evaluation and improvement
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Step 1: Problem formulation Step 2: Setting objetives and overall project plan
Step 3: Model building
Step 4: Data collection
2.6 Simulation Software-Automod (Banks 1998, Pegdent ET AL 1995) This software from AutoSimulations, Inc., has general model-building features, including the specifications of processes, resources, loads, queues, and variables. AutoMod software is a very powerful tool for describing material handling systems. Automated guide vehicle (AGV) and path-guide transporters, conveyors, bridge cranes, AS/RS’s, and power and free devices can be defined rapidly. The animation capabilities include true-to-scale three-dimensional graphics, rotation, and tilting, to mention a few. A CAD-like drawing utility is used to construct the model. A separate utility option is AutoStat. It provides simulation warm-up capability, scenarios management, confidence interval generation, and design-of-experiment capability. AutoView is a post-processor that provides threedimensional walkthrough capability for presentation-quality animation.
Step 5: Coding
No Step 6: Verified? Yes No Step 7: Validated? Yes Step 8: Experimental desing Step 9: Production runs and analysis
Step 10: More runs?
Yes
• Medium-term models (might run for up to a few simulated years) are used to evaluate equipment plans and scheduling and operating policies. • Long-term models (might run for up to a few simulated decades or more) are used to evaluate long-range mine plans, look for development bottlenecks, and so forth.
2.7 Software Considerations
No Step 11: Document program and report results
Input Considerations CAD Translation
Step 12: Implementation
Figure 1: Flowchart, Used to Guide a Model Builder in a Simulation Study (by Bank, Carlson, and Nelson 1996)
Importing a File
Table 2: Uses and Benefits of Simulation in Underground Mining (Banks 1998, Brunner 2001)
Exporting a File
Uses
Benefits
Analysis of proposed capital expenditures. Analysis of operating procedures. Analysis of plans and schedules. Understanding and communication of system behavior. Day-to-day decision making.
Supporting processes such as services, materials, and repairs. Material handing including muck movement by vehicle; bin flow with ore passes, crushers, hoppers, and other intermediate storage; and hoisting and removal. Trucking operations. Operator training.
Syntax Debugger
Along with the many different objectives possible in underground mine modeling, the areas of focus can also vary widely. Some general areas of focus include (Banks 1998, Brunner 2001): • The development process • The production process The time frame of the model depends, in part, on its intended use (Banks 1998, Brunner 2001). • Short-term models (which might run for up to a few simulated months) are used to evaluate operating policies, to directly schedule the operations, or to assess the impact of exceptional conditions.
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Input Data Analysis Capability
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If there exists a CAD drawing of the static background, a CAD translator will take a CAD drawing and convert it into the drawing system used by the simulation. These provide the capability to import a data file for use in the simulation. The output file will be used as input to a spreadsheet for drawing business graphics beyond or different from those generated by the simulation software. This should be easily understood, consistent, and unambiguous. Even the best of simulation analysis makes mistakes or commits logical errors when building a model. The debuggers assist in finding and correcting those errors in the following ways: - The simulation can be monitored as it progresses. - Attention can be focused on a particular area of the simulation or a particular entity. - Values of selected model components can be observed. - The simulation can be temporarily suspended, or paused, not only to view information, but also to reassign values or redirect entities. The ability to determine whether input data can be described by a statistical or mathematical distribution.
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Processing Considerations Speed
Run-time flexibility
Random Variety Generator
Reset
Attributes and global variables
Programming (Custom logic representation)
Portability
When there are many entities in a system, the software speed should not degrade to the point of slow motion. An example of this feature is the scenario generation. In this case, with some prodding, the simulation software will automatically generate alternative possibilities for simulation (i.e.; of scenario generation allows input data to vary over a range) There are about 12 statistical distributions that are commonly used in simulation. Most, but not all simulation software has the ability to generate random varieties using these 12 distributions. For steady-state analysis, it is important to have the ability to reset the statistics that have been collected to zero. This reset is accomplished without clearing the entities that are currently in the system. Attributes are local values available to the entity processing that attribute, and global variables are available to all entities. The question is how many of each of these is available. The ability to mimic custom logic accurately to any desired degree of detail usually requires some type of internal programming capability or underlying language. It is an absolute necessity for modeling complex problems or systems in order to build highfidelity model. This means the software can be run on various classes of computer without changes in the software.
Collection of Desired Mathematical Expressions Custom performance Expressions
Write to a File
could become rather a large stack of paper. At the other extreme is a database that contains all these output in an organized fashion. This allows the specifications of measures of interest to the modeler. Does the software allow the analyst to define and create new or custom measures of performance for a model? Does the software allow data, events, or system variables to be written to a file whenever desired? This feature allows the analyst to later import the file into spreadsheet or database programs for further customized analysis or manipulation.
Environment Considerations Ease of use
Ease of Learning
Quality of Documentation
Animation Capability (3D)
Stability
Support
This is important to some, not important to others. The power of the software is probably much more important than ease of use. This is important to the casual user, not so important to the frequent or continuous user. Often, documentation is so impossible to understand that users refuse to read it. Contextsensitive and useful on-line help is an advantage. Not all animation is created equally. Consider the ease of development, the quality of the picture, the smoothness of movement, and the portability for remote viewing. Is simulation software their primary business or just a sideline? Does the vendor offer adequate technical support and access to the system?
Output Data Components of the Path-Move System Standardized Report
Customized Report
Business Graphics
Data Base Maintenance
462
Examples of standardized output measures are the average number in queue, average time in queue, etc. The software can produce these and other values automatically or upon request. These are tailored presentations such as those that would be shown to a manager. The format can be set by the simulation analyst. The software can have the ability to generate bar chart, pie charts, and histograms that are of such high quality that they can be shown to managers and included in reports. One possibility is to collect a stack of paper output representing the replications from each scenario. However, this
Guide paths
Transfers
Control points
Vehicles
Santiago Chile, 22-25 August 2004
Guide paths represent routes that are taken by vehicles or people in a system. They may consist of a number of segments. Guide paths can be one or two directional Transfers are a connection that joins two segments of guide path. For a vehicle to move from one segment to another, the two segments must be connected by a transfer. Control points are locations at which vehicles can pickup or set down loads in the system. These can be located anywhere on a path. Vehicles transport loads from one location to another by following a path in the path-move system. Massmin 2004
Vehicles can be defined and grouped by type and can differ in velocity, capacity, and the time required to pickup and set down loads in the system. Vehicles can also be assigned different attributes based on the type of loads they are carrying; for example, an LHD has a different rate of acceleration or velocity if it traveling empty or with a load.
3 APPLICATION OF SIMULATION TO PLANNING AT ESMERALDA In 1997, panel caving with pre-undercutting was introduced in the Esmeralda sector. Since that time, sector production has increased from 250 tpd to about 40,000 tpd. Although there are similarities between pre-undercut and conventional panel caving, there are also a number of significant differences. It has taken the mine some time to learn both about the pre-undercut method and about its application to extract primary ore under the relatively high stress conditions which are present. There has been little opportunity to try and optimize the planning and sequencing operations but this is now in order as the daily production rate is scheduled to rise to 45,000 by year 2005. This will be the first improvement opportunity evaluated by simulation. The simulation example will show how the utilization of the available production area can be maximized. 3.1 Simulation Model DevelopmentI Mine Layout For simulation, the production, haulage, and undercut levels will all be included since they interact directly. Of these, the most important is the production level because its advance rate controls: • the increase in the area of extraction, • the movement of the front on the undercut level, and • the need to incorporate new drifts on the haulage level.
Figure 3: Plan View of the Production Level. This is represented by the Path – Mover.
Table 3: Extraction Rates Imposed by the Geomechanical Group Primary Ore Extraction Ranges (%)
Initial Caving (m/d)
Steady Caving(m/d)
0–5 5 – 10 10 – 15 15 – 20 20 – 25 25 – 30
0.05 0.07 0.08 0.10 0.13 0.16
0.10 0.13 0.15 0.17 0.20 0.24
The different model components for the example are shown in Figures 2 and 3. Ore Flow The simulation model is used to study flow rates and capacities. The flow of ore is represented by the height of the ore column (Figure 4). The flows are subject to the restrictions imposed by the Geomechanical Group. This means that the extraction rates are defined so as to try to control the breakage of the solid ore column (Table 3).
Figure 4: A 3D Representation of the Mining Model Showing The Production Level, Haulage Level, and Ore Column Height.
Problem Formulation Production area availability and utilization values under conventional and pre-undercut conditions are presented in Table 4. Table 4: A comparison of Production Level Areal Availability and Utilization by Mining Panel Caving System Figure 2: Section View Showing the Draw Bell between the Undercut Level and Production Level. This is represented by a QUEUE Massmin 2004
Conventional Pre-Undercut
Santiago Chile, 22-25 August 2004
Availability (%) Production Level
Utilization (%)
50 > 90
45 to 50 65 463
As can be seen, the physical availability of the preundercut panel caving production area is very high. This is due to the better rock mass conditions resulting from the development of the production openings in a de-stressed area. Although the utilization of the area is higher than in conventional panel caving (65 percent versus 45 to 50 percent), it can be improved. Today the main challenge associated with the pre-undercut variant is to improve the utilization of the available production area. • Areal utilization is defined as "the total area with extraction divided by the total open area, expressed as a percentage" • Physical availability is defined as "the total area for extraction over the total open area, expressed as a percentage" The objective is to maximize the production from the available production area. In this evaluation, one determines the available physical area by multiplying the total production area by the physical availability. This available area should them be utilized to as high a degree as possible to minimize the specific development and to maximize the production efficiency. 3.2 The case Study Model The Model considers the utilization of the production level area. This involves LHD movement on the production level, the filling and emptying of ore passes, and train movement on the haulage level from ore pass to ore bin.
The current production area is 300m by 150 m in extent. consists of: 317 draw points 15 streets 2 train loops 15 ore passes
Three simulations will be made with the model. Simulation 1: The current mine (LHD and train) schedule with 65 percent area utilization. Simulation 2: Increase the utilization of the current available area to 80 percent. This is done by increasing the number of LHD’s. It is accompanied by increase in the production rate. Simulation 3: Increased production efficiency while maintaining the current areal production rate and the number of LHD units. Model Building The overall model consists of five "processes" or modules. These are: PRODUCTION AREA, LOAD, LHD, ORE PASS, and TRAIN.
• • • •
Data Collection The data needed are divided into: Schedule of the production plan LHD distribution schedule The number of trains An ore pass file
Simulation 1a – Schedule by Shift A particular shift schedule is simulated and the area utilization is determined. The production rate and the utilization of the LHD’s and the trains was compared with the actual information from the mine. The difference between the actual LHD productivity and the average productivity from 7 simulation runs is 13 tons/shift or 1.8 percent. For the trains the difference is 85 tons/shift or 3.6 percent and for the area utilization the difference is 0.5 percent. Essentially, there is no difference between the simulated and actual conditions. An output report presents all of the information from the analysis. The information for one shift is represented in Table 5. Table 5: Model – Case 1 Validation. Productivity
Actual Simulation Shift 1 Shift 2 Shift 3 Shift 4 Shift 5 Shift 6 Shift 7 Average 718 2450 65
687 2250 61.4
883 3000 67.2
757 2350 66.1
777 2335 66.8
692 2166 62.2
700 2190 62.8
739 2270 64.9
705 2365 64.5
Simulation 1b– Schedule by Month A second simulation was performed based upon the monthly schedule. The actual (or planned) mine operation by month including area utilization was compared to that arising from the simulation. Two simulation runs were made. The first was to compare the actual mine production for the months April and June 2002 with the output of the model. The second run was made comparing the output fo the model to that planned for the months of August 2002 and December 2002 as included in the program for year 2002. The actual/planned and simulated production rates and the LHD and train utilization values were compared. The results are shown in Tables 6 and 7. Table 6: Model Case 1 – Mine Production Rate 2002 Year 2002
Actual Production Rate April June
Production (t/d)
22,370
23,350
Planned Production Rate August December 26,909
25,717
Table 7: Model Case 1 - Simulated Production Rate for 65 Percent Area Utilization Year 2002
Model Verification In this step, model verification means the determination as to whether the computer implementation of the model is correct. In this case, verification is done using animation. In the simulation analysis one can detect an action that is illogical. The verification is made by observing the condition 464
3.3 Model Validation-Simulation 1 In this step two simulations are run using historical extraction data from the mine. The first model is run by shift and the second is by month. The required LHD and train data are the cycle time values (the time to load, speedloaded, speed-empty, time to dump, etc)
LHD( tons/Shift) Train(tons/Shift) Area Utilization (%)
The respective capacities are: • LHD = 7 t (7.3 yd3)/cycle • Ore pass = 400 t • Train capacity = 400 t
It • • • •
of ore column extraction. This means examining how the ore column is broken during the extraction time.
LHD (t/d) Number of LHD’s Train haulage(t/d) Open Area (m2) Area of Utilization (m2) Area of Utilization (%)
Santiago Chile, 22-25 August 2004
Simulation results for comparison to the data from the mine April June 22,598 11 22,191 94,175 61,214 65
23,599 11 23,134 96,520 62,738 65
Simulation results for comparison with the planned program for 2002 August December 26,911 13 26,374 105,507 68,580 65
25,745 12 25,672 111,370 72,390 65 Massmin 2004
The difference between the actual and the simulated production rates for April is 223 t/d, for June it is 230 t/d. In both cases the simulation yields a smaller production than the actual. This is because the actual production is compared with the simulated output production recorded on the haulage level and the train takes the load from the ore pass only if it is full. When comparing the planned production versus that simulated for August, the difference is 522 t/d less than the plan. For December the simulated production rate is 25 t/d more than the plan. 3.4 Model Result – Simulation 2 In this simulation, the area utilization will be increased from 65 percent to 80 percent. This is done with by increasing the number of LHD’s assigned to the same area as was used in simulation 1 (CVB – Open Area). By adding LHD’s there is a natural increase in the production rate. The output data from the model are summarized in Table 8. The output report shows the possibility of increasing the monthly production rate by increasing the number of LHD’s in production. The next step in the evaluation is to perform an economic analysis. Table 8: Simulated Production Rate Based Upon 80 Percent Area Utilization. CVB – Open Area. Year 2002
April
June
August
December
LHD (t/d) Number of LHD’s Train haulage (t/d) Open Area (m2) Area of Utilization (m2) Area of Utilization (%)
26,317 12 25,704 94,175 75,340 80
29,524 14 28,815 96,520 77,216 80
33,830 16 33,507 105,507 84,406 80
32,117 15 31,536 111,370 89,096 80
3.4 Model Result – Simulation 3 In this simulation, the areal utilization will be increased to 80 percent while maintaining the same production rate and number of LHD’s as indicated in simulation 1. In achieving this, the total area in production can be less thereby improving the production efficiency. The output data are summarized in Table 9. Table 9: The Results for an Areal Utilization at 80 Percent and the CVB - Production Rate Year 2002
April
June
August
December
LHD (t/d) Number of LHD’s Train haulage (t/d) Open Area (m2) Area of Utilization (m2) Area of Utilization (%)
22,526 11 22,210 76,518 61,214 80
23,267 11 22,909 78,423 62,738 80
26, 908 13 26,815 85,725 68,580 80
25,694 12 25,615 90,488 72,390 80
The output report indicates the possibility of increasing the production efficiency while maintaining the production rate and number of LHD units constant. The results demonstrate the importance of good management when generating the production schedule because with less area in development and construction it is possible to achieve the same production rate. This generates budget savings for the company every year. 4 SUMMARY AND CONCLUSIONS • To meet the combination of geometrical, extraction and operational constraints in pre-undercut panel caving, Massmin 2004
much greater attention to planning and sequencing is required of pre-undercut panel caving than of conventional panel caving. Simulation assistance is required to achieve an optimum result considering all of the different factors. • With respect to the Case Study, one of the major advantages of pre-undercut panel caving over conventional panel caving is the much higher physical availability of the production area ( >90 percent as compared to about 50 percent). This is due to the fact that the development and construction work on this level has been done under largely de-stressed conditions. In conventional panel caving, the development and construction on the production level is done under the existing high in situ stress conditions. These stresses are then increased even higher by the abutment stress associated with the passage of the undercutting front. This means that the production level can be severely damaged and in need of repair even before actual production begins. Repair and rebuild of the openings is an on-going task in conventional panel caving. • With the high physical availability of draw points in the pre-undercut variant, one can choose several different production strategies. • Three alternatives were considered: Alternative 1: was the production strategy in use today. It has a production level area utilization factor of 65 percent which means that 65 percent of the available draw points are in use at any one time. The simulation yielded results very close to those achieved in the mine. Alternative 2: the same physical production area as in Alternative 1 was assumed but the utilization factor was increased to 80 percent. This was accomplished by adding more LHD’s to the fleet. The result was an increase in the areal production capacity by 20 percent. Alternative 3: was to maintain the same number of LHD’s as in Alternative 1 but to increase the utilization to 80 percent. The result was that the same production as obtained in Alternative 1 could be achieved from a significantly smaller area. The overall conclusion from the Case Study simulation was that it is possible to increase the areal utilization of the production level from 65 percent to 80 percent. In doing so one can either get the same production from a smaller area than today or a higher production from the same area as today. Although an economic calculation has not been made at this point, it is certain that the increased draw point utilization will result in very substantial cost savings.
REFERENCES • Baiden, G.R. (2001) "TeleminingTM System Applied to Hard Rock Metal at Inco Limited" Underground Mining Methods, ed. SME 2001, pp 671 – 712, USA • Banks, J., ed. (1998) "Handbook of Simulation: Principles, Methodology, Advances, Applications, and Practice", John Wiley, New York. • Banks, J., ed. (2000) "Getting Started with AutoMod", AutoSimulation, Inc, Bountiful, UTAh 84010. • Banks, J., ed. (2001) "AutoMod User’s Manual v 10.0", Brooks Automations, Inc. AutoSimulations Division, Bountiful, UTAH 84010. • Barraza, M., and Crorkan, P. (2000) "Esmeralda Mine Exploitation Project" Proceeding of the MassMin 2000, pp 267 – 277, Brisbane, Australia. • Barraza, M., San Martin, J., Crorkan, P., and Bustamante, S. (2000) "Grupo de tarea planificación mina Esmeralda", PL-I-096/2000, Internal report, Codelco Chile División El Teniente.
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• Betancourt A., Silva, M., and Valdivia, C. (1999) "Resultados de instrumentación mina Esmeralda’, PL-I021/99, Internal report, Area Ingeniería de Rocas, Codelco Chile División El Teniente. • Brunner, D.T. (2001) "Simulation of Underground Mining Operations" Underground Mining Methods, ed. SME 2001, pp 705 – 679, USA. • Bullock, R. and Hustrulid, W. (2001) "Planning the Underground Mine on the Basis of Mining Methods" Underground Mining Methods, ed. SME 2001, pp 29 – 48, USA. • Cavieres, P. (1999) "Evaluación de los métodos de explotación en mina El Teniente" Internal report, Area Ingeniería de Rocas, Codelco Chile División El Teniente. • Codelco Chile División El Teniente, (2000) "CASO BASE VIGENTE 2001 – Unidad de Gestión Autónoma Mina – Concentradora", Internal report, Subgerencia Minco. • Diering, T. (2000) "A Block Cave Design and draw Control System" Proceeding of the MassMin 2000, pp 469 – 484, Australasian Institute of Mining and Metallurgy, Brisbane, Australia. • Hollberg, K.F., Graehl, D., and Despain, L. (2002) "Application of Simulation Technology to Underground Production", SME Annual Meting, Phoenix, Arizona, USA. • Jofre, J., Yanez, P. and Ferguson, G. (2000) "Evolution in Panel Caving Underground and Drawbell Excavation, El Teniente Mine" Proceeding of the MassMin 2000, pp 249 – 260, Australasian Institute of Mining and Metallurgy, Brisbane, Australia. • Jofre, J., and Blondel, J. (1993) "Geomecánica conceptual Proyecto Esmeralda", Internal report, Departamento de Estudios y Métodos, Codelco Chile División El Teniente. • Karzulovic, A. (1998) "Evaluación Geotécnica Métodos de Socavación Previa y Avanzada", Estudio DT-PE-99-003, A. Karzulovic & Asoc., Chile. • Karzulovic, A. (1997) "Caracterización Geomecánica Rocas Proyecto Esmeralda", Estudio DT-PE-97-001, A. Karzulovic & Asoc., Chile. • NCL, (2000) "Simulación Capacidad de Producción Proyecto Esmeralda", Internal .report, Codelco Chile División El Teniente.
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• Pegden, C. D., Shannon, R. E., and Sadowski, R. P. (1995). Introduction to Simulation Using SIMAN, 2nd edn, pp 8-24 (McGraw Hill: New York). • Rojas, E. Molina, R., and Cavieres, P. (2001) "Preundercut Caving in El Teniente Mine, Chile" Underground Mining Method, ed. SME 2001, pp 417 – 423, USA. • Rojas, E., Molina, R., Bonani, R., and Constanzo, H. (2000) "The Pre-Undercut Caving Method at El Teniente Mine", Proceeding of the MassMin 2000, pp 261 – 266, Australasian Institute of Mining and Metallurgy, Brisbane, Australia. • Rojas, E., Cavieres, P., Dunlop, R., and Gaete, S. (2000) "Control of Induced Seismicity at El Teniente Mine", Proceeding of the MassMin 2000, pp 775 – 781, Australasian Institute of Mining and Metallurgy, Brisbane, Australia. • Rojas, E., Barraza, M., Bonani, A., Morales, A., Munoz, R., Pasten,O., and Morales, M.A. (2000) "Grupo de Tarea Esmeralda Mediano y Largo Plazo, Situación Sector Fw", PL-I-005/2000, Internal report, Codelco Chile División El Teniente. • Rech, W. D. (2001) "Henderson Mine" Underground Mining Methods, ed. SME 2001, pp 397 – 403, USA. • Rech, W., Keskimaki, K. W., and Stewart, D. S. (2000) "An Update on Cave Development and Draw Control at the Henderson Mine", Proceeding of the MassMin 2000, pp 495 – 505, Australasian Institute of Mining and Metallurgy, Brisbane, Australia. • Schriber, T.J. (1991) "An Introduction to Simulation Using GPSS/H", John Wiley & Sons New York. • Sturgul, J.R. (2000) "Mine Design – Examples using simulation" SME. • Sturgul, J.R. (2000) "The Use of Simulation and Animation for modeling Underground Mines", 3rd National Conference of Underground Constructions. • Sturgul, J.R.(2000) " Using Animation of Mining Operations as Presentation Models" Mine Planning and Equipment Selection, Panagiotou & Michalakopoulos (eds) 2000 Balkema, Rotterdam . • Sturgul, J.R.(1998) "Advances in Simulation and Animation for the 21st Century", Third Regional APCOM Symposium.
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Extraction rate: As an index of effectiveness Francisco Carrasco J., Víctor Encina M., Soledad Maass V. Instituto de Innovación en Minería y Metalurgia S.A., IM2
Abstract Extraction rate in Block/Panel Caving mining method is analyzed as the so called Critical Technological Dimension that represents a whole index of the effectiveness of the process, in this case the mining method. Extraction rate is critical for mining business because it limits the production capacity in a given ore body. It can be distinguished three rates of extraction in Block/Panel Caving as follows: Propagation rate of extraction during caving propagation period, Full rate of extraction, when all material is broken and Mean extraction rate representing the global performance of the process considering the total active area, both under propagation and full extraction of broken material. This paper presents the relations of dependence among those three extraction rates with a discussion of the impact of rock mass condition affecting propagation rate and the effect of fragmentation in full rate of extraction of broken material. Also, it is presented an application to Conventional Block/Panel Caving Method using LHD, whose results set a structural limit for its mean rate of extraction [t/m2-day]. This analysis includes the effect of interruptions of flow in draw points (hungup frequency) and restrictions imposed by the batch system of extraction (LHD).
1 INTRODUCTION Development of innovative technological projects requires an insight of technological evolution of processes under study. Toward this, it is convenient to identify the main focus and targets to direct research and development efforts. This work intends to adapt the concept of Critical Technological Dimension (CTD) to get a new tool for management of innovation in mining industry. CTD is defined as a numerical index that reflects the technological effectiveness and/or competitiveness of any process under evaluation. CTD has to have an aggregated character to represent the whole process, be universal to be applied by/to any process user/application and variable in time to allow monitoring its evolution. CTD concept is applicable to any process; being underground mining method only a particular one, however, this work is dedicated exclusively to Block/Panel Caving. Present work validates the use of mean extraction rate as Block/Panel Caving CTD, and useful application of this concepts are presented to actual state of the art in order to outline its potential for innovation in underground mining methods. Conventional LHD Block/panel caving technology is deeply analyzed to emphasize how CDT can guide us to look for high impact technological innovation.
• Better advance taking from high grade blocks, allowing in practice, an advance of fine ore output. • In the case of new mines, lower time is required to reach full production.
2 MEAN EXTRACTION RATE AS CTD
Mean extraction rate results from a combination of several factors, including mine design and material handling or other unitary operations technology, and restrictions imposed by the ore body or originated in management practices. Two operational conditions controlling the ore extraction can be distinguished: extraction done while caving process is ongoing, also called propagation phase; and extraction done when all material is already broken after propagation phase, when full extraction rate can be applied. Propagation phase is the period when it is not convenient to extract more material than the "ground" delivers as caving phenomena converts in situ material in broken material. It is a common practice to accept that this process finishes when 30% of material has been extracted. During propagation phase, it is necessary to operate at an extraction rate Tp, named propagation extraction rate, whose value is usually much lower than the full extraction rate Tf. In full extraction rate stage, restrictions imposed in propagation stage are progressively abandoned and consequently, other process control enters into play, as the occurrence of hangups that interrupt the flow through drawpoints.
For Block/Panel Caving methods, mean extraction rate expressed in units of t/m2-d, has been identified as a CTD, that is, a measure of global process effectiveness. CTD strictly refers to technological key factors that determine the business value. For this analysis, we will define the mean extraction rate (Tm) as the ratio between the total capacity of production (t/d) and active area (m2) needed to obtain that production. Business impact is clear when greater rate of extraction can be achieved as listed below: • Greater productivity from active area can be achieved, which means that equal active area yields greater production.
2.1 Full Extraction Rate Analysis In conventional Block/Panel Caving system with LHD, a production module is usually composed of a production drift with 8 to 10 drawpoints each side and one ore pass. This means that each LHD extracts the ore from any of those 16 to 20 drawpoints and dumps it to the ore pass. If the bucket capacity is 6 tons and the mean transport distance is between 50 and 60 m, as it is frequent in different mines of Codelco Chile, LHD productivity ranks at 200 t/h, equivalent in the best case to 3,600 t/d (18 h/d). As a result, for one module composed by 16 points with 225 m2 of influence each, (drawpoint spacing 15x15 m), production will be carried out from 3,600 m2, meaning a maximum structural capacity of 1.0 t/m2-d.
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469
Even in conditions of maximum operational capacity, the mean extraction per drawpoint is 225 t/d, with 1.125 h/d of LHD operation, so, drawpoints utilization is less that 5%. This low utilization, which is remarkable by itself, becomes even lower due to ore does not draw regularly through drawpoint, but intermittently due to hangups. In fact, depending on the size of broken ore, flow is intermittent due to hangups even when dealing with "fine material". To face this situation, operational experience indicates that most convenient method to hold LHD productivity is to allow them shift the drawpoint each time the flow is interrupted, and shift the drift every time that module has completed 10 hanged drawpoints. This because secondary blasting crew can accomplish 10 unhangups per shift, so under this work organization, almost two modules are needed per LHD, therefore a lower extraction rate will be achieved because the extra area needed to face hangups. Following the logic of this work organization for ore extraction operation, the impact on the extraction rate can be calculated for different ore quality and drawpoint spacing. Results are presented in Table 1: Table 1: Impact of hangup frequency Case A: Conventional 6 t LHD, 16 drawpoints per drift, 10 unhangups per shift size
hangup time for frequency drift change h
very coarse
80
4
0.40
coarse
160
8
0.57
0.76
medium
320
16
0.73
0.97
1.35
fine
640
32
1.12
1.57
very fine
1280
64
-
-
drawpoint spacing in m 15x15 13x13 11x11
1.70 1.00
1.33
1.86
Table 1 shows that when material is medium sized the possible achievable rate is 73% of ideal condition without hangups. Extraction rate decreases because every time when a change of drift is required, another one has to be available, so to obtain the same production more active area is needed. Drawpoint production capacity is independent on its area of influence, so extraction rate increases when drawpoint spacing is small because module total area becomes smaller. To evaluate the impact of unhangup performance, a hypothetical exercise was done reducing unhangup time to one half of the standard, getting Table 2 results. From Table 2 it is concluded that if unhangup could be done in a half of standard time, extraction rate in medium sized rock could increase since 73% to 84% of the ideal extraction rate, that is, approximately 15% more than with the standard unhangup performance. Another way for increasing the extraction rate is to increase LHD bucket capacity. Results of an exercise with 12 t LHD bucket capacity are presented in Table 3. In this case, the ideal rate (without hangups) is duplicated; however extraction rate in medium sized rock only grows 56% compared to 6 t bucket (Table 1). Then, to duplicate the extraction rate, considering the actual situation with hangups, not only LHD bucket capacity has to be duplicated, but also unhangup time should be 470
size
hangup time for frequency drift change
extraction rate t/m2-day
t per hangup
h
very coarse
80
4
0.57
coarse
160
8
0.73
0.97
medium
320
16
0.84
1.12
fine
640
32
very fine
1280
64
-
-
no hangups
drawpoint spacing in m 15x15 13x13 11x11
1.22
1.57 1.70 1.78
1.00
1.33
1.86
Table 3: Impact of increase LHD bucket capacity Case A.2: 12 t LHD bucket, 16 drawpoints per drift, 10 unhangup per shift size
hangup time for frequency drift change
extraction rate t/m2-day
t per hangup
h
very coarse
80
2
0.50
coarse
160
4
0.80
1.07
medium
320
8
1.14
1.52
fine
640
16
very fine
1280
32
-
-
extraction rate t/m2-day
t per hangup
no hangups
Table 2: Impact of better unhangup productivity Case A.1: Conventional 6 t LHD, 16 drawpoints per drift, one half of standard unhangup time
no hangups
drawpoint spacing in m 15x15 13x13 11x11
1.94
2.13 2.70 3.13
2.00
2.66
3.72
done in one half of standard time, as it is shown in Table 4: From analysis on extraction rate at full extraction rate for conventional LHD technology, it can be concluded that: • With flat infrastructure (drawpoint spacing of 15x15 m and 13x13 m) in medium sized rock, extraction rates have a structural limit of 0.7 to 1.0 t/m2-d respectively. • Former figures can be increased to 1.3 t/m2-d if 11x11 m or less drawpoint spacing is used, for which infrastructure design should be inclined or a Macrozanja type. • A technological improvement of unhangup operation to reduce the standard time to one half only implies 16% extraction rate increase.
Table 4: Duplication of extraction rate Case B: 12 t LHD bucket, 16 drawpoints per drift, one half of standard unhangup time size
hangup time for frequency drift change
extraction rate t/m2-day
t per hangup
h
very coarse
80
4
0.80
coarse
160
8
1.14
1.52
medium
320
16
1.45
1.94
fine
640
32
very fine
1280
64
-
-
no hangups
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drawpoint spacing in m 15x15 13x13 11x11
2.24
2.70 3.13 3.40
2.00
2.66
3.72
Massmin 2004
• Double LHD bucket capacity implies only 56% increase of extraction rate • To duplicate current extraction rate in all cases, both LHD bucket capacity and unhangup performance should be duplicated. In such a case, extraction rate Tf should reach at most 1 to 2 t/m2-d when flat infrastructure and wide drawpoint spacing design is applied in medium sized rock, and when small drawpoint spacing (11x11m and Macrozanja one) extraction rate could raise to less than 3 t/m2-d. 2.2 Mean Extraction Rate Analysis As formerly noted, mean extraction rate Tm represents the link with business because it expresses the whole production capacity per unit of active area. Given that propagation occurs while the first 30% of any column is extracted, and in the remaining 70% full extraction rate is used, the following relationship can be established: Time in days needed to extract a column of ore:
time =
0, 3 * T 0, 7 * T + Tp Tf
0, 3 * T f 0, 7 * Tp
7 T f = * n * Tp 3
(2)
Ap * Tp + Af * T f AT
Tm =
1 0, 3 0, 7 + Tp T f
(3)
At the limit, when full extraction rate is as large as wanted, it is noticed that mean extraction rate has an asymptote, which is proportional to extraction rate of propagation phase:
lim v 2 → ∞ (Tm ) =
Tp 0, 3
(9)
AT : Total area of system in m2 Ap : Propagation area in m2 Af : Full area in m2 With (10)
From equations (8), (9) and (10) it is deduced that mean extraction rate Tm can be expressed as:
Tm =
or:
(8)
Where "n" factor also represents the relation that should exists between the area at propagation stage and the area at full extraction rate. Mean extraction rate will be also given by:
AT = Ap + Af
T
(7)
or:
(1)
Then, mean extraction rate Tm in t/m2-d is given by:
0, 3 * T 0, 7 * T + Tp Tf
n=
Tm =
T = tonnage per m2 of column TP and Tf = Extraction rate at Propagation and Full production phases respectively in t/m2-d.
Tm =
Where "n" factor is defined in such a way those times of extraction in propagation and full extraction rate time, could be the same, so:
10 * n * Tp
(11)
3 * ( n + 1)
It can be observed that Tm is asymptotic to 10/3 of Tp, figure that is consistent with the one encountered in equation 4. In this way, if one takes a reasonable value Tp=0.40 t/m2-d, the maximum possible Tm will be 1.32 t/m2d. Figure 1 shows values obtained for Tm and Tf in t/m2-d (main axis) according to different values of n and for propagation rate Tp=0.40 t/m2-d. Productivity of drawpoints in t/h (secondary axis) is also plotted (dotted line).
(4)
A condition of production stability has to be set to assure that any time the mine will produce the same amount of ore. Toward this is necessary that each time an area is exhausted, an equivalent one where its propagation period has finished have to be ready to start extraction at full production rate.
propagation _ time full _ time
0, 3 * T Tp
0, 7 * T Tf
(5)
(6)
Condition of stability is verified when: propagation time = n * full production time Massmin 2004
Figure 1: Relationship between mean extraction rate and full extraction rate for Tp=0.40 t/m2-d Santiago Chile, 22-25 August 2004
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It can be observed that: • Case A: For conventional system, is enough to maintain an area ratio on the order of n=0.76 to be able to replace opportunely the exhausted area. Full extraction phase reaches extraction rates on the order of 0.7 t/m2-d (as mentioned earlier) so, mean extraction rate would be close to 0.57 t/m2-d. • Case B: if conventional LHD system could be improved to achieve full extraction rates on the order of 1.45 t/m2-d, as in the case of Table 2, mean extraction rate should increase to 0.80 t/m2-d, for which it will be required an area ratio close to n=1.55. • Case C: In an hypothetical system able to reach a full extraction rate on the order of 3 t/m2-d, which accordingly to earlier analysis is not reachable with conventional LHD system, mean extraction rate could raise to 1.0 t/m2-d, with an area ratio higher than n=3.25. 3 CTD ANALYSIS
• A combination of those factors will give a characteristic average output per drawpoint and the correspondent full and mean extraction rate. For example, if an LHD producing 200 t/h is serving 16 drawpoints, the maximum output will be 12.5 t/h per drawpoint (225 t/d per drawpoint ). This value could decrease when considering all those aspects formerly mentioned. • Returning to analysis of Figure 1, technological output needed for each case are: - Case A: Output slightly lower than 10.5 t/h per drawpoint, a value that is consistent with the one analyzed. - Case B: Output of 22.5 t/h per drawpoint. - Case C: Output close to 45 t/h per drawpoint, that is, 4 times the conventional output. Considering the maximum technological output associated to materials handling and a hypothetical situation in which rate of propagation is twice the limit set on Figure 1, give us Figure 3:
In practice, in mines it is common to observe mean extraction rate lower than those presented in Case A (Table 1) mainly due to: • The influence of propagation rate, which is the factor that controls the process, as shown in Figure 2. In fact, propagation rate average at first 30% of extraction could in practice be lower to 0.40 t/m2-d for primary material, being even closer to 0.30 t/m2-d or smaller in some cases.
Figure 3: Relationship between mean extraction rate and full extraction rate for Tp=0.80 t/m2-d
Figure 2: Mean extraction rate curves for different propagation rate. Note: It is assumed there is no technological limit associated to material handling or another operational factor.
• Once restrictions due to caving propagation are over, full extraction rate is mainly controlled by productivity of material handling system (Secondary axis in Figure 1), however, many factors are also involved in final full extraction rate as, among others: - Secondary size reduction done in drawpoints and/or in ore passes. - Grizzly and ore pass dimensions. - Productivity of size reduction equipment (pick hammers or crushers). - Productivity of main transport. - Operational interference (undercut and new area adding). - Availability and preventive maintenance and repairing of equipment and productive areas.
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• In case A, attainable mean extraction rate with conventional technology is at most on the order of 0.7 t/m2-d, which is equal to full extraction rate that this technology can reach. This happens despite "ground" can delivers more than that. • On the other hand, if an improvement of material handling system (Case B) were possible in order to increase 105% the maximum output of Case A, improved technology could reach mean extraction rates on the order of 1.15 t/m2-d. • Notice that in Case B mean extraction rate increase is only near to 65%, that is far from being proportional to the effort required to improve conventional technology. • In Case C, a new technology as Continuous Mining or Macrozanjas or other is considered, attainable mean extraction rate is close 1.63 t/m2-d, that is, 2.3 times higher than obtained with current technology. • In Case C, it is clear that conventional LHD technology (even the improved one) could hardly attain the required rates. In an hypothetical exercise where propagation rate is completely unrestricted (Figure 4), the new propagation rate will be fixed by production capacity of material handling system and will be equal to mean and full extraction rate. It is clear in this case that the grater the material handling system capacity the greater the advantage in mining process.
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• Material handling technology for conventional and future caving method. • Capacity for adding new area. • Production and development management. Finally, present work has introduced a different approach to innovation management in mining industry by means of relatively simple analysis driven by business effectiveness indexes, to focus the efforts of research and development on those aspects that actually governing the mining business behavior, instead of impulse partial improvements increasing productivity of current technology, that only tends to give us marginal results in mining process output. ACKNOWLEDGEMENTS
Figure 4: Extraction rate by % extraction of varies technologic scenarios Tech1: Conventional technology with propagation limit Tech2: Conventional technology without propagation limit Tech3: Conventional technology improved without propagation limit Tech4: Alternative technology without propagation limit
Another interesting element to analyze is the behavior of area ratio "n" that can be a management tool, which allows the owner to choose the better system capacity compatible with operational complexity. The greater the n factor the greater the operation complexity. Also, as can be seen in Figure 2, mean extraction rate increments due to n increments are lower than mean extraction rate increment due to jump to a different curve corresponding to a greater rate of propagation. One can guess that a borderline value for area ratio should be between n=3 and n=4. On the other hand, an increase of the mean extraction rate implies necessarily that production area exhausts faster. Hence, new area adding capacity could become a technological restriction of mean extraction rate. 4 CONCLUSION The main conclusion of this analysis is the relevance of borderline that caving propagation process imposes to rates of extraction. Considering that mean extraction rate is the one effectively linked with profitability of business; one observes that it makes little sense to have a technology capable to reach high full extraction rates after propagation phase, if for being able to use it, a high area ratio is required, even though that mean extraction rate will not have a proportional increase. In fact, it was seen in the hypothetical situation when restrictions imposed by caving propagation were eliminated, current material handling system does not take advantage from that condition and consequently it becomes the critical technology that determines the productivity of the whole system. With current technology, the expectation for increasing the productivity of the method is relatively limited, because it will be blocked at currently levels, not exceeding 0.7 to 1.0 t/m2-d. If new technology could be able to get 2.0 t/m2-d (like Macrozanjas) or 3.0 t/m2-d (like Continuous Mining concept) as a full rate of extraction, expected profit would be commanded by the ability to raise the rate of extraction during caving propagation phase. Future analysis will focus in finding new technological elements not included in this work, as well as determining the controlling factors of each of mentioned technological borderlines as: Massmin 2004
Authors are grateful to all their colleagues of IM2 that helped them during the development of this work. Also, the authors want to acknowledge the permission given by Codelco Chile to publish this technical paper. REFERENCES • A. Karzulovic & Asoc., 1998. Evaluación geotécnica métodos de socavación previa y avanzada, División El Teniente, Codelco Chile. Internal Report. • Bartlett, P. and Croll, A., 2000. Cave mining at Premier Diamond Mine, in Proceedings MassMin 2000, pp 227234. • Diaz, G. and Tobar, P., 2000. Panel Caving experiences and Macrotrench- an alternative exploitation method at the El Teniente Mine, in Proceedings MassMin 2000, pp 235-248. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2001. Estudio sistemas traspaso de mineral, División El Teniente, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2001. Técnicas de manejo de materiales en el nivel de producción, División Salvador, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Diseño conceptual para Minería Continua, División Salvador, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Estudio de metodología de acondicionamiento de macizo rocoso para hundimiento, División Andina, Codelco Chile. Internal Report. • IM2 Instituto de Innovación en Minería y Metalurgia S.A., 2003. Estudio de minería alternativa en Tercer Panel, División Andina, Codelco Chile. Internal Report. • Maass, S, 2003. Análisis de competitividad de métodos de explotación de minas mediante indicadores tecnológicos, 118 p. Final Thesis to obtain Degree in Mining Engineer, Universidad de Chile. • Metálica Consultores S.A., 2001. Ingeniería básica del sector Inca Extensión Oeste, División Salvador, Codelco Chile. Internal Report. • NCL Ingeniería y Construcción S.A., 1998. Desarrollo conceptual método de explotación Macrozanjas, División El Teniente, Codelco Chile. Internal Report. • Rojas, D., Molina, R., Bonani, A. and Constanzo, H., 2000. The pre-undercut caving method at the El Teniente Mine, in Proceedings MassMin 2000, pp 261-266. • V.E Ingeniería Ltda., 1999. Informe de ingeniería conceptual proyecto Extensión Inca Oeste, División Salvador, Codelco Chile. Internal Report. • Waissbluth, M, 2000. Gestión de la innovación tecnológica. Cátedra de Gestión y Economía Minera, Depto. Ing. Minas, U. De Chile. Notes.
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Controlled draw in block/panel caving Francisco Raña, Mauricio Telias, Instituto de Innovación en Minería y Metalurgia S.A., IM2 Mario Vicuña, Superintendencia Planificación, El Salvador Division, Codelco Chile
Abstract One of common paradigms in Block/Panel Caving set uniform draw as a rule to delay dilution entry, this is based on a simplified description of the problem, that assumes ore-waste interface is an horizontal surface and that ore value distribution is uniform. This work presents, through theoretical analysis and simulation exercises applied to actual cases, that it is possible to set non-uniform extraction schedules not only delaying dilution, but getting high value ore earlier. The effect is achieved using a Dynamic Gravity Flow Simulator (DGFS), which is able to replicate any sequence of extraction and allows knowing the broken ore configuration at any time. By using DGFS and an special algorithm of optimization, developed at IM2, it is possible to get a significative improvement of global benefit using different sequences of draw, each one adapted to each ore body and mining layout.
1 INTRODUCTION
2 DYNAMIC GRAVITY FLOW SIMULATOR
The goal of modeling gravitational flow on granular materials have been pursued long before mining industry implemented block or panel caving as a mining method. However, its massive use has contributed heavily to the development of important studies on the matter. The most relevant development is the one achieved by D.H. Laubscher, who generate a mixed grades predictive model based on the behavior of interactive flow, known as Volumetric Model. Additionally to the studies about gravitational flow, arise new alternative solutions for solving the problem of premature dilution. Nowadays, uniform draw is considered as the best alternative; however, it could be far from the best choice. Considering uniform draw as the best solution is a simplification because it considers that ore-waste interface is an horizontal surface and that all the ore has the same value. Nevertheless, experience shows that deposits are not homogeneous and the ore body limit is just an economic consideration. This work introduces two essential planning tools. First of all, it is considered a dynamic model for gravitational flow problem, where every action in the present causes an effect in the future; based on that, it was developed a gravitational flow simulator which can represents the deposit configuration, in every stage of drawing.
The model proposed for representing gravitational flow is very simple and based on mathematical modeling studies. It works as follows: the deposit is represented by preferably a small cubic block model (5x5x5 meters). Every block takes as its main variable the density (proportional to the tonnage of ore material) and as a secondary variable the ore grade inside the block. There are countless secondary variables, as geomechanic or geologic data, but only economical significant variables were used for this study.
WASTE ORE
Figure 1: Simplification of the problem. Secondly, it proposes an algorithm that finds a sequence of extraction, which maximizes the global profit by considering each block with a different value. 474
Figure 2: When a block diminishes its density, the upper blocks contribute with material.
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Drawpoints are modeled as blocks with capacity to get empty, that means to became blocks with density zero. Process get started when drawpoint got empty and then refilled with upper blocks material. The contribution of every block is previously defined and becomes a parameter that has to be calibrated. Blocks that contribute to refill the drawpoint diminish their density, so they have to be filled by their respective upper blocks, repeating the process up to the upper level, generally, surface. During the filling process of blocks, there is a mix of secondary variables, in the case ore grade. Resultant grade is calculated as a weighed average, in this manner, every time that a drawpoint is draw it is possible to determine its exact secondary variables value (grade). The model can be used for replicate the history of extraction, determining for each drawpoint the grade obtained after each draw. 3 DGFS APLICATION
Figure 4: Result of replicating the real sequence in Inca Central, with the Volumetric Model.
DGFS was tested with real data from Salvador copper mine of Codelco Chile Corporation. There were made tests in two different sectors, Inca Central and Inca Norte, both of them mined with Panel Caving method.
Figure 5: Result of replicating the real sequence in Inca Norte, with the Volumetric Model.
Figure 3: Layout of the drawpoints in the analyzed sectors.
The background information was a block model, which provides density and grade of copper up to surface. Also, there was available the drawpoints extraction history and grade sampling record of both Inca Central and Inca Norte sector which have been already exploited. Having this data it was possible to replicate the exploitation step by step, as it was actually achieved, as it is frequent, in a different way than scheduled. Volumetric Model Validation As a reference, the results obtained by replicating the actual exploitation using the Volumetric Model were compared. The results obtained were not as satisfactory as it was supposed to be, getting more than 20% error in ore recovery estimation. In fact, actual ore extraction was more than 20% of ore predicted by Volumetric Model when it is applied to actual sequence of extraction. Massmin 2004
The explanation for these results is basically that the Volumetric Model is only applicable when a quasi-uniform exploitation has been performed. In actual case, there are drawpoints that are early closed and others with over exploitation. In the last case, the most probable is that overexploitation is not only vertical but presents a horizontal movement that this model is unable to represent. DGFS Validation The solution for obtaining satisfactory results implies considering a dynamic model. The results obtained doing the same test using DGFS are very promising, getting less than 5% of total error on ore recovery. In this case the model adapts itself to the extraction history and it considers the actual draw in every drawpoint mixing the upper material grade following the actual exploitation characteristics. 4 OPTIMIZATION ALGORITHM Using DGFS it is possible to know the deposit configuration in terms of secondary variables value in every moment. In practice, we can control every drawpoint status, determining the copper grade value (or another variable) in the next extraction and also have an effect on the sequence
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Figure 6: Result of replicating the real sequence in Inca Central, with the DGFS.
reasoning above and the non-linearity of the problem makes necessary the use of heuristics. The heuristics based algorithm generates a list of drawpoints with the amount of material to be extracted from each one of them. It works as follows: at every moment (everyday or every shift) it is determined the expected benefit each every drawpoint. This expected benefit is an average of the benefit generated by each one of the upper column blocks, weighed by the probability of been extracted through that drawpoint. The geometry of the upper column and the weighing of every block are parameters to be determined for each deposit in particular. In practice, it can be assumed that the material moves around successive ellipses and the probability of being extracted decreases exponentially with the distance from the drawpoint. The advantage of this heuristics is that not only determines the drawpoints with a better benefit but also the extraction combinations that contributes to generate a greater global benefit, bringing forward the extraction of higher grade ore and delaying dilution. 5 ALGORITHM CONCEPTUAL ANAYSIS The optimization algorithm above introduced reverts the paradigm of uniform draw as the best choice for extraction. This, because such process only happens under special conditions (homogeneous deposit and constant height of mineralization).
WASTE WASTE
ORE
WASTE
ORE
ORE
Figure 7: Result of replicating the real sequence in Inca Norte, with the DGFS. Figure 9: Different cases of geometries in ore bodies. To demonstrate this, it was performed a theoretical study about draw behavior considering different geometries of ore distribution and a typical layout of drawpoints. In each case, it was compared an uniform draw against a draw proposed by the algorithm.
Figure 8: Every block makes a contribution to the expected benefit of drawpoint. In this way, dilution can be detected in advance.
Figure 10: Drawpoints layout
of extraction, which is the most important. Changes in sequence may means bringing forward the closing of points or extending overexploitation. The complexity of the
Case A: Horizontal Geometry This is a typical deposit geometry. Despite the intuitive proposing of a uniform extraction sequence as the best
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The comparison was done between the ore recovered before dilution percentage of every drawpoint were above 50%
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solution, results of the algorithm, shows a different sequence. Algorithm proposed result is a consequence of the fact that central drawpoints could have a greater interaction than lateral drawpoints. Therefore, it is better draw them at a lower rate in order to control premature dilution.
Case C: Intermediate Waste This is not a very common geometry in massive deposits. Nevertheless, it is possible to find it in some cases, for example, when outstanding material from older mined sectors on upper levels, or ore bodies having high grade in lower and upper levels separated by a low grade zone. It is important to analyze the algorithm behavior under these circumstances. As we can see, the best draw strategy is not easily determined. Intuitively uniform draw should have a good performance, although ore recovering would be notably influenced by grade difference among the layers.
Figure 11: Average percentage of activity in the drawpoint during the mining stage. Case A Quantitative result shows that global recovering for uniform draw is 85%, whereas, for controlled draw, recovering is 91%. Case B: Inclined Geometry In this case, conventional strategy was to have a uniform draw in a first stage, followed by a second stage only extracting from higher column drawpoints (right). Algorithm proposes a solution where, at the beginning only higher columns are extracted (drawpoints located to the right) and then following stage, continue with an extraction like case A. Briefly, the algorithm suggests a sequence of extraction where the ore-waste interface is moved up to an horizontal status, where it applies the solution already known.
Figure 13: Average percentage of activity in the drawpoint during the mining stage. Case C The result proposed by the algorithm is very interesting. At the beginning supposes the extraction like an horizontal geometry deposit but when intermediate waste material gets the drawpoints, it only extracts from some of them. The reasoning is absolutely understandable, it assumes that dilution is unavoidable but extraction from less drawpoints implies a lower extraction of the waste material and a faster flow from the upper high grade ore. In few words, once the intermediate waste gets the drawpoints and high grade ore is detected over itself, the most advisable is a sequence that privileges the entrance of the diluting material (in this case corresponding to high grade ore). The roles get inverted and the sequence get adapted to the change. There is a random choice of the points that have to be overexploited, because there is not big differences among them. The uniform draw got a global recovering of 55% and controlled draw, 72%. In all cases above, the aim is to prove that despite that a uniform draw gives good results, it is always possible to have a better draw strategy. Although these are fictitious examples, reality is just the combination, in different proportions of those exposed cases.
Figure 12: Average percentage of activity in the drawpoint during the mining stage. Case B
6 SCHEDULING ALGORITHM APLICATION
It may seem that both extraction sequences are very alike, however, the order, in these cases, do alter the result, delaying dilution much more when draw is controlled. The global recovering result for uniform draw is 77% whereas controlled draw result is 89%.
In order to verify the true potential of scheduling proposed algorithm, it was applied on mine sectors presented before (Salvador mine). The same basic parameters applied in actual mining were considered:
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• Fixed tonnage for every year of exploitation. • Opening of new areas identical to the reference case. • Controlled rate of extraction. Results obtained show that using the algorithm as sequence generator it would have been possible increase near 20% of ore recovery compared with actual results. The greater recovery obtained is result of bringing forward the extraction of high grade ore, which, in particular for these sectors, correspond to outstanding high grade material in upper levels, that were not mined before.
We have to be kept in mind that the given values correspond to benefits on ideal cases, where it has not been taken into account operational problems of the mining method. So, they can be considered as the "ceiling" value. However, the benefits margin is quite great, to recommend the use of this scheduling method in future plans. 7 CONCLUSIONS Mining paradigm of uniform draw as "the best solution for delaying dilution" is based on a simplification of the actual problem. Depending on the deposit, there are non uniform strategies allowing greater ore recovery while delaying the dilution entry. DGFS, developed by IM2, provides the current status of drawpoints in every moment of exploitation and it can represents, according to obtained results, the behavior of the gravitational flow. As DGFS can be also a good simulator of ore draw in Block/Panel Caving mines, optimization algorithm provides a solution for scheduling depletion in a realistic way. ACKNOWLEDGEMENTS The authors thank Codelco Chile Division Salvador that supported all the steps of this research and provided data from its exploitation to perform the validation. REFERENCES
Figure 14: Result of applying the algorithm in Inca Central.
• Alfaro, M, 2000. Modelamiento Computacional Predictivo del Flujo Gravitacional, Proyecto FONDEF 1037, Santiago de Chile. • González, G, 1999, Estudio del Comportamiento de un Material Granular Mediante Modelos Computacionales. Memoria Universidad de Chile, Santiago de Chile. • Goles, E y Peña, S, 1996, Modelamiento y Simulación del Flujo Gravitacional, Centro de Investigación Minero y Metalúrgica (CIMM), Santiago de Chile. • Imenitov, V, Abramov, V and Gorbunov, V, 1969, Probability Distribution for a Stochastic Model of the Motion of Ore During Discharge, Moscow Mining Institute. • Yongjia, W, and Xinguo, L, 1981, Numerical Simulation of Ore Drawing, Northeast Institute of Tecnology.
Figure 15: Result of applying the algorithm in Inca Norte.
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A draw control system for scheduling production in block caving David Rahal, Research Scholar, Martin Smith, Senior Lecturer, JKMRC, University Of Queensland
Abstract Block caving is an underground mining method that is suitable for low grade mineral deposits. One of the biggest factors in determining the success of a block caving operation is the implementation of a rigorous draw control system. This paper presents a scheduling system that allows the development of long term, tactical draw plans. The system is based on the integration of a mixed integer linear programming scheduler (MILP) with a dynamic resource database. The draw schedules generated are optimised for all production periods simultaneously. This life-of-mine optimisation offers the advantage of having the production in each period being optimal with respect to all other periods. Once this tactical plan has been implemented, actual production is used to update the panel contents before subsequent scheduling cycles. The results presented in this paper show that the MILP model can be used for mine planning and tactical scheduling in block caving mines.
1 INTRODUCTION The JKMRC has developed the Integrated Draw Control System (IDCS) to provide the basis for optimal draw practice and to minimize the impact of disruptions in the production cycle. Draw control, as referred to herein, relates to the sequencing, scheduling and rates of draw from block and panel cave drawpoints. While the basic requirements of good cave management are well understood, they have been difficult to implement. IDCS is designed to schedule production from drawpoints as close as possible to the desired production targets while satisfying operational and geotechnical constraints. Over the life of the panel, the draw schedule provided by IDCS will maintain a smooth draw of the ore in order to minimise early dilution entry and thereby maximise panel recovery. IDCS does not chase NPV at the expense of operational or geotechnical restrictions. In fact, production targets exceeding operational or geotechnical limits are not allowed. IDCS provides a rational approach to cave management consistent with current draw control practice. Cave scheduling can be broken into two major components, short-term and a long-term. Short-term scheduling is aimed at guiding production towards the monthly plan. If variations between the planned targets and the actual draw are recorded, the daily scheduling system adjusts subsequent calls so that the monthly plan is not compromised. Thus, shortterm scheduling consists of an iterative process of monitoring draw, comparing to plan and correcting if needed. This is the full extent of draw control at most operations and is insufficient for guiding life-of-mine production. Long-term (monthly to annual) scheduling aims at meeting life-of-mine objectives such as retaining the integrity of the cave while maximizing recovery or cash flow. Production targets set by long-term draw scheduling should dictate short-term and operational planning requirements. The solution provided by IDCS includes the tonnage to be drawn from each drawpoint by scheduling period, as well as the deviation from the ideal production target. The objective of optimisation in IDCS is to find a sequence of solutions as close to the ideal geotechnical draw profile and overall production requirements as possible. Therefore, IDCS is primarily designed to be implemented as a tactical rather than a strategic mine planning system, i.e., basic strategic Massmin 2004
decisions such as annual production rates, cave footprint, progression of the undercut and production system capacities have already been determined. Still, as will be demonstrated, IDCS provides the means to evaluate the feasibility of strategic objectives and to compare the relative value of alternative mine plans. The original concept for IDCS was derived from a draw control scheduling system initially developed by De Beers (Guest et. al, 2000) as shown in Figure 1. In it’s current version, IDCS includes several components: an Access database for the various production scenarios and a Mixed Integer Linear Programming-based (MILP) draw optimisation module. IDCS also includes the option of cave simulation and material mixing as a component of draw optimisation. As envisaged in Figure 1, a fully integrated system would also include production/vehicle tracking and short-term scheduling down to the level of a shift using the monthly draw target coming from the MILP. Integration of short and long term draw control within the framework of IDCS is planned for future releases.
Figure 1: De Beers scheduling system schematic (after Guest et. al, 2000). 2 DRAW OPTIMISATION Draw optimisation is implemented in IDCS as an MILP. An MILP is a Linear Program having binary variables controlling
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the availability of drawpoints for production in a given scheduling period. It consists of linear equations describing constraints on production and an objective function that seeks to minimize the sum of the period deviations (the sum of the penalty weighted block deviations from ideal draw), the deviation from the production targets and from external sources such as dumps. The critical draw control constraints regulating long-term production in the IDCS are: • Relative draw rate between adjacent drawpoints – Ensure interactive draw by limiting the draw rate between adjacent drawpoints. Interactive draw aims to pull neighbouring drawpoints simultaneously to promote equal draw down over the entire footprint. Controlling draw rates between adjacent drawpoints avoids early waste ingress by maintaining a smooth ore/waste interface. • Minimum draw rate - Eliminate point loading by avoiding ore recompaction in the drawpoints. • Maximum draw rate (Maturity Rules) – Limit the maximum draw rate to ensure that the cave advance rate does not exceed the undercut rate. Also, the draw rate is chosen so that the secondary fragmentation process within the cave (induced by mechanical interaction of the caved material) reduces secondary breakage by blasting. • Waste limit – Waste content in the plant feed cannot exceed a percentage during a given period. This constraint is imposed because processing plant efficiency drops dramatically with increasing waste percentage. • Materials handling and Capacity constraints – production cannot exceed the capacity of the materials handling system including drawpoints, tunnels, orepasses, etc. through to stockpiles and other accumulation areas. Capacities can also be applied to the processing plant both in terms of tonnages by material category and as allowable grade ranges. The equations defining the objective function and constraints are composed of variables for the sequencing of drawpoints, drawpoint production, upper and lower bounds on drawpoint production, and the flow of various material grades and classes through the materials handling system. The constraints define the values that these variables can take in any feasible solution while the objective function drives the values these variables take towards optimality. The objective function takes the form of a Goal Program in which there is a balance between multiple, and possibly competing, objectives. In IDCS there is a balance between maintenance of an idealized ore/waste interface or Ideal Profile and obtaining the desired production targets. As described in the following section, this Ideal Profile is defined as a set of contours in the IDCS GUI that describes the desired draw down of the ore across the panel as a depletion surface. The premise is that deviation from this Ideal Profile will result in premature waste entry and stress concentration resulting in loss of reserves and drawpoints. Constraints in IDCS relate the desired depletion level into the tonnage of production required from a drawpoint in a given period that will maintain an even draw down as defined by the Ideal Profile. Deviation from this ideal drawpoint production is represented by a deviation variable in the objective function. Other deviation variables represent the deviation of total production from user specified production targets. 3 DRAW CONTROL INTERFACE The interface to the draw scheduling system is a C++ GUI that interfaces with ILOG CPLEX‘ (or AMPL‘) and Microsoft Access‘. Access was selected as the production database because it is a portable database that is available on most mine sites. However, links between the GUI and other database software can be forged because the GUI extracts 480
data using SQL queries via ODBC connections. The scheduling system is based on the concept of draw scenarios containing panel data for different operating conditions. Creating a draw scenario consists of the following steps: • Importation of block names, coordinates, tonnage data, and column heights. • Enter period definition (length in months and days, target production, economic discount factors), maturity rule definition, and rock type definition. • Create the nodes and links in the materials handling system (tunnels, orepasses, haulages, etc.) An example of this is shown in Figure 2 where a crusher is being connected to the shaft. • Define the ideal draw profile as a series of relative depletion lines. The ideal chevron profile shown in Figure 3 was drawn in one of the three possible modes available to the user. • Apply a weighted distance algorithm to calculate the relative depletion levels for each block • Emulate ideal production by depleting the panel. The depletion algorithm determines the levels of ideal depletion for each block in each period based on the production targets specified in the period definition. This information is used with the remainder of the data to generate the MILP formulation. • The model is then solved using the CPLEX MILP solver and the optimum draw schedule is output for review. An example of the results that can be obtained using this system are presented below for the BA5 panel of the De Beers Cullinan Diamond Mine.
Figure 2: Links in the materials handling system (BA5 panel).
4 CASE STUDY The BA5 cave, opened in 1990, has served as a major testing ground for block caving. It has a lift height of 130 m with ore drawn on the 630 m production level using both electric and diesel LHDs. The average monthly production was on the order of 150,000 tons for the months between January 1996 and April 2000. The panel layout is an offset herringbone as shown in Figure 4. Each production drift on the 630 level is bisected by barricades and ventilation barriers which separates the panel into North and South ventilation districts. These barricades restrict LHD movement such that each end of
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the drift is treated as a separate tunnel for production scheduling. A schematic of the BA5 materials handling system can be seen above in Figure 5. Rock is loaded at each drawpoint and transported to a nearby orepass by LHD. The material then gravity flows to the haulage level where it is transported by train to underground crushers. The crusher product is subsequently transferred by conveyor to the production shaft for hoisting to the surface. This materials handling system (also shown in plan view in Figure 2), along with the BA5 data, was used to demonstrate how the system can be used in mine planning and for evaluating the feasibility of production plans. Its use as a tactical planning tool for the recovery of a poorly developed panel has been discussed in Rahal et al, 2003. An important concern in planning an operation is the effect draw strategy has on total production. This effect is explored by considering the results of the following four scenarios: • Control, Baseline draw strategy in which all production control constraints are active. • Maturity, Illustrates how draw maturity rules limit production when no relative draw rate (RDR) constraints are applied. • Reversed, Highlights how the system can be used to determine effect that direction of advance has on production. • Relative, Tightens the limit on the allowable RDR between adjacent blocks.
Figure 3: Ideal depletion profile definition.
The basic schedule definition for the trials and the important differences between each trial and the Control scenario are presented in Tables 1 and 2. It can be seen that each schedule contains 35 periods with an "ideal" production target of 150,000 tons. Table 1: Schedule summary for all trials. Schedule Length Schedule Duration Days Per Month "Ideal" Production Target Maturity Rules Rock Density Mean Virgin Reserve Per Block Mean Reserve Of Super Blocks (15 largest of 307 blocks)
35 periods 186 months (15.5 y) 20 150,000 tons/month 3 Levels 2.62 tons/m3 76,340 tons 214,446 tons
Table 2: Scenario constraint parameters.
Figure 4: BA5 production level.
Control Maturity Relative Reversed
Profile Direction
Lower RDR
Upper RDR
Normal a Normal a Normal a Reversed b
0.25 * 0.40 0.25
4.0 * 2.5 4.0
a Chevron advance from North-West to South-East. b Chevron advance from South-East to North-West. * Constraint dropped from optimisation.
Figure 5: Materials handling system schematic. Massmin 2004
A breakdown of the number of months in each period can be seen in Figure 6. Table 1 also makes reference to a set of "super blocks" which contain roughly three times the average virgin reserve. This abnormally high reserve causes these fifteen blocks to deplete more slowly than the
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remainder of the blocks in the panel. It will be shown how this slower depletion rate interacts with the RDR constraints to have a significant effect on panel production. Figure 7 shows the shape of the maturity rules production profile. Note that the ratio between the minimum (101 mm/day) and maximum (404 mm/day) production rates is 4.0 (and conversely 0.25). The allowable production difference between adjacent drawpoints, the lower and upper relative draw rate ratios (0.25 and 4.0 respectively), were set to reflect this characteristic of the maturity rules.
Figure 6: Duration of each period in months.
Figure 7: Three level maturity rule profile. The total production achieved in each scenario compared to the "ideal" production target can be seen in Figures 8 and 9. Figure 8 shows that all of the scenarios initially lag behind the ideal production target. The Control, Maturity, and Reversed schedules require a production ramp up of ten periods (12 months) before achieving the ideal production target for the remainder of the panel life. This delay in achieving full production is not unexpected since the maturity rules restrict total production until the blocks are greater than 6.5% depleted. In contrast, the Relative schedule only achieves the cumulative production target towards the end of the schedule, nearly 114 months later than the other trials. The deviation from the ideal panel profile (shape) for all four trials can be seen in Figures 10 and 11 where the 482
deviation is plotted as the sum of squared error (SSE). The deviation has been expressed as SSE because the total positive and negative deviations for all blocks in the panel for each period must be represented. The deviation values are squared to prevent positive and negative values from cancelling each other out. Figure 10 is rescaled in Figure 11 to show the SSE for the Control, Maturity, and Reversed trials. A comparison of the Control and Maturity schedules shows that the maturity rules are a major factor limiting total production. The cumulative production profiles for both are similar. However, Figure 11 shows that the two have different profiles with time. Both trials deviate from the ideal during production ramp up. The difference occurs in Period 10 when both achieve their ideal production target. The Maturity trial assumes the ideal panel shape (zero deviation) while the Control trial fails to achieve the target profile until Period 26. The difference between the two occurs because of the presence of a number of "super blocks" in the panel. These blocks each contain nearly three times the average block reserve (Table 1) and as a result they are depleted at a slower rate than the remainder of the panel (e.g. removing one thousand tons from a super block depletes it only one third as much as removing a similar amount from an average block). Because these blocks deplete more slowly, they restrict, and are restricted by, their relative draw rate limit with adjacent blocks. This effect is illustrated in Figure 12 where the percentage of the total deviation generated by these fifteen blocks has been plotted. A comparison of the Control (Figures 3 and 14) and Reversed (Figures 13 and 15) trials show the following: • Super blocks are major contributors to the total deviation in both trials after Period 10, • The super blocks contribute on the order of fifty percent of the total SSE until they are fully depleted, • The RDR constraints interact with the super blocks to limit production across the panel, and • The interaction between the RDR and panel production is related to the larger footprint and slower depletion rates of the super blocks. The first three trials illustrated the effect that the RDR constraints have on the cave profile when combined with production from super blocks. What remains to be shown is whether or not the RDR achieves its goal of maintaining a smooth profile throughout the panel life. A comparison of the Control and Relative trials shows that the SSE deviation reflects a smooth profile in which the super blocks and their neighbours lag behind the desired depletion levels. Figures 14, 16, 17 and 18 show the amount of material remaining in each block as a percentage of the original content. Figures 14 and 16 are the panel profiles for Period 7 of the Control and RDR trials, respectively. Figures 17 and 18 show the same information for Period 20. In both pairs of figures it can be seen that distortions of the panel profile are tied directly to the location of the super blocks (larger areas). However, the reminder of the panel maintains the ideal chevron profile. It can be seen that tightening the allowable lower and upper RDR limits from 0.25 and 4.0 to 0.4 and 2.5 has lowered production and increased the profile deviation in the Relative trial (Figures 8 and 9). The primary cause of this reduced production is the delayed maturity of the super blocks. It can be seen in Table 3 that most of the super blocks reached maturity (404 mm/day) one period later in the Relative trail. The most dramatic of these increases (Blocks 5, 6, and 7) had the greatest effect on total production because their peak production, along with that of their neighbours, was delayed for between 21 and
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Figure 8: Total cumulative production, Periods 1 to 10.
Figure 11: Expanded section of Figure 10.
Figure 9: Total cumulative production, Periods 10 to 35.
Figure 12: Contribution to total SSE of the fifteen super blocks (Control Period 31, Maturity Period 11 and Maturity Period 12, are outliers. They are a large portion of an insignificant deviation from ideal targets).
Figure 10: Deviation from ideal plan as expressed by sum of squared differences. Figure 13: Design of reversed panel profile. 96 months (Period 4 to Period 13 and Period 5 to Period 27). 5 CONCLUSIONS The four scheduling trials presented in this case study have shown that both the draw maturity rules and the Massmin 2004
relative draw rate constraints have a significant impact on production rates. The draw maturity rules tend to limit production early in the life of the panel. Once sufficient drawpoints have matured, the ideal panel profile can be maintained.
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Table 3: The delay in super block maturity caused by tightening RDR limits.
Figure 14: Panel profile, Control, Period 7.
Super Block
Control
Relative
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15
PERIOD04 PERIOD04 PERIOD04 PERIOD04 PERIOD04 PERIOD05 PERIOD05 PERIOD05 PERIOD06 PERIOD05 PERIOD12 PERIOD12 PERIOD13 PERIOD17 PERIOD31
PERIOD05 PERIOD04 PERIOD05 PERIOD06 PERIOD13 PERIOD27 PERIOD22 PERIOD10 PERIOD08 PERIOD06 PERIOD13 PERIOD12 PERIOD13 PERIOD16 PERIOD31
Figure 15: Panel profile, Reversed, Period 7. Figure 17: Panel profile, Control, Period 20.
Figure 16: Panel profile, RDR, Period 7. Figure 18: Panel profile, RDR, Period 20. The effect of the RDR constraints is more complex. When ore blocks are uniform in tonnage and area, RDR limits that are proportional to the initial and mature draw down rates (101 and 404 mm/day versus 0.25 and 4.0 lower and upper bounds) will not have a significant effect on the panel profile. However, when large "super blocks" are present, production is restricted by the RDR. The lower production levels occur because the larger blocks take longer to mature and they restrict the production 484
from their neighbours. This lower production level is subsequently transferred across the panel because adjacent blocks limit production from their adjacent blocks and so on. Tightening the allowed RDR exacerbates this effect. In planning production where the ore blocks have a wide range of contained material and cross sectional areas, draw
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control policy can mitigate the effect of non-uniform block contents by relaxing the relationships between super blocks and their neighbours. This can be done in two ways, by ignoring the relationship during the optimisation (drop the relationship from the set of relationships) or by increasing the allowable lower and upper limits on the RDR. Either of these are suitable if it is accepted that super blocks generally occur on the periphery of the panel and must therefore lie outside of normal cave operations. The examples of draw control given herein demonstrate a variety of applications for IDCS and the impact of geotechnical constraints on production. Production targets are commonly set based on the capacities of the materials handling system (LHDs) and the number of operational drawpoints, without sufficient consideration of draw constraints based on sound draw control practice. Adhering to constraints such as drawpoint maturity and relative draw rate limits between adjacent drawpoints will initially limit total production, but the consequence of violation of these constraints has been shown in practice to lead to reduced recovery at best and loss of the panel itself at worst. The case study has demonstrated the impact of the parameter values used in the RDR and maturity rules constraints. While these constraints correspond to geotechnical requirements, their impact on production ramp up demonstrates the sensitivity of production to these draw control parameters. The MILP methodology provides the means to find a balance between poor draw control and overly constrained production. It also demonstrates the need for continued research into flow and caving mechanics. The MILP optimisation technology incorporated into IDCS also provides the means to consistently quantify and compare various design and production strategies. An example is given of two different draw (or "Ideal") profiles showing the potential advantages of a reversal of the undercut and production direction. By creating alternative production or design scenarios, the life-of-panel production and recovery results can be used to evaluate alternative undercuts, rates of advance, panel footprints and production level layouts.
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Ongoing research includes development of an SLC version of IDCS and a version that is integrated with cave simulation and mixing. The structure and function of the SLC version is similar but optimisation seeks a draws front across all sublevels that optimises either cash flow (in a sequential optimisation across a series of planning periods) or NPV (in a multi-period formulation). The version of IDCS that includes cave simulation also uses sequential optimisation as opposed to the multi-period formulation described in this paper. It has the advantages of accounting for the non-linear relationship between draw and the state of the caved material and enables the optimisation process to account for vertical variability in material properties and both vertical and lateral mixing from period to period. ACKNOWLEDGEMENTS Funding for this project was provided by the sponsors of the International Caving Study (ICS II). The sponsoring companies are De Beers, Anglo American Technical Services, Rio Tinto Technical Services, CODELCO Chile, Sandvik Tamrock, LKAB, Newcrest Mining and WMC. Special thanks are due to De Beers for the contribution of the BA5 panel data and to the JKMRC Draw Control Team including Alan Cocker and Kai Riihioja. REFERENCES • Guest, A.R., van Hout, G.J., von Johannides, A. and Scheepers, L.F., 2000. An Application of Linear Programming for Block Cave Draw Control. MassMin 2000, 461-468. • Rahal, D., Smith, M.L., van Hout, G.J. and von Johannides, A., 2003. The Use Of Mixed Integer Linear Programming For Long-Term Scheduling In Block Caving Mines. Application of Computers and Operations Research in the Minerals Industries, 123-132.
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Combining long term scheduling and daily draw control for block cave mines Tony Diering, Principal Consultant, Gemcom Software International Inc.
Abstract Daily draw control is an integral component of the cave management at an operating block cave mine. It requires the routine gathering of actual production data such as tonnages and status per draw point as well as sampling information. This information is required to be stored and processed as well as being made available for long term production scheduling. Some of the problems which have been encountered and overcome include the following: Allowances for time delays in receiving daily actual data and the need to generate an order for the next day; the need to plan for the next month before the current month is completed as well as the need for smooth data flow from one month to the next. Sole reliance on Excel for such data manipulation is extremely error prone and risky. This paper describes an effective methodology which has been established in the PC-BC and CMS programs. A key component of this system is the effective manipulation of the large amounts of data generated on a daily basis. A second important component is the assigning of responsibilities for the effective maintenance of this data in the long term. The system is fully integrated into a SQL database for added data accessibility and security. The transition between the daily draw order cycle and the keeping track with the long range plans can then be established and reconciled. Benefits of the new methodology are discussed.
1 INTRODUCTION In principal, daily draw control in a block caving operation should be a simple process. An operator is required to set a tonnage target to be mined from each active draw point on a daily (or shift) basis. Draw points which are lagging behind should have a higher tonnage order, while draw points which are overdrawn should have a lower (or zero) order. However, experience has shown that the full implementation of a daily draw order system is much more complex. The objectives behind the draw control and draw order process can be divided into two components, long term and short term. The long term objectives are to mine in such a manner as to meet the long term mine plan, while the short term objective is largely focused on dealing with geotechnical stability and with other practical operating considerations. This paper presents the methodology and challenges which were encountered during the development and implementation of a daily draw order system. The system is called CMS and resides within Gemcom Software’s PC-BC system. CMS has been implemented and is operational at the DOZ mine of PT Freeport Indonesia and is soon to be commissioned at the De Beers Finsch mine in South Africa. The following overall aspects of the problem are discussed below: • The planning loop. Long term, daily and shift planning • Data collection and storage • The Catch up concept • Draw control. • Reporting • Some lessons learned 2 LONG TERM, DAILY AND SHIFT TIME SCALES The objectives of long term plans for block caving are similar in many respects to those for other mining methods. 486
Nevertheless the operational and geomechanical particularities of the mining method add several constraints to the process of planning and scheduling the block cave. A few relevant aspects of long term planning related to this paper are the following:: • Overall cave shape • Controlled opening and development of draw points • Maximizing NPV to the extent possible subject to the various geotechnical and other constraints • Generating monthly plans which can be fed to the daily draw control system Key functions of the daily draw control are: • Meeting the monthly target. • Adapting the draw rapidly to compensate for localized over or under-draw situations. • Being responsive to safety considerations (For example: wet muck draw points). • Being responsive to geotechnical considerations such as over-stressed draw points. • Being responsive to operational considerations such as draw belling, tunnel capacities due to secondary blasting and maintaining even daily production. • At a shift level, to enable dispatch of the LHDs to draw points in an orderly manner and to best achieve the daily draw order, with respect to both draw compliance and total tonnage produced. 2.1. From long term to daily draw order The key data required for daily draw orders from the long term plan are as follows: • Total tonnages for the month (separated into total, production, undercut and development tons) • Tonnage targets for each draw point for the month • Maximum allowable draw rates per draw point per day day which differs from the draw rate used in the long term schedules and includes availability and utilization of the draw points
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The data required for routine generation of the daily draw order can be broken into categories shown in Table . Table 1: Data requirements for daily draw control management. Totals / Other
By draw point
Setup
Shift details Priority info Strategy "rules" User status codes Report format info
Draw point areas Report positions
Occasional
Min buckets/shift Ave. dpt availability Panel status
Best HOD Best HOD tons
Monthly
Total tons targets Development tons Production tons New draw points
Target tons Past tons PRC limits
Target tons Control dates
Order tons Actual tons Latest assays Latest status Dpt priorities Special order dpts
GEMS Profiles
SQL workspace
Daily
Storage
• Total tons mined per draw point to date • Draw point closure strategy which might be linked to a NPV and/or a geomechanical sequence. • Total maximum tonnage per draw point (referred to as Best HOD tons) Note that there are no direct NPV or grade targets sent from long term to daily. If the daily draw meets the monthly target, then the overall grade and NPV goals of the long term plan will be met. The above statement highlights one of the overall strategies in setting up the CMS system. That is to clearly identify and isolate the different responsibilities for the overall process. We do not want daily operations personnel trying to second guess the financial strategies in the long term plan, or for the daily draw order to presume to know how best to dispatch the LHDs to each draw point etc.
In the first implementation of the CMS system at Freeport, much of the information was stored in Excel sheets. These had the advantage of allowing for easy data entry and visualization, but obviously lacked the data security of a SQL database. In a project with De Beers Finsch mine, the CMS system was upgraded to utilize the GEMS database (from Gemcom) running on the SQL Server platform. The current system is much more effective in producing daily orders as well as management of the transition from one period to another. 4 CATCH UP PROCESS The overall strategy is that if the daily draw order can reasonably meet the monthly targets, that the monthly target, in turn, will drive the cave shape towards the long term plan. Usually, this will not be achievable in a single
2.2. From daily draw order to shifts. The key data required for each shift is as follows: • A list of active draw points • A tonnage target for each draw point. This is specified as buckets • A tonnage tolerance for each draw point (called Extra tons) • Overall draw point status (as available at the time the order is generated) • deally, some indication as to the priority for the draw point 3 DATA COLLECTION AND STORAGE In order to have an efficient system, we need ready access to the appropriate information. In order for the system to be robust, we need to be able to get by under a number of conditions where we have insufficient, late or erroneous data. There is little point in telling a draw control operator that he cannot continue because of a lack of data, when there is a deadline to generate an order. So, we have to work around these types of exception as best possible.
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Figure 1: Example of progress with ten draw points illustrating the process of catch up.
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month, but may require several months, depending on the extent of deviation from plan. The process of steering the cave shape back to plan is called "Catch up" This is shown schematically in a 10 draw point example representing a cross section (Figure 1). In general, the longer term constraints on minimum and maximum draw point tonnages will limit how quickly draw shape can be restored. Figure 1 illustrates how the initial irregular tonnage profile is smoothed out as quickly as possible as permitted by the maximum tonnage per period constraints. If the maximum tonnage per draw point is reduced (say from 300 per period to 250 per period), then it takes longer for the smoothing effect to be completed. This is shown in Figure 2 below.
Figure 2: With reduced maximum tonnages, the speed at which catch can occur is reduced.
5 DRAW CONTROL A methodology has been devised to define the daily tonnage that is to be taken from each draw point, in order to meet the monthly target. If an objective function can be defined, then it is possible to use Linear Programming principals to derive a solution. A rating system has been set up for each draw point. "Points" can be allocated to each draw point based on the following: • Extent of over- or under-draw relative to the monthly target • Number of days idle • The forecast grade of the draw point • Other factors such as whether the draw point is in drawbell status • Safety or other special operational requirements By using a points system, it is possible for each operation to assign relative degrees of importance to the above components. Once the points have been allocated, we can also assign the minimum and maximum tonnages allowable per draw point in a shift. The maximum tonnage is derived from the overall PRC (Production Rate Curve) for each draw point. The minimum tons (if non-zero) is set as a minimum number of buckets per shift which is considered reasonable for each operation. The maximum tonnage per draw point per day also has to be adjusted for the expected monthly availability for the draw point, as well as some further adjustments for draw point which are heavily over- or under-drawn. 488
Then, a draw order can be assigned to each draw point, starting with those draw points with the highest "points" rating. Further adjustments are required to ensure that the draw order can achieve the daily production target set by the draw control officer. This is described in more detail by Samosir, Brannon and Diering (2000). 6 REPORTING If the right data is stored in an efficient database, then the flexibility to generate useful production and management reports increases dramatically. The CMS system currently generates the following reports on a daily basis: • Detailed draw order by draw point (an example is shown in Table 1) • Summary of draw order with tunnel and draw point type summaries (an example is shown in Table). • Comparison of Actual vs Order for previous day. • Compliance summary (Refer to the example in Table 2). Summary of draw point by "condition" as follows: - Overpull - Underpull - Hang up too long - Over 100% drawn, but still active - Other : For example, no monthly target set in LTP - Inactive too long. (ie Draw point is active status, but tons are not being taken from the draw point) - Hang up or temporary closed for too long. - Low grade warnings • Month to date summaries for Order and Actual tons and draw point Status. • Month to date summaries for draw point assay samples. (E.g. Cu and Au) • Detailed report by draw point for internal use by Cave management department. • Summary report for use by cave managememt department. • Draw point closure analysis considering the following: - % draw (vs Best HOD tons) - HOD for neighbours and this draw point - Grades, # low during last 30 days, latest and PC-BC slice file (reserve) - Current metal prices and shut-off grade - Rock type • Draw point sampling recommendation. Draw point may be sampled for any of the following: - Tons since last sample - Days since last sample - Check up unusually high or low grade samples - Check up on unexpectedly large change in grade. 7 SOME LESSONS LEARNED 7.1. Draw point priority Determination of draw point priority is an interesting challenge. There are various alternatives and each is likely to have different appeal to different mines. The draw point priority rating system in CMS tends to assign priorities to draw points in the following order: 1. Safety draw points. Usually wet muck draw point which are required to be mined regularly to avoid moisture build up. 2. Geotechnical considerations. Any draw point showing signs of stress or high convergence gets high priority. 3. Draw bell status. It is important to give these draw points high priority to avoid delays to the development and undercut process. 4. Under-drawn draw points. 5. "Normal" draw points. 6. Over-drawn draw points (in various categories).
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Table 1: Example of a detailed draw order
Table: 3 Example of a draw order with tunnel and draw point type summaries
Panel 15 Draw Point Status To Date Order Extra
P15-01W CP 125189 0 0
P15-02W CP 139245 0 0
P15-03W P15-04W CP CP 0 0 0 0 0 0
P15-05W P15-06W CP CP 0 0 0 0 0 0
Panel 16 Draw Point Status To Date Order Extra
P15-01W CP 180162 0 0
P15-02W CP 0 0 0
P15-03W P15-04W CP CP 0 195409 0 0 0 0
Draw Point Status To Date Order Extra
P15-01W A 155334 0 0
P15-02W CP 168611 0 0
P15-03W P15-04W CP A 179781 173907 0 0 0 0
P15-05W P15-06W A A 173457 155869 0 3 0 0
Panel 17 P15-05W P15-06W A A 156208 147859 0 0 0 0
Panel 18 Draw Point Status To Date Order Extra
P15-01W A 152146 0 0
P15-02W SP 160769 0 0
P15-03W P15-04W SP A 163615 170552 0 0 0 0
P15-05W P15-06W A A 152393 147669 0 0 0 0
Table 2: Example Compíance Report.
7.2. Hang up / secondary breakage priorities It turns out that the same drivers for the draw point priority for a daily draw order can also be applied to assign hang up clearance or secondary breakage priorities. If a draw point has been idle (hung up) for too long, then the priority to clear the hang up increases. Of course, when these priorities are sent to the shift level, additional shift based considerations may need to be applied, such as how many draw points in a panel are hang up and the overall crew availability etc. Massmin 2004
7.3. Adaptive vs shared shift orders There are two basic strategies which can be followed at the shift level: • The daily order is divided equally amongst each shift. This is probably preferable in a multi-shift environment which is not fully automated. Doing this provides each shift with a more or less equal chance of meeting their targets. However, the second and subsequent shifts are not responsive to over or under-draw from the previous shift. • The second (or third) shift must be adaptive to the results of the first shift to best try to meet the daily order. This is the system which will be used at Finsch and it will be highly automated. This seems to be a better approach, but there is an acknowledged draw back which needs to be managed. The issue is that there may be a tendency to mine the easier draw points in the earlier shifts and to leave the harder ones for later on. This leads to there being a higher risk of not making daily total tonnage targets. Another factor is that if there is a large number of draw points active, it may not be practical to try to mine every draw point every shift. 7.4. LHD Dispatch sequence Given that a draw point priority or rating system is already included in the CMS draw order generation routine, it makes sense to use that same rating system to try to dispatch the LHDs in an optimal manner. This was tried at Freeport DOZ with some success. It makes sense to try to mine from those draw points which are furthest behind, so that they have the best chance to catch up. However, if production continually starts from the "hardest" draw points, the risk that the total draw order (in terms of tons) will not be met increases. At Freeport, during the initial tonnage build up, when daily tonnage was important, tons were generally scheduled from one side of a panel to the other. The starting side was the more favourable ground. Once full production was achieved, LHD dispatch on the basis of a priority rating was successfully implemented. There is a assumption in the above that the LHD dispatch system is sufficiently advanced to be able to facilitate this type of optimization.
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7.5. Keep it simple Perhaps the most important lesson learned in implementing CMS was to keep things simple wherever possible. This is achieved in part by keeping long term planning considerations out of the daily draw order as far as possible and only sending the information necessary to generate the daily order. The CMS system has a number of scaling parameters which allow the relative importance of various parameters to be adjusted to suit each site. 8 CONCLUSIONS The CMS system has been successfully developed and implemented at Freeport DOZ block cave. It is currently being implemented at De Beers Finsch Mine for production start up during 2005. Initial development was done on site at Freeport. This was very useful in allowing the day to day exceptions to be studied and accommodated in the system. At the same time, the interface between the daily draw order and the monthly and longer term planning was streamlined and is now operating smoothly at Freeport.
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Although hard to measure directly, there seems little doubt that the current system has been proven to be effective in facilitating a smoother operation with obvious benefits to the bottom line. ACKNOWLEDGEMENTS The author is grateful to PT Freeport Indonesia and De Beers Finsch Mine for their assistance in the development of the CMS system. Thanks are also due to Gemcom Software International Inc. for time and funding to complete the work and submit this paper. REFERENCES • Diering, T, 2000. PC-BC: A block cave design and draw control system, Proceedings MassMin 2000, Brisbane, pp. 469-484. • Samosir, E, Brannon C and Diering, T, 2000. Implementation of Cave Management System (CMS) Tools at the Freeport DOZ Mine. Proceedings MassMin 2004, Santiago, Chile.
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Status of draw control practice and waste Management at Cullinan diamond mine Gert van Hout, De Beers Consolidated Mines, Technical Support Services, Operations Geotechnical Stephen Allen, Mark Breed, John Singleton, De Beers Consolidated Mines, Cullinan Diamond Mine
Abstract Cullinan Diamond Mine (previously known as Premier Mine) currently employs two mechanized block caves and plans another extraction level at approximately 900 m below surface. As with all major block cave operations, challenging situations and difficult events occurred during the implementation phase, unforeseen throughout the project feasibility study. The panel caves in the kimberlite pipe have experienced problems of various types that caused deviations from the original planned mining sequences. In combination with (the traditional) production pressures, numerous problems resulted in not achieving good Draw Control practice. A drastic change to the weekly Draw Control planning was introduced with the buy-in from production personnel and mine management. In conjunction with the improved Draw Control practice, an innovative but simple waste determination process has been developed and implemented. Both the Draw Control and the Waste Management at Cullinan Diamond Mine are currently based on ‘back to basics’ principles, straightforward and not controlled (yet) by impressive computer programs. This paper describes some of the geotechnical and practical related difficulties that were encountered during the implementation phase and how those were tackled with varying degrees of success. The importance and impact of Draw Control and Waste Management on the overall mine performance is illustrated.
1 INTRODUCTION
risks dictating Draw Control at many other sites are not discussed.
Cullinan Diamond Mine (previously called Premier Mine) started mining diamonds in 1903. The kimberlite pipe, the largest in South Africa is cut by a flat dipping 75 m thick gabbro sill at approximately 400 m below surface, shown in figure 1, together with the position of current mining blocks. Mining above the sill was initially open cast mining, later long hole benching (early 60’s) and block caving (late 60’s). Below the sill, resources in the BA5 and BB1E mining blocks are currently exploited by retreat panel caving. BA5 and BB1E presently have a combined production of approximately 13,000 tons per day. In BA5, mining started in 1988; 130 m below the gabbro sill and this mining block has a current life of mine until 2005. It is anticipated that the BB1E, where production started in 1996 approximately 230 m below the sill, will cease operations in 2009. In 2005, the mine plans to start producing kimberlite ore from the BB1E Advanced Undercut Cave providing interim tons until the Centenary Cut (previously referred to as C-Cut) commences mining at approximately 900 m below surface and starts production in 2009. Draw Control in the current block caves BA5 and BB1E has always been regarded as strategic (Rood & Bartlett, 1994, Bartlett & Nesbitt, 2000, Nesbitt & Vorster, 2000). A great deal of effort and capital have been spent over the years on infrastructure, computer programs and other tools to monitor the drawn volumes from the cave blocks. There are several reasons to understand why Draw Control is considered important, not only at Cullinan Diamond Mine but also at most caving operations worldwide.
2.1. Avoid premature gabbro dilution Draw Control is very crucial in avoiding early ingress of the overlying gabbro waste as this premature dilution would reduce the overall ore recovery and shorten the life of cave. Right from the Feasibility Study stage for BA5 and BB1E onwards, fragmentation of the kimberlite was predicted to be coarse and that of the overlying gabbro sill to be fine (Rood and Bartlett, 1994). At Cullinan Diamond Mine, it is anticipated that a recovery of 85 percent of the in situ ore can be achieved if proper Draw Control is exercised. Poor draw practice results in a much lower ore recovery rate as drawpoints will be forced to close earlier than anticipated for two possible reasons. Firstly, cut-offs due the influx of gabbro into the drawpoints adjacent to ‘overdrawn’ drawpoints occurs much earlier than planned. Secondly, inconsistent draw practice stimulates migration of material over considerable vertical and horizontal distances, inducing the premature mixing of waste with ore, especially as gabbro fragments are fine and would thus move through the column quickly, (Bartlett, 1998). To compound matters, it was found that the kimberlite fines percentage (a function of kimberlite accelerated weathering) was underestimated at the time of initial block cave feasibilities in the mid 80’s (BA5) and early 90’s (BB1E). 2.2. Minimise gabbro into plant headfeed The gabbro not only sterilises the drawn ore but also causes problems at the Dense Media Separator in the diamond recovery plant due to its high specific density.
2 VITAL REASONS FOR DRAW CONTROL Cullinan Diamond Mine hasn’t experienced serious mud rushes or seismicity and therefore these two supplementary Massmin 2004
2.3. Reduce risk of ore recompaction Lack of good Draw Control results in static columns of cave rock, resulting in ore re-compaction, generating point
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2.5. Optimal ore fragmentation Uneven draw or drawing at too rapid a rate may lead to very coarse fragmentation, poor grade control and severe waste dilution. As mentioned before, coarse ore fragmentation and fine waste fragments induce rapid gabbro waste movement from the top, through the column, to the drawpoints, thereby sterilising the drawpoint prematurely and creating problems at the recovery plant. It has often been observed that drawpoints with coarse fragmentation were subjected to draw, far above the call allowed by the maturity rules. If material is drawn too quickly, there is not enough ‘residence time’ for the kimberlite material: there is a lack of mechanical interaction between the ore fragments and insufficient communition of primary fragments as material gravitates down the cave column.
Figure 1: Diagrammatic plan and section of the kimberlite pipe showing geology and mining blocks.
loads on the extraction level that damage the production infrastructure. In kimberlite, if the draw of ore occurs more than three months after the undercut has taken place, the material in the drawbell and above the major apex may compact. This ore re-compaction creates uneven loading conditions on the extraction level or excessive localised stresses that adversely affect tunnels. This becomes apparent as support integrity worsens or when tunnels collapse. Experience in the BA5 and BB1E caves shows localised re-compaction of kimberlite becoming a major issue when the drawpoint was closed for rehabilitation for more than six weeks. Re-opening of compacted drawpoints requires a high level of secondary blasting, further damaging the brow area and in some cases it has taken as long as eight weeks to re-open a drawpoint. 2.4. Maximum Draw Zone Interaction In both the BA5 and BB1E, the offset herringbone drawpoint layout (figure 2) was used with the drawpoint spacing across the major apex ranging between 24.2m to 24.7m, possibly creating interaction problems across the major apex in some areas as the Isolated Draw Zone (IDZ) varies from 9m to 11m, depending on the kimberlite rock type. Figure 2 shows the zones of draw at the three different draw interaction modes according to Laubscher (2000). If drawpoints draw in isolation, (the horizontal section of) the zone of material affected by this draw can be approximated by a circle with a diameter equal to the IDZ. If there is even draw between the two drawpoints within the same drawbell, the affected area has an elliptical footprint, one and a half times larger than the IDZ. When there is interaction between adjacent drawbells, the zone of influence enlarges another one and half times. 492
Figure 2: Interaction modes (a: isolated, b: interaction within drawbell and c: interaction across minor apex)
Large blocks cause high hang-up frequencies in the drawpoints, create problems in the ore handling system and have a negative effect on the productivity as well as on the operating costs. Removal of drawpoint hang-ups may result in long down-time and are costly because of the secondary breaking requirements. Oversized ore blocks in the drawpoints must be reduced to reasonable sizes that can be handled by the load-haul-dump (LHD) machines. Back analysis of hang-up data during the period between July 1998 and May 2000 (Rahal and Smith, 2000) revealed that in any given shift, 34 percent of the available
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drawpoints are hung up and that the vast majority of these hang-ups were cleared within one day. 3 THE HISTORY OF DRAW CONTROL SYSTEMS EMPLOYED AT CULLINAN DIAMOND MINE To impose Draw Control on the current mining blocks, various systems and software packages were implemented. Some systems had more success than others. A comprehensive Draw Control system should at least consist of an integrated system of three major components: an accurate and dynamically updated ore resource database, a reliable vehicle monitoring system and a planning system. 3.1. Ore Resource database The very first version of PC-BC, developed for Cullinan Diamond Mine in 1988, was cumbersome to use: it took a long time (2 to 3 hours) to deplete all drawpoint loads on a daily basis and it did not have the fine graphics displayed by the present version. After some time, mine personnel moved towards spreadsheet type applications to store daily production data from the drawpoints. Later, a Microsoft Access database was employed to store the drawn volumes in combination with the status of each drawpoint. A user interface was then developed to present this data in a graphical format and to allow management to extract comprehensive summary reports. This application called BLOCINFO, also permitted the user also to reconcile the drawpoint production figures with the tons recorded by the treatment plant. In 1999, MinRAS an SQL based product that also contained the mixing algorithms, enabling the Mineral Resource Manager to derive a more accurate and auditable ore reserve statement (Guest et al, 2000). 3.2. Vehicle Monitoring System Reconciling tons, calculated from the loads that are recorded by the vehicle monitoring system (VMS) with the tons from the weightometers at the plant is essential in any good planning or effective Draw Control program to obtain a correct and representative mineral resource database. Various VMS options have been trialled at Premier including a beacon system using micro-wave technology (Nesbitt & Vorster, 2000) followed by a gyroscopic based monitoring system. None of the systems were a hundred percent satisfactory due to technical problems, as well as resistance from production personnel who saw it as a management policing tool. A combination of the radio-based voice communication (utilising the leaky feeder to establish contact between control room at surface and underground) and a manual recording based information system, currently in use, yields the best results. However, it requires production crews to buy into Draw Control. As Cullinan Diamond Mine has different size LHD units, there is an issue with the average bucket factor applied to derive ‘tons mined’ from ‘buckets loaded’ at each drawpoint and until recently, the total tons hoisted were consistently larger than the value based on the recorded buckets. 3.3. Planning System Up until 1998, Production planning at Cullinan Diamond Mine was founded on empirical geotechnical guidelines (Bartlett and Nesbitt, 2000) and did not take into account the full effect of resource and equipment availabilities. The empirical guidelines were derived from extensive block cave experience at Cullinan Diamond Mine and have been described by Bartlett, 1998. A more pro-active approach was initiated in 1998 when the weekly planning also incorporated mining constraints Massmin 2004
such as LHD availability, ore pass capacity, haulage and hoisting capability. The next draw control related concept established was that of the ‘ideal depletion surface’. A predetermined ideal depletion sequence up until the life of draw determines the short-term (week) draw schedule in that the plan attempts to come as near to the ideal depletion state as possible. At any stage, mining is in function of the ideal depletion profile and the call per drawpoint is influenced by its draw history (corrective call when there was poor draw in the past). Planning on a monthly and weekly basis was also carried out using spreadsheets. These files produced satisfactory output as it catered for the input of all parties involved in Draw Control. It was rather cumbersome to use but has served its purpose to stimulate interaction between geotechnical, mining and mineral resource management departments. Regular meetings were held between these departments to agree upon the monthly production plan produced by these excel files but adherence to this plan could be improved. 4 DISCUSSION OF CURRENT DRAW CONTROL PRACTICE 4.1. The Integrated Draw Control System In 1999, the ‘Integrated Draw Control System’ (IDCS) with a Mixed Integer Linear Programming (MILP) component described by Guest et al , 2000 was introduced on mine. The long and medium term schedule program based on the MILP was originally developed at Koffiefontein Mine for its Front Cave (Hannweg and van Hout, 2001), and is currently being used in the BB1E Advanced Undercut at Cullinan Diamond Mine. This scheduling module is the first planning optimisation tool in block cave operations that is able to optimise over the life of mine as well as over multitime periods. It incorporates all geotechnical, mining and financial constraints, in a unique way. Long term production planning for the older BA5 and BB1E blocks is still done by the use of spreadsheets and does not incorporate all ore flow constraints or geotechnical rules, except for the draw rate limits. The reason for the implementation of the excel files instead of MILP is threefold. Firstly, those panel caves are considered too depleted to optimise according to the principles within the MILP. Secondly, the main focus of those caves currently is avoiding tunnel collapses occurring from vertical loading and keeping porosity in the areas that hasn’t experienced destructive stress levels. And lastly, the MILP version allowing remote access only became available late 2003. The current Draw Control system at Cullinan does not cater for planning based on maximising NPV as the grades in the ore columns are set to an average column value and the mixing algorithms do not cater for vertical mixing within an ore column. The choice of an average column grade is justified, as there is generally very little resolution in the assumed grade profile vertically across the massive orebody. An optimal (and in South Africa legally prescribed) plan for all mining operations should be based on maximum ore recovery, thus maximising tons instead of maximising NPV, the last being a method relying rather on grade and revenue per carat. Block cave mining is a massive mining operation where principles of selective mine planning, based on financial parameters, cannot determine the production plan. These principles may constrain the schedule but an optimal plan is primarily based on geotechnical considerations. The mixing algorithms mentioned in section 3.1 were developed on site to simulate the ore movements within the cave and the parameters were calibrated successfully: the predicted time of drawpoint closure in the BA5 cave was within one month from the actual date of closure. These
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mixing algorithms are still being used but it is expected that REBOP,(Pierce et al, 2004), will be used as the standard tool to update the ore resource database. This tool should then also provide an accurate production waste ingress and grade profile (or carat production graph). 4.2. The back to basics approach Towards the end of 2002, a major drive from management, the ‘back to basics’ principle occurred, not only in terms of ‘keeping it simple’ but of increased adherence to the basic Draw Control rules outlined in the Cullinan Diamond Mine Code of Practice, (Bartlett, 2003) and ameliorating interaction between Draw Control and Production Departments. Some of the simple but critical procedures for a successful Draw Control Practice are discussed in this section. With issues of interaction being inherited from design (see section 2), the main aim of the current BA5 and BB1E Draw Control system is to draw equally throughout the caves with all drawpoints in production at one time and to minimise dilution from the gabbro sill. Future optimising codes such as the MILP are seen to be not applicable simply due to the maturity status of the BA5 and the BB1E and past Draw Control practices. The key to the whole short-term Draw Control process (figure 3) is the accumulation, storage, processing and presentation of data. The key relationship is between Draw Control and Production. The culmination of the whole process is presented at the weekly meetings that are held where all parties "strategise" key loading and tunnel or drawpoint rehabilitation patterns around Draw Control. Production and short-term planning understand that Draw Control needs alignment to the official annual production plan, therefore by applying correct Draw Control principles long-term cave management and production targets can be achieved.
system daily (Bartlett & Nesbitt, 2000). This information, in combination with daily production data, enables the Draw Control engineer to analyse the effect of the draw rates and draw history on the frequency and type of hang-up as well as waste entry parameters. Extraction rate limits at Cullinan Diamond Mine have hence been determined in terms of tonnage per day using a maturity classification, based on the accumulative production or life of draw from a drawpoint. Four classes were established and the associated draw rates increased from a maximum of 50 tons per day for a new drawpoint to a maximum of 200 tons per day for a mature drawpoint. As can be seen in figure 4, since the end of 2002, adherence to Draw Control has improved significantly (the average deviation in 2000 was approximately 73%). The graph displays the deviation (actual production - planned), averaged over all BB1E drawpoints with the weighting factor being the planned tons. A value around 10% is considered to be very good. A Production Factor (van Hout & Guest, 2000), based on this data has not been implemented at the mine as it is felt that average weighted deviation is adequate enough to communicate how well Draw Control is adhered to.
Figure 4: Weighted Average Deviation (datapoints and trendline) from the weekly Draw Control plan for BB1E cave.
Figure 3: Process flow chart on Draw Control at Cullinan Diamond Mine. Draw Control Observers gather the occurrence and type of drawpoint hang-ups and record it into the Draw Control 494
Cullinan Diamond Mine personnel are confident that the present recording, storage and presentation of actual mined tonnages in the MinRAS is satisfactory. Initially, the actual production per drawpoint was entered into the system on a weekly basis. This data is now imported on a shift to shift basis, in combination with other information associated with the drawpoint status (produced volume, hang-up type, waste content, remedial support status, etc.). Presentation of the draw data is usually in graphical format, accessible for all people involved in Draw Control to get a clear idea of the mined tonnage profile over a user defined period as well as of the actual drawpoint status. Figure 5 shows an example of the weekly Draw Control plan that has been derived on a Friday afternoon, after consultation between Draw Control and Production. Copies of this sheet are given to the Draw Control Officers (who enter it into MinRAS) and Production Mine Overseers (who distribute it to the Shift Bosses). The Mine Overseers also write this information onto ‘Draw Control loading’ whiteboards and compare the planned tons with the actual production data on a shift to shift basis. Listed below are some of the more important Draw Control production principles developed with the "back to basics approach" that are vigorously implemented with the aim of achieve correct block cave management:
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• Drawpoints closest to the orebody perimeter (highly depleted or not), are continually loaded, at a reduced rate if necessary. This allows for porosity in the caves and prevents the movement of groundwater from the contact areas to the centre of the ore body. • A strict minimum (120 tons) and maximum (1,200 tons) weekly call per drawpoint is implemented with a maximum daily call of 300 tons. This prevents the loading of all planned production from a drawpoint in a single shift, with no loading during the remainder of the week. • A target for the weighted average deviation (figure 4) has currently been fixed at ten percent with acceptance from both production crews and senior management. • It is strived to achieve equal draw across the major apex and continuous production per drawpoint throughout the week. This ensures achieving a maximum zone of influence (as in figure 2c). • Frequent interactions between Draw Control and production to communicate clearly the availability, waste content and status of drawpoints to anticipate correct loading calls. • Reducing the closure/maintenance time from six to three weeks where possible, to avoid re-compaction.
5 WASTE CALCULATION ON SURFACE 5.1. Current practice For the last five years the official measurement of global waste percentage has been determined by taking a sample of approximately 50 kilograms with a plough sampler per shift on the Sortex tailings stream. The size fraction on this belt is -65mm, +32mm. The sample is washed and handsorted into three different piles: internal waste, external waste and kimberlite. The piles are then weighed and their relative percentages calculated. Different tests have been carried out to ascertain that this process of waste determination is appropriate as the following underlying assumptions could be questioned: • The waste percentage in the -65mm to +32mm range is representative for the total size distribution of the headfeed. • One sample of 50kg material per shift is an adequate representation of the entire volume fed into the plant during that shift. • The waste percentage does not depend on the person who performs the test. The first test involved the waste analysis in the +4mm size distribution classes. It is impossible to quickly and accurately recognise rock types in fragments smaller than 4mm by manual visual methods. Results for this analysis were very similar to those of the -65mm to +32mm range. It can therefore be concluded that waste is evenly distributed across the different size fractions and the current fragment size range is adequate. During the second test, samples were taken every 15 minutes from the -65mm +32mm stream. The results from this test did not indicate that the current practice of one sample per shift needed to be adjusted. The last test consisted of an identical series of samples given to four different laboratory assistants. As can be seen in figure 6, results may vary depending on the lab assistant but would not justify the extra cost of additional personnel.
Figure 5: Weekly Draw Plan for the BB1E block cave.
The implementation of above principles should dramatically improve fragmentation across the caves, creating further reductions in secondary blasting and the possible achievement of monthly "production" targets at acceptable waste percentage. A secondary function of Draw Control was developed in Feb 2002 and consists of a qualitative assessment of the physical state of drawpoints. The system involved a monthly rating of each drawpoint on the basis of condition, stress damage, water and LHD damage. The ratings are transferred onto mine plans for future analysis. Fragmentation data in line with Laubscher’s, (2000), Rock Mass Rating classification were added to the data collection. The data is currently used in back analysis for the fragmentation analysis programs in order to model future block caves within the same ore body. Another Draw Control function is to maintain waste levels at twelve percent or below. Past sampling practices did not record levels of gabbro in each drawpoint. A waste management system developed by the Mineral Resource Management department assists in the prediction of waste tons and is discussed below.
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Figure 6: Analytical bias due to different lab assistants 5.2. Research and future technologies Research was undertaken to identify technology that could recognise waste accurately in - and possibly remove it from - the ore flow within a short time span. Providers of technology based on Microwave Attenuation, Infra red, Laser Induced Fluorescence, Natural Gamma Emission were approached and three different optical sorters underwent testing and extensive assessments.
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All applications were successful in differentiating between waste and kimberlite to some degree on surface (conveyor belts) but none were suitable for underground application. Optical sorter technology, widely used in the food and glass industries, proved to be the most effective method in recognising and ejecting waste from an ore stream. The ability to eject waste from an ore stream offers obvious, additional advantages in ore processing, allowing improved flexibility, an increased ore extraction ratio by recovering more diamonds from previously considered non payable drawpoints, lower crushing costs and ultimately improved revenues. 6 WASTE CALCULATION UNDERGROUND To manage waste effectively, an accurate means of measuring the waste percentage in each drawpoint must be established but the following factors complicate this process: • Dust, generated by LHD’s, makes visual observations extremely difficult. • Washing the ore in the drawpoint with water to get rid of the dust settled on the muckpile can result in rapid disintegration of the ore, thereby biasing the sample taken to determine the waste content. • There are different types of waste and different percentages of waste within different size fractions of ore, making the derivation of an average waste value rather difficult. • The finer the fragment size the more difficult it is to differentiate between the different waste and kimberlite types. The ore fragment sizes depend largely on the rock type and the maturity of the cave. At Cullinan Diamond Mine waste is classified as either internal or external waste. Internal waste is that which slumped back into the pipe during volcanic emplacement of the kimberlite pipe and consists mainly of felsite, Waterberg quartzite, norite and metasediments. The internal waste also includes barren late stage carbonatite dykes. In the metallurgical process these "floats" are separated from the ore stream via a process of dense media separation, as these rock types have a lower specific gravity. External waste consists of the country rock norite and the gabbro sill. These two rock types are mineralogically similar and can differ only slightly in texture. The norite enters the cave through the boundary drawpoints when country rock blocks detach, slump down, and cuts off or sterilizes parts of the resource It also enters the ore flow through the tipping of waste development into ore passes. External waste fragments are known in the metallurgical process as "sinks". Their specific gravity is similar to that of diamond bearing kimberlite, and it therefore reports to the concentrate of the dense media separation. Before 2002, the waste content was measured by visually estimating the waste in each drawpoint on a daily basis. Loss of historical waste data combined with the inaccuracy of visual estimates made it impossible to compare waste estimates from underground with the surface measurement of waste. The test work described in section 5 showed that the sampling methodology used on surface yields adequate waste percentage values. It was therefore decided to change the underground waste determination process from the visual estimation to one similar as applied on surface. This involves samples taken from a drawpoint, transported to a lab analysed and results are entered into a database. 7 WASTE MANAGEMENT UNDERGROUND AND BENIFTS THEREOF As mentioned above, the waste percentage of the headfeed needs to be kept below a critical level, above 496
which the plant recovery would be less optimal. Waste at Cullinan Diamond Mine has become increasing challenging over the last few years, as the current block caves become older and the reserves in the block are depleting fast. Keeping the average waste content below a threshold has historically been addressed by stopping underground drawpoints with waste content higher than 15% but this has sterilised large portions of the kimberlite resource. Waste problems can be controlled to some degree by blending, ore from various sources underground and on surface can assist to draw higher waste drawpoints longer, thereby maximising extraction. Prediction and control of waste tons is a task performed by Draw Control. Multiplying drawn tons per drawpoint with a waste percentage and adding this for all drawpoints across the two caves represents a value that indicates the expected waste tons. Initially (2002) there was an extremely close match between the predicted waste tons and the actual recorded waste tons. But of late, differences of 4% 5% have been observed, with the predicted values based on the underground results always exceeding the on surface determined results. Benefits of the current waste management system are listed below: • The greatest benefit from a mineral resource perspective is that more accurate forecasts and estimates can be made. Financial contribution of drawpoints can be established and assist in the decision making process, together with geotechnical factors, whether or not drawpoints or tunnels need to be closed. • When production planning is done, the expected waste tons can be calculated for each draw scenario. On the basis thereof, optimal draw and waste percentage can be derived. The estimated waste percentage is of great value to the metallurgical department as unexpected and excessive high density waste negatively affects recovery efficiencies. • A third benefit lies in the improved understanding of the migration of gabbro sill material as the cave depletes. The drawpoint waste data recorded since 2002 is still too scarce but correlation between the column depletion status and waste percentage will be analysed to determine waste ingress curves that can be used in future ore depletion planning. 8 SUMMARY AND CONCLUSIONS Cullinan Diamond Mine has experienced Draw Control challenges, resulting in early waste ingress and serious loading onto production tunnels. After a drive from senior management for ‘back to basics’, Draw Control adherence and ground conditions have improved drastically. Several Draw Control procedures are outlined in this paper. Waste is and always will be a mining and treatment constraint requiring constant management to ensure optimum extraction of ore. The introduction of a system that recognises and can eject waste from the ore flow will reduce the constraint of waste on Draw Control and treatment efficiencies. Cullinan Diamond Mine and the greater De Beers Group are vigorously pursuing the implementation of optical sorting technology, having completed extensive testing. A full feasibility study of the project is being undertaken. ACKNOWLEDGEMENTS The authors are grateful to all their colleagues who helped and supported them during the development and implementation of this work, in particular AR Guest, M Preece, C Baltus, HP Grobler and PJ Bartlett. The authors acknowledge the permission of the Director Operations and
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the Cullinan Diamond Mine, General Manager to publish this technical paper. REFERENCES • Bartlett, P.J. 1998. Planning a mechanised cave with coarse fragmentation in kimberlite. PhD. Thesis. University of Pretoria, South Africa. • Bartlett, P.J. & Nesbitt, K. 2000. Draw Control at Premier Mine. Proc. MassMin200, Brisbane. Vol 1: pp, 485-493. • Bartlett, P.J. 2003. Block Cave Code of Practice for Cullinan Diamond Mine. • Guest, A.R., van Hout, G.J., von Johannedis, A & Scheepers, L.F. 2000. An application of linear programming for block cave Draw Control. Proc. MassMin2000, Brisbane. Vol 1: pp, 461-468. • Hannweg, L.A. & van Hout, G.J 2001. Draw Control at Koffiefontein Mine. Proc. VIth International Symposium on Mine Mechanisation and Automation, SAIMM, 2000, Vol 1:pp, 93-96. • Laubscher, D.H.L 2000. Block Cave Manual.
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• Nesbitt, K & Vorster, J.A. 2000. Premier mine Draw Control and underground vehicle monitoring and control system. Proc. VIth International Symposium on Mine Mechanisation and Automation, SAIMM, 2000, Vol 1:pp, 93-96. • Pierce, M., van Hout, G.J. & Singleton J 2004. Draw Control of the BA5 cave Cullinan Diamond Mine: Back Analysis with REBOP. Proc. MassMin2004, Santiago. • Rahal, D.C. & Smith, M.L. 2000. De Beers Site Visit, unpublished internal report. Corporate Head Quarters, De Beers Consolidated Mines. • Rood, H.R. & Bartlett, P.J. 1994. Mechanized Caving at Premier Mine. Proc. XVth CMMI Congress, Johannesburg, SAIMM, 1994, Vol 1: pp, 219-225. • van Hout, G.J. & Guest A.R. 2000. Production Factor and Draw Control Factor, presentation to the International Caving Study. • Malope, P. 2001, Waste distribution in different size fractions of the kimberlite ore at Premier Mine, unpublished internal report, Premier Mine, De Beers Consolidated Mines.
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Draw point analysis using a marker trial at the Perseverance Nickel Mine, Leinster, Western Australia Bradley Hollins, Mining Engineer, WMC Resources Ltd., Olympic Dam, S.A. (formerly of Perseverance Mine) John Tucker, Production Engineer, WMC Resources Ltd., Perseverance Nickel Mine, Leinster, W.A.
Abstract An important challenge when operating a sub level cave (SLC) is to optimise draw point performance against metal recovery and ore grades. WMC’s Perseverance Nickel mine (Leinster, Western Australia) has recently improved metal recovery, whilst maintaining consistent grades from its SLC operations. To improve the recoveries, a series of steps was followed culminating in the implementation of a formal draw marker trial. This draw marker trial deployed 1762 cement filled steel markers across a series of 126 conventional production long holes. The aims of the trial were to determine draw ellipse shape, ascertain the nature and extent of any interaction of draw, and analyse the effect of fragmentation on draw performance. By doing this it was planned to assess the nature of material flow within the cave so that draw and hence cave extraction horizons can be optimised.
1 INTRODUCTION Perseverance nickel mine is part of WMC Resources Leinster Nickel Operations. Leinster is approximately 645km northeast of Perth in Western Australia. The Leinster nickel deposit is located within the Agnew-Wiluna greenstone belt at depths of 375m to +1120m (Figure 1).
Mining of this deposit is by sublevel cave (SLC) mining method. Several smaller nickel lenses/pods are mined using a variety of open stoping mining methods. In 2003 the SLC produced >1.5Mt. at 1.93%Ni., with an additional 118kt. at 2.29%Ni mined during stope development. Current proven and probable SLC reserves total 14Mt at 1.77%Ni. (Cooper, 2004). Forecasting of tonnes and conducting resource reconciliations in the SLC have proven difficult in the past. This was in part due to ore stocks being left in higher levels of the mine, following a planned ‘drop down’ strategy (triggered by an inflexion zone in the orebody), and a change from tonnage to grade based cutoffs (Wood et al., 2000). This paper discusses improvements in the understanding of flow behaviour and in predicting the draw performance within various areas of the SLC at Perseverance mine. 2 DESIGN PARAMETERS
Figure 1: Perseverance Regional Geology In 2003, 42,000 tonnes of nickel-in-concentrate were produced from the Leinster nickel concentrator and dryer. Ore was sourced from Perseverance underground and Harmony open pit mines. The Perseverance underground mine consists primarily of a disseminated nickel orebody.
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In 2000, two major reviews of the sublevel cave performance were undertaken at Perseverance (Bull, 2000; Rosengren and Scott, 2000) from which several design and operational changes were made. The most notable change was a reduction in the centre-to-centre crosscut spacing from 17.5 to 14.5 metres. This was based on the theory of interactive draw as shown in Figure 2 below. Essentially the theory supports a uniform draw down of the cave material, reducing early dilution ingress, and hence improved metal recovery. The theory of interactive draw also required the introduction of strict bogging practices for the SLC. Interactive bogging involves retreating individual levels with an aligned cave front and cycling the bogging of drawpoints between adjacent crosscuts on the level. Perseverance ‘rules’ were developed where bogging from individual crosscuts was limited to a maximum of 300 tonnes per cycle (with 100 tonne the optimum) and an absolute maximum of 600 tonnes per shift. These measures were implemented in an attempt to avoid surging of waste ingress. The shift bogging limits have also assisted in providing a more
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Figure 2: Independent vs Interactive Draw (Bull, 2000)
consistent hoist performance by ensuring that problems in an individual cross-cut are dealt with in a timely manner rather than having all the adjacent cross-cuts stop and wait. Prior to managing by interactive practices a difficult crosscut could be left for days or even weeks before being addressed. Some of the key design parameters currently used in the SLC are summarised in Table 1 and shown in Figure 3. Rings between adjacent crosscuts are staggered by 1.5 metres (looking in plan view) to avoid drilling into holes from adjacent crosscuts. The drill holes are drilled past the ring design tonnage outline in an attempt to fragment the material between adjacent crosscuts.
Figure 3: Cross section of 9715 crosscut (XC) 25 and 9715 XC27 ring designs
Table 1: Typical Design Parameters Parameter
Value
Blasted drive width
5.1 metres
Blasted drive height
4.8 metres (with flat backs)
Level interval
25 metres, floor to floor
Blast ring rump
150 forward
Apparent ring burden
3.0 metres
Ring toe spacing
3.2 metres
Explosive type
Powerbulk VE, density 1.0
Blast hole size
102 mm
Shut off grades (grade at which bogging is stopped and next ring prepared for firing)
0.9% below forecast tonnes; 1.2% forecast tonnes to design tones; 1.4% above design tonnes
Cut off grade
0.9% nickel
Production ring tonnage
3,500 tonnes (approximate)
SLC ground support
75mm of fibrecrete; mesh (floor to floor) pinned with 2.4 metre long split sets; an additional 50mm of fibrecrete; rings containing 15 debonded gewie bolts (3 metre long) on 1.5 metre burdens.
3 HISTORICAL DRAW MARKERS Steel sets were installed in sections of the 10030 level of Perseverance mine. Although these steel sets were intended Massmin 2004
Figure 4: Movement of steel sets and red sand through the Perseverance mine for ground support rather than draw markers, they have also provided an opportunity to follow the movement of large structures over multiple levels. Figure 4 shows the movement of these steel sets along with reported sightings of red sand (which was originally used as backfill in old stopes). Movement of the steel sets and sand has shown that the ‘drop down’ strategy pillar has broken up as planned. In addition, the steel sets are known to have moved 270 metres while deviating less than three degrees from a
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vertical trajectory. The path of the red sand is less well understood as there were three adjacent stopes filled on the 10000 level. In 2001, unsuccessful attempts were made to use old bogger tyres to study the flow of material through the cave. Following this attempt, six concrete filled drums were bolted into the sidewalls of the 9850mRL prior to production blasting. These drums are yet to be recovered. The SLC is now down to the 9740mRL. 4 DRAW MARKER TRIALS A more formal approach to studying the flow of material through the Perseverance SLC was commenced in 2002. The aim of these draw marker trials was to further improve metal recovery through gaining an understanding of the flow characteristics in the SLC. The approach undertaken was based on similar trials conducted at Ridgeway Gold Mine (Power, 2002-2004). The markers used in the trial were constructed of steel pipe 250mm long and 45mm in diameter. This marker size is consistent with measurements of fragmentation indicating that around 90% of material is smaller than 0.4 metres (Passmore, 2002). Each marker had an individual identifier welded on the side and was filled with concrete to simulate the ore density. The markers were finished with a spider at the top to centralize it during insertion in the marker hole and a ‘red cap’ at the base to keep it in position. The trials were conducted in five separate crosscuts on three different levels of the mine. A total of 1762 markers were installed at one-metre intervals. Half of the 126 marker holes were grouted after the markers were installed. By varying the number and pattern of marker rings between blast holes, the depth and width of draw could be studied. After firing of production blast rings, markers were primarily retrieved manually when ore was transferred by loader to a designated inspection area. This method accounted for 53% of markers found. During the period of markers being recovered detailed records of the ring material and visual flow characteristics were collected for analysis at a later stage. Secondary means of marker collection included inspecting the draw point rill, bogger bucket and crosscut. As a last resort, a magnet located adjacent to an underground conveyor belt was used to collect any remaining markers. When markers were collected using the magnet (20% of the markers found), an estimate was made of the tonnes that would have been extracted when the marker reported to the draw point. During the trials, there was no evidence of interactive draw behaviour in any of the trial areas (figure 5). The maximum width of draw measured through the recovered draw markers was 11.5 metres (+/- 1 metre). This means a zone of blasted material located between crosscuts and at the toes of blast holes did not report to any of the draw points from which the material was fired. During the trials, the measured size and width of this zone varied, however generally 35% of the blasted tonnes did not report to the crosscut from which they were fired or any of the adjacent crosscuts. As more than 100% of design tonnes were bogged from the marker trial areas, material must have been travelling to the draw point from outside the blast envelope. To study the depth of draw, up to three marker rings were positioned between each apparent blast burden (3 metres). On the 9715mRL crosscuts, markers were recovered from positions up to 2 metres in front of the ring being fired (figure 6). There was no difference in results between areas where the markers were grouted into position and areas where they were not grouted. The trials indicate that markers placed at the level above can flow into the draw point within bogging of 20% of the 500
Figure 5: Cross section looking west showing the marker trial results
Figure 6: Long section (looking north) showing marker trial results blasted (design) tonnes. This is consistent with early observations of barren ultramafic and felsic dilution entry when crosscuts above were bogged to waste. At the time of writing this paper, markers were starting to appear in draw points on levels below the trial areas. The second stage of marker recovery confirms the observations of vertical piping of material from above (figure 6). Recovery has been from markers positioned directly above the crosscut in an area between the extraction draw zones of adjacent cross cuts on the level above.
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The recovery of markers behind a freshly fired ring is consistent with brow break back records throughout the mine. At Perseverance, typical wear at the brow is approximately 1.5 metres and in some cases up to 6 metres of the brow has worn away during bogging. Brow wear and shearing of blast holes occasionally means that remedial work including cleaning out blast holes, redrilling blast holes or slashing (heavily dumped drill rings) through existing blast holes is necessary. At the time of writing, 540 markers had been recovered. The percentage of markers recovered ranged from 21% in 9740XC24 to 53% in 9715XC27. While this percentage of markers recovered might seem low, the pattern of markers recovered indicates that the collection methods were successful. Rather, the low percentage of markers found to date from the 9740 level trials is believed to be due to the production decisions made as a result of poor ground conditions in the trial areas. In one of the areas, two production rings (six marker rings) were fired together. In the other crosscut, slashing rings were drilled and fired through the area containing markers. While unplanned events, the slashing and multiple ring firings actually contributed substantially to the understanding of material flow at Perseverance. In the 9715 level trials, 20% of the markers were recovered from behind the ring being bogged. Therefore, without firing multiple rings together as occurred on the 9740 level, a true indication of the depth of draw in front of a ring would not be gained. When two rings were fired together, markers were initially recovered within one metre of the brow. In the crosscut where slashing was drilled through marker rings, the flow of markers was preferentially distributed around the position of the slashing holes. Detailed draw analyses were undertaken for all the marker trial rings. The results showed dilution entry points occurred between 11 and 25% of the tonnes extracted. Two of the crossucts which had early dilution entry points were affected by initial blasting related issues. Figure 7 shows the recovery curves of the marker trial rings.
Figure 8: Dilution Draw Curves showing the affect of hangups on recovery
5 PREDICTING OVERDRAW In 2003, reconciliation of material extracted against the tonnes fired at that point in time was undertaken. This showed that 94% of the tonnes bogged, 86% of the grade and 81% of the nickel metal contained had been recovered (Wood, 2003). Perseverance mine has historical records of draw point observations at intervals ranging from 50 to 300 tonnes. These observations provide detailed information on all aspects of the draw point performance. Using this information, block modeling of historical production information was undertaken and is regularly updated. With this block model and statistical analysis of the historical data, the impact on metal recovery of changes in design (including crosscut spacing, level intervals and ring burden) and varying draw point performance factors (such as hangup frequency, rock size and bogging rate) have been investigated. One of the best ring performance prediction tools was found to be the nickel grade left in the levels immediately above a draw point. Areas where metal was left above (indicated by higher last nickel grade calls) tend to perform better than areas mined to the cutoff grade of 0.9% nickel. This result was not unexpected and is consistent with the vertical nature of the draw observed during the marker trials. Using plots of the last nickel grade call recorded on a level (Figure 9), the extent of a dilution blanket on new levels can be studied in advance. CONCLUSIONS
Figure 7: Draw Recovery Curves from the marker trial rings The analysis of the data confirmed a correlation between hang-ups in the drawpoint and waste surging (Figure 8). From visual observations of the hang-ups it appeared that waste ingress was coming from the front of the ring as opposed to dilution from above the blasted ring. When a hang up did occur, markers from deeper, wider, and lower down in the blast envelope were retrieved. Massmin 2004
Metal recoveries at Perseverance are comparable with those achievable using an open stope mining method. The flow behaviour of material in Perseverance SLC is now better understood, and future planning will make use of this information to assist in optimising the extraction process. The discipline and implementation of ‘rules’ for firing and bogging practices associated with interactive draw have had a substantial positive impact with regards to metal recovery. The theory of interactive draw and the associated interaction of material between cross-cuts does not appear to be true. The ongoing recovery of markers below the initial trial areas will add to our knowledge, and allow ongoing incremental improvements in the modeling of flow behaviour.
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ACKNOWLEDGEMENTS The authors are grateful to all their colleagues at Leinster Nickel Operations for their assistance and interest shown during the trial, in particular the support from Chris Stone, Geoff Booth, and Andrew McDonald is greatly appreciated. Also, the authors want to acknowledge the permission given by WMC Resources Ltd to publish this technical paper. REFERENCES • Bull, G, 2000. WMC Leinster Nickel Operation, Sublevel cave review. Internal report, September 2000. • Cooper, A, 2004. Ore Reserves December 2003, Leinster Nickel Operation. Internal report, January 2004. • Passmore, A, 2002. Sub-Level Cave Fragmentation Study, WMC Resources – Perseverance Geology Internal Report, October 2002. • Power, G, 2002-2004. Personal Correspondence. • Rosengren, K, Scott, A, WMC Resources, Leinster Nickel Operations Sublevel cave, Review of Drilling and Blasting Operations, Internal Report, October 2000. • Wood, P L, Jenkins, P A, and Jones, I W O, 2000. Sublevel Cave Drop Down Strategy at Perseverance Mine, Leinster Nickel Operations, in MassMin 2000 Conference Proceedings, Ed. Chitombo G., ISBN 1 875776 76 9, AusIMM, Melbourne, pp. 517 - 526. • Wood, P, 2003. Perseverance mine sublevel caving, internal presentation, October 2003.
Figure 9: Block model slice of 9780 level showing last nickel grades
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Chapter 13
Caving Propagation & Subsidence
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Continuous modelling for caving exploitation Octavio Araneda Osés, Sergio Gaete Becerra, Gerencia de Recursos Mineros y Desarrollo, El Teniente Division, Codelco Chile
Abstract Based on the mineral stock balance equations, relations for the active area and production capacity are deduced; these equations are solved in closed form. The models represent the mining dynamics in terms of extraction and development rate. The models allow to estimate capacities and its variations, in tonnage and cash flow. The lost due to approximation error is overcompensated by the obtained simplicity and the insight. The models do not replace the detailed final calculations, but allow wider analysis saving detailed case evaluations.
1 BASIC MODEL FOR CAVING EXPLOITATION Planning and improving planning requires a representation of the processes easy to understand faithful enough and easy to use. We will present a model class that ,in the authors opinion, fulfills the requirements. Let us consider a sector with constant height h, undercutting rate y extraction velocity , the production area satisfies the following relation
d A = Vhdt −
T dt ph
In this case the area evolves according to AT phv (t −tm ) vh ph ph A (t ) = 1 − e e v AT v ph ph h 1 − e v A (t ) = AT phv (t −tm ) vh ph ph 1 e − e v
t ≤ tm t ≤ tm
The production rate is T = νA ,then The exploitation is designed whit a constant steady area Ar less than the maximum, to maintain it constant A = 0 is required and
vA d A = Vhdt − dt ph
% r vA = v%h ph
The production area satisfies the differential equation
A+
vA = vh ph
Since
Vh ph v v%h vh < v% v% Ar
tT
The solution for Aq is similar for the basic model vq t vh ph A (t ) = − 1 − e ph 3vq
If Aqr , Alr are the steady state areas, and k = 3 , we have
v%h =
Model including broken and unbroken material
Let us consider a sector with solid height hs and a height of broken material above the solid, let pq the broken material density. A simplified model for caving propagation can be built assuming that the process propagates filling the cavities generated by extraction. let Vq the extracted solid volume, the broken volume, the balance equation is
ph
=
3vl Alr 2 p ( h + hs )
h vl Alr = 2 1 + s h
vq Aq
The steady state production capacity is
T = vl Alr + vq Aq 2h T = vq Aq 3 + s h
pq ( vq + ve ) = pVq Vq =
3vq Aq
pqVe p − pq
In term of the steady undercutting rate
If caving propagates vertically let the extraction height, the limit for the breaking process, the base area
v%h =
B ( hq − he ) =
pq Bhe
3vq Aq ph
2h T = phv%h 1 + s 3h
p − pq
he p − pq 1 = = h p k
The production capacity is strongly dependent on the undercutting rate.
When the whole column is broken, so the extraction required to complete the process is
OPERATION VARIATIONS EFFECTS
he p − pq = h p
Undercutting rate variation If the undercutting rate decreases this results in a flow production lost and the steady state starting is delayed.
The broken material density varies between 1.8 and 2.1 , the in situ material density is 2.7, hence
Let us suppose that since ti and during ∆t the undercutting rate is vh d vh , during the period the breaking area evolves according to
2 0.6 he 0.9 1 = ≤ ≤ = 9 2.7 h 2.7 3
vA% & A% = vh − dvh − ph
In practice the column is assumed totally broken when the extraction is one third of the in situ column.
The original area satisfies
vAq A& q = vh − ph
Let Aq , Al , vq , vl the area and extraction velocity for the breaking zone, the broken height over the in situ material. The breaking area satisfies
v A Aq + q q = vh h p k vq Aq k Aq + = vh ph The broken area satisfies
Al = 506
vq Aq k ph
−
vl Al pq ( h + hs )
So the difference δA = A – A satisfies
d A& = dvh −
vd A ph
The complete solution is
phdv h 3vq d A (t ) = phdvh 3v q Santiago Chile, 22-25 August 2004
3vq ( t −ti ) 1 − e ph t ≤ ti
3vq 3vq ∆t ( t −ti−∆t ) ph 1 − e e ph
t i < t ≤ t i + ∆t t > t i + ∆t Massmin 2004
The steady state starting time satisfies
A ( tr + dtr ) − Ad ( tr + dtr ) = Aqr by definition tr > + tr + ∆t , then the variation is
dv 3vq ti 3vq ∆t ph dt r = log 1 + h e ph e ph 3vq vh The general rule for calculating the effects of a variation is to compare the differential equation with and without the variation and to derive a equation for the difference. Present values calculation The continuous version for present value for a flow ƒ(t) is ∞
L ( f ) = ∫ f ( s ) e − as ds
Assuming zero initial condition for Aq , Al and applying to the differential equations, we obtain
L ( Aq ) =
L ( v∆h ) 3v a+ q ph
3vq 3vl L ( Al ) = a + a + 2 p ( h + hs ) ph
(1 − e ) + 3v Av 0 ar
L ( v∆h ) = vh
q
a
qr
ph
e0 ar − e0 aT a
h q vq 3vq hl vl − ch L ( v∆h ) + v= 3vq ph 3vq v 3 1 a + a + ph a + 2 p h + h ph ( s )
0
Let η q , η l , c h , the net revenues per ton and the developing cost per square meter the net flow is
u ( t ) = h q vq Aq + hl vl Al − ch v∆h The present values is
V = h q vq L ( Aq ) + hl vl L ( Al ) − ch L ( v∆h ) vh 3v q tr < t < tT v∆h = ph 0 t > tT We recall that Aq , Al satisfy
vA A& q = vh − q ph 3vq Aq 3vl Al − A&l = 2 p ( h + hs ) ph If we denote ƒ the flow derivative then L (ƒ) = ƒ(0) + αL(ƒ)
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Since all the relations are linear the present value variation is the present value of the variation, for example to evaluate the impact of decreasing de undercutting rate before the steady state we only need to calculate
0 dv∆h dv∆h = 0 3vv A vh − q qr ph 0
t ≤ ti ti < t ≤ ti +∆t ti + ∆t < t ≤ tr t r < t ≤ t r + dt r t > t r + dt r
3v A e − oar − e − a(tr +dtr ) + vh − q qr ph a hv 3v hl vl − ch L (dv∆h ) dv = q q + q 3vq ph 3v 3v1 a + a + q a + ph 2 p ( h + hs ) ph L (dv∆h ) = dv∆h
e
oai
− a ( t1 + ∆t )
−e a
CONCLUSIONS The continuous models for a caving exploitation allow to build explicit formulae for the evolution of mining parameters and their variations, additionally explicit formulae for the present values were built. This way the economical evaluation is simplified and the analysis ability increased.
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Propagation of a caving zone, A case Study from PT Freeport, Indonesia T. Szwedzicki, E. Widijanto, F. Sinaga, PTFI Freeport Indonesia
Abstract Time Domain Reflectometry was used to delineate changes in size, shape, cave ratio and to define a caving rate of progressive caving over the Deep Ore Zone Mine at PT Freeport Indonesia. The mine was designed to block cave 600 m with subsequent caving to the surface making the total caving height in excess of 1200 m. TDR cables were monitored and analyzed against rock mass properties, undercut sequence, production rate and total production. It was found that the cave rate was related to the rock mass rating (RMR) and on average varied from 0.15 to 1.10 meters per day with an average for three years being 0.6 meter per day. From a time and space distribution of breakage it was inferred that caving took place in cycles with relatively large caves-in randomly progressing in isolated areas. The period between cave-ins depended on rock mass properties and varied from one to six weeks. It was also found that large scale rock mass fracturing took place periodically forming a dilation zone around the cave. Fracturing extending from tens to hundreds of meters took place in more than six months intervals.
1 INTRODUCTION PT Freeport Indonesia block caving mines are located within one geological complex and are divided into three vertically stacked ore bodies which have been mined by Gunung Bijih Timur (GBT) Mine, Intermediate Ore Zone (IOZ) Mine and Deep Ore Zone (DOZ) Mine (Fig. 1). The GBT Mine was in operation from 1980 to 1993 and produced 60 million tonnes of ore. The GBT cave (300 m high) reached the surface in 1986 (Barber et al, 2000). The IOZ mine started in 1994 and ceased operations in 2003 producing over 50 million tones of ore. The IOZ cave superimposed on the GBT cave. The vertical distance between production level of the GBT and IOZ mines was 200 m. The DOZ mine started in 2000 and by the end of 2003 produced 22 millions tones. The total amount of ore mined from the three mines exceeded 132 million tones. The area of the DOZ undercut was 79,300 m2 and the perimeter of the undercut was 1.6 km (January 2004). The designed height of draw of DOZ mine was over 600 m with the total cave height from the production level to the surface being in excess of 1200 m. The shape of the DOZ mine caving zone was ellipsoidal with the short axis being 200 m and 400 m (width and length of the undercut) and the long axis being 600 m (height of draw). The DOZ cave zone merged with the GBT and IOZ cave and reflected on the surface in 2003. Fig. 2 shows surface reflection of the GBT, IOZ and DOZ block caving mines in January 2004. Caving reflection on the surface termed subsidence area was more than 1 km wide and 1.2 km long with the total area estimated to be 1.2 km2. Area of caving influence i.e. within a crack line limit was 1.7 km2 (January 2004). Caving and subsidence monitoring using TDR cables was carried to: • define a draw rate • determine an air gap between broken ore and the crown pillar above, and to • determine the shape of the caving zone, caving rate and the effect of caving on rock mass fracturing. The shape of the caving zone was crucial from a draw control point of view. The caving rate and rock mass fracturing was important to determine the effect of caving on surface facilities such as main exhaust fans and access roads. 508
Fig. 1. PTFI’s Underground Mines Complex
Fig. 2 Subsidence area over block caving mines
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2 GEOTECHNICAL CONDITIONS AT THE DOZ MINE The Deep Ore Zone (DOZ) mine is situated within the Ertsberg East Skarn system (EESS) and consists of skarn assemblages locally intruded by variably altered Ertsberg Diorite. Ertsberg diorite forms the footwall with forsterite skarn, magnetite-forsterite skarn, magnetite and DOZ Breccia and marble in the hangingwall (Coutts et al, 1999). Along the footwall the main diorite body was intruded by the skarn producing local alternation which resulted in poor ground conditions. The DOZ Breccia zone forms a lenticular zone that can be traced continuously across the western half of the DOZ mining block. The zone contains both diorite and skarn fragments within clay-carbonate matrix. Ground conditions within the EESS system are highly variable. Within good to very good ground conditions there are elongated zones of very poor ground conditions characterized by low strength, low core recovery and low RQD values. The values of the Uniaxial Compressive Strength (UCS) and Rock Quality Designation (RQD) together with Rock Mass Rating (RMR) classification are given in Tab. 1.
measure a location of a break along the cable. When caving intersects the cable, the cable forms an open circuit modifying the internal impedance of the cable. Consequently the breaking point can be identified in a space by reading the distance from the collar to the intersected location. The accuracy of the reading is estimated between 5 to 10 m. TDR cables were installed on the DOZ undercut level (1200 m below the surface), the IOZ production level (900 m below the surface) and from the surface (Rachmad, 2002). At IOZ and DOZ, 33 and 70 cables were installed respectively, making the total length of monitored cables to over 14,000 m. At the DOZ TDR holes were drilled up dip at an angle varying from 10 to 30 degrees. At the IOZ TDR holes were drilled sub horizontally, with the dip angle varying from 10 to 30 degrees. Fig. 3.
Tab. 1. Geotechnical classification of major rock types, Deep Ore Zone Mine. Rock type
UCS [MPa]
RQD [%]
RMR
Forsterite Skarn
127
84
Very good
Forsterite - Magnetite Skarn
56
67
Fair
Magnetite Skarn
97
71
Good
Diorite
111
80
Good
Breccia 2
22
40
Very poor
Breccia 4
41
45
Poor
Marble - Sandstone
22
65
Poor
Caving and fragmentation is a function of geotechnical properties. The fragmentation distribution for the East side of DOZ is shown in table 2. Table 2. Average fragmentation in the DOZ east Block size (m)
Percentage (%)
Cumulative percentage (%)
2.00
17.4
100.0
Fig. 3. 3-D view of the TDR cables located around the DOZ cave zone In each location, the TDR cables were installed in an array of 8 – 15 stations in a fan like manner to cover a large area so that the cave progression could be determined, Fig. 4. On the DOZ undercut level the TDR cables monitored cave initiation and horizontal progression of caving and fracturing. On the IOZ production level the TDR monitored vertical and horizontal propagation of the DOZ caving zone. The TDRs installed from the surface aimed at assessment of vertical propagation of the cave zone. However, till January 2004, the caving zone did not progressed far enough to the West to be detected by the cables installed from the surface.
The rock mass across the DOZ layout consists of diorite (brittle, massive and competent) rock at the south part of the layout, breccia (plastic and weak rock) in the middle of the layout, and marble (friable, low strength) at the north side of the layout. 3 TDR MONITORING Time Domain Reflectometry (TDR) monitoring is widely used in mining especially for monitoring of cave propagation (O’Connor et al, 1994). Monitoring of TDRs breakage provides information on timing and volume of the rock mass affected by cave propagation either by fracturing or caving. TDRs consist of coaxial cables that allow to identify and Massmin 2004
Fig. 4. Plan view of the TDR cables location around the DOZ mine caving zone
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4 PROGRESSION OF CAVING For analysis of the obtained results three series of TDR cables were selected. Two series were installed at the bottom part of the cave in the south and north part of the ore body and the third one at 330 m above, at the IOZ production level. Results obtained for a period of over two years were analyzed i.e. from the first detected TDR breaks surrounding the DOZ mine until the cave back propagated to the surface. Results of TDR breaks were calculated as the absolute length of cables, their horizontal component and their vertical component. The vertical component of the breaks of the cables installed at the IOZ mine were used to determine propagation of the cave to the surface. The horizontal component of the breaks of the cables installed at the DOZ mine were used to determine lateral cave propagation. The absolute length of the breaks were used to determine the extend of individual collapses of ground. 4.1 Vertical progression of the cave Nine out of the 33 cables installed at IOZ were used to estimate the vertical cave progression of DOZ. The original length of the cables varied from 290 m to 470 m. The absolute length of broken off cables varied from 10 m to 120 m. Out of 21 measured cable breaks, four were longer that 60 m and it could be assumed that these breaks represented rock mass fracturing i.e. formation of a dilation zone. Four cables broke at the length 25 to 60 m. The remaining 13 breaks were less that 20 m.
of the mine with the purpose of monitoring the effect of the advanced undercut sequence on weak and plastic breccia rock.
Fig. 6. TDRs break vs. monthly production at the North side of a caving zone 4.2.1 Horizontal progression, South Side Thirteen out TDR cables that were installed at the south side of the DOZ undercut level reported breaks. In total 28 breaks were reported and analyzed. Fig. 7 shows the absolute length of broken cables in time. There were 17 breaks shorter than 30 m, four between 30 m and 60 m and seven longer than 60 m. The average length of the broken cables was about 36 m.
Fig. 5. Periodical break of the TDR cables positioned over the progressive cave zone During caving cycles the TDR cables were broken in series over a period from two to four months. Fig. 5 indicates that the caving process was cyclic. The periods of caving activities were intertwined by periods of stability, lasting from a month to over two months. Values of the vertical components of the length of broken cables vs. monthly production of the panels underneath the progressive cave are shown in Fig. 6. It can be concluded that it was not the production rate which controlled cave progression but rather the total volume of removed rock. 4.2 Horizontal progression Two sets of TDRs cables were installed on the undercut level of the DOZ East. The first set was installed at the South side of the layout with the purpose of monitoring the effect of the advanced undercut sequence on competent diorite rock. The second set was installed at the North side 510
Fig. 7. Periodical break of the TDR cables at the South side of the DOZ Mine. TDRs breakage progressed cyclically. The cycles lasted from a day to about a month, with periods between the breakages being about three weeks. Values of the horizontal component of the length of broken cables and position of the advancing cave line vs. monthly production from the drawpoint in the immediate vicinity are shown in Fig. 8.
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The intensity (the average period between breaks of the cables) and the length of the breaks decreased with time. There was an indication that caving progressed to the east of the ore body (in the direction of the undercut). After undercut passed the over the cables, the cables did not indicate any ground movement or fracturing. Production rate from the draw points on the north side of the ore body was kept steady at the same level and caving was progressively induced, Fig 10.
Fig. 8. TDR breaks and advancement of the undercut line vs. monthly production at the south side of the caving zone
The breakage indicating rock mass fracturing took place in front of the approaching caving face. 4.2.2 Horizontal progression, North Side Eighteen cables were installed at the undercut level at the North part in breccia zone (weak and plastic rock) of the DOZ mine. The length of cables varied from 120 to 160 m. The cables faced the caving front approaching from the southwest. Fig. 9 shows the absolute length of broken off cables in time. In total 20 breaks were recorded. Two cables broke off at a distance of 50 – 60 m indicating the large-scale movement possibly across faults or shear zones and 18 breaks shorter than 30 m with an average length of a broken off cable being 15 m. Fig. 9 shows the absolute length of the breaks in time. The cables broke cyclically, as at the South side with cycles of ground movement / caving lasting for a period of four weeks. Periods of rock mass stability varied from two to six weeks.
Fig. 10. TDR breaks and undercut line vs. monthly production at north side of DOZ mine
5 ROCK FRACTURING AROUND THE CAVE ZONE Fig. 8. and Fig. 10. show the position of the undercut in relation of to the breaks of the cables. Analysis indicated that some of the TDRs installed at the sides of the DOZ mine broke in front of the undercut line and some behind. Breaks located in front of the undercut line indicated rock mass fracturing in advance of the created void. As the result of that fracturing a dilation zone was formed. Breaks located behind the undercut line indicated rock mass disintegration by caving Rock mass fracturing started immediately over the undercutting front and advanced in front up to 50 m. In a few cases the final cracks / fractures were established in line with the undercut line. However the most of TDRs revealed that caving behind the advancing face still took place after stoping undercut and the cave line was "catching up" with the undercut line. After the undercut lines reached their limits and advance stopped, some stress adjustments in the rock mass still took place with formation of new fractures continued up to six months, Fig. 10. In a number cases, a few TDR cables broke at the same time. With a wide coverage of the cables and large length of breakages, it can be assumed that the volume of cyclically fractured rock mass might have exceeded million of tones. In a case of caving the volume could easily exceed tens of thousands tones. 6 CAVE RATIO
Fig. 9. Periodical break of the TDR cables at the North side of the DOZ Mine. Massmin 2004
Cave ratio is defined as the ratio of the cave back height and the height of draw. The cave rate was estimated based on TDR breaks and production data. The result proved that the cave ratio was not constant but increases as the cave propagates.
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Fig. 11. indicates that for the cave height less than 200 meters, the range of the cave ratio is between 1 and 4, and at the cave height more than 200 meters the range is between 4 and 6. Cave growth or cave rate dictates the production rate. When caving slows or stops the production must slow down or stop. It was found that cave rate was related to the class of rock mass (RMR): • for marble cave rate was from 0.25 to 1.10 m per day • for magnetite skarn was from 0.15 to 0.95 m per day, and • for forterite skarn was from 0.08 to 0.30 m per day.
depended on rock mass properties and mining activities and in some cases lasted for a few weeks. The periods of caving were intertwined by periods of rock mass stability. In poor ground conditions an average length of broke cables was 15 m and frequency of caving was higher. In good to very good ground conditions the average length of broken cables was 36 m and the frequency of caving was sparser. The results of TDR monitoring proves that most of the caving happens in relatively large fall of ground. The volume of cave in rock as well as time intervals between fall of rock did not depend on a production rate but rather on the total volume of production. Long cable breaks indicated rock mass fracturing and formation of a dilation zone rather than caving. Cables installed at the upper part of the mine indicated that the dilation zone immediately above the void could reach 120 m. The TDR cables installed at the perimeter of the cave zone indicated that fracturing of the rock mass could take place in front of progressing undercut to the distance of 50 m. Fracturing in some cases continued even after undercut stopped progressing. Propagation of rock mass fragmentation and caving above the undercut depends on rock mass properties. The calculated average cave rate (for a period of three years) was 0.6 m per day. The TDR cables installed in different geotechnical domains showed that the rate varied from 0.08 m very good ground conditions to 1.1 m in poor ground conditions. 8 REFERENCES
Fig. 11. Change in cave ratio with caving height For breccia and diorite cave rate was not determined. For a period of three years, starting from the first cave after reaching the hydraulic radius, the cave propagated vertically for estimated 645 m i.e. the average cave rate was 0.6 m per day 7 CONCLUSIONS An analysis of breaks of the TDR cables proved that caving took in large periodical falls of ground at various parts of the mined out void. Periods of active caving
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• Barber J, Thomas L, Casten T, 2000, Freeport Indonesia’s Deep Ore Zone Mine. Proc. AusIMM MassMin 2000 Conf. Brisbane, 29 Oct – 2 Nov. • Coutts BP, Susanto H, Belluz N, Flint D, Edwards A, 1999, Geology of the Deep Ore Zone, Ertsberg East Skarn System, Irian Jaya. Proc. AusIMM PACRIM Conference, October 10-13. • O’Connor KM, Wade LV, 1994, Application of Time Domain Reflectometry in the Mining Industry. Proc. Symp on Time Domain Reflectometry in Environmental, Infrastructure, and Mining Applications. Northwestern University, Illinois, September 17-19. • Rachmad L. Sulaeman A. Cave Management practices at PT Freeport Indonesia’s block caving mine. Proc. NARMS-TAC, (ed Hammah at al). Balkema, 2002
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Implementation of cave management system (CMS) tools at the Freeport DOZ Mine Eddy Samosir, Project Engineer, Strategic Planning, Freeport-McMoRan Inc Charles Brannon, Manager, Strategic Planning, Freeport-McMoRan Inc Tony Diering, Principal Consultant, Gemcom Software International Inc
Abstract A daily draw order system called CMS has been successfully implemented at the DOZ mine to assist with the cave management process and to help manage the daily data generation and storage requirements. An effective interface between CMS and the Modular Dispatch system used for LHD dispatch has been developed. Monthly cumulative figures are used to deplete the draw column reserves, providing the ability for long-term plans to be adapted accordingly. Management of the cave shape requires effective control of the ore pulled from the drawpoints. This depends on the rock type, the drawpoint condition, the height of draw (HOD), the cave profile, the water content, and many other constraints. The overall process requires effective underground monitoring of drawpoint status, fragmentation, wet muck draw points and interaction between different panels. Related forecasting tools that have been implemented include PCBC applications that produce schedules of ore grades, fragmentation, and rock type distribution.
1 INTRODUCTION The Deep Ore Zone (DOZ) Mine is in the Ertsberg Mining District in Papua, Indonesia. The operation is run by P.T. Freeport Indonesia (PTFI) under contract to the Republic of Indonesia. The PTFI project site is located approximately 4o-6'S latitude, 137o-7'E longitude (Figure 1), in the Sudirman Mountain range of Papua, the eastern most province of Indonesia which occupies the western half of the island of New Guinea. The ore deposits, discovered in 1936 and then acquired and developed by PTFI beginning in 1967, are located approximately 96 kilometers north from the southwest coast, between elevations of 2900m and 4000m above sea level. Access to the project is through the PTFI portsite of Amamapare on the Tipoeka River, and from the international airport of Timika, some 43 kilometers north Amamapare. The mine site is 118 kilometers from Amamapare. An access road to the mine project site connects the portsite to the mill, passing by the Timika airport en route. DOZ is a copper-gold skarn deposit located on the northeast flank of the Ertsberg diorite intrusive body. It comprises the lower elevations of the East Ertsberg Skarn System (EESS). The EESS outcropped on surface at about
Figure 1. Location of PTFI’s mining operations. Massmin 2004
4000 meters, and the DOZ lift of the EESS is located on the 3100 meter level. Current operations in the district include the Grasberg open pit (200,000 tpd ore) and the DOZ block cave mine (40,000 tpd). The DOZ mine is a mechanized block caving operation. The eight cubic yard loader is utilized in DOZ production. In addition, the oreflow system, due to the coarse nature of the fragmentation, utilizes truck loading from chutes filled from ore passes from the extraction level, and direct dumped into a 54 inch gyratory crusher. The DOZ is the third lift of the block cave mine that has exploited the East Ertsberg Skarn complex since 1980, and design and operation has benefited from the previous experience gained while mining the upper lift (GBT) and the intermediate lift (IOZ). There are four main levels at the DOZ mine, from top to bottom they are; undercut level, extraction level, exhaust level, and the truck haulage level. An advanced undercutting system is employed at DOZ. 2 CAVE MANAGEMENT SYSTEM (CMS) 2.1 Draw Control at DOZ Draw control at the block cave operations has become progressively more sophisticated over time as the operation has proceeded through the successive lifts of the mine. Effective draw control has many goals: • minimize the dilution • prolong the drawpoint life • control convergence at a safe level • control water influx and wet muck • maximize ore recovery Today a system is in place at the DOZ mine that addresses these goals by fully integrating production forecasting, draw control, cave management, draw order, and compliance to the draw. The Cave Management System (CMS) is a sub-system of the PC-BC© block cave planning software. The end results from CMS include the daily draw orders, which are exported directly to the dispatch system which in turn interfaces with the LHDs. The LHD operator is able to read the draw order for the current
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drawpoint from the Modular Dispatch system (Prasetyo, et al, 2004). CMS thus provides a seamless two-way connection. The following sections will discuss the procedures for implementing the CMS system. 2.2 PC-BC to CMS PC-BC is a program used by Freeport for studies and activities ranging from pre-feasibility through to detailed daily draw control (Diering, 2000). The ultimate basis for the daily draw order is the production forecasts generated in PC-BC. The production forecast schedule is run in PC-BC in "catch-up" mode. The catch-up concept provides a mechanism for the draw order to be adjusted so that the actual tonnes can be brought back in line with the longer term target cumulative tonnages. One of the PC-BC main objectives with catch-up mode is to smooth the cave back profile from month to month. The benefit of the smoother cave back is to avoid a premature ground water inflow or dilution due to a cave spike or "chimneying" created. The cumulative tonnes to date, the latest active drawpoints, and the opening sequence are the input into this run to create a medium term plan. The long term planner will run PC-BC and generate a production schedule in which the monthly target tonnage is generated using the "catch-up" mode. The target tonnes are stored in a SQL database table for shared access by PC-BC and CMS users. CMS users will then retrieve the target tonnes bucket for developing the draw orders. Figure 2 is a simple diagram showing the work flow for getting a new target tonnage for CMS from PC-BC at month end. A monthly draw order is calculated at the beginning of each month. The order is based upon the total tonnage requirements but also considers a drawpoint priority which
Figure 2: PC-BC to CMS interaction is in turn based on how long a drawpoint has been idle, the percentage drawn for the month, allowable draw rate and draw point grade, and geotechnical input. The draw order emphasizes an even draw and respects the overall cave development strategy and long term plan. 2.3 CMS Data Process At the beginning of the each month, the CMS is set up with the actual historical data (i.e. the monthly development tonnes and grades through the previous month). The daily actual tonnage per drawpoint is downloaded from the Dispatch system and from regular LHD reports of
Figure 3: CMS data processing flowsheet. 514
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Figure 4: Production Rate Curve adjustment due to stopped temporarily condition
the development and undercut tonnage. The Modular Dispatch system will record all of the data from the loader at the drawpoint, while the other tonnage will be reported manually by the operator for production from undercut muck or development muck to the CMS operator for input into the Modular system manually (Prasetyo, et al, 2004). The sum of all these tonnages together is calculated as the actual daily tonnage. The CMS operator also inputs the daily drawpoint assay data as an actual grade into the CMS which allows for the assay to be compared against the model grade in the PCBC "slice file". The slice file is essentially a block model of the deposit but with tonnes and grade allocated to individual drawpoints in slices, as opposed to blocks in a model. The slice file grades are derived (per drawpoint) from the geological block model. Another input into the CMS is the drawpoint status. This status is one of many CMS parameters used to determine the buckets order per drawpoint. These parameters are shown in Figure 3 below. Beside the inputs, as shown above, within CMS there are other inputs that need to be completed while running through the process, such as current date, total order for the day, etc. Sometimes the daily order will vary from the average remaining tonnage of the month, for instance on crusher preventive maintenance (PM) schedule, oreflow system PM schedule, production demands, etc. For these particular types of cases, a "manual tonnage adjustment" is utilized that allows CMS to force a higher or lower daily draw order than its calculated remaining tonnage. 2.4 The Principal of the Daily Order Calculation The basic driver for the draw order calculation is the required tonnage target for the month. If the drawpoint is hung up for a period of time, or if another temporary production interruption occurs, the production rate curve (PRC, inches/day) will be adjusted for the subsequent daily order. The conceptual diagram in Figure 4 illustrates the philosophy. After a period of time in which a drawpoint is behind target production, the PRC is adjusted by CMS, within limits, to allow the drawpoint to catch up to the monthly target. Massmin 2004
2.5 Drawpoint Priority When generating the daily draw order, drawpoints are categorized into groups of descending importance or priority as follows: • Wet Muck The highest drawpoints priorities to be pulled are those that are wet. Wet drawpoints must be pulled in order to minimize wet mud rush hazards by continually removing water form the caved muck, and to minimize spreading of the wet areas to additional drawpoints. The daily or monthly target for the active wet muck drawpoints is set up before start of the month. But for the new wet muck drawpoint, the daily or monthly target needs to be adjusted as soon as it is classified as wet. • Convergence/Geotechnical Stability The next highest priority drawpoints are those related to the panel drift convergence. Unfavorable convergence is often related to the mucking history. By prioritizing the mucking plan, the stress is more evenly distributed and convergence mitigated. As with the wet muck drawpoints, the unfavorable convergence drawpoints must be mucked out regardless of whether month-to-date tonnage has been achieved or not. • Drawbell Newly blasted drawbells are pulled hard until production reaches the first 6,000 tonnes, the theoretical blasted quantity of material in the drawbell. This is the next priority to provide enough space to propagate the cave. • Underpull Drawpoints The PRC for the underpull drawpoints need to be adjusted from day to day in the draw order to catch the tonnage target for the month. • Normal Drawpoint These will have a regular order limited by daily maximum PRC. • Ovepull Drawpoint The overpulled drawpoints, those where the month to date production is above the plan, in the normal condition
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Figure 5: Drawpoint type based on CMS priority hierarchy
will typically be ordered as idle. These drawpoints can be pulled only if there is insufficient tonnage available from other drawpoints due to hung up or other temporary stop condition. Figure 5 below shows the drawpoint priority hierarchy for the conditions described above. 2.6 Output CMS produces a daily draw order, as well as tabulations of compliance to the draw, drawpoint status, comparisons of actual grade versus model grade, etc. Figure 6 shows an
example of part of one daily draw order summary. The output highlights drawpoints that may require attention by the operations group, such as drawpoints that have been idle too long, newly hung-up, newly released, etc. When CMS is run, an ascii data file of the draw order is created automatically. Then with one "import" command this file is imported directly into the Modular system and ready to use as a daily draw order. This daily draw order will be available at the LHD’s using the Dispatch System.
DOZ-Draw Order (3 Shifts) Panel
Bucket
WMT
DMT
%Cu
ppm Au
P26 P27 Sub-total Develop Total
144 27 4.269 4,269
16.27 305 48.240 2,986 51,225
1.557 292 46.162 2,857 49,019
0.82 1.07 1.15 0.80 1.13
0.62 0.45 0.74 0.30 0.71
Description Dry Coarse Dry Fine Wet Coarse Very Wet Coarse Draw Point DrawBell Wet Fine Very Wet Fine Overpull Special Order Total Active DPTS
Status
Description
A B C D D/P DB E F OP SP
DP_id require action Temp closed draw points-Idle too long P13-04W P13-06W P16-02E
214 5 0 2 0 2 0 0 0 18 241
Hang up (A) Temp Close (A) Hang up (B) Temp Close (B) Close Permanent Temp Close (D) Temp Close (E) Temp Close (F) Temp Close (C) Hang up (C) Total Close DPTS
Status A/H A/T B/H B/T CP D/T E/T F/T C/T C/H
0 1 0 0 18 1 0 0 1 0 21
Daya idle
%Drawn
Status
New Hangup
New released
16 7 20
0% 0% 0%
CP A/T CP
P21-07W P19-10W
P18-009E P19-09W P17-11E
Figure 6: Daily Draw order summary. 516
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Drawpoint closure tracking is another output. This analysis is based on the current height of draw (HOD) compare to its best HOD and the average grade for the month. If the actual grade for a month is consistently lower than the reserve grade, CMS will recommend the closure of the drawpoint. Final decisions on closure involve other criteria in addition to grade, such as geotechnical (convergence) considerations. Production tonnes and grade daily outputs of the actual and the reserve grade, results from CMS. The outputs are produced in various formats that can highlight individual drawpoints, panels, or mining areas. The month to date tonnage and historical grade information is also available from the CMS, and once finalized and reconciled becomes the official production data. 3 QUALITY CONTROL AND RESULTS 3.1 Field Inspection To ensure that the draw order is executed and reported well, a field inspection of the drawpoints is conducted daily. This field inspection takes place at the same time that the draw control crew collects drawpoint samples for assay. The muckpile in the drawpoint is marked with the date by spray paint when the draw control personnel take the grab samples. This is a mark reviewed on the following day by the draw control crew. The crew will make sure that there is no tonnage reported from the drawpoint if the mark is still in the drawpoint. If there is production reported, then the draw control supervisor will report this to the production supervisor and remove the recorded tonnage from the modular mining. The marked drawpoint also alerts the draw control crew for over-idle drawpoints so that they can be brought to the attention of the operations staff for action. 3.2 Compliance To evaluate the mucking performance to the daily order, the draw compliance is analyzed daily. The draw compliance is a degree of expectation of how close the actual tonnage is to the daily draw order. Compliance is calculated on individual drawpoint basis rather than the total tonnage. Perfect compliance would give a value of 100%. The formula below shows how the compliance is calculated: % Compliance = 1 - (S abs (order-actual)/S order) Where, % Compliance is degree of the expectation Σ absolute (order-actual) is sum of absolute order to actual Σ order is sum of the order There are other ways this calculation could be done, but experience has shown that this formula works well for DOZ. DOZ compliance shows a continual improvement over time since it has been used. Figure 7 below shows the historical compliance from January 2002 through December 2003 on a monthly basis. Initially, the compliance target was set to 60%. Gradually, the target was increased by 5% per month, up to the current 90% target. Some months had worse compliance compared to the previous month. Usually there was a reasonable explanation, i.e. unexpected loading point problems or unexpected longer hung-up period, etc. In the beginning the compliance implementation was not so easy. The habit of production personnel was to produce as much ore as possible, regardless of the draw order. Now, however, there is a better general understanding of the benefits of good compliance, so that the draw order process is well followed right through to the dispatch system. Massmin 2004
Figure 7: DOZ historical compliance to the draw order. 3.3 Results One of the objectives of the CMS is to maintain as consistent and smooth a cave back as possible. Figures 8 and 9 illustrate the monthly progress of the draw profiles for two more mature panels in the mine. Each curve represents successive months. Panel 16 East, for example (Figure 8), at the beginning drawpoint 3E had excessive draw, relative to adjacent drawpoints. As the draw order was controlled over the last 10 months, the cave profile gradually smoothed out and height of draw variances controlled. Panel 18 East (Figure 9) is similar, except that the draw profile was maintained smooth since initiation of mining and was well maintained. 4 SUMMARY The CMS was installed at the DOZ mine in late 2001. Since that time the overall draw compliance at DOZ has improved significantly. The man-hours required to generate the daily order is greatly reduced, and more importantly, the draw order and compliance to the draw is now more independent of individual operator preference. Finally, the real benefit of the CMS system is in the improved draw profile which will help to reduce dilution, improve draw point stability, maximize ore recovery, and greatly assist development of accurate production forecasts.
Figure 8: Panel 16 East cave height of draw over time.
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5 REFERENCES • Diering, T, 2000. PC-BC: A block cave design and draw control system. Proceedings MassMin 2000, Brisbane, pp. 469-484. • Diering, T, 2004. Combining long term scheduling and daily draw control for block cave mines. Proceedings MassMin 2004, Santiago. • Prasetyo, et. al. Use of the Modular Dispatch System to Control Production Operations at the DOZ Block Cave Mine. Proceedings MassMin 2004, Santiago
Figure 9: Panel 18 East cave height of draw over time.
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Development and measurement of the subsidence zone associated with SLC mining operations at Perseverance – WMC, Leinster Nickel Operations Duncan Tyler, Geotechnical Manager, Nickel Business Unit, Andrew Campbell, Geomechanics Engineer (Planning), Leinster Nickel Operations, Stephen Haywood, Senior Surveyor, Leinster Nickel Operations, WMC Resources Ltd, Western Australia
Abstract The Perseverance mine is located 15 km north of Leinster and 370 km north of Kalgoorlie, Western Australia. The mine produces approximately 40,000 tonnes of nickel per annum with the majority of the ore extracted from the sub level cave (SLC). All development and infrastructure is located in the hanging wall and current mining is at the 9715mRL (805m depth). The mine has ongoing reserves for a further 12-years of extraction, with extraction of the SLC scheduled to 9400m RL (1122m depth). The first signs of surface failure as a result of underground extraction were observed in December 1995. The ongoing use of the SLC mining method has lead to wide spread discontinuous subsidence and toppling failure in the area surrounding the previously completed open pit. The proximity of the mine’s surface infrastructure to the ore body makes it likely that almost all major surface infrastructures will ultimately be subjected to some level of cave related displacements. This paper discusses the observations made using the cave monitoring systems at Perseverance and discusses the potential impacts of surface subsidence on critical mine infrastructure.
1 INTRODUCTION The Perseverance mine is located 15km north of Leinster and 370km north of Kalgoorlie. The deposit comprises a disseminated nickel sulphide resource of approximately 45Mt nominally grading at 2% nickel and several smaller orebodies comprising narrow veins of massive nickel sulphide grading to 8%. Since 1994, the majority of the disseminated orebody has been mined by the sub-level caving method. The cave broke through to surface in 1995. This paper documents the progression of the cave front and the methods of monitoring employed. 2 MINING OPERATIONS The Perseverance orebody was discovered in the 1970s and was initially mined by the Agnew Mining Company between 1978 and 1986 when the mine was mothballed due to difficult ground conditions and a low nickel price. The lease was purchased by WMC Resources Ltd in 1989 and an open cut pit was established above the existing mine workings (Perseverance pit) whilst underground development and rehabilitation continued in preparation for full scale underground mining. The Perseverance pit was completed at a depth of 190m in 1995. Most of the nickel in the Perseverance resource is contained in the ultramafic hosted disseminated ore body which, apart from remnant mining around old stopes, WMC Resources has mined exclusively by sub-level cave (SLC) methods since 1995. However, the disseminated ore body at Perseverance has not been mined exclusively by means of SLC. Historically, several mining methods have been used, as shown in Table Massmin 2004
1. The majority of the critical infrastructure required for the operations was sited prior to the development of the Perseverence pit and the use of SLC mining methods. Table 1: Mining timeline for the Perseverance main disseminated ore body Date
Depth, m
Mining Method
1989 to 1995
0 to 190
Open Cut
1978 to 1986
190 to 375
Post Pillar (AMC)
June 94 to June 1998
375 to 500
Sublevel Cave
N/A
500 to 600
Un mined (temporary pillar)
June 97 to Present
600 to 805
Sublevel Cave
3 GEOLOGY AND GEOMECHANICS 3.1 Geology The Perseverance nickel deposit is situated in the Archaean Yilgarn Craton of Western Australia. The main disseminated ore body occurs within ultramafic komatiite and dunite rocks set in the intensively deformed eastern part of the Agnew-Wiluna greenstone belt. The nickel mineralisation occurs as massive and disseminated sulphides hosted by ultramafic-serpentinite lithologies (Barnes et al, 1988). The Perseverance deposit is bounded to the west by deformed felsic meta-sedimentsand to the east by barren ultramafics. The disseminated ore body is defined by a one per cent Ni grade boundary and is typically about 80m wide
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(east-west) and 150m long, see Figure 1. The dip of the disseminated ore body is sub-vertical towards the west with an inflection zone between the 10100 and 9900mRl’s (420m to 620m below surface) where the dip flattens to around 45∞ before steepening again below the 9900mRl (Figure 2). The ground surface is taken as 10520mRl.
3.2 Geomechanics The vertical stratigraphy can be summarised as 80m of sapprolite, a 40m transition then fresh rock below. The indicative UCS for the fresh meta-sediments is 150MPa and for the ultramafics 90MPa. Stress measurements have been undertaken down to 9400mRl using HI cells and hydraulic fracturing (AMC 2004). The results of the measurements within the hanging wall felsic metasediments can be summarised as: Sigma 1 = 0.0642z MPa, dip 22∞ towards 112∞ Sigma 2 = 0.0367z MPA, dip 28∞ towards 010∞ Sigma 3 = 0.0250z MPa, dip 53∞ towards 235∞ Z is the depth below original surface in metres. There are several major structural features close to and passing through the ore bodies at Perseverance. These include a proterozoic dolerite dyke that dips north at approximately 60∞, a gabbro dyke that passes between the Progress and 1A ore bodies, and a hanging wall sheared contact between the main disseminated ore bodies and the felsics, see Figure 1. Seven joints sets are identified in the felsic waste, three of which are combined to create rhombic blocks: J1 80∞ / 286∞ Foliation, occ faulted J2 73∞ / 200∞ Planar joints, occ sheared J3 22∞ / 196∞ Planar joints, often faulted
4 SURFACE CAVE CHARACTERISATION The surface deformation due to caving can be defined into three zones; the cave zone, the fracture zone and the stable zone. At Perseverance, each of these zones displays characteristic modes of failure as described below.
Figure 1: Geology plan on 9700mRl.
• Cave Zone Defined in terms of absolute deformation and rates of movement. The current cave front along the western margin is essentially defined by a 20m high scarp. Survey prisms within the cave zone have generally moved at a rate >25mm/year, with prisms moving at up to 1500mm/year. The interface with the fracture zone generally represents a discontinuity both physically and in terms of deformation rates. • Fracture Zone Defined in terms of absolute deformation and rates of movement. Cracks up to 50mm wide are evident but these are often discontinuous. Prisms within the fracture zone generally move at a rate of 10 to 20mm/year. • Onset of the stable zone Onset of the stable zone is defined in terms of rate of movement. GPS stations tend to be used and the rates are typically < 10mm/year.
5 CAVE MONITORING
Figure 2: Geology section looking north. 520
Since 2000, significant resources have been allocated to measuring and monitoring the surface expression of the cave and its effect on the surface/sub-surface infrastructure. This was driven by the acceleration of obvious surface movement towards critical items of infrastructure including the main shaft, F93 vent shaft and southern vent shaft (SVS). Prior to this only qualitative methods had been used to monitor the cave. The most enduring of these has been photogrammetry, which can be traced back to the early 1980s. A basic manual prism monitoring program was established along with surveying a number of points using Real Time Kinematic (RTK) GPS located around the pit. Santiago Chile, 22-25 August 2004
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Neither derived any useful data that could be used for future reference. Since 2000, rigorous monitoring programs have been implemented, including an extensive GPS program capable of recording true movements down to ±5mm and survey prisms with an accuracy of ±20mm. The types and effectiveness of monitoring systems currently in use at Perseverance are shown in Table 2 and discussed below.
Table 2: Appraisal of monitoring systems Element
Strengths
Limitations
Photography
Low CapEx./OpEx. Quick, easy, instantaneous record
Non quantitative Image only
Surface mapping
Low CapEx/OpEx. Quantitative, details detected.
Time consuming Interpretational
GPS
High precision & accuracy in X,Y & Z planes. Millimeter detection. Early warning
High CapEx./ OpEx. Extensive setup required – Long term. Intensive monitoring
Survey prisms
Mod CapEx./OpEx. Program easy to setup - fully automated Centimeter detection
Limited Accuracy Atmospheric conditions
Inclinometers
Low Cap Ex./Op Ex. Continuous data with depth
Reading error Discrete X, Y points
Aerial Photogrammetry
Low CapEx. Cover large areas quickly
High OpEx. Limited resolution (quoted ±200mm)
Laser & tape extensometer
Low CapEx./OpEx. High accuracy
Limited coverage Localised result
Precise leveling
High CapEx. Early warning Sub millimeter detection
Low OpEx. Intensive monitoring Weather variant
5.1 Qualitative surface & underground monitoring • Photography Photographs have been taken progressively of the open pit, starting in 1995. They are taken from specific locations on a routine basis. No effort has been made to ortho-scale them. • Surface Mapping Since July 2000 the surface effects of caving have been mapped on a six monthly basis. New or continuing cracks are marked and surveyed using RTK – GPS and added to a composite map. • Underground Mapping Since June 2000 the effects of the cave on 3 Level (3L) infrastructure (10030mRl) has been mapped on an approximately annual basis. This is recorded on a composite underground level map.
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5.2 Quantitative ground monitoring • GPS A series of sixty six GPS stations are strategically placed outside the cave zone in a series of nominally radial lines. Stations are also concentrated around the SVS. The GPS points are used to monitor deformation up to 100mm. A high resolution Leica SR530 GPS receiver and AT 503 "Choke ring" antenna combination is used to record point positions. Each point consists of a concreted solid stainless steel shaft with a machined 5/8th inch thread to force center the antenna. Two control base stations located 4km apart enable dual differential solution for the roaming rover unit. Each reading takes approximately 30 minutes, after which the receiver is moved to the next station. The point accuracy obtained using this method has been found to be in the order of ±5mm in the horizontal plane and ±10mm in the vertical plane. • Prisms Monitoring Sixty one prisms are located within the western and southern margins of the cave and fracture zones. These are surveyed on a daily basis, via an automated Leica Total Station TCRA1100. The instrument is linked via radio, to a PC based software program (QuickslopeTM) located at the survey office several kilometers away from the instrument station. The system has an average sighting error of ±20mm in both the horizontal and vertical planes, with a distance error of ±2mm. The prisms are positioned in areas where displacements are expected to be >100mm/yr, many are now well within the cave zone. All readings automatically have reference corrections applied to them via the software for atmospherics. • Bi-axial inclinometer A single 80m long bi-axial inclinometer was comissioned near to the SVS in January 2003. This instrument is read on a monthly basis. • Aerial Photogrammetry This has been undertaken 6 times since the cave broke through: 1995, 1998, 2001, 2002, 2003 and 2004. Initially fly-overs were not undertaken with measurement of subsidence in mind and the resolution on contour plots was 500mm. Since 2001, flyovers allow subsidence resolutions of 200mm. This information is otho-scaled and digital terrain models created. 5.3 Quantitative surface infrastructure moniting • Laser & tape extensometer These methods have been used to monitor differential foundation movements. • Precise leveling This is conducted on all the critical infrastructure within the potential region of the fracture zone. This includes the main shaft, F93 vent shaft and SVS. A Leica NA3003 precise level, capable of measuring to ±0.01mm is used in conjunction with fixed invar strips and a 4m bar-coded staff. Individual traverses are performed around each of the structures to detect settlement. A linking traverse connects each of the structures to a dual benchmark (common reference point). The quality of results allows detection of sub millimeter movements. • Multi point extensometers Wire extensometers have been installed in deep boreholes to monitor key strucures. 5.4 Quantitative underground monitoring • Multi point extensometers Rod and wire borehole extensometers have been installed at various levels in the hanging wall to monitor the cave development.
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Figure 3: Surface cave history with mined footprint to 1997. 522
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Figure 4: Surface cave history with mined footprint 1998 to 2003. Massmin 2004
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Figure 5: Prism monitoring data, prism 17 6 CAVE HISTORY 6.1 Surface The history of SLC mining and cave propogation at the surface is summarised in Figures 3 and 4. These plans are based on interpretation of air photographs, walk-over surveys and GPS / prism data. SLC mining commenced on the 10145mRl in August 1994 and the cave broke through into the floor of the pit in 1995. The initial cave break-through was closely related to the mined foot print of 1994/95. In the following two years the cave zone moved rapidly to the south and west before extending north west. The fracture zone expanded ahead of the cave zone at a rapid rate. By 1998, a distinct nose remained in the northwest; this corresponded to an unstable wall feature in the pit that had been buttressed during pit mining. Interpreted expansion of the cave zone in 1999 prompted increased monitoring. Cave zone advance slowed significantly on the western margin after 2000 but acclerated to the south over the period 2001-04. The western limit of the fracture and cave zones have remained largely unchanged since 1999 and 2001 respectively. This is consistent with the westerly dip of the orebody steepening to near-vertical below 9920mRl (extraction completed in 1998) and increased mining towards the south. Mining is currently advancing on the 9740, 9715 and 9690mRl’s (780 to 830m depth). In general there appears to be a one to two year lag between mining levels and the development of associated surface damage. However, there is a direct relationship between mining and cave movement in the cave zone. During 2003 a shaft re-fit was carried out between 30 Sept and 31 Oct leading to a significant reduction in SLC production. During this period, movement as recorded by prisms in the cave zone almost ceased. This response to reduced mining was detected within 48-hours (see Figure 5). However, corresponding prism movement in the fracture zone was unaffected. Buttressing of the F93 shaft area has been undertaken since 2001 by tipping mine waste into the north western 524
quadrant of the pit. To date, approximately 1.4 Mt of waste has been tipped. The limit of the stable zone to the west of the pit has not been established conclusively. GPS stations over 500m from the pit show movements of 5mm/year towards the cave. Fine cracking of less than 2mm has been observed parallel to the cave front 300m to the west. To date the SVS inclinometer has not indicated any lateral movement. However, a GPS point directly on it indicates 15mm movement toward the cave to date. The discrepancy
Figure 6: Cave zone development on 3 Level.
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is attribuited to inclinometer casing failure at 80m depth, which is just above the fresh rock interface. The has been no evidence of wall instability on the east wall of the pit. 6.2 Underground Observation of caving underground is difficult as production levels are generally barricaded shortly after completion of mining. However, access is possible on 3L to a ventilation drive that was developed to the west of the cave zone in 2000, prior to the collapse of 3L access drive. Multi-point wire extensometers were installed from the vent drive to monitor cave development. Visual assessment can also be made on the north and south ends of the cave zone. The interpreted 3L cave zone is shown in Figure 6. Expansion of the cave zone over the period 2001-04 is between 10 and 25m. The least expansion has been to the west and is greatest at the northen and southern ends, consistent with surface cave deformations over that time.
GPS stations are routinely replaced with prisms once the total movement exceeds 100mm and/or when access for reading is deemed to be unsafe. The GPS stations are then relocated. New automated GPS stations are to be trialled with solar panel power supply and modem connection to the base station. Gyroscopic surveys of the shaft guide rails are proposed. Biennial reviews of the cave zone will be carried out by external consultents. 9 CONCLUSIONS
mRl
Factor
MRMR
Cave Angle
10520 – 10400
0
30
35°
10400 – 10360
0.2
37
52°
10360 – 10330
0.5
39
64°
10330 – 10200
3.2
42
80°
10200 – 10030
11.8
40
83.6°
• The implementation of robust surface and underground monitoring systems has allowed routine measurement of both the cave and fracture zone development associated with the Perseverance SLC. • Monitoring should ideally commence prior to development of the cave as the data gleaned from the systems is vital to allow for long term strategic planning. • Development and implementation of monitoring systems is time consuming and requires site champions to ensure their long term survival. • At Perseverance there is a one to two year lag between level completion and its impact on the surface. However, prisms in the cave zone indicate that there is negligible lag between SLC production and movement in this zone. This suggests elastic continuity between the surface and the production horizons through the caved material. • Cave and fracture zone development is not impacted by rainfall events. • The cave zone was defined by a Laubscher cone to the north and west in July 2002. This work will be reviewed and updated in May 2004. • At present, the main shaft is not expected to be impacted by the cave zone when SLC production reaches 9400mRl.
10030 – 9960
17.2
39
84.5°
ACKNOWLEDGEMENTS
9960 – 9900
22.9
37
85.2°
9900 – 9760
40.8
29
87.0°
9760 – 9715
47.5
29
87.5°
9715 – 9615
64.5
29
88.4°
The authors are grateful to all their colleagues at the Perseverance Mine who provided aid and assistance to this paper. The authors acknowledge the permission given by WMC’s Nickel Business Unit to publish this paper.
9615 – 9490
89.0
29
89.5°
REFERENCES
9490 - 9415
104.3
30
89.9
7 EMPIRICAL ANALYSIS & PREDICTION Review of Laubscher cave cones was undertaken by SRK Consulting in 2002 and shown is in Table 3. Table 3: Laubscher Cave Angles (after SRK, 2002)
The method was calibrated to both site monitoring data and cave observations for the west and northern cave zone. It is noted that the east wall of the pit is stable and that a line projected up along the east edge of the mined areas to the surface suggests that the cave angle is at least 75∞ in the ultramafic. The SRK report indicated that the main shaft will not fall inside the cave zone when scheduled SLC extraction to the 9400mRl is completed.
• AMC Consultants Pty Ltd, 2004. Rock Stress Measurement – 9590 Progress Incline. AMC report 103127, issued to WMC Resources January 2004. • Barnes, SJ, Gole, MJ, Hill, RET, 1988. The Agnew Nickel Deposit, Western Australia: Part I. Structure and Stratigrphy. Economic Geol., 1988 Vol 83., 524-536. • SRK Consulting, 2002. Assessment of possible subsidence zones – Perseverance SLC 6 to 11 level, Hanging wall inclusive. SRK report WM212, issued to WMC Resources July 2002.
8 FUTURE MONITORING & ANALYSIS Additional monitoring is planned for areas around the main shaft and the F93 vent shaft. This will consist of 200m deep inclinometers sited to the south east of each shaft.
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526
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Chapter 14
Rock Mass Pre - Conditioning
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Hydraulic fracturing used to precondition ore and reduce fragment size for block caving Enrique Chacón, Victor Barrera, El Salvador Division, Codelco Chile Rob Jeffrey, CSIRO Petroleum, Australia Andre van As, AVA Mine Geotechnical Services, Australia
Abstract A full-scale experiment was carried out in a block of primary ore from the 2600 level of Salvador mine to investigate the use of hydraulic fracturing to precondition the ore for caving. A reduction in the size of rock fragments produced during caving is the main concern in this orebody. Ten monitoring boreholes were drilled around a central injection hole at the site to allow fracture pressure, growth rate, opening and stress change to be measured. Seismic cross-hole velocity mapping was carried out before, during and after the hydraulic fracturing and two of the fractures were mined and mapped on the undercut level below the site. The site was extensively characterized by conducting overcore stress measurements and microfrac breakdown measurements. Rock property and geological data was obtained from the mine database. From direct measurement of fracture intersection with monitoring holes the fracture growth rate and orientation was determined. Stress change monitoring was used to measure the maximum stress change caused by the fractures and helped in orienting the fractures, which were sub-vertical. Microseismic events at 1 to 3 per second were recorded during the growth phase of the fractures and cross hole seismic measurements made before and after the treatments revealed a significant change in compressional wave velocity, corresponding to a change in rock strength.
1 INTRODUCTION The Salvador block caving copper mine, a division of Codelco, is located in the northern part of Chile, in the Atacama desert, about 800 km north of Santiago. Much of the ore extracted from the mine over past years has come from the oxidized zone of the porphyry copper orebody. However, much sulfide primary ore remains in the deposit and is of mineable grade, but the panel caving operations result in a significant amount of oversized rock blocks that, in turn, cause production delays and increase mining costs. Therefore, methods to precondition the primary ore in order to increase fragmentation during mining are being investigated by the mine. Salvador mine therefore offered the International Caving Study II a site on the 2600 level of the Inca east sector of the mine where the project described in this paper could be carried out. Figure. 1 shows a 3D view of part of the Salvador mine with the shaded cube corresponding to the experiment site. HF02 was the injection borehole and was drilled downward into the 100 x 100 x 100m block of ore beneath the 2600 level. A caving operation was approaching the site from the east, as indicated in Figure. 1. 2 DESCRIPTION OF SITE The injection borehole, HF02, was drilled to the east at 59 degrees dip. Around HF02, ten monitoring boreholes (also nominal 60 degree dip to the east) were drilled from this area of the mine previously used as an underground garage and workshop. A grid-work of shotcreted tunnels, about 5m x 5m in cross section, made up the garage and allowed easy access for the drilling and pumping equipment and for instrumentation activities. Figure. 2 contains a plan of the experimental area, showing the tunnels and collar positions of the boreholes used. Overcore stress measurement sites are also shown in this figure as OC-01 etc. Massmin 2004
Figure 1: Overview of site at Salvador mine. The contact between the oxidized and primary ore cuts across the site just south of overcore stress measurement holes OC3 and OC2. Stress measurements, made in the primary ore at OC4, were therefore relied on to site the boreholes. In addition, a vertical HQ-size hole was drilled (HF01 in Figure. 2) and, during March 2002, five hydraulic fractures were placed in it at depths of 27 to 49 m below the collar. This work was undertaken to test the operation of the new fracturing pump and to verify the orientation of the fractures formed so that HF02 and the monitoring holes drilled around it were correctly positioned. The fracturing work in HF01 produced fractures that intersected the garage tunnels 10 m to the east of HF01. Acoustic scans of HF01 showed nearly vertical fracture traces that left the hole with
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Figure 2: Plan view of site on 2600 level showing injection and monitor holes. a west dip. As a result, the monitoring and injection holes at the main experimental site were drilled dipping at 60 degrees to the east. This orientation allowed more consistent isolation of the fractures formed using inflatable straddle packers. Figure 3 shows one of the post-fracture acoustic scans of the injection borehole HF02. The trace of the hydraulic fracture on the borehole wall can be clearly seen. The fracture has initiated, because of stress concentrations around the hole, as an axial fracture and then rotated to align with the far-field stress. The strike of the fracture is north-south, based on the trace in the borehole and the fracture is turning from the borehole dip of 590 E to a dip of about 750 W (based on intersection data and stress change interpretations). As the fracture realigns itself with the stress field, a number of en echelon branches formed. These multiple fractures increase the fluid friction losses as the fracturing fluid is pumped through them, but they are expected to only extend for a short distance near the borehole. The initiation along the borehole axis also gives the fracture an S-shape for a few meters near the hole, which in turn means the two wings of the fracture are offset slightly above and below the depth of the straddle section. Development of such offsets results in less control of the exact placement of the fractures along the hole and the effective spacing between fractures may vary. Some fracture treatments may open existing natural fractures that are favorably oriented at the borehole. In these cases, the offset that is suggested by the trace in Figure 3 may be small or non existent. The fracture traces observed in HF02 were fairly consistent and would be expected to produce similar offsets from the straddle, implying that the effective fracture spacing away from the hole may not be severely affected by the offsetting. Modification of the straddle packers to allow the pressure to load the borehole axially and notching of the borehole ahead of fracturing could be implemented to better control the fracture initiation and effective spacing. 3 FRAGMENTATION Primary fragmentation is determined by the rock structure and strength interacting with the stress acting during caving (Brown, 2003). Hydraulic fracturing can modify all of these main controlling parameters. Hydraulic fracturing forms new fractures through the rock mass, connects and weakens existing natural fractures, and modifies the strength (by pressurizing and resulting shearing) of natural fractures and joints in the rock on either side of the main hydraulic fracture path. The hydraulic fractures extend as 530
Figure 3: Acoustic scan log of the fracture at 57.5m in HF02. single connected features for 30 to 50 m from the injection point, or for farther if higher injection rates or more fluid is used in each treatment. The in-situ stress field is changed by the fracturing, as reflected in part by microseismic activity, and by residual opening locked into the main fracture channel. Shearing of natural fractures will tend to reduce stress differences while the locked in deformation can increase the minimum principal stress and, eventually, lead to changes in the hydraulic fracture preferred growth direction. In the rock mass at Salvador, the mine through data collected and discussed later in this paper, clearly demonstrates that the hydraulic fractures cross most veins and filled joints and form new fractures through unbroken rock. Interactions are with weaker sheared joints which can be taken to form the primary block size in the ore before the introduction of hydraulic fractures. Therefore, the spacing between the hydraulic fractures should have a strong effect on final fragment. 4 HYDRAULIC FRACTURING Hydraulic fractures are initiated in a borehole by pressurizing an isolated section of the hole. The fracturing fluid, which can be water or a thicker water-based gel, is injected into a straddle section between two inflatable packers. The pressure in the hole induces tensile stresses along the hole axis that eventually exceeds the tensile strength of the rock, initiating a fracture there. This fracture is extended by the fluid entering and pressurizing it. Alternatively, natural fractures that intersect the borehole may be opened by the fluid injection. In some cases the hydraulic fracture may follow along a succession of natural fractures, extending them and linking them up into a continuous fracture plane. In either case, the hydraulic fracture tends to grow in the plane of the maximum and intermediate in-situ principal stresses and to open against the minimum principal stress. The pressure to extend the fracture must exceed the minimum principal stress magnitude. In addition, an excess
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pressure above the minimum stress magnitude must exist that produces fracture opening and stress concentration at the leading edge of the fracture that are sufficient to break the rock as the fracture extends. 3.1 Fracture Treatments An electric-driven triplex pump capable of pumping 215 liters per minute at up to 35 MPa pressure was used to produce the fractures at Salvador. The injection string consisted of BQ drill rods with an inflatable open-hole straddle packer used to isolate 0.5 m long sections of the hole for each treatment. Fractures were initiated in borehole HF02 at depths down the borehole of from 50 to 70 m below the collar. The two fractures that were mapped during mining were placed 112 and 118m below the collar. The treatments produced fracture intersections with monitoring holes located to the north and to the east of the injection hole, but not in monitor holes located south and west of HF02. Fracturing fluid was usually noted coming from the tunnel floor and walls 15 to 30 m east of HF02, consistent with a 70 dip of the fractures to the west. The fracture treatments are summarised in Table 1. A typical treatment record is contained in Figure 4 for fracture treatment 1. In addition to the treating pressure and injection rate, the plot shows pressure response at monitor holes M1 and M8. The rapid rise in pressure at these holes, located about 25m from HF02, is associated with arrival of the hydraulic fracture and provides a measure of the fracture growth rate and of the pressure in the fracture at that point after arrival. Extensometers in monitor holes M2 and M7 measured average fracture opening of 0.3 to 0.5 mm during the water treatments and up to 0.5 to 1.0 mm for crosslinked gel treatments. Numerical fracture models predict fracture openings that are larger by about a factor of two compared with these measurements. The extensometers used measured deformation over 10 and 20m intervals and the elasticity of the rock between the anchors is expected to have reduced the deformation measured by up to 30 percent, which does not explain the smaller than expected opening measured. In addition, the shut-in pressures measured up and down the injection hole revealed a stress gradient that favors fracture growth up dip over the down dip direction. Asymmetric growth of the fractures would move the maximum fracture opening up dip, which should produce more opening at M7 and less at M2 than what would be measured for symmetric fracture growth. The measured displacements did follow this trend. The displacements are consistent with opening width in opening mode fractures and provide fracture growth data and residual opening information.
3.2 Fracture Growth Intersections with extensometers and piezometers grouted into monitor holes M2 and M7 and with holes monitored with packers (M1 and M8) provide a direct measure of the fracture growth in the rock mass at Salvador. Figure 5 contains a summary of the fracture growth data from several treatments. A power law curve fits the data and provides an empirical fracture size versus time relationship that can be used to design fracture treatments in this rock mass. The growth relationship is: (1) R = 7.56t0.55 The relationship given by equation 1 is for water fracturing fluid injected at 215 lpm. A numerical model that fits the growth data can be used to consider the effect of changing the fluid, injection rate or injected volume on the fractures produced. 3.3 Fracture Stress Change and Orientation Stress change monitoring cells were installed in BQ boreholes located about 7m north and south of HF02. Two ANZI stress change cells (Mills, 1997) were installed in each hole at about 40m from the collar. During each of the fracture treatments, the stress change sensed by these cells was logged every 15 seconds. Analysis of this data gives the 3D stress change tensor induced by the hydraulic fracture. The maximum stress change component is a reflection of the pressure inside the hydraulic fracture and also provides information about the fracture orientation since this stress change vector points toward the plane of the fracture. A more detailed presentation of the stress change results is contained in Mills and Jeffrey (2004). The stress change analysis indicated sub-vertical fractures were formed with a general north-south strike and dipping at 75 to 80 degrees to the west. This orientation agrees with the directly observed intersections of the fractures with the tunnel to the east of HF02. Planes were fitted through the intersection data recorded at the monitor holes around the site using a least squares method. Table 2 summarises the orientation obtained from this fit to the monitor hole intersections, which agree with the stress change orientation data. The magnitude of the maximum principal stress change, during the injection, was in the range of 0.5 to 1.0 MPa for the water based fractures and 1 to 2 MPa for the crosslinked gel fractures. These stresses decreased with time after shut-in but limitations of the instrumentation logging system did not allow for monitoring this decay for more than a few hours. The change in maximum and intermediate stress recorded is consistent with sub-vertical fractures dipping to the west and propagating as opening mode fractures rather than shear fractures. The mine through observations confirm the fractures formed are opening mode. 4 MINE THROUGH OF TWO FRACTURES
Figure 4: Summary of fracture 1 at Salvador. Massmin 2004
Two hydraulic fractures were placed in HF02 at depths of 111.5m and 117.5m from the hole collar. These fractures were therefore at the depth of the undercut level. Red plastic chips were mixed into the crosslinked gel fluid used for the fracture at 117.5m and yellow plastic was mixed into the linear gel fluid used for the fracture at 111.5m. During the treatment, crosslinked gel containing red plastic was seen leaking from an overcore stress measurement hole on the extraction level, indicating this fracture had intersected the end of this short hole. This intersection point is consistent with the dip of the red propped fracture mapped on the undercut level. Red plastic containing gel Santiago Chile, 22-25 August 2004
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Figure 5: Measured fracture growth in Salvador rock mass.
also leaked from the zone on the 2600 level located 120m east of HF02. In January 2003, a tunnel was driven near the bottom of HF02 on the undercut level, intersecting the two plasticpropped hydraulic fractures. Figures 6 and 7 contain drawings of the fractures mapped on the north and south walls of this tunnel. Both the yellow and red propped hydraulic fractures were mapped. Because much less yellow plastic was mixed with the lower-viscosity linear gel fluid, the trace of this fracture was only visible on the south wall near the borehole. The borehole was not directly intersected by the tunnel but is estimated to lie less than 0.5 m into the rock on the south wall of the tunnel.
Table 1: Summary of fractures at site. Fracture
Straddle top (metres)
Fluid
Vol. (m3)
1
57.55
water
6.3
2
66.55
water
6.5
3
51.55
water
5.1
4
51.55
xlinked
3.6
5
54.55
Linear
4.9
6
60.55
water
4.2
7
63.55
xlinked
3.8
8
117.55
xlinked
5.5
9
111.55
linear
5.2
10
57.55
xlinked
4.4
In contrast to the shallower fractures at the preconditioning experiment site, these two fractures both dipped slightly to the east. The change in dip with depth may be the result of stress rotation caused by interaction with the approaching cave front. The east dip also caused the two fractures to be closer to each other than otherwise. A 250 E dip results in the 6m separation along the hole direction becoming a 0.5m perpendicular separation between the fractures. The red propped fracture was mapped on both the south and north sides and across the back of the tunnel. Both fractures grew through most veins and natural fractures they encountered without any visible interaction or offsetting of their path. Some more weakly cemented natural fractures produced minor offsets in the hydraulic fracture path. For example, the fracture near the roof line on the north side (Figure 7) caused an offset of about 20mm in the hydraulic fracture path. This natural fracture had undergone shear deformation previous to the placement of the hydraulic fracture. The hydraulic fracture mined at Northparkes (van As and Jeffrey, 2002) interacted with most of the natural fractures it crossed and contained many small and several larger offsets in its path. Mineral filling and cementing is less prevalent or absent in the rock mass at the Northparkes site and comparison between the Salvador and Northparkes mined fractures demonstrates the effect of natural fracture shear strength and permeability on the type of crossing interaction that develops. As the natural fractures become stronger in shear and less permeable, few offsets and other interactions develop when a hydraulic fracture grows toward and through them. In contrast, the widely spaced more permeable and weak shear zones at the Salvador site, which were spaced at from meters to tens of meters apart, interacted strongly with the hydraulic fractures, In some cases, the hydraulic fracture growth was stopped in that direction by these shear zones. Average spacing of major structures such as these should be determined at preconditioning sites so that the spacing
Table 2: Fracture strike and dip from intersections.
532
Fracture
Strike azimuth
Dip (degrees)
1
02
81 W
2
343
90
3
07
75 W
4
07
75 W
5
10
78 W
6
356
85 E
7
02
82 W
Figure 6: Fracture trace on south side of tunnel.
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growth and serves to verify the fracture size used in the design. 5 SEISMIC DATA Microseismic events were recorded during several of the treatments using 12 hydrophones and a seismograph system. Event locations cannot be determined from this type of logging, but the number of events during different phases of the treatments was measured. During the initiation of the fracture, about 20 events per second were recorded. Later, during the fracture growth stage, this rate dropped to 1 to 3 events per second. Even at this lower rate, the 20 to 25 minute long treatments were associated with several thousand microseismic events each. Cross hole seismic surveys were carried out between holes M9 and HF01 before and after the fracture treatments in HF02. Figure 8 shows the before and after compressional wave velocity for the rock between M9 and HF01. The velocity change was 2 to 4 milliseconds over the 60m distance, reflecting a significant change to the rock mass strength. 6 CONCLUSIONS
Figure 7: Fracture mapped on north side of tunnel. between the injection boreholes can account for the possibility of impeded fracture growth across them. It is also useful to plan for monitoring of intersections with pre-drilled injection holes around the borehole being treated. This monitoring provides direct feedback of fracture extent and
Ten hydraulic fracture treatments were carried out at the Salvador preconditioning experimental site. Two of these fractures were mined on the undercut level to obtain information about fracture geometry and interaction with natural fractures in the rock mass. Monitoring of the other eight fractures included measuring pressure and opening, and stress change during each treatment. Cross hole seismic surveys were carried out before and after the fracturing. Stress changes of up to 2 MPa were measured. The stress change data also provided a measure of the fracture orientations and indicated they were sub-vertical, dipping about 75 degrees to the west. This orientation was supported by intersections with monitoring holes and with the tunnels on the 2600 level and is in agreement with the
Figure 8: Cross hole seismic survey geometry and velocity changes measured. Massmin 2004
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stress field measured during the characterisation phase. Hydraulic fractures 40 to 50m in radius were produced by injecting 4,000 to 6,000 liters of fracturing fluid at 215 litres per minute. Growth occurred to the north and east of injection hole HF02 with only limited growth to the south and west. A conductive fault or shear zone is believed to have acted as a barrier to growth to the south and west. The mined hydraulic fractures on the undercut level crossed most pre-existing joints and natural fractures. Small offsets in the fracture path occurred at weaker natural fractures. Most of the hydraulic fracture path was through fresh rock. Cross hole seismic data demonstrated a large change in seismic velocity occurred as a result of placing 8 fractures in a 20m zone of the 60m ray path. A significant reduction of rock mass strength can be inferred to have occurred in the treated zone. ACKNOWLEDGEMENTS The authors acknowledge the International Caving Study II (ICS-II), Salvador mine and CSIRO for support of the work described in this paper.
534
REFERENCES • Brown, E.T., 2003. Block Caving Geomechanics, JKMRC Monograph Series in Mining and Mineral Processing 3, University of Queensland. • Mills, K.W. 1997 In situ stress measurement using the ANZI stress cell. Proceedings of the International Symposium on Rock Stress, 149-152. Rotterdam: Balkema. • Mills, K.W. and Jeffrey, R.G. 2004. Remote high resolution stress change monitoring near hydraulic fractures, Proceedings of the MassMin 2004 symposium, Santiago. • van As, A., Jeffrey, R.G., Chacon, E., and Barerra, V. 2004. Preconditioning by hydraulic fracturing for block caving in a moderately stressed naturally fractured orebody, Proceedings of the MassMin 2004 symposium, Santiago. • van As, A. and Jeffrey, R.G. 2002 Hydraulic Fracture growth in naturally fractured rock: Mine-Through mapping and analysis. NARMS-TAC 2002, Mining and Tunnelling Innovation and Opportunity. Pp 1461-1469.
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Preconditioning by hydraulic fracturing for block caving in a moderately stressed naturally fractured orebody Andre van As, AVA Mine Geotechnical Services, Australia Rob Jeffrey, CSIRO Petroleum, Australia Enrique Chacónn, Victor Barrera, El Salvador Division, Codelco Chile
Abstract Preconditioning by hydraulic fracturing is a tool with the potential to modify the caveability and fragmentation of an orebody prior to and during mining. A field-scale experiment was carried out in the Lift 2 orebody at Northparkes E26 mine near Parkes, NSW to measure hydraulic fracture growth and assess the affect of the fracturing on rock mass strength and in-situ stress. The 9700 level exploration drill drive located near the top of the Lift 2 orebody provided an ideal site for the experiment. Eight fan arrays of NQ-size exploration drill holes had already been drilled laterally and down into the ore from this level. These holes provided excellent access for placing fracture-monitoring instrumentation across the hydraulic fracture growth path. Site characterisation work included, geology, rock properties, overcore and micro-frac stress measurements, and seismic velocity profiling. In addition, a microseismic monitoring array was extended to the site by adding to the existing mine microseismic array. Cross-hole seismic measurements were conducted before, during and after the hydraulic fracturing period. In addition, the fracture geometry and hydraulic fracture volume were also remotely monitored using eighteen sensitive tiltmeters. Stress change associated with the fracturing was recorded using three ANZI cells located 15 to 40 m above the hydraulic fracture plane. When combined with previous data from mining and mapping a hydraulic fracture placed into the country rock ahead of development tunnels, the experiment provided a comprehensive data set to define and understand hydraulic fracture growth in naturally fractured rock. The effect of hydraulic fractures on rock mass caveability and resulting fragment size during caving is being assessed using this data. This paper summarises the measurements made and important results obtained that will assist in future design and implementation of hydraulic fracturing as a preconditioning tool for cave mining.
1 INTRODUCTION In the past the use of cave mining has been limited to massive ore bodies characterised as having a fairly uniform grade distribution, large aerial extent (i.e. foot print) and relatively weak rock mass strength (such as kimberlites). In recent years several low-grade, massive ore bodies have been discovered at depth, which due to low metal prices, can only be exploited profitable through the adoption of the block cave mining method. Unlike the typical caving ore bodies of the past, many of these new deposits comprise highly competent rock masses and their geometrical dimensions (i.e. foot print and block height) are often dictated by their economic viability. Hence, the rejuvenated interest in cave mining of more competent ore bodies has forced the industry to reassess the suitability of current caveability predictive tools for stronger rock masses whilst simultaneously investigating the prospect of rock mass preconditioning and its potential to enhance the caving characteristics of these deposits. The successful hydraulic fracture cave inducement program conducted at Northparkes Mines has lead to a considerable interest in the technique with respect to its application to cave mining. van As and Jeffrey (2000) have demonstrated the technique to be an extremely cost effective means of cave inducement and state that it also has the potential to be applied to cave preconditioning, ultimately reducing the risks associated with caving hard rock orebodies. Recent site experiments were conducted at both Northparkes and Salvador Mines, partially supported by the Massmin 2004
International Caving Study (ICS II), and formed a major component of the ICSII research. The experiments were specifically designed to measure hydraulic fracture growth in competent and well jointed rock masses and examine their influence on the rock mass. This paper describes the Northparkes Mines site experiments, provides an overview of the findings and discusses the implications for cave preconditioning. 2 WHAT IS HYDRAULIC FRACTURING? Hydraulic fracturing involves isolating a section of a borehole, often by means of a straddled packer system. Once sealed, fluid is pumped into this straddled area causing the fluid pressure to rise, generating tensile hoop stresses along the axis of the borehole which eventually exceed the tensile strength of the rock, causing a fracture to form. The orientation of the fracture plane is defined by the orientation of the minimum principal stress whilst the propagation of the fracture into the rockmass will continue as long as the pumping rate exceeds the rate of fluid loss into the rock and as long as the pressure in the fracture exceeds the far-field minimum stress magnitude.. 3 PRECONDITIONING EXPERIMENTAL SITE Two site experiments were conducted at Northparkes Mines, the first involving mapping a hydraulic fracture by mining it and the second, the main preconditioning experiment, involved a multiple fracture treatment within a well instrumented region of the proposed E26, Lift 2 block cave.
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The Mine-Through experiment involved placing two hydraulic fractures well ahead of an advancing tunnel. The fracture fluids contained fluorescein and plastic proppants which enabled the fractures to be visually traced and inspected upon mine through. The objective of minethrough experiment was to map the hydraulic fractures, recording their orientation, residual width and interaction with the natural jointing. More detail on this experiment can be found in the paper by van As and Jeffrey (2002). The main preconditioning experiment involved creating several hydraulic fractures at various intervals within a region of the proposed E26, Lift 2 block cave. The site location was selected primarily due to its abundance of diamond drill holes which would allow for a dense array of monitoring instrumentation to be installed in and around the hydraulic fractures. A single HQ hole, drilled between these existing monitoring holes, was used as the ‘injection’ hole along which all of the hydraulic fractures were placed.
indicate a horizontal major principal stress of approximately 40 MPa whilst the minor principal stress is subvertical with a magnitude of around 10 MPa. It should be stated however that these measurements are slightly different to those measured elsewhere in the mine and this is thought to be attributed to the presence of the diorite sill. 5 MONITORING AND INSTRUMENTATION Hydraulic fractures were propagated from a central injection hole out into the rock mass, intersecting the adjacent monitoring holes as the fracture grew. The arrival time and pressure provided by the intersections measured the fracture growth rate, pressure and orientation, including their variation over time. The effectiveness of different fracture fluids and proppants in inducing changes to the rock mass was also investigated. The experiment site was located on the 9700mRL level, which is 250m above the Northparkes E26 Lift 2 extraction level. The 9700mRL level was previously used as an exploration drilling level and hence contains numerous NQ diamond drill holes, drilled from eight cuddies spaced at 25m centres along a north-south oriented drill drive (refer to Figure 2). Each cuddy contains a fan of holes drilled with varying dips yet with roughly the same bearing. The injection hole used to place the hydraulic fractures was located centrally such that all of the surrounding holes could be effectively used to monitor the hydraulic fractures. The instrumentation was grouped into near-field and far-field monitoring systems with most of the near-field instrumentation installed in the drill holes nearest to the injection hole. The far-field instrumentation included the seismic monitoring system, the downhole seismic system and an array of tiltmeters located throughout the mine. The positioning of these instruments is illustrated in Figure 3, and is summarised as follows: • The injection hole was an HQ hole drilled from cuddy 5 with a dip of 560 towards 1180 east. • ANZI stress cells were installed in a hole directly above the injection hole with the cells located between 15 – 25m from the nearest fracture to 40 – 48m from the furthest fracture. • Piezometers and extensometers were installed in 4 holes, one hole immediately below the injection hole in cuddy 5,
Figure 1: Location of the preconditioning experiment site with respect to the Northparkes block caves. 4 GEOLOGY AND GEOTECHNICAL The Lift 2 orebody generally comprises four main rock types. Central to the orebody is a quartz monzonite porphyry body around 70 metres wide, to the east is a biotite quartz monzonite which is a massive equ-granular unit that comprises the greater proportion of the rock mass within the Lift 2 block. To the west of the monzonites is a volcanic sequence, dominantly porphyritic lavas, through which a coarsely crystalline, diorite sill has intruded. The diorite sill measures up to 80 metres thick, dips into the orebody at approximately 35 to 45 degrees and is the main unit in which the preconditioning experiment was conducted. In general the rock mass is well jointed and rock type is the dominant control on variations in rock mass properties. The rock mass has been classified using Laubscher’s MRMR system, Laubscher, (2000), with the Rock Mass Rating (RMR) ranging from a minimum of 41 in the volcanics to a maximum of 64 in the biotite quartz monzonite. The in situ stress measurements conducted at the experiment site 536
Figure 2: Plan of the 9700 mRL level, showing the site layout and drill holes.
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•
•
•
•
•
two in the steepest holes in cuddy 4 and one in the steepest hole in cuddy 6. Packer systems, each containing down-hole pressure transducers, were lowered down four holes. Single packers were installed 30m from the collar in one hole at cuddy 6, one in cuddy 4 and one in cuddy 3. A straddle packer system with a 25m long straddle was installed in D284 in cuddy 6 with the top of the straddle located initially at 81.5m. A geophone used as a receiver for the down-hole seismic measurements (during the fracturing process) was lowered down the steepest hole in cuddy 3. All of the remaining drill holes were used simply as open holes. Any intersection of the hydraulic fractures with these holes resulted in water flowing from the hole collars and provided useful information on both the geometry of the fracture and its propagation rate. The mine microseismic monitoring system consisted of 16 tri-axial accelerometers and five uni-axial geophones. Sampling rates for all accelerometers was set at 6 KHz so as to ensure the detection of extremely small events. Finally, an array of 19 tiltmeters were installed, 13 were located on the 9700 mRL level, 5 were located on the 9800mRL level and one was located in a conveyor drive on the 9770mRL level.
Table 1: Summary of hydraulic fracture preconditioning treatments at Northparkes. Fracture
Straddle top (metres)
Fluid
Vol. (m3)
1
79.76
water
11.6
2
79.76
water
6.4
3
99.36
water
7.9
4
93.76
water
10.6
5
96.56
water
9.3
6
110.56
water
10.3
7
79.76
xlinked gel
10.8
13
96.56
gel/water
12.5
tiltmeter monitoring and from the stress change cells. Analysis of the tiltmeter data also provides information about the fracture volume while the stress change monitoring provides information about the stress change induced by each fracture and the type of fracture growth (opening or shear). 6.1 Fracture Pressure Analysis Analysis of the treating pressure response can be used to determine the general mode of fracture growth, the amount of fluid lost from the fracture during the treatment via leakoff, the pipe friction and fracture entry loss, and the minimum principal stress magnitude. For example, fracture 13 (Figure xx) has been analysed in detail. Figure xx contains a Gfunction plot of the falloff data from which the pressure associated with closure of the natural fractures and of the hydraulic fracture have been determined. Closure of the natural fractures causes a decrease in the fluid loss rate. The closure of the hydraulic fracture is a measure of the minimum principal stress in this part of the rock mass. A value of 19.2 MPa for the closure stress is indicated.
Figure 3: Isometric view of the 9700 mRL level depicting the instrumentation layout
6 MONTORING RESULTS The hydraulic fractures were placed using a straddle packer tool that exposes 0.5m of the open hole to fracture pressure. A down hole pressure transducer is located at the top of the top packer in this tool and measures the fluid pressure as it enters a 20mm diameter mandrel tube that carries it into the straddle section. A small pressure drop, therefore, occurs between the transducer and the open hole section. Table 1 provides a summary of the 8 fracture treatments carried out during the preconditioning experiment. Fractures 9 through 12 are not listed since injection was limited to a few litres per minute and a higher rate treatment was not carried out for these injections. Intersections were recorded for most treatments with several piezometer, extensometer, and packer monitoring points. The fracture orientation was obtained from the Massmin 2004
Figure 4: Summary of fracture 13 showing bottom-hole pressure, injection rate, and pressure in monitor hole D284.
The pressure declined throughout the treatment which is an indication of continued fracture growth with an approximate radial geometry. The sharp increase in pressure a D284 corresponds with arrival of the hydraulic fracture at this monitor hole and provides a direct measure of the fracture growth rate. The pressure in D284 after the fracture arrival gives a measure of the pressure in the hydraulic fracture at this location away from the injection point. However, loss of fluid from the hole into other fractures acts as a pressure regulating mechanism which
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means the recorded pressure is only a lower bound on the pressure in the fracture itself.
Figure 6: History match of fracture treatment 1. fracture is predicted by the model to grow to a radius of about 30m and a maximum width at the injection hole of 2.2 mm. Figure xx shows the model-predicted shape of the fracture, with the center of the fracture moving east of the injection point
Figure 5: Pressure falloff after shut-in of fracture 13.
The efficiency of a fracture treatment is defined as the ratio at any time of the injected volume to the volume stored in the fracture itself. Based on analysis of the falloff data, the fracture efficiency at shut-in for fracture 13 can be estimated as (Nolte, 1989)
η
Gc =
,
(1)
2 + Gc where Gc is the value of the G-function at fracture closure. Using this relationship, an efficiency of 52 percent is found. The fracture closure pressure implies a minimum principal stress of 19.2 MPa which is about 7 MPa higher than the minimum stress measured by overcoring at the 9700 level tunnel. Fracture 13 is 84m vertically lower than the elevation of the overcore test, which is expected to result in an increase in the sub-vertical minimum stress by about 2.2 MPa. Furthermore, the fracture site is located inside the diorite dyke which may be acting as a stress concentrator because of its higher elastic modulus. In contrast to these effects, there is a decreasing minimum stress trend across the 80 to 110m interval fractured. The ISIP (Instantaneous Shut-In Pressure) values recorded after each fracture show a gradient in the minimum principal stress exists at this site. The ISIP values are upper limits for the minimum stress but will reflect the change in minimum stress with position. The ISIP measured gradient in stress is minus 0.06 MPa per meter into the hole. Such a strong gradient is expected to significantly affect the hydraulic fracture growth, causing the fractures to grow more in the direction of lower stress. Hence, the fractures created are expected to have grown along their strike direction and more to the east than to the west. The existence of the measured stress gradient is consistent with a stress shadow caused by the overlying Lift 1 abutments and mined area (see Figure 1). Fracture 1 at the site has been history matched using a pseudo 3D hydraulic fracture model. Figure 6 contains a plot comparing the modelled and measured data for this treatment. Measured pressure and fracture size are shown on this plot with a reasonable match obtained. A minimum stress of 20 MPa and a rock modulus of 77 GPa were used in the model, perhaps reflecting the stiffer diorite dyke. The 538
Figure 7: Plan view of fracture 1 showing biased growth to the east.
6.2 Tiltmeter Monitoring of Fractures An array of 19 tiltmeters was deployed to monitor the hydraulic fractures formed in the rock mass. Each tiltmeter measures the horizontal gradient of the vertical displacement at its location. Tiltmeter monitoring of hydraulic fractures is a commercial service in the petroleum industry and Pinnacle Technologies provided the tiltmeters at the site. As a hydraulic fracture grows and opens, it induces displacements in the surrounding rock. A tiltmeter located remote from the fracture will measure the tilt associated with these displacements. Analysis of such tilt data is an inverse modelling problem. The goal of the inverse modelling is to find a fracture with opening and orientation that produces tilts at all measurement points that are the best fit to the measured data. The instruments at Northparkes were located more than three fracture radii from the hydraulic fracture which means only fracture volume and orientation can be determined (not fracture length, width, and height) (Lecampion et al., 2004).
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The best fit line, using a power law model, is given by: (2) L = 8.15t 0.46
Table 2: Tiltmeter analysis results. Frac
Depth (metres)
Dip (deg)
Strike (deg)
Efficiency %
1
80
18 E
57
29
2
80
16 W
136
30
3
99.6
26 E
27
56
4
94
23 E
32
47
5
96.7
23 E
32
51
6
110.7
11 E
80
49
7
80
11 E
71
30
13
96.7
17 E
59
34
Results from analysis of the tilt data are given in Table 2. The Pinnacle and CSIRO analysis of the data gave essentially similar orientation results and CSIRO results are shown. The fractures were found to be subhorizontal and dipping to the east in all cases except for fracture 2 which was interpreted to dip to the west. However, fracture 1 and 7, at the same depth as 2 were both found to dip to the east. The azimuth strike of the subhorizontal hydraulic fractures is less well defined by the tilt data than is the dip. A strike to the NE is consistent with the overcore stress data for the site and with local failures in tunnels observed nearby the site. The fracture volume determined from the tilt analysis allows the fracture treatment efficiency to be calculated. The efficiency listed in Table xx is the volume of the fracture at the end of pumping divided by the total fluid volume injected. The numerical fracture model history match provides an estimate of fracture treatment 1 efficiency of 26 percent, which compares well with the 29 percent calculated from the tiltmeter analysis. Tiltmeter monitoring of the fractures provides a reliable method to remotely determine the fracture orientation and can be used to provide rapid feedback during preconditioning operations. Fracture growth rate and size must still be determined by direct measurement at monitoring boreholes combined with fracture modelling. 6.3 Fracture Monitoring by Borehole Instruments. Packers were used to monitor pressure in four holes and grouted-in extensometers and piezometers were used in four other holes at Northparkes. These instruments provided data on fracture growth rate, pressure and opening. Figure 8 contains a summary of the fracture growth data obtained from the direct intersection data.
Figure 8: Measured fracture growth at site. Massmin 2004
where L is the fracture half-size (radius for a circular fracture) and t is pumping time in minutes. Hydraulic fracture growth for ideal geometry models follows a power law behaviour which is motivated using a power law fit to the measured data. Equation 2 can be used to determine a pumping time that will produce fractures of a desired size at Northparkes. This empirical approach can be improved by fitting individual fracture treatments. A numerical model with parameters that fit the average growth curve can be used to vary the injection rate and volume as part of a preconditioning design exercise. Extensometers measurements recorded 0.15 to 0.2 mm of opening during treatments with water and 1 to 1.9 mm of opening for the crosslinked gel treatment. Measured maximum pressure response at monitor holes was 19 to 20.4 MPa, which is about equal to the estimated minimum stress at each depth. It appears likely that other fractures in each monitor hole opened at about this pressure and acted as pressure regulators. These pressure measurements then serve to confirm the minimum stress value determined from falloff data and provide a lower limit for the pressure in the fracture. 6.4 Stress Change Monitoring Stress change was measured during each treatment using three ANZI stress change cells (Mills, 1997) installed in a BQ-size borehole drilled just to the north and above the injection hole. Each cell contains 18 strain gauges on an inflatable packer that is cemented to the borehole using a special epoxy. More detail of the use of these cells can be found in Mills et al. (2004) in these proceedings. The stress changes displayed consistent responses to the start and end of injection for each fracture treatment. The magnitude of the maximum stress change observed varied from 0.5 to 1.4MPa for distances of 15-40m from the fracture plane during the treatments using water. The maximum stress changes measured for the cross-linked gel treatment varied from 2.3 to 3.3MPa at 15-25m from the fracture plane. Water treatments resulted in a residual stress change, measured several hours after the treatment, of about 0.5MPa (0.2-0.6MPa). The cross linked gel treatments resulted in a residual stress change of about 1.5MPa (1.02.1). These residual stresses are expected to have decreased further with time but logging system signal drift did not allow for longer term monitoring. The maximum stress change vector can be used to help determine the orientation of the hydraulic fracture. East dipping subhorizontal hydraulic fractures were consistent with the orientation of these stress change vectors provided the hydraulic fractures grew more to the east than west. This fracture growth mode is supported by the ISIP measurements and stress gradient discussed in section 6.3 above. The numerical modelling of fracture 1 took account of this measured stress gradient, with the result that the fracture grew more in the east direction as shown in Figure 7. 6.5 Micro Seismics The flurry of seismic activity associated with hydraulic fracturing was unexpectedly high and in some cases the intensity of the activity lead to communication bottlenecks in the acquisition system resulting in a loss of data. In addition the system rejection rate (i.e. the number of associated waveforms defined as noise) proved significantly higher during the injection period, where a large number of associated waveforms appeared to contain more than one event, making processing ambiguous and consequently leading to the rejection of
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the events. Nevertheless a large number of events were recorded and processed leading to the following findings, De Beer and White (2003): • Seismic activity tended to lag breakdown. • The spatial distribution of seismic events do not seem to ‘cloud’ along a discrete fracture zone but exhibit a large degree of scatter, thus suggesting that the fracturing results in stress redistributions and inelastic deformations over a large volume (hundreds of metres from the source). This finding alone holds considerable implications for preconditioning and particularly the benefits of hydraulic fracturing as a preferred technique, but may in part arise because of multiple events close to the fracture being rejected by the system. • There was a continuation of elevated seismic activity for approximately 10 days after the hydraulic fracturing. • The calculated Apparent Volume (measure of inelastic deformation of the rock mass) increased substantially during fracturing and continued to do so for up to a week after fracturing. This high deformation rate has since proved similar to that experienced during the Lift 2, block cave undercutting process. • The calculated Apparent Stress (measure of stress change in the rock mass) revealed large stress changes induced to the rock mass although small when compared to those associated with undercutting. 6.6 Down-hole Seismics Down-hole seismic measurements of the treated rock mass were conducted before, during and after hydraulic fracturing. Measurements taken during fracturing revealed that the travel-time increased, over a distance of 91 metres, by 2.0 ms (ie 6.7 %), De Beer and White (2003). Thus the concept of a ‘fracture zone’ where the fracture fluid penetrates, opens, shears and propagates surrounding fractures stemming off the primary conduit fracture/s seems to provide a more plausible explanation than a discrete fracture. However in contrast to this observation, the propped fracture at the mine through site was mapped as a primary single fracture 1 to 2 mm wide at most locations. Measurements conducted after fracturing continued to show a reduction in velocity around the fracture zone although not as pronounced as those measured during fracturing. The implications of these low velocity regions is expounded further in subsequent sections. 7 INDUCED ROCK MASS CHANGES Quantifying induced changes to the rock mass as a consequence of hydraulic fracture preconditioning can be demonstrated through various rock mass classification schemes. The parameters most notably affected are :• joint frequency (through the introduction of new hydraulic fractures), • joint condition (through the breaking of rock-bridges, induced shear and the introduction of pressurised fluids), and • the stress regime (through increasing pore pressures, the opening and closing of fracture systems, shearing along fractures and a reduction of in-situ stress differences ). Although much effort has been made to measure the changes to each individual parameter it is difficult to attribute exact changes to specific parameters. Thus calculating a preconditioned rock mass rating from individual parameters, although instructive, is in most cases speculative. However, quantifying the cumulative affects induced to the rock mass can be reliably accomplished through the application of seismic data, namely seismic profiling and microseismic emissions. Several researchers 540
have demonstrated strong correlations between rock mass characteristics and seismic P-wave velocities (VP) and have subsequently derived equations which can be used to verify and validate both individual rock mass parameters and their cumulative affects. Barton (1991) established a correlation between Q and VP, for hard rock below 500m depth where: Vp = 5.0 + 0.5logQc
(3)
Other relationships established between VP and rock mass parameters include those with RQD and fracture frequency, e.g. Sjogren, et al. (1979), Palmstrom (1996), and Tanimoto and Ikeda (1983). In general, VP decreases with increasing fracture frequency, decreasing RQD, increasing porosity, decreasing density and increasing fluid content. Results from the hydraulic fracturing preconditioning experiment at Northparkes Mines revealed a decrease in the post fracturing VP of around 15%. This equates to a increase in the fracture frequency of between 5-8 fractures/m, a drop in RQD of around 25-30 and a reduction in Q from around 3.8 to 1.2. A similar change could be demonstrated using the MRMR system where speculated changes to the various indices reduced the IRMR from around 58 to 46. It must be emphasised however that these induced changes are dependant on the character of the initial rock mass to begin with, i.e. a highly fractured, weak rock mass is less likely to be influenced by preconditioning than sparsely fractured, homogenous hard rock. Boadu (1997) suggested that once a rock mass has fractured such that the VP has decreased by approximately 25% of its intact rock value, it becomes insensitive to further increases in fracture density. Thus with respect to preconditioning this finding implies that there is a critical fracture frequency and\or joint condition value below which preconditioning is ineffective. However, in terms of improving the caving characteristics of a rock mass one may argue that a rock mass insensitive to preconditioning is one which does not require preconditioning. 8 DISCUSSION AND CONCLUSIONS. The fracture treatments produced subhorizontal fractures that extended from 30 to 50m from the injection point. The orientation of the fractures formed was consistent with the measured stress field. Growth of the fractures was affected by stress gradients, lithology, and structure in the rock mass. Borehole monitoring of fracture growth was the only method tested that directly verified the size of the fracture created. Microseismic monitoring should be able to be used for this purpose with modifications to event logging and analysis. Tiltmeter monitoring provided the best data on fracture orientation, which was confirmed by stress change monitoring after accounting for non-symmetrical fracture growth effects. Stress change monitoring provides a direct measure of the induced stress around the fractures and can be used to discriminate between shear and opening mode fracture growth. There is no doubt that the measured changes induced to the rock mass as a direct result of hydraulic fracture preconditioning are significant and should consequently be realised through a reduction in primary fragmentation, improved caveability and enhanced caving rates. Figures 9 and 10 demonstrate examples of the anticipated effects of preconditioning calculated for the Northparkes experimental region, from either the introduction of discrete fractures or a fractured zone, as discussed in the preceding section. The next stage of preconditioning research should therefore focus on acquiring cave performance data that can be used to validate these rock mass changes.
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Unfortunately the mining industry, with a few exceptions, has yet to invest in adequate cave propagation and fragmentation monitoring tools and resources from which cave performance can be evaluated
Ltd for their work on the seismic profiling and microseismic monitoring, Ken Mills for his analysis of the ANZI cells, Pinnacle Technologies and Brice Lecampion for their analysis of the tiltmeter data and Tim Fergusson, Kevin Quinlan and Anthony Coleman for all their work installing and monitoring the instrumentation and operating the pumping equipment.
REFERENCES
Figure 9: The effects of preconditioning demonstrated on Laubscher’s Stability Chart (2000).
Figure 10: The effects of preconditioning demonstrated on primary fragmentation predictions.
• Barton N, 2002. Some new Q-value correlations to assist in site characterisation and tunnel design, International Journal of Rock Mechanics & Mining Sciences 39 (2002) 185–216 • Laubscher D H, 2000. The Block Cave Manual, Internal Report: International Caving Study (ICSI). • Tanimoto, C., Ikeda, K., 1983. Acoustic and mechanical properties of jointed rock. Proceedings of the 5th International Congress for Rock Mechanics, pp. 15–18. • Palmstrom A (1996) Characterizing rock masses by the RMI for practical rock engineering. Part 1: The development of the rock mass index (RMi). Tunnelling Underground Space Technol ll:l75-188. • Sjógren B. Ovsthus A, Sandberg J (1979) Seismic classification of rock mass qualities. Geophys Prospect 27(2):409-414 • van As, A. and Jeffrey, R.G. 2000. Hydraulic fracturing as a cave inducement technique at Northparkes. In Proceedings of MassMin 2000 Conference pp 165-172. • van As, A. and Jeffrey, R.G. 2002 Hydraulic Fracture growth in naturally fractured rock: Mine-Through mapping and analysis. NARMS-TAC 2002, Mining and Tunnelling Innovation and Opportunity. Pp 1461-1469. • De Beer, W and White, H. (2003), Final Report: Characterisation of the hydrofracture process using seismic techniques, Northparkes Mine (ICSII). Internal Report: International Caving Study ICSII. • Mills, K.W. (1997) In situ stress measurement using the ANZI stress cell. Proceedings of the International Symposium on Rock Stress, 149-152. Rotterdam:Balkema. • Mills, K.W., and Jeffrey, R.G. (2004) Remote High Resolution Stress Change Monitoring near Hydraulic Fractures, Proceeding of the MassMin 2004 Symposium, Santiago. • Lecampion, B., 2004. Mapping hydraulic fractures from tiltmeter measurements at Northparkes E26 mine, CSIRO Petroleum confidential report No. 04-015, March, 2004. • Nolte, K.G. 1989. Fracturing-pressure analysis, Ch. 14 in Recent Advances in Hydraulic Fracturing. J. Gidley et al. editors, Monograph 12, SPE, Richardson. • Boadu, F.K. 1997. Fractured rock mass characterization and seismic properties: Analytical studies, Journal of Applied Geophysics 36, 1-19.
ACKNOWLEDGEMENTS The authors would like to acknowledge Northparkes Mines, CSIRO Petroleum and the International Caving Study II for supporting this research. Thanks also to iGeo
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New Vision in Caving Mining in Andina Division, Codelco Chile Jorge Sougarret, Luis Quiñones, Ricardo Morales, Reinaldo Apablaza, Gerencia de Recursos Mineros y Desarrollo, Andina Division, Codelco Chile
Abstract Andina Division has a world class ore deposit, with relevant reserves of competent rock. Their mining may be performed in two ways: one, by conventional mining methods; another, by trying to innovate a different mining option. Competent rock mining by conventional method shows low productivity at high costs. In this regard Andina Division has decided to reconsider caving mining in order to make it more attractive at a long term period. A question comes to our mind, how can we operate over the rock mass previously to the caving in order to get a more manegeable final product? Under this scope, we are introducing the experience of Andina Division in the use of a preconditionning technique of the rock mass by blasting over a competent rock mass. This technique should enable us to develop in a close future from the present traditional mining method to a high productivity mining method.
1 INTRODUCTION División Andina, disposes of potential reserves at a world level. Approximately 70% of these are located within a competent rock environment. This fact should influence the production capacity of the future mines, comparing it to the present mining from low to medium competent rocks. This influence shold be reflected mainly in the folowing aspects: - Low caving of the rock mass. - Low speed in caving propagation. - Thick fragmenttion. - Less productivity. - Operational cost increasement. So far, the efforts to face the challenges shown by competent rock mining, have been oriented to affect the mining designs, aiming to strengthen the pillars by increasing the extraction network and incorporating large
sized equipment, investing in relevant resources in order to measure the phenomena involved in this mining, like the demonstration of efforts and the occurrence of seismic events. In this environment, mining engineering areas have worn out efforts and in a certain way they have inhibited the actions that may bring a different point of view from the knowledge reached in this new reality. In fact, the idea to check the factors involved in competent rock mining arises. At first, this has produced a technological break in the way of facing competent rock mining. This sort of new paradigm for mining, puts into operation the concept of operating on the rock mass before its mining and work on it at a point such as to count on a material that may allow to take up again the standards of a low cost and large volume mining. Therefore, reach a high economical afficiency. We can visualize that the point is to go back to the beginning of the concept of the mininf method by block caving. This means to take up firmely
Figure 1 542
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Figure 2 the utilization of the gravitational flow in the different stages of the extraction systems. 2 DEVELOPMENT This project can be found in the program of Technological Innovation in Underground Mining from Codelco. It was developed by Andina Division and the Mining and Metallurgy Investigation Institute IM2, Filial of Codelco, Chile. It is important to emphasize that this operational process on the rock mass is focused in identifying the relevant factors and then affect them through an action that may provoke a weakening of the competent condition. Figure 1 This way, through the action of explosives, it is intended to achieve the weakening of the rock mass. It is not expected to be fragmented, as it happens in a "forcing" operation through free face blastings. It rather leaves the "processed rock mass" in the proper conditions so that the next coming caving may favour the caving propagation. Figure 2
Figure 3. Massmin 2004
The experimental place selected, emerges as a strategic measure, to perform a previous process on the rock mass within a demarcated area, so as to insure its caving. This place is located in the Southwest extreme of the Third Panel under the following characteristics: Experimental area RMRL Veins frequency Lithological Units
: : : :
7000 m2 65 3 – 4 (ff/m3) Granodiorite Río Blanco and Granodiorite Breccia Río Blanco
Height of the Rock Column Competent
:
100 to 150 meters.
This effort for technological innovation should develop a process that should have a maginal cost, regarding the technological alternatives available nowdays. Within the operational aspects, the project faced the lack of experience at a national level in the requirement of drilling in an upwards direction from an underground work, lengthes over 120 m in a diameter over 5 1/2" .
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543
Another relevant challenge concerned the explosives cargo in an upwards direction. This explosive must guarantee the maximum efficiency in the reconditionning process and it must be safely loaded in vertical drillings over 100 m length. For this purpose a group of tests at a pilot scale were performed. The main purpose was to determine the principal parameters that rule the blasting process in a confined environment, calibrate the numerical models of the blasting process, visualize and measure the interaction effect of the stress waves, check and calibrate the electronic blasting system selected, and improve the blaster system designed for this process. The criteria used to design the preconditionning blasting was to maximize the energy available from the waves of stress. This indicated to measure a multiple blasting strategy along the column of explosives and select a sequency of start that would allow a wave-wave interaction and/or a wave –crack interaction, and finally for safety reasons, direct the front of reinforced waves towards the zones distant from the mining substructure.
Amount of Explosives Length of the Cargo Length of the shots Area to be covered Time of the blasting
: : : : :
29500 Kg. 85 m 100 y 112 metros 7000 m2 26 ms
Evaluation of Preconditionning of the Rock Mass. Fragmentation The evaluation of this project was focused in the fragmentation, which is the most representative variable to measure the impact, in order to settle a comparisson base. The evaluation of the fragmentation, consisted on a photographic sample of the draw points in the sector located in the extreme south of the streets of preconditionning, as well as in a competent mineral sector not preconditionned located in the streets 61 and 63. Figure 6.
Figure 6. These lifts show the following results: Figure 5 Interaction wave-crack. In the process of understanding the interaction mechanisms of the stress waves produced by the blasting of a column of explosives, and besides configurating the optimus blasting strategy, it was necessary to face an exhaustive study to introduce a numerical tool that could allow to simulate at its best accuracy the blasting phenomenon in terms of propagation and interaction of the stress waves. This was achieved based on an important number of controled experiences that enabled to calibrate and validate the numerical tool introduced. (FLAC 2D and FLAC 3D). As an explosive, an emulsion especially formulated to fulfill the requirements of this process, diminishing its density to1.15 gr/cm3, in order to reduce the hydrostatic load was selected. The optimal distance between the starting points was determined by a numerical simulation for the emulsions used in this project. The starting points must be placed every 8 m along the column of explosives. Regarding the starting time between them, the most convenient determination should be that every detonation point in the column should start simultaneously. 3 INDUSTRIAL TEST Area processed Amount of shots Diameter of the shots Explosive of the Column System of Initiation 544
7000 m2 19 5 1/2" Emulsion RS-5 Density: 1.15 gr/cm3 : electronic blasting machine
: : : :
Total tonnage sampled Nº of draw points sampled Streets evaluated
: : :
1.000.000 32 4
Figure 7 shows the average curves obtained for each production street, bearing in mind that the experimental sector was formed by four production streets, 69, 71, 73, 75 into which streets 71 and 73 are located in the centre of the preconditionned zone. Additionally to this graphic a representative curve of competent rock not preconditionned was included as a reference. Note that in relative terms, the curves associated to the streets 69 and 75 are more displaced towards the referring curve, and represent the limits of the experimental area. The representative size D80 for the preconditionned rock, according to Figure 7, varies in a range from 0.70 m to 0.80 m, which is approximately 50% of the size of D80 of a not preconditionned material. Figure 8 shows a typical digital image of the preconditionned material of the experimental sector. According to the results obtained, it is necessary to mention in this part, that, there is a 20% of primary rock which exceeds the value of the result desired of 0.70m or 0.80 m Nevertheless, the remarkable fact is that no hangings in height were recorded, as it frequently happens in not preconditionned materials. The obstructions recorded were just at a floor level by the interaction of 3 or 4 rubble stones with sizes that could be handled and reduced without important operational problems. The image was obtained from CP-73, Trench 22-W, after approximately 18,000 Tons passed by the draw point. . The size D80 of this image is 0.83 m, nevertheless, there is one sole rubble stone whose
Santiago Chile, 22-25 August 2004
Massmin 2004
Figure 7. commonly cover a big part of the draw point, generating in these cases, hangings at different heights. This brings difficulties while trying to achieve its reduction. Figure 9 shows a typical digital image obtained from the not preconditionned sector. In this case the value of the size D80 is 1.82 m, and the maximum visible size is 2.3 m. Other measures correspond to the reduction of the burnings and the light blasts in the draw points. This diminished in 50%.
Figure 8.
Figure 10.
Figure 9.
largest size is 1.6 m, corresponding to 20% of the oversize expected. On the other hand, opposed to the characteristic that shows a preconditionned competent rock, the not preconditionned material is usually found in big blocks that Massmin 2004
Figure 11.
Santiago Chile, 22-25 August 2004
545
Another indicator sampled, coresponds to the entry of the dilution of this preconditionned area. This is shown in Figures 10 nd 11. As it may be noted, a delay in the entry of the dilution regarding similar areas in competent rock takes place.
•
4 CONCLUSIONS. The main results of this experience correspond to: - The preconditionning technique of a rock mass is possible. - The results of this project show the possibility to get important reductions in the fragmentation expected. - It helps the caving of a competent rock mass. - It delays the entry of dilution. - At the beginning of the year 2005, Andina Division starts its application in a massive way in the extreme North of the III Panel. Besides, it will evaluate aspects that have not been deeply discussed, such as the handling of extraction speed and the impact over the abutment stress. - It is important to emphasize that the application of this technique will require a higher control and accuracy in the planning and operation of this additional front, which is connected to fronts of development, construction, caving and extraction.
•
•
•
•
•
There is no doubt that according to these data, the expectations and the potentiality of this technique will lead to a higher efficiency in the productive system, generating mechanization alternatives that will favour the productivity of mining by Panel Caving.
•
REFERENCES
•
• Amadei, B., and O. Stephansson, Eds. (1997) Rock Stress and its Measurement. London: Chapman & Hall. • Whittaker, B.N., R.N. Singh, and G. Sun, Eds. (1992) Rock Fracture Mechanics, Principles, Design and Applications. Amsterdam: Elsevier. • Persson, P., R. Holmberg, and J. Lee, Eds. (1994) Rock Blasting and Explosives Engineering.Boca Raton: CRC Press. • Ouchterlony F., Prediction of crack lengths in rock after cautious blasting with zero interholedealy. FRAGBLAST – International Journal of Blasting and Fragmentation 1(1997):417-444 • Stagg, M.S. and J.N. Michael, Influence of blast delay time on rock fragmentation: onetenth-scale tests. Bureau of Mines Information Circular IC 9135, Surface Mine Blasting proceedings: Bureau of Mines Technology Transfer Seminar, Chicago, IL, April 1987. • Liu, L. And P.D. Katsabanis, A numerical study of the effects of accurate timing on rock fragmentation. Int. J. Rock mech. Min. Sci. Vol. 34, No. 5, pp. 817-835, 1997. • Stagg, M.S. and S.A. Rholl. Effects on accurate on accurate delay on fragmentation for single-row blasting in
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•
•
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a 6.7 m (22 ft) bench. Proc. 2nd Int. Symp. Rock Fragmentation and Blasting, eds W. Fourney and R. Dick. Soc. Exptl. Mech., 1987, 210-223. Chiappeta, R.F., Precision detonators and their applications in improving fragmentation, reducing ground vibration and increasing reliability – A look into the future. Fourth High-Tech Seminar, Blasting Technology, Instrumentation and Explosives Applications, Nashville, Tennessee, U.S.A. 20-25 June 1992. Pyrak-Nolte L. J. And Nolte D. D. Frecuency dependence of fracture stiffness. Geophys. Res. Lett. 3(19), 325-328 (1992). Chacón E., E. Lagos, y L. Quiñones, Plan Maestro Proyecto Pre-Acondicionamiento Macizo Rocoso Primario, Informe interno División Andina / IM2 (2000) Chacón E., y V. Barrera Plan Maestro Proyecto PreAcondicionamiento Macizo Rocoso Primario, Informe interno, División El Salvador / IM2 (2001) INFORME FINAL PROYECTO IM2 18/99 – 073/01 - 251 - Revisión A Jeffrey R. Hydraulic Fracture Test at El Salvador Mine Internal Draft Report, División El Salvador / IM2 (2002) Blair D., Minchinton A. (1997), "On the damage zone surrounding a single blasthole",International Journal of Blasting and Fragmentation (1): 59-72. Daehnke A., Rossmanith H.P., Schatz J.F., "On dynamic gas pressure induced fracturing", International Journal of Blasting and Fragmentation (1): 73-97. Hustrulid W. (1999), "Blasting principles for open pit mining. Volume 2. Theorical foundations", A.A. Balkema, Rotterdam, 1013p. Itasca Consulting Group (1998), "FLAC: Fast Lagrangian Analysis of Continua, Users manual, Version 3.4", Minneapolis, Minnesota. Kouzniak N., Rossmanith H.P. (1998), "Supersonic detonation in rock mass – Analytical solutions and validation of numerical models – Part 1: stress analysis", International Journal of Blasting and Fragmentation (2): 449-486. Liu L., Katsabani P. D. (1997), "A numerical study of the effects of accurate timing on rock fragmentation", Int. J. Rock Mech. Min. Sci. Vol. 34, No 5, pp. 817-835 Lorig L. (1997), "Fundamentals of mining geodynamics", SIMIN ’97 (Proceedings of the Conference, August 1997), Universidad de Santiago de Chile, Departamento de Ingeniería de Minas, Santiago, pp 161-174. Minchinton A:, Lynch P. M. (1997), "Fragmentation and heave modelling using a coupled discrete element gas flow code", International Journal of Blasting and Fragmentation (1): 41-57. Rossmanith H.P., Uenishi K., Kouzniak N. (1997), "Blast wave propagation in rock mass – Part I: monolithic medium", International Journal of Blasting and Fragmentation (1): 317-359.
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Remote high resolution stress change monitoring of hydraulic fractures K.W. Mills, SCT Operations Pty Ltd, Australia R.G. Jeffrey, CSIRO Petroleum, Australia
Abstract This paper describes the use of strain gauge based borehole instruments to monitor stress changes associated with the creation and extension of hydraulic fractures in massive rock strata at Northparkes Mine in Australia and Salvador Mine in Chile. This work was conducted as part of the International Caving Study ICSII. These instruments proved very sensitive to the stress changes induced by the hydraulic fractures close to the fracture plane. Analysis of the stress changes observed allowed the fracture orientation and non-symmetric fracture growth to be constrained sufficiently that a clearer insight into fracture behaviour could be obtained at both sites, particularly when combined with other observations. Recognition of the elastic stress reorientation about an opening mode hydraulic fracture has proved to be an important element in the interpretation of stress change monitoring data. The nature of the stress reorientation is useful in discriminating between opening and shearing mode fracture growth. A technique of identifying a range of possible solutions of fracture orientation and non-symmetric fracture growth consistent with the stress changes observed on multiple instruments has been developed. Unique definition of fracture orientation from the stress change instruments is possible if the instruments are sufficiently distributed relative to the hydraulic fracture plane.
1 INTRODUCTION This paper describes the use of strain gauge based borehole instruments to monitor stress changes associated with the creation and extension of hydraulic fractures in massive rock strata at Northparkes E26 Mine in Australia and Salvador Mine in Chile. The stress change monitoring described in the paper was one of several monitoring systems used to measure the behaviour of hydraulic fractures at two field sites as part of the International Caving Study ICSII. The stress change monitoring described is a new method to monitor hydraulic fracture growth to obtain stress change and fracture orientation information. Further details of the other work conducted at the two sites are presented elsewhere in these proceedings (van As et al. 2004, Chacon et al. 2004). The operation of the stress monitoring instruments, their in situ calibration and analysis procedure are common to both sites and these are described first. The installations, results and implications for the hydraulic fracture behaviour are then described for each site.
Figure 1 (a): Epoxy cement being applied to one of the two ANZI cells installed in each hole.
2 INSTRUMENTATION AND ANALYSIS PROCEDURE ANZI stress cells are strain gauge based stress change monitoring instruments. Their operation is described in detail by Mills (1997). Each instrument comprises eighteen electrical resistance strain gauges of various orientations. The instrument is internally inflated using air pressure to press the strain gauges into contact with the rock until an epoxy cement coating applied to the outside of the instrument has cured. Figure 1 shows a photograph of one of the instruments during installation at Salvador Mine. ANZI stress cells are able to be tested in situ prior to the commencement of monitoring, and subsequently if required, to check their correct operation and determine the equivalent stiffness of the rock into which they were installed. This process gives a field calibration that takes into account cable lengths, temperature effects and the data logging system. The internal pressure of the instruments is incremented in stages from the initial set pressure and back again while the instruments are being continuously logged. In each Massmin 2004
Figure 1 (b): Installation of the instrument into the monitoring hole case a high initial set pressure was necessary to inflate the instrument against the hydraulic head in the water filled
Santiago Chile, 22-25 August 2004
547
borehole plus provide sufficient pressure to bond the strain gauges to the rock. Figure 2 shows the pressure test results for one of the instruments installed at Northparkes as an example. The pressure test shows that the instrument is operating correctly. The six circumferential gauges go into tension and the axial gauges going slightly into compression as the pressure was incremented. Any individual gauges that are identified as being not properly bonded to the rock or bonded across joints can be ignored in the subsequent analysis.
Figure 2: Pressure test results - NPK7. The thick black line in Figure 2 indicates an equivalent stiffness or modulus for the instrument. The monitoring results have been analysed using this equivalent modulus. It should be recognised that the equivalent modulus can not be determined with a high degree of precision because of the low strain magnitudes involved. The effect of an error in the modulus reflects on the magnitude of the stress changes determined but not the orientation of the stress tensor. The strain changes measured on the data logger during each hydraulic fracture treatment were recorded at a resolution of 1µV and a signal noise level of about 5µV with occasional electrical interference spikes of 100µV. The data was prepared, before analysis, in four stages. A reference gauge located within each instrument but not subject to any strain changes was used to eliminate systematic strain changes in the cable and data logger system. Electrical spikes were then removed from the record and replaced with the strain values from the previous scan. A triangular, moving average filter of two minute duration was applied to the smooth out random variation. The stress changes were then determined every 10 minutes by averaging strains over a four minute interval. The determination of the stress field uses a standard multiple linear regression analysis. Multiple strain readings are analysed statistically to give a best fit estimate of the stress field. This process is standard for reduction of borehole strains to determine stress changes. An important characteristic of this process is that a statistical correlation between redundant strain gauges gives an indication of the confidence that can be placed in each result.
Figure 3: Location of stress change monitoring instruments at Northparkes.
3 STRESS CHANGE MONITORING AT NORTHPARKES MINE
Three ANZI stress cells were installed in a single BQ hole drilled from Drill Cuddy 5 on 9700 Level. The instruments were installed at depths of 83m, 76.3m and 69.7m respectively in a borehole dipping at 39° from horizontal at an orientation of 112°-115°GN. The hydraulic fracture treatments were conducted in a hole collared 0.75m from the stress monitoring hole, dipping 56° and oriented at 119°GN. Table 1 summarises the timing of the hydraulic fractures and the injection details. Injections not listed either did not result in breakdown or were associated with reopening existing fractures to allow intersections in monitoring holes to be located. The stress cells were not monitored during these injections. The three stress monitoring instruments are located 1448m above and 5-20m laterally from the initiation point of the hydraulic fractures. The full three dimensional stress field had previously been measured at the site using ANZI stress cells and the overcoring method of stress relief (Mills 2002). The results of these measurements indicated that at this site the major horizontal stress is dipping 5° at 132°GN with the minor principal stress dipping 67° at 235°GN. The vertical stresses are slightly less than the anticipated weight of overburden consistent with the proximity of the site to the extraction in the overlying block cave.
3.1 Site Description Figure 3 shows the locations of the stress change monitoring hole, the instruments and the hydraulic fracture treatments conducted at Northparkes Mine.
3.2 Results Figures 4 and 5 summarise the results from one instrument (NPK5) for two treatments, Fracture 1 and Fracture 7. Both these treatments were conducted at the
548
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Table 1: Summary of the Hydraulic Fracture Treatments Frac
Date
Injection Period
Straddle Interval (m)
Injection Fluid
1
25/09/2002
11:09-11:45
79.76-80.26
Water
3
27/09/2002
12:56-13:37
99.36-99.86
Water
4
28/09/2002
17:42-18:12
93.76-94.26
Water
5
29/09/2002
10:36-11:08
96.56-97.06
Water
6
29/09/2002
16:19-17:02
110.56-111.06
Water
7
1/10/2002
9:50-10:34
79.76-80.26
x-linked gel
13
3/10/2002
9:24-10:02
96.56-97.06
water/broken gel
Figure 4: Stress changes measured by NPK5 during Hydraulic Fracture 1 at Northparkes. same location in the injection borehole, one using water as the injection fluid and the other using cross-linked gel. Each of the figures is plotted on a consistent time base. Figure 4a is a record of the hydraulic fracture treatment. Figure 4b is a record of the strain readings on all the strain gauges after pre-processing to remove electrical spikes and application of the moving average filter. Figure 4c shows the vertical stress change calculated for the strains averaged over a four minute interval every ten minutes. Figure 4d shows the horizontal stress change components plotted as they would appear if projected onto a horizontal plane at the same stress scale as the vertical stresses. These stress monitoring results show a high degree of internal correlation with the correlation coefficient approaching 1.00 once the stresses begin to change. Other characteristics that increase confidence in the results are: • The timing of the stress changes correlates closely with the start and end of the hydraulic fracture treatments. • The orientation of the stress changes is consistent across all three instruments. • The alignment of the stress vectors appears to be broadly consistent with the expected orientation of the hydraulic fracture. Massmin 2004
In each treatment there are four clearly defined stages. There is a steady state of essentially zero stress change before the treatment starts. The vertical stress increases almost immediately after each treatment commences. Changes in strain are apparent within one minute of the commencement of pumping. A peak is reached and the stress change induced in the rock remains steady. The peak is typically in the range 0.7-1.4MPa for the water treatments at a distance approximately normal to the fracture plane of 15-30m. When this distance is greater than about 30m, the peak vertical stress is consistently lower at 0.5-0.8MPa. In the gel treatment (Fracture 7), the peak vertical stress indicated is 2.3-3.3MPa with the stress change decreasing with distance away from the fracture plane. Once pumping stops, there is a gradual decay in pressure. The rate of the pressure decrease reduces over time in the form of a classical decay curve. Logging typically only continued for 1-2 hours after pumping stopped. By the time logging is discontinued, the residual vertical stress is in the range 0.4-0.6MPa for the water treatments and 1.12.1MPa for the gel treatment.
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549
Figure 5: Stress changes measured by NPK5 during Hydraulic Fracture 7 at Northparkes. Table 2 summarises the maximum principal stress orientation for the last stress change calculated prior to the cessation of pumping for each treatment that was monitored. These measurements give an indication of the hydraulic pressure in the fracture and the orientation of the hydraulic fracture. The observation that the major stress change is subvertical, while the horizontal stress changes are small by comparison, is consistent with the hydraulic fractures forming on a sub-horizontal plane. A plane of this orientation is consistent with the measured in situ stress field at the site and further corroborated by borehole intersection and tiltmeter measurements. 3.3 Fracture Orientations Based on Stress Orientations The orientation of the major stress change vectors is expected to provide an indication of the dip and dip direction of the hydraulic fractures once the fracture becomes large relative to the distance of instruments from the fracture initiation point.
Figure 6 shows the compressive principal stress vectors that would be expected from modelling of a hydraulic fracture in an elastic, isotropic, homogeneous half space. Close to the centre of the hydraulic fracture, the compressive stress change component is oriented normal to the fracture plane. However, toward the fracture tip and at greater distance from the fracture plane, there is a tendency for rotation of the stress change component outward and therefore away from the fracture plane. The hydraulic fractures at Northparkes are estimated to have grown to a maximum radius of 30 to 50m. Assuming a 40m radius fracture, the stress cells are located on the diagonal shown in Figure 6 at the time of maximum fracture extent (assuming the hydraulic fractures are subhorizontal). A process of matching the stress changes measured with the stress changes modelled allows a number of admissible orientations to be determined. The number of admissible orientations based on the stress change measurements is a function of the spatial relationships of the stress change
Table 2: Summary of Measured Stress Changes at Nearest Time Period Analysed Prior to Maximum Extent of Each Hydraulic Fracture Being Reached Frac
550
Time
NPK5
NPK6
NPK7
Stress (MPa)
Dip /Brg
Stress (MPa)
Dip /Brg
Stress (MPa)
Dip /Brg
1
11:40
1.41
74/244
0.93
81/212
0.69
86/253
3
13:30
1.27
75/103
1.08
64/114
0.50
60/117
4
18:10
1.43
80/65
1.34
71/100
0.78
68/109
5
11:00
1.28
77/73
1.22
67/98
0.61
66/105
6
17:00
0.84
64/133
1.38
53/125
0.57
51/123
7
10:30
3.27
67/291
2.90
77/301
2.25
77/261
13
10:00
1.59
86/122
1.67
71/113
0.92
72/123
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Massmin 2004
Table 3: Admissible Hydraulic Fracture Orientations on a Plane Striking 117°GN Based on Stress Change Measurements
Figure 6: Compressive stress vectors about a horizontal hydraulic fracture. instruments to the hydraulic fracture, the size of the hydraulic fracture and the symmetrical or otherwise growth of the hydraulic fracture about the initiation point. Table 3 summarises some of the admissible orientations indicated by the stress change measurements for each of the hydraulic fracture treatments. These orientations are based on the assumption that the fracture plane passes through the injection point (i.e. there are no bypasses through other holes or fracture cross-over). While there are other combinations of fracture size, fracture orientation and non-concentric growth that would fit the stress orientation data from each measurement, other independent information narrows the possibilities. For instance, a fracture radius of about 40m is indicated by intersection and modelling data, so this is assumed as a first pass. The fracture orientation that would fit the stress orientation assuming concentric fracture growth is shown in the first instance. The fracture orientation that would fit assuming some non-symmetric growth of the hydraulic fracture about the injection point is also shown for a fracture plane that is dipping slightly to the east. The orientations that would be consistent with other different sized fracture are also shown. The results presented in Table 3 suggest that in general the stress changes measured support an east dipping fracture only if there is some possibility of non-concentric fracture growth. Only the stress changes observed during Fracture 7, the crosslinked gel treatment, are consistent with an east dipping fracture without there being a requirement for some non-concentric (or possibly noncircular) fracture growth. In a pre-mining stress environment, the vertical stress would tend to increase with depth and therefore the preferred fracture growth direction would be to the west (up dip in an east dipping fracture), although it is recognised that the effect would be small in a shallow dipping fracture. However, at the trial site, the extraction of the Lift 1 orebody above is expected to have significantly lowered the vertical stress to the east of the trial site. In fact, measurements of fracture shut-in pressure, after each treatment at the site show a reduction in the shut-in pressure with increasing depth down the hole. This data implies a stress gradient exists with lower stress occurring down dip and to the east (van As et al., 2004). Therefore, it is considered quite likely that a hydraulic fracture would preferentially grow toward this lower stress (i.e. in an easterly direction). Massmin 2004
Frac No
Assumed Radius (m)
Fracture Offset To East 1 (m)
Plausible Fracture Orientation (°)
12
40 40
0 8
7°W 2°E
3
40 40 25
0 26 21
50°W 10°E 3°E
4
40 40 25
0 26 20
42°W 7°E 10°E
5
40 40 25
0 26 17
47°W 7°E 10°E
6
40 40 60
0 24 0
No fit possible 30°E 60°W
7
40 40 60 25
0 14 0 No fit possible
7°E 29°E 11°E
13
40 40 25
0 20 7
34°W 5°E 50°W
1 The fracture offset is a measure of the degree of nonsymmetric growth of an assumed circular fracture relative to the injection point along the plane of the fracture. 2 There is significant component of dip out of the projection plane for Fracture 1 so the dips in the plane are less meaningful. If only concentric fracture growth is assumed, then the stress change orientations from all the deeper treatments (all those except Fracture 1 and Fracture 7) are not consistent with east dipping hydraulic fractures. The data would only be consistent with west dipping fractures at dips of between 34° and 50°. There does not appear to be any other observations that support fracture growth of this orientation so the implication is that the water fractures did not grow symmetrically about the injection point. Independent measurements using tiltmeters (van As et al. 2004) indicate that fractures formed were sub-horizontal with an east dip, but the size and shape of the fractures cannot be determined independently from the tiltmeter data. Fracture 7 is not particularly sensitive to the size of the hydraulic fracture and this fracture appears to have grown concentrically about the injection point. The stress change data supports a fracture radius of about 40m, but the fracture radius could be as low as about 30m (no match was possible at 25m) or upwards of 60m. The concentric growth is thought likely to be a consequence of the higher viscosity of the cross-linked gel fluid. Principal stress change magnitudes shown in Table 3 indicate that the water injections generate smaller stress changes in the rock mass than did the gel injections as would be expected. It should be noted that the hydraulic pressure in the fracture may be slightly greater than
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551
indicated because of the distance the monitoring points are from the plane of the hydraulic fracture. 4 STRESS CHANGE MONITORING AT SALVADOR MINE CHILE 4.1 Site Description Figure 7 shows the layout of the monitoring site at 2600 Level in Inca East sector at Salvador Mine. The locations of the stress change monitoring holes, the instruments and the hydraulic fracture treatments are shown. Four ANZI stress cells were installed in two BQ boreholes. These instruments were installed at depths of 37.39m and 39.90m in borehole S1 and 41.13m and 44.31m in borehole S2. Hole S1 dips at 59° from horizontal at an orientation of 85°GN. S2 dips 58° from horizontal at an orientation of 89°GN. The hydraulic fracture treatments were conducted in a hole HF02 collared midway between the stress monitoring holes. This hole also dips 59° and is oriented at 87°GN.
Figure 8: Stress changes measured by CODI during Hydraulic Fracture 10 at Salvador Mine (Chile). Table 4 summarises the timing of the hydraulic fractures and the injection details. Figure 7: Location of stress change monitoring instruments at Salvador Mine (Chile) 552
4.2 Results Figure 8 summarise the stress changes measured during Fracture 7 on COD1.
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Table 4: Summary of the Hydraulic Fracture Treatments Frac
Date
Injection Period
Straddle Interval (m)
Injection Fluid
1
26/11/2002
10:50-11:20
57.55-58.05
Water
2
27/11/2002
14:44-15:15
66.55-67.05
Water
3
28/11/2002
11:28-11:52
51.55-52.05
Water
4
29/11/2002
11:13-11:30
51.55-52.05
X-linked gel
5
29/11/2002
14:22-14:45
54.55-55.05
Linear gel
6
30/11/2002
10:56-11:16
60.55-61.05
Water
7
1/12/2002
13:07-13:25
63.55-64.05
X-linked gel
8
2/12/2002
16:12-16:41
117.55-118.05
X-linked gel
9
3/12/2002
13:13-15:00
111.55-112.05
Linear gel
10
4/12/2002
12:08-12:31
57.55-58.05
X-linked gel
The stress monitoring results show a high degree of internal correlation with the correlation coefficient approaching 1.00. The observation that the major stress change is subhorizontal, while the vertical and other horizontal stress changes are small by comparison, is consistent with the hydraulic fractures forming in a sub-vertical plane. There is corroborating evidence from other measurements made at the site that this is indeed the case. The hydraulic fractures at Salvador Mine are estimated to have grown to a maximum radius of approximately 40 to 50m based on the length of injection and the timing of various intersections. The principal stress changes are expected to be approximately normal to the plane of the hydraulic fracture when the fracture is at maximum extent. Table 5 summarises the maximum principal stress orientation for the last stress change calculated prior to the cessation of pumping. These measurements give an indication of the maximum hydraulic pressure in the fracture and the orientation of the hydraulic fracture. No stress changes were perceptible in the results for Fractures 8 and 9. These treatments were conducted at a much lower horizon for other purposes. The correlations on some of the results from COD4 were too low to give meaningful indications of the stress orientations. Assuming that the maximum stress changes are oriented approximately normal to the plane of the hydraulic fractures, the measured stress changes indicate that all the hydraulic
fractures except Fracture 7 are dipping to the west (normal to the plane oriented at 87°GN) at about 15° from vertical. The fracture plane orientations appear to be much more consistent in the vicinity of stresscells COD1 and COD2 than in the vicinity of the other two instruments. Nevertheless, the average orientation indicated by COD3 and COD4 is still essentially the same as indicated by COD1 and COD2, and the scatter may be a result of more variable behaviour in the rock mass in the vicinity of COD3 and COD4. However, other measurements indicate only limited growth of the hydraulic fractures occurred to the south of HF02 (Chacon et al., 2004) The different orientation observed in Fracture 7 is apparent in all four instruments. Fracture 7 was a new hydraulic fracture using cross-linked gel as the injection fluid and it would appear from the stress monitoring results that it grew predominantly in a northerly direction toward COD1 and COD2. Fractures 1, 2, 3 and 6 are all water treatments and the maximum stress change from these treatments are all closely aligned dipping 12° at 80°GN ± 4° in the vicinity of COD1 and COD2. Fractures 4, 5 and 10 also show similar alignment to each other. These are either linear gel treatments (Fracture 5) or cross-linked gel treatments injected into pre-existing hydraulic fractures. These tend to be aligned more easterly with the stress changes dipping 13° at 90°GN. Fracture 7 is a cross-linked gel treatment injected into a previously
Table 5: Summary of Measured Stress Changes (_1) at Nearest Time Period Analysed Prior To Maximum Extent of Each Hydraulic Fracture being Reached Frac
Time
COD1
COD2
COD3
COD4
σ1 MPa
Dip /Brg
σ1 MPa
Dip /Brg
σ1 MPa
Dip /Brg
σ1 MPa
Dip /Brg 4/80
1
11:20
0.47
11/80
0.93
12/84
0.85
12/90
1.0
2
15:10
0.95
14/79
0.97
11/84
0.54
8/78
-
3
11:50
0.97
15/85
1.2
16/89
0.90
22/110
1.6
6/113
4
11:30
1.8
15/90
2.1
13/94
1.6
11/103
2.1
3/107
5
14:40
1.5
15/92
1.7
12/96
0.39
23/61
0.98
29/75
6
11:10
0.92
17/81
0.90
13/84
0.67
29/73
0.46
12/56
7
13:20
0.59
49/85
0.91
43/86
0.34
37/50
0.42
42/38
8
16:40
0
-
0
-
0
-
0
-
9
15:00
0
-
0
-
0
-
0
-
10
12:30
1.8
12/89
2.0
9/92
1.7
7/87
2.0
9/71
Massmin 2004
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untreated section of the borehole. This fracture dips at 30° from vertical with the normal aligned at 87°GN. The average results for all four instruments indicate westward dipping hydraulic fracture that dip at an average of about 15° from vertical. 5 FRACTURE GROWTH AND STRESS CHANGE The stress change around a hydraulic fracture can be obtained from analytical solutions for certain fracture geometries. The stress change around a circular or pennyshaped fracture that is uniformly pressurised (Sneddon, 1946) has been used in this section to compare the modelled and measured stress change. If the fracture growth is assumed to occur symmetrically about the injection point, then the stress change at the location of any instruments is easily found as the fracture grows from an initially small radius to a radius much larger than the distance separating the instruments from the injection point. The stress change for such symmetric growth is found by tracking along a straight line running from the centre of the model fracture outward at an angle to the fracture plane that passes through the location of the instrument. This line is shown in Figure 6. When the fracture is small relative to the distance to the instrument, the stress change corresponds to points located at large r/R on this line. Conversely, as the fracture grows r/R decreases.
Figure 9: Stress changes measured and modelled for Fracture 10 at Salvador Mine (Chile).
The stress change can be calculated along such a line and both the stress change magnitude and orientations can be compared to the measured stress changes. Such a calculation has been carried out for Fracture 10 at Salvador and is shown in Figure 9. The relative distance r/R has been translated into a time by using the fracture growth relationship established by direct measurement of fracture growth at this site. The growth relationship is:
R = 7.56t 0.55
(1)
Equation 1 can be rewritten as: 1
R 0.55 t = 7.56
6 CONCLUSIONS The stress changes observed are consistent in terms of timing and magnitude with the commencement and cessation of pumping in all the hydraulic fracture treatments. The stress changes orientations are consistent across the instruments at both sites when elastic stress distributions about hydraulic fractures are taken into account. At Northparkes, the initial stress monitoring results indicate that the hydraulic fractures have formed in a subhorizontal plane, consistent with the in situ stress measurements made at the site using the overcoring method of stress relief. The orientations of the hydraulic fractures are not able to be uniquely defined using only the stress monitoring information because of the locations of the instruments relative to the fracture plane and the directions that the fractures have grown. Nevertheless, the stress change monitoring constrains the possible fracture orientations to only a few possibilities. Using other information, these possibilities are further constrained to give a unique result. The measurements indicate that all the hydraulic fractures dip gently to the east at an orientation consistent with the in situ stress field at the site, but the fractures have grown non-symmetrically in a down dip direction consistent with the stress geometry expected about the overlying block cave. The magnitude of the stress change measured is consistent with the nature of the injection fluid. For water treatments, the magnitude of the stress change observed was about 0.5-1.4MPa for distances of 15-40m from the fracture plane. For the cross-linked gel, the stress changes measured were 2.3-3.3MPa at 15-25m from the fracture plane. The pressures locked into the fracture at the completion of the treatment were also reflective of the nature of the injection fluid. In water treatments, the residual stress change observed several hours after the treatment was complete was typically about 0.5MPa (0.2-0.6MPa). For the cross linked gel treatments, the residual stress change was about 1.5MPa (1.0-2.1). At Salvador Mine, the initial stress monitoring results indicate that the hydraulic fractures are generally westward dipping at an average dip of about 15° from vertical. A detailed model of the stress change associated with Fracture 10 produced a good fit to the measured data and illustrated the sensitivity of the orientation of the maximum stress change vector to fracture growth relative to the position of the stress change monitoring instruments. ACKNOWLEDGEMENTS
(2)
The model results are calculated in terms of r/R, where r is the distance between the injection point and the location 554
at which the stress change is measured. Figure 9 compares the measured and modelled data after applying equation 2 to obtain a time from the start of fracture growth for the modelled data. The model assumes the fracture strikes north-south and dips at 80o to the west. A uniform pressure of 2MPa was used inside the model fracture to obtain the fit shown. Pressure falloff after shut in is not modelled. The fit to the stress magnitudes is reasonable considering the assumptions used and the orientation fit to the orientation of the maximum stress change is quite good and also shows the sensitivity of this measured parameter to fracture growth and flow back. The magnitude of the secondary principal stress change, which corresponds to the north-south stress change, is overestimated by the model suggesting the Poisson’s ratio used for this calculation may be too high.
The fieldwork described in this paper was conducted as part of the International Caving Study (ICSII) with additional support provided by Salvador Mine and Northparkes Mine. The analysis of the stress change data undertaken to date
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has also been supported by CSIRO Petroleum and SCT Operations Pty Ltd. The authors gratefully acknowledge the support of all of these groups. REFERENCES • Chacon, E, Barrera, V, Jeffrey, RG, and van As, A, 2004. Hydraulic fracturing used to precondition ore and reduce fragment size for block caving, Proceedings of the MassMin 2004 symposium, Santiago, August 22-25. • Mills, KW, 1997. In situ stress measurements using the ANZI stress cell, Proceedings of the International
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Symposium on Rock Stress, edited by Sugawara and Obara, Kumamoto 7-10 October 1997, published by A.A. Balkema. pp 149-154. • Mills, KW, 2002. SCT Report to Northparkes Mine In Situ Stress Measurements – 9700 Level and Lift 2 Decline. • Sneddon, IN, 1946. The distribution of stress in the neighbourhood of a crack in an elastic solid, Proceedings Royal Soc. Of London, 229-260. • van As, A and Jeffrey, RG, 2004. Preconditioning by hydraulic fracturing for block caving in a moderately stressed naturally fractured orebody, Proceedings of the MassMin 2004 symposium, Santiago, August 22-25.
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Chapter 15
Seismicity & Rock Bursts
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Design and implementation of seismic monitoring systems in a block-cave environment Hylton White, Willem de Beer, Hamish White, iGeo Ltd, P O Box 58-598, Greenmount, Auckland, New Zealand Andre van As, AVA Mine Geotechnical Services, Kenmore, Queensland David Allison, Northparkes Mines, Australia
Abstract Seismic monitoring provides quantitative information on the rock mass response throughout the various stages of the block-caving lifecycle. This includes the monitoring of: • seismicity which may effect the stability of excavations during the early development and construction phase, • pillar stability and abutment stresses during the undercutting phase, • seismically active structures, • cave propagation and, • subsidence effects (i.e. cave angle) as the cave walls break back, possibly influencing surrounding excavations. Due to the significant role that seismic monitoring can play during the various phases of the cave lifecycle, it is imperative that mine development and seismic system installation be done in parallel. Consequently, system design must be undertaken very early in the mine-planning stage, with close interaction between mine planners and seismologists to ensure adequate access for the system installation. System design must consider all aspects of the evolution of the mine, including relocation of subsystems as requirements change, as well as shadowing effects of the cave itself. This includes sensor selection (geophone and/or accelerometers), sensor placement (which controls location accuracy and system sensitivity), communications and cabling requirements, hardware and software selection, and system documentation and reporting procedures. In particular, it is imperative that documentation systems (both during installation and operation) provide a traceable and auditable record of the seismic system performance and operational status in addition to recording the mine seismicity. This is of paramount importance in the mining environment, as system performance (availability of seismometers and sensors) can influence the perceived levels of seismicity. Such systemrelated effects must be taken into account to ensure the accurate quantitative interpretation of seismic source parameters related to the mining activity and rock mass behaviour.
1. INTRODUCTION
Seismic monitoring allows one to quantify the rock mass response to mining, and can yield the following information:
Seismic monitoring provides a direct method for rock mass characterization during the various stages of mining in a blockcaving environment. Section 2 of this paper discusses the information that can be gleaned from seismic monitoring at the different stages of mining. Section 3 discusses the system design and implementation principles essential to achieving a robust monitoring environment, both from a Health and Safety, and a rock mass characterization perspective. 2. WHAT DOES SEISMIC MONITORING ACHIEVE? Specific objectives of seismic monitoring in a block-caving environment include: • Monitor seismic events related to induced stress concentrations on the undercut level, extraction level pillars and major excavations in a block cave mine at different stages of mining. • Monitor seismic events related to cave propagation and the definition of the active seismogenic zone above the cave back. • Monitor seismic events related to preconditioning or possible cave induction. Massmin 2004
2.1. Location and magnitude of seismic events At the very least, seismic monitoring indicates the magnitude and location of seismic events within the mine. This is the most common application for seismic monitoring; however magnitude is at best a summary number that conveys the relative intensity of a seismic event. 2.2. Largest Likely Event Seismic event magnitudes can be quantified in terms of the Gutenberg-Richter Distribution (GRD). In equation form this reads: LogN (≥ mL) = a – bmL
(1)
N (≥ mL) is the expected number of events greater than local magnitude mL, and a and b are constants (Gutenberg and Richter (1949)). Equation (1) states that a seismic event larger than magnitude mL is 10b times more likely to happen than a seismic event with magnitude larger than mL + 1. Typically, b U1.5. Therefore, an event of magnitude would be times more likely to occur than an event of magnitude
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mL ≥ 0 would be 101.5 U 31 times more likely to occur than en event of magnitude mL ≥ 1. The value where N = 1, is the (extrapolated) one event with expected largest magnitude. In other words, if everything else stays the same, this provides a measure of the possible "largest" event to be expected within the mine.
Figure 1: Gutenberg-Richter distribution of seismic activity. The extent of the linear best fit line (mL = 2.6 and mL = 0.5) provides the magnitude limits in which seismic events are accepted for further analysis.
2.3. Rock mass stiffness Apart from magnitude, a seismic event can be further quantified in terms of numerous "source parameters". Among these are the Radiated Seismic Energy and the Seismic Moment . It is possible to fit a straight line to the energy-moment distribution: LogE = dLogM + c
(2)
Seismic Moment provides a measure of the deformation associated with a seismic event, while Radiated Seismic Energy provides a measure of the stress release at the source. Therefore the -value can be viewed as a modulus or stiffness (Mendecki et al (2000)). The slope of the GRD (b-value) indicates the proportion of small to large events and is also an indicator of system stiffness. An increasing d -value together with an increasing b-value indicates a stiffening regime, where stress release is by comparatively more small fractures. Conversely, a decreasing d-value together with a decreasing b-value slope may indicate a softening regime, where stress release is by comparatively more large fractures. Figure 2 shows the changes in b-value, d-value and activity rate as a function of time for an 11 months period of monitoring a block cave. The overall character of the rock mass response in Figure 2 has a cyclical pattern of ‘stiffening’ followed by ‘softening’, mimicking the phases of mining. Initiation of the undercut began at the very end of February 2003. Table 1 describes the rock mass behaviour after the undercut was initiated.
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Table 1: Rock mass behaviour during undercutting Period mm/yy
Observation
03/03
d - and b -values oscillate, not necessarily in phase: First changes in stress redistribution
04/03 – 05/03
Substantial increase in d - and b values, and activity rate. Increase in rate of cave development. The initial redistribution of stress in the intact rock surrounding the cave causes the stiffness to increase.
05/03 – 09/03
d - and b -values decreasing. Reduction in stiffness accompanied by events having comparatively greater seismic moments (i.e. larger deformations).
10/03
d - and b -values increasing, coincident with ramp-up in undercut firing.
11/03 – 02/04
d - and b -values decreasing together with activity rate as undercut firing ceases.
2.4. Time series Several other source parameters can be used to monitor the rock mass response to mining. These include the Cumulative Apparent Volume (CAV) and Energy Rating. The CAV provides a measure of the deformation within the rock mass. It is intuitive that generally the greater the amount of seismic activity, the more deformation there will be. This is evidenced in Figure 3 by increasing CAV coincident with peaks in activity rate. Energy rating gives a measure of the amount of energy released by a specific seismic event having a given moment Table 2: Characterisation of rock mass response based on comparison of energy rating and CAV as a function of time (derived from Mendecki (1996) and Mendecki et al (2000)). Energy Rating
Rate of change of CAV
Characterization
Increasing
Little or no change / decreasing
"Brittle" or very stiff
Increasing
Constant, higher rate of change
Creep / soft: Linear regime where stress release is accompanied by concomitant deformation
Decreasing
Increasing
Plastic: Successively smaller stress drop – events are accompanied by increasing deformation
Decreasing
Little or no change
Quiet or stress build up. To be compared with system availability
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Figure 2: Example plot of b-value, d-value and activity rate for an 11 month period. All trends have been derived from seismic events larger than for the whole mine volume of interest. In this example, undercut advancement firing began on 27 February 2003.
Figure 3: A plot of activity rate, CAV and energy rating for the undercut and extraction volume of interest, for a period of 17 months. The activity rate and energy rating trends have been calculated in a moving 14-day window, in increments of 12 hours. The minimum magnitude ‘filter’ is set at . as compared to the average energy radiated by events of that moment. An increase in energy rating suggests larger than expected stress releases are being observed for a given amount of deformation. Conversely, decreases in the energy rating are indicative of lower than expected stress releases. In the most basic terms CAV is a strain related parameter, whilst energy rating is a stress related parameter. Essentially, the relationship between the trends of CAV and Massmin 2004
energy rating as a function of time indicate different regimes of rock mass response (see Table 2). 2.5. Post-firing activity analysis Post firing activity analysis shows: • The decay time of seismic activity after firing sessions in a volume surrounding the undercut and extraction levels as a method for statistically quantifying the re-entry time into the mine;
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Figure 4: Activity rate (per 10 minutes) for 6 hours after undercut advancement firing, comprising 666 events. The curve fitted to the decay is approximated by .
• The difference in decay time of seismic activity after blasting between phases of mining activity. An example is shown in Figure 4, where the activity rate is plotted vs. time. Here a large number of blasts have been stacked together, to give a statistically relevant indication of activity rate immediately after blasting. This can be used to determine safe re-entry times.
2.6. Peak particle velocity and magnitude Since magnitude is merely a summary parameter by which to characterize a seismic event, it is useful to find a relationship between the damage potential of a seismic event and its magnitude. In practice, this means finding a relationship between the magnitude and estimated peak particle velocities (PPVs) at different distances from the event. A direct mathematical relationship between the
Figure 5: Plot of relative changes in stress from a hollow inclusion cell ( , , ) including the energy rating trend, for a period of two weeks. There appears to be a correlation between decreases in the energy rating slightly preceding an onset of changing stress, indicated by the thick black vertical lines. 562
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Figure 6: Plot of anchor node displacements for an extensometer for a period of three months, including CAV and energy rating. Only events larger than have been used in creating the CAV and energy rating trends. magnitude of a seismic event and PPV does not exist. However, it is possible to establish an empirical relationship between the energy, moment and the PPV.
a strength degradation (decreasing energy rating indicating lower stress release through seismic activity) in the seismically active zone.
2.7. Blast characterization Where and when firing takes place, it is possible to evaluate • Conformance to blast design; • The effect of the blast on nearby mine infrastructure.
3. SYSTEM DESIGN AND IMPLEMENTATION FOR A ROBUST MONITORING ENVIRONMENT
At minimum, data from sensors in the firing district and those close to infrastructure provide waveforms, PPVs, Peak Particle Accelerations and ground motion response spectra. 2.8. Seismic and geotechnical data sets Seismic sensors can be installed remotely from active mining areas – consequently it is practical to install a seismic system in the early stages of mine development. On the other hand, geotechnical sensors (e.g. extensometers and hollow inclusion cells) have to be installed after development as their measurements focus on the nature of the rock mass response surrounding the mine infrastructure. Consequently, geotechnical sensors will generally be installed much later than seismic sensors. It is now possible to integrate the data from seismic and geotechnical datasets in quasi-real-time. An example of this is shown in Figure 5 and Figure 6, where outputs from hollow cells and extensometers respectively are integrated with the seismic time series. 2.9. Analysis During the undercut progression, the advance of the undercut could be followed very clearly by visualising induced seismicity. Activity followed the line of advance, clustering mainly in the crown of the cavity. An initial redistribution of stress in the intact rock surrounding the cave caused the stiffness to increase and was followed by an apparent relaxation to an equilibrium phase. In turn, it appears as if this was interrupted by a ramp-up of development. Combination of seismic and geotechnical data points towards stress increases in undercut pillars correlated with Massmin 2004
3.1. Design System design must be undertaken very early in the mineplanning stage, with close interaction between mine planners and seismologists to ensure adequate access for the system installation. System design must consider all aspects of the evolution of the mine, including relocation of subsystems as requirements change, as well as shadowing effects of the cave itself. The system specification must ensure that (at all stages of mining) the required accuracy of event location and sensitivity (i.e. minimum event magnitude that can be reliably recorded) is achieved. It is possible to model both the sensitivity and the location error. Sensitivity can later be checked by referring to the Gutenberg Richter Distribution once sufficient events are recorded. In the example in Figure 1, the GRD can be seen to taper off for events having. 3.2. Installation System design for block caving must provide for future shadowing effects by the void. It is therefore necessary (as far as possible), to surround the ore body with sensors. This poses several challenges: • Several holes of hundreds of metres in length need to be drilled; • Sensors have to be lowered to the required depths and • Grouting performed in stages, with sensor tests at every stage. 3.3. Sensor selection Accelerometers have higher sensitivity than geophones at higher frequencies. The frequency at which accelerometers become more sensitive than geophones is usually of the order of a few hundred Hertz. Very small seismic events associated with caving have high frequency content and require accelerometers for accurate measurement.
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Nevertheless, it is common to have mixed networks comprising (for example) triaxial accelerometers (for source parameter calculation) and uniaxial geophones (for improved location accuracy).
than a specified amount (typically 3% of the average hypo central distance). Accepted events must also approximate expected values of moment and energy for their given local magnitudes.
3.4. Communications The communications infrastructure is potentially the single most expensive subsystem in a seismic monitoring system. Consequently, the design must contain sufficient redundancy to allow relocation of seismometers as required during the various stages of mining. In particular, it must be possible to accommodate unforeseen requirements.
4. SUMMARY
3.5. Documentation Because of limitations imposed by mine development, it is usual to phase the installation of sensors as mining progresses. The consequence of this is that system sensitivity changes with time as more sensors are installed, causing an apparent rise in activity rate. Documentation systems are put in place which ensures that apparent levels of seismic activity can be correlated easily with seismic system status and performance are essential. 3.6. Operation It is imperative that documentation systems (both during installation and operation) provide a traceable and auditable record of the seismic system performance and operational status in addition to recording the mine seismicity. This is of paramount importance in the mining environment, as system performance (availability of seismometers and sensors) can influence the perceived levels of seismicity. Such system-related effects must be taken into account to ensure the accurate quantitative interpretation of seismic source parameters related to the mining activity and rock mass behaviour. During normal mining operations, parts of the seismic system may be offline temporarily due to maintenance or because of inadvertent damage. Apparent activity rates will thus rise and fall depending on the system status. Careful management of the data is required to prevent false conclusions being drawn about the activity rate related effects During monitoring, an adequate amount of quality verification needs to be conducted. This includes regular (at least daily) detailed checks of system status. Any changes to monitoring parameters should be accompanied by a detailed analysis of possible consequences, and properly recorded in the system documentation. Data quality verification is undertaken at the processing stage. All accepted events must have a location error of less
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4.1. The outputs of seismic monitoring From a short term perspective, seismic monitoring enables up-to-date information to be provided on the state of the rock mass response. If these results are to be used as a means of determining procedures underground then the information must be available to mine staff in a readily updated form that presents the important facets of rock mass response quickly and clearly. The long term benefits of seismic monitoring include providing a knowledge base concerning the nature and pattern of rock mass response to mining in mass mining environments. These results form the constraints and empirical limits for numerical models of the caving process. As modelling of the mechanics of caving becomes more complex more detailed empirical information recorded from active mines will be necessary to verify these. 4.2. System design and implementation Seismic monitoring systems are by their nature extremely complex distributed systems. Careful evaluation of system issues is required when analyzing data, to ensure that correct conclusions are drawn regarding the rock mass response to mining. The scale and resolution of monitoring required in a mass mining environment means that system design and implementation must be an integral part of the mine development process. ACKNOWLEDGEMENT The authors wish to acknowledge the support provided by Northparkes Mines in the preparation of this paper. REFERENCES • Gutenberg, B. and C. F. Richter 1949. Seismicity and the Earth, Princeton University Press, Princeton. • Mendecki, A.J., ed. 1996. Seismic Monitoring in Mines. Chapman and Hall, London. • Mendecki, A. J., G. van Aswegen and P. Moutford 2000. Chapter 9: A Handbook on Rock Engineering Practice for Tabular Hard Rock Mines. A. Jager and J. A. Ryder, Ed., SIMRAC, Johannesburg.
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Seismic monitoring of block cave crown pillar – Palabora Mining Company, RSA Stefan Glazer, Palabora Mining Company, RSA Neil Hepworth, Somincor, Portugal
Abstract With mining operations taking place concurrently in the open pit and underground, it was considered prudent in mid 2002 to establish a 200m high crown pillar, or exclusion zone, between the base of the open pit and the cave back. This pillar had two functions, firstly safety of the open pit operations due to potential instability of the pit walls as a result of caving and secondly to prevent rapid ingress of rainfall water into the underground workings before adequate protection measures were in place. By the end of 2002 this pillar had become fractured and de-stressed to the extent that there was a hydraulic connection between surface and underground despite the indicated cave back being about 200m below the pit. This paper describes the history of this pillar as revealed by seismic data analysis, which became the principal means of monitoring the cave progress when other means and instrumentation were lost..
1 INTRODUCTION Initiation of the block cave undercut at a depth of approximately 1200m below surface, commenced at the end of 2000 and by the end of 2002 the area of undercut had reached 60,000m2 which is considerably more than that predicted to achieve caving. Despite significant rockmass fracturing around the undercut and in the cave back above the undercut, which by definition is associated with seismicity, up to the end of 2002 Palabora was not experiencing a seismic hazard in the working places. Seismicity was taking place remote from the production and development beneath and immediately around the cave footprint. By mid 2002, when the critical hydraulic radius of the producing area was reached and the caving process was initiated, the situation changed, as more of the seismic events started to concentrate on the geological discontinuities. This clustering of events was investigated and the seismically active faults and dykes were identified. At that time the depth trend of the larger seismic events was that the large events were migrating upwards away from the undercut level and the working places. As from the beginning of 2003 this situation changed and the depth trend of these relatively large seismic events was no longer upwards. During 2003 Palabora mine experienced several large seismic events with associated damage in the workings, which meant that Palabora become a seismically active mine. The aim of this paper is two fold. Firstly to document the observed changes to the character of the seismicity that took place and secondly to relate these changes to the mining and development operations in order to understand the likely cause of the changes. PALABORA MONITORING FACILITIES The Palabora copper body is an elliptically shaped, vertically dipping volcanic pipe. The pipe measures 1400m and 800m along the long and short axes. The ore reserves are proven to a depth of 1800m below surface. Copper grades of approximately 1% are found in the central core of the ore body and decrease gradually towards the peripheries with no sharp ore-waste contact. Three main rock types host mineralization. Trangressive and banded carbonatites form the central core of the ore body and are Massmin 2004
made up of magnetite-rich sövite with minor amounts of apatite, dolomite, chondrodite, olivine and phlogopite. Barren dolerite dykes with a steeply dipping northeast trend are present and account for about 8% of the ore reserve (Calder et al, 2001). The average uniaxial strength of the carbonatites is about 120 MPa. Dolerite is a strong, brittle rock with a uniaxial strength of up to 320 MPa. Adjacent to the major faults, dolerite is extensively jointed and locally weathered with marked reduction in strength to around 80 MPa. The underground mine exploits the ore below the open pit using mechanized block caving. The undercut level is at elevation –800m (1200m below the surface) and approximately 400m below the open pit base. The production level with its drawpoints is located 18m below the undercut. The Palabora cave back monitoring facilities during 2002 consisted of a digital seismic network and four open holes. In addition there were five TDR holes. By the end of 2002 all the open holes were lost due to movements along the dykes and faults. In fact the only reliable measurements of the cave back position and the expansion void height were achieved from one open hole located in the open pit. The other open holes were lost due to hole dislocations before the cave back had reached them. On the other hand the TDR holes measurements were very successful in confirming the cave back position and assumed swell factor. Seismic data had been used to estimate the cave back position and to chart the progression of the cave back relative to the tonnage of ore drawn. By the end of 2003 the Palabora seismic network consisted of 21 recording stations located around the cave on several levels. Nine of these stations are located on the production and development levels. Four stations are situated in the open pit. The eight remaining ones are located in the old and abandoned Exploration Shaft and in a deep borehole. The latter were installed as part of the network upgrade to give better coverage close to the pit base. After completing the network upgrades the area of maximum sensitivity and location accuracy of the network encompasses the entire production area of the mine. The network accuracy is important not only for locating events, but also for calculating the source parameters. From the
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data interpretation viewpoint it is important to know the weighting that has to be attached to each set of data. At Palabora the cave monitoring devices installed in boreholes were lost with time not only due to rockmass response to mining, but also because of interference from the mining activities themselves. As it is not always convenient or possible at a later stage, to install additional TDRs or drill new open holes, the seismic network might be the only cave back monitoring option remaining after some point in time. It is the Palabora experience that it is important to fully implement the seismic system right from the outset with some redundancy to allow for the inevitable equipment downtime and losses. 2 FRACTURING MECHANISM The tonnes extracted from the cave result in the cave back propagating upwards, towards the surface. At Palabora the fracturing zone is located about 60 to 80m above the cave back with an aseismic zone of already fractured rockmass between the cave and the ongoing fracturing (Figure 1).
the second mode can be of magnitude up to 2.1. The first mode of events is associated with rockmass fracturing immediately ahead of the undercut and in the propagating cave back. The second mode of events locate at geological discontinuities some distance from the mining. Events of this second mode are connected with stress redistribution ahead of the undercut abutment and around the cave. Due to their locations and amounts of emitted energy these events are considered a negative feature of the block caving and result in seismic hazard (Gibowicz and Lasocki, 2001). On the other hand seismicity of mode one events is the principle manifestation of the cave progress and can be used to monitor and manage the cave and thus are considered as a positive feature. The progression of the cave generates fractures in the more competent (intact) rock immediately ahead of the fracture zone, which changes the rock properties and lowers its load carrying ability. As the cave back approaches the fractured rock it will yield under the increased load and increased shear movement between the blocks of rock will create further propagation of fractures. The cave progression will also result in the breaking of asperities and other locking mechanisms in the fractured rockmass, creating a favorable environment for shear movement and more growth of the fracture zone around the cave back. In addition to the fracturing around the cave back resulting in mode one seismicity, there is another mechanism of rock failure that is taking place. This mechanism is aseismic deformation, which is the process of either extension of already existing shear fractures, or creation of new fractures with little or no seismicity. While the asperities on the joint surfaces are being broken down there will be sliding and opening of the joints. This firstly does not produce significant seismic energy during movement and secondly inhibits propagation of seismic energy through this zone of fractured rock. The seismic and aseismic deformation of the rockmass around the cave will result in stress redistribution, which in turn will lead to further fracturing. These fracturing mechanisms are very similar to the one described for destress blasting used to reduce the seismic hazard for underground excavations (Rorke and Brummer, 1990). 3 CROWN PILLAR MONITORING
Figure 1: Aseismic zone around the cave As the thickness of this aseismic zone during 2002 was relatively constant over several months, the location of this zone could be used to monitor the cave back position. This was by assuming that the increase in the average elevation of recorded seismic events followed the increase in elevation of the cave back. This progression of the aseismic zone was also used to estimate the natural cave expansion rate. If the cave was pulled at a rate above this, then an excessive expansion void would occur increasing the risk of air blasts. Below this rate the expansion void would not develop sufficiently and would impede cave progression and possibly reduce fragmentation. This natural cave expansion rate value was compared with extrapolated natural expansion rate of mines with lower rockmass quality and then used as the Palabora protocol target draw rate. Other seismic information recorded by the end of 2003 confirmed that this target rate was a reasonable estimate and that because of the lower cave draw rates, there was no significant expansion void. At the beginning of 2004 it is expected that the cave break through into the open pit will be slow and that there will be no significant air blast hazard. Seismic data recorded at Palabora displays bimodal patterns typical for mine induced seismicity (Gibowicz and Kijko, 1994). While for the events belonging to the first mode the maximum magnitude is in range 0.0 to 0.5, the events of 566
With mining operations taking place simultaneously in the open pit and underground, it was considered prudent in mid 2002 to establish a 200m high exclusion zone, or crown pillar, between the base of the open pit and the cave back. This crown pillar had two functions, firstly to minimize the risk to the ongoing operations from destabilizing the pit walls by the cave undercutting them and secondly to provide a barrier to prevent rapid ingress of rainfall water into the underground workings before adequate measures to control the water were in place. Analysis of the seismicity recorded in this pillar was carried out in mid-February 2003, when the crown pillar was estimated to be 180m thick. For several months preceeding, a very low amount of seismic activity was observed above the cave back maximum elevation. Initially it was thought that this low amount of seismic activity above the cave back was because of poor network configuration, but in January 2003 additional seismic sensors were installed in the open pit, and this did not change this pattern of seismicity. Figure 2 shows all of the seismic events recorded during the month following the installation of the additional open pit sensors. The cave profile in Figure 2 is for February 2003. The cave back maximum elevation is about –600m and the open pit minimum elevation is –417m, so the crown pillar is 183m thick at worst, with the exclusion zone top just above the pit base. At this time ramp scavenging operations are taking place well above the pit base and were not considered at risk.
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Figure 4: Cumulative energy and seismic activity.
Figure 2. The 200m exclusion zone.
Figure 3 shows a volume above the February 2003 cave back elevation with all seismic data recorded between January 2002 and mid- February 2003. There are only 77 such events.
Figure 5 presents the time histories of the energy index and the moment fragmentation index. Interpretation of these plots is as follows: the increase in the values of the energy index at the end of June 2002 indicates the start of the fracturing process with associated seismicity. At this stage the rockmass is assumed to be sufficiently intact to be able to emit seismic energy. By the end of October 2002 this process is completed and the energy index plot flattens.
Figure 4 shows time histories of two independent parameters of seismicity for the rock volume in Figure 3 the cumulative seismic energy and the monthly seismic activity rate. The activity rate shows a sharp increase during the period August to October 2002, but due to two relatively large events (February and December 2002) the trend of the cumulative energy is not so clear. Still there is a visible increase during September 2002 followed then by a more flat shape by the beginning of 2003.
Figure 5: Energy and moment fragmentation index time histories.
Figure 3: Volume above the cave.
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The rockmass is fractured and further fragmentation results in seismic events that emit only very small amounts of energy. The moment fragmentation curve indicates the same process. Energy index is a function of seismic moment and energy, while moment fragmentation index is cumulative seismic moment normalized over a rockmass volume. For this reason these two parameters are related and in general should indicate the beginning and the end of this same process. The energy index difference between its background and maximum values is very small. This indicates that the actual seismic fracturing process was in already relatively fractured rockmass. It was concluded that this small amount of recorded seismicity could be accounted for by the essentially aseismic deformation taking place in the rockmass volume below the open pit and above the cave.
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Figure 6 shows the monthly seismic activity rate and its average depth. This data is recorded in a rockmass volume centered around the middle of the February 2003 cave footprint.
4 CONSEQUENCES OF THE CROWN PILLAR FAILURE. The premature fracturing of the crown pillar has changed the stress distribution around the mine on a regional scale. This failed rock restricts the passage of horizontal stresses through it. As indicated by Figure 7 this has tended to increase the vertical stress relative to the horizontal stress acting on the subvertical structure in, and immediately adjacent to, the cave zone. In consequence Palabora become a seismically active mine and during 2003 experienced several damaging seismic events, of which the first took place already in mid January 2003.
Figure 6: Monthly activity and average depth.
From July 2001 to December 2002 the average elevation of seismicity recorded above the cave migrated upwards at a rate of about 15m per month, from about –750m to –460m. This upper limit is assumed to be at the depth of the fractured zone below the pen pit, which would not emit or transmit seismic energy for the same reasons as the aseismic zone. For this reason the seismicity could not migrate any higher in the area below the pit base. The average location of seismicity during 2003 is characterized by monthly average depth increases, but with a new characteristic, namely the increase of the horizontal migration of seismicity around and away from the pit base. The second trend displayed by this figure is the change in the number of events recorded per month. There is a distinct decrease in seismic activity from October 2002. These two trends in depth and activity rate could be best explained by the aseismic zone around the cave reaching the fractured zone located below the open pit base sometime during the last quarter of 2002 Based on the above it was concluded that the 180m thick crown pillar between the top of the cave and the base of the open pit is fractured and de-stressed. The majority of fracturing that took place in 2002 was aseismic deformation. A contributing factor to the early failure of this pillar is the geometry of the dolerite fractures and of the carbonatite joints. At the time it was made clear that this rock volume is fractured but has not caved. This differentiation is important for the stability of the pit walls and the safety of the open pit operation, as the crown pillar continues to act as a relatively stable foundation for the pit walls. For the underground operations, however, there will be a substantial increase in permeability for both situations. New protocols and actions had to be put in place to account for the water hazard. At the beginning of March 2003 the town of Phalaborwa experienced its first strong rain fall of the season. Over 60mm of rainfall was recorded for the open pit. The recent water management study examined a range of values from 1.00E-02 m/s, 1.00E-04 m/s, but the assumed permeability was 1.00E-03 m/s, which would result in the first rainwater arriving at the production level in 4.4 days. In reality the rainfall water was observed in the underground excavations after only six hours – close to the most pessimistic prediction. Fortunately the rainfall was only 20% of that for the 100 year designed storm. This rainfall has verified that the crown pillar is fractured, as evidenced by seismic analysis. 568
Figure 7: Redistribution of stress. Comparison of seismicity recorded during 2002 with seismicity recorded during 2003 shows very significant differences. The stress change influenced not only the small scale jointing in the rockmass, but also the large scale subvertical geological features close to, and in the cave zone. The decrease in horizontal stress reduced confinement on the planes and increased shear movement, which is confirmed by the seismic data. Analysis of the source parameters of seismic events recorded in 2003 indicates a significant increase in the shearing component, in addition to a major increase of released energy per moment. Another important change is nearly a three-fold increase in relatively large seismic events. There remains a significant difference between the seismicity recorded on the west and east sections of the mine. The seismicity on the west side is much deeper and closer to the development and production levels than seismicity on the east side. This is attributed to the more vertical cave profile on the west as well as the progress of the cave footprint towards the Mica Fault, which is a major structure. For this reason seismic hazard precautions were issued for the west side of the mine in June 2003. Additional support was installed as deemed necessary. During the last quarter of 2003 further stress redistribution took place due to the cave shape. The west side of the mine is experiencing low stresses, while there is a significant stress build-up in the east side of the mine. In consequence, as from September 2003 the east side of the mine is experiencing more seismic events with the potential for damage than the west side. This is indicated by the difference in the seismic source parameters of events recorded in the west and east sides of the mine. The slow development of the cracks in the open pit, which started to be apparent from February 2003, confirmed the crown pillar was indeed fracturing and de-stressing by the end of 2002. The slow development of the cracks and the absence of seismicity indicated that the rockmass at the bottom of the pit is in a process of relaxation. The ramp
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mining in the open pit was successfully and safely continued until the beginning of November 2003. The first wedge failure in the pit wall close to the bottom of the pit took place at the end of November 2003. Observations of the open pit in January 2004 indicate that a second wedge failure can be expected soon. According to January 2004 estimations it is calculated that the cave back is less than 40m from the base of the pit and production rates are indicative that there is no expansion void above the cave. The absence of significant instability of the lower pit walls tends to confirm that the progression of the cave is gradual with no large collapses into high expansion voids. 5 CONCLUSIONS Mine induced seismicity is both site and time dependent. Site dependent means that the seismicity depends on properties like depth, ore body geometry, geology and mining methods and their unique interdependency. This interdependency is dynamic and practically never repeats itself in either time or space. As a result, the mine induced seismicity characteristics defined for one mine, or its part, at a specific mining stage are not necessarily applicable to other mines, or even to the same mine in the future. The failure of the crown pillar and the consequent redistribution of stress and seismicity is a very good example of how and when the mine induced seismicity can change. The understanding of mode one seismic events is important for understanding the processes of the block cave mining. Mode one seismicity is a natural process indicating that the cave is progressing and should not only be monitored in a passive way to confirm the cave progress, but should be also used in an active way to manage the cave development. The space and time distributions of mode one seismicity, as well as the changes of their source parameters with time, are directly associated with what is happening in the rockmass around the cave and underground mining infrastructure. By nature, mode one seismic events are relatively small events and as such difficult to detect and record. For this reason seismic networks installed at mines using the caving method must have good sensitivity and accuracy. From Palabora experience it seems that the minimum sensitivity would be all events from magnitude about –2.0 and above with five stations. The XYZ location accuracy should be within 10m. Several applications of seismic monitoring at Palabora have proved that it is a very valuable and useful tool for cave mining. Apart from the monitoring and management of the cave progress it was used to solve the following problems: • Allowed continuous underground operations during blasting operations in the open pit. In 2002 the network allowed analysis of the ground motions at the development level resulting from the blasts in the open pit. A maximum charge per delay was established that resulted in much lower PPV values than those expected to result in damage to the underground excavations. The increased underground production time on its own from this single application of the network resulted in recovery of all capital expenses connected with purchasing and installation of the network.
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• Evaluation of swell factor. This estimate was done by comparing the cave back profiles with different applied swell factors with the recorded space distribution of seismicity. The assumed swell factor was then confirmed by the TDR measurements. • Recording the initiation of the caving process when the critical hydraulic radius was reached. This is important information for comparison with the theoretical considerations and future design. • Estimation of the natural cave expansion rate. This allows control over the size of the expansion void, which is important for maintaining the correct cave profile and reducing the risks inherent with too large an expansion void. • Establishing the expansion void existence. This is done by establishing the relationship between the cave production and the resulting seismicity. This is an indirect assessment, but very useful when confirmed by other geotechnical observations. • Stress distribution around the cave and underground excavations. • Seismic hazard monitoring. The real challenge for the future is full application of mine seismology for the purpose of cave management. This can be achieved only with proper network configuration and very good quality of recorded data. It must be also understood that mine seismology, as any other geophysical method, on its own cannot solve all problems. It must be supplemented by other data. At this stage it is planned to implement seismic passive tomography as an additional tool for cave monitoring and management. In this way information about the seismic source parameters will be supported by data relating directly to the conditions of the rockmass. ACKNOWLEDGEMENTS The authors wish to thank the Palabora Management for granting permission to publish this paper, especially Keith Calder, now at North Parkes, for his support and encouragement in the early stages. REFERENCES • Calder, K, Townsend, P, Russell, F,2001 Palabora Underground Project. Underground mining methods: Engineering Fundamentals and International Case Studies ( Ed: W A Hustrulid and R s Bullock) Society for Mining, Metallurgy and Exploration, Inc. Colorado, USA, pp 405-409 • Gibowicz, S, and Kijko, A, (1994), An Introduction to Mining Seismology, Academic Press, New York • Gibowicz, S,J, and Lasocki, S (2001) Seismicity Induced by Mining: Ten Years Later, Advances in Geophysics, Vol 44, Academic Press, New York. • Rorke, A, J, and Brummer, R, K, (1990) The Use of Explosives in Rockburst Control Techniques, Rockbursts and Seismicity in Mines, (Ed: Fairhurst)pp. 337-385, Balkema, Rotterdam.
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Strain energy control for the Big Bell longitudinal sublevel cave John Player, PhD Student, Western Australian School Mines
Abstract The Big Bell longitudinal sublevel caving operation had a production tonnage of 1.8Mtpa prior to the onset of seismic activity in 1999. During 1999 and 2000, substantial damaging seismic events resulted in a suspension of mining operations in late 2000. Major mine redesign work such as; extraction principles, development location, ground control systems, and man access to working areas, were undertaken to develop a safe and stable seismic environment. This was achieved in mid 2002 once all design principles were implemented. Due to rising costs from the lower production rate of 0.7Mtpa the mine closed in mid 2003.
1. INTRODUCTION Big Bell had a history of rockbursts from 1999 until mine closure in 2003. During this period the mine had to develop suitable mining front geometry, development locations, ground control systems, and automated equipment usage, to minimize the potential hazard to the underground work force. Other papers by Player 2004a, Player 2004b, Barrett and Player 2002, Turner and Player 2000, Player 2000, Sandy and Player, 1999 describe the rockmass properties, ground response, mining environment, and mining methodology. These papers were written over a number of years when the ground response to mining changed from non-seismic to seismic. They detail the progressive understanding of strategic factors involved in combating mine seismicity. This paper will examine the driving factor for the mine seismicity and the performance of the hangingwall cave. In particular, it updates the approach to mining the orebody published by Turner and Player 2000. 2. SEISMIC GEOTECHINCAL ISSUES It is difficult to determine when a mine changes from a non-seismic to a seismic environment. The relative induced stress compared to the rockmass strength, and how the rockmass stores and releases the energy are some of the most important controlling factors. How the rockmass releases strain energy from stress change is a function of; the rock matrix, discontinuities present, the regional / mine wide loading system, and the amount of work that the rock does before and after failure. The loading system is considered to be influenced by the mining front rate of advance. Beck 2000, takes the approach, "…seismic induced events can be described from numerical modelling utilising Mohr Columb slip criterion for events on discontinuities and the relationship between for seismic events not associated with deviatoric mechanisms". To undertake this work requires quality processed seismic events from the mine micro-seismic monitoring system. The method is applied as a back analysis of known mining steps and associated seismic activity. It is possible to allocate seismic criteria to future mining steps or sequences, if it can be assumed that the seismic criterion remain consistent. Brady and Brown 1994, provide the following definitions, "Rockbursts arise from unstable energy changes in the host
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rockmass for mining….Energy changes in a mine domain arise from generation and displacement of excavation surfaces and energy redistribution accompanying seismic events…the strain energy changes which arise from the way in which surface forces are applied as part of the mine structure" are considered to be the dominate source of energy. According to Brady and Brown 1994, two factors need to be considered in relation to energy changes, "increase in static strain energy occurs in areas of stress concentration (stress concentration occurs around all underground openings)….sudden excavation of a surface causes an energy imbalance in the system". This can lead to a dynamic stress component related to the volume of material that is rapidly removed / blasted. It should be noted that local rock fracturing around openings would consume some of the released energy. The rapid excavation of a mining void applies a dynamic stress to the ground closest to the excavation then spreads into the rockmass. When this encounters a static stress field concentration the sum of the stresses can result in damage to the rockmass by exceeding the rockmass strength. For mining environments, the energy released from the excavation of the rock can be considered as an index for the potential local degradation of the rockmass. The degradation, or yielding, can either by non-violent or violent (i.e. bursting conditions). Hence the energy released by an excavation can be used as a design principle. It is better to have a consistent energy release rate from a mining geometry and sequence, rather than one which has sudden energy changes. This can be succinctly explained by the illustration of three different excavation sequences in Figure 1 from Whittaker et al 1992. Sequences one and two have a rapid energy release during the excavation cycle. This makes the sequences more susceptible to a rockburst than sequence three, which has a regular energy release. Sequences one and two could be related to; • a mining geometry utilising primary and secondary open stopes (with pillar removal and fill compared to a pillar less sequence with the use of paste fill), • or a longwall mining face (that uses an arrow head rather than a flat face), • or the release of strain energy along a fault (clamping a fault and mining towards it compared to releasing the fault and moving away).
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Figure 1: Example energy release rate
Strain energy release into the mine environment is controlled by the displacement of the excavation surface and the creation of new surfaces. The mechanisms will depend on the loading system stiffness and rockmass strength to stress ratios. The creation of the hangingwall cave has a very high potential energy release. The cave effectively increases the excavation size and available energy to the mining environment from the formation of new surfaces. Rather than a mining environment that has controlled displacements of the hangingwall, from fill placement or partial extraction by leaving pillars. The additional energy needs to be accounted for as part of the mine design. The Big Bell operation utilised rapid advancing longitudinal mining front with ‘limit retreat pillars’ from 1997 to 2001, Barrett and Player 2002. The use of pillars and a flattening of the mining front, promoted a non-uniform energy release. Intuitively the performance of the hangingwall strata (particularly from rockmass variations between strata), would be expected to influence the caving process in conjunction with; mining geometry, dimension, and rate of advance of the mining front. These factors probably contributed to rapid changes in the energy release in the rockmass, similar to sequence one or two of Figure 1, rather than the gradual release of sequence three. Brady and Brown 1994, define the potential energy of the loading system upon the rockmass as unstable, when it will lead to unstable deformation i.e. seismic events from an instability source. There are two modes of unstable rockmass deformation that lead to instability and mine seismicity. Method One typically involves the crushing of rockmass in pillars, and around excavations at both development and stope scale. These events are modelled by relationships of the monitored seismic events in a process that is explained by Beck 2000 and Wiles 2004. Method Two involves the slip of structures, which could be natural, or mining induced. Slip on a structure can be defined by the Mohr Columb criteria, Beck et al 1997. Beck 2000, assesses both modes of rockmass deformation from monitored micro-seismic events for the period April to August 2000 at Big Bell. The criterion does not describe the potential magnitude for a seismic event, but rather defined when conditions exist for a seismic event to occur. This was based on previously monitored seismic events. At Big Bell five criteria were developed for and Mohr Columb modes of rockmass instability. The above approach relies on the results of micro-seismic monitoring, stress modelling and rockmass properties. It is unlikely to be sensitive enough to provide why one location will burst within a zone that meets a criteria, as opposed to Massmin 2004
another location. However, the approach should be able to describe at which zones rockbursts won’t occur. Work by Wiles 2002, and Wiles 2004, examines a combination of modelled rockmass properties, released energy and loading system stiffness, to determine conditions for rockburst occurrence from the back analysis of previous bursts. The reliance of back analysis techniques to determine future rockburst potential implies early rockbursts will be unexpected hence good mining practices must always be applied during the life of a mine. Good mining practices should use suitable geometry and sequences that redistribute the induced stresses as uniformly as practical. This is the first step in controlling a potential rockburst problem and should be used in combination with an advance rate that controls strain energy buildup in the rockmass. There are a number of tactical issues that need to be considered when evaluating the geotechnical needs of a mine once a strategic decision has been made to assess seismic activity; • When does seismicity present a problem, or is there already a problem? • Can the seismicity be represented or restricted to a rock unit, mine sequence, mine geometry or development location? • Which personnel are available to assess the seismic activity including processing of the data, recording mine geometry change and maintenance of the seismic system? There are specific issues relating to geotechnical properties and mine seismic that are worth further expansion (Section 2.1 to 2.3). 2.1. Rockmass, Local and Global Issues The following list of local and global rockmass items have shown to have an that influence the geotechnical seismic potential and properties; • Many attempts at determining a rock burst index from rock samples have been made. However, the factor that influence a potential rock burst have complexity in their geological and non-geological controls, • Rock mass characterization across the mine environment, • Rock mass stiffness and post peak properties, • Rock P and S wave velocity calculation then undertaking a site survey to provide greater accuracy for the seismic system, • Regional structures that will influence stress, can be a source of seismic activity. They could also damp seismic energy transmission from an event, and • Insitu stress levels and background seismic activity. 2.2. Mine Sequence and Geometry Issues The following list of mine sequence and geometry items have shown to have an influence on the geotechnical seismic potential and properties; • The stress path from mining sequence, Beck and Sandy 2002, examines ground performance changes with different loading conditions, • Extraction rate from areas of the mine, a number of South African reports refer to square meters exposed, some operations with an extended life have been able to develop critical benchmarks. This forms the basis of the Energy Release Rate principal that is discussed by Brady and Brown 1994, Brink et al 2000, Spottiswoode et al 2000, and Spottiswoode and Milev 2002, and • The presence of pillars in the mine layout, and the overall geometry of the mining front. Both of these influence the hangingwall response in a sublevel cave operation.
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2.3. Development Location and Support Criteria Issues The following list of development location and support criteria have shown to have an influence on the geotechnical seismic potential and properties; • Drive orientation and location relative to geological structure and the stress fields, both insitu and mining induced, • Support / reinforcement installed, yieldable reinforcement, strong surface support that won’t fail, quality integration of the two systems, and • Appropriate design criteria for the determination of support and reinforcement. 3. REGIONAL STRUCTURES AND SEISMICITY At Big Bell the examination of damaging seismic events showed that the graphitic shear acted as a dampening barrier. This observation applied for events occurring within the ore zone or the footwall amphibolite, Figure 2. It was not established whether the graphitic shears reflected energy away due to fractures, or whether the brecciated material reduced the amplitude of a wave travelling through the shear zone. Damage only occurred within the zone that the event sourced. Damage would not cross the lode graphitic shear to influence closer parallel development but rather occurred along strike, up and down dip, to effect other high stress areas. Improvements to the seismic monitoring array in 2001 and 2002 clearly identified seismic events occurring on the graphitic shears. The implementation of additional triaxial sensors and a higher number of accepted triggers dramatically reduced the scatter in location of processed events about the lode graphitic shear. Due to the nature of the brecciated graphitic shear, it was possible that recorded seismic events occurred very small distances off the shear along parallel fractures, local splays, or rock asperities in the shear. Far field footwall activity, often resulted in events felt on the surface. These occurred on what has been termed the Big Bell Fault. Grinceri 2002, proposed this as being the
footwall contact between the granite and the mine sequence. Distinct activity was also located in the hangingwall about the granite contact. Both areas were several hundred meters to one thousand meters from the nearest workings. 4. MINE SEQUENCE AND GEOMETRY Sequence and mine geometry effect stress redistribution around mine openings. A poor initial mining geometry and / or poor mining sequence may bring about unstable pillars, extensive linear mining fronts, or ‘pendants’ being developed. These shapes can lead to very high localised stress. If a longwall mining sequence is not correctly progressed, then an increased proportion of the mining front will be more highly stressed compared to a good geometry. Morrison and Beauchamp 2002, explore geometry and seismic proneness principals. A numerical approach can also be in conjunction with seismic data. 5. ROCKBURST HISTORY Barrett and Player 2002, fully describe the history of rockbursts from 1999 to 2002. Damaging seismic events from 1999 to 2000 had a reduced association to specific production blasts. This most likely occurred because of a change to flatten the mining front. As a consequence of the flattened geometry any particularly firing had less influence on the stress field. The flattened geometry contributed to the complete mine front becoming highly stressed rather than just an abutment stress around each production heading. Figure 3 is a long section through the cave profile showing the damaging event locations (each event is represented by an individually shaded rectangle), and change in mining fronts from February 1999 to July 2000. Mining in the lower grade southern extension was undertaken, from September 2000 to December 2001 as shown in the right side of Figure 4. Figure 4 also shows, has the change in mining profile in
Figure 2: View in cross section showing damage transmission from rockbursts. 572
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Figure 3: Change in mining front with major rockburst locations February 1999 to July 2000
Figure 4: Change in mining front with major rockburst locations from February 2002 to July 2003 Massmin 2004
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Figure 5: Sequence of stope extraction to achieve good mining front angles from
the lower levels with corresponding damaging seismic events from December 2001 to July 2003. Each damaging seismic event area is again shown by individually shaded rectangles. The principal damage morphology was similar for all moderate to significant damaging seismic events. The observed principal mechanism was shear rupture along foliation planes and intact failure of the rock on the footwall of the rupture plane. Occasionally minor damaging events exhibited different morphology. An assessment of the principal damage mechanism indicated that a specific foliation place was not responsible for these events, rather any number of foliation planes from one meter into the hangingwall shoulder of the drive, to halfway across the drive. The breakout location was most likely controlled by local factors such as; drive orientation, the presence of the foliation plane, rock mass properties, and the stress field. Numerical modelling of the mine geometry did not distinguish significant differences in stress concentration between the southern and northern ends of the mine. However, seismic activity and the rockburst damage were different between the northern and southern ends. The largest damaging events of local magnitude greater than 2.0 only occurred north of 3750mN. The papers by Barrett and Player 2002, and Player 2004a provide detail on the ground control system utilised at Big Bell to control rockburst damage. Survey’s of contained rockbursts showed displacements of 300mm to 700mm in the ground control system.
6. THE IMPORTANCE OF MINE FRONT GEOMETRY ON CAVING During a review of access development in 2000 the mining front angle was also examined. The mining front was flattened during 1999 and 2000 due to; • the production rate exceeding the required development rate, • failure to open up the 535 slot in a timely fashion, • damage from seismic events restricting development rate, and • use of limit retreat pillars (Barrett and Player 2002) at the final cross cut location. 574
Limit retreat pillars are ore strike pillars from one level to the level above at the last cross cut location. Figure 3 shows an example of a limit retreat pillar. Ore strike development occurred from the crosscut to the end of the ore zone, where a slot was put up and the stope / cave was then retreated to the pillar, which was mass blasted. The flattening of the mining front changed the mining induced stress inturn modifying the ground response and seismic activity. This change was only realised in hindsight through analysis of a sufficiently large seismic event database. Comparative seismic activity analysis should only be undertaken on similar mining geometries for a mining operation. The changing geometry of the mining front made it difficult to correctly evaluate the changes in the seismic response to mining. Differences in the seismic response should be expected when there is a variation in the mining geometry. With hindsight, the data processed in 2000 was influence by mining decisions that were made many months if not a year before hand. The establishment and maintenance of a favourable mining front angle was a key criterion in the management of seismic activity for the lower levels of the mine from 2002. The planned scheduled sequence to achieve a good geometry from March 2002 is shown in Figure 5. This required a slow production rate. The established mining front and mining front angles at October 2002 are shown in Figure 6. The schedule from March 2002 could not be maintained exactly due to ground control problems and additional tonnage draw at the northern end of 535 and 560 levels in the middle of 2002. However, the 45∞ angle was maintained at the northern end of the mine as shown in Figure 6. This was due to the higher seismic activity level and magnitude that occurred at the north when compared to the south. Seismicity was considered to be strongly influenced by the hangingwall conditions. The hangingwall conditions included more competent rockmass that would have altered the local loading system and potentially stored additional strain energy from the caving process. The caving process was also likely to be less regular.
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Figure 6: Base of Cave Geometry October 2002 The 45∞ angle for mining front had been determined to be the ideal angle, for it was expected to minimise seismicity. This was mainly due to the horizontal and vertical lengths to distribute mining induced stresses being maximised. The 45∞ angle was maintained at the northern end until the mining sequence required the 535 and 560 stope numbers 78 and 80 to be extracted to enable the opening of the 585 slot. As expected, this increased seismic activity and resulted in deteriorating ground conditions at the northern end of 535 and 560 levels. The 585 level was opened slower than scheduled, due to blasting and geotechincal problems. Blasting problems, sub-optimal slot design resulted in bridging. Geotechnical, oversize development and levels less than 25m apart resulted in increased ground damage. These delays required ore from the southern side of the mining to be sourced with out excessively flattening the mining front angle. A mining front angle of 20∞ was observed to be satisfactory in mid 2002, and as such mining continued with this angle. The satisfactory performance of the flatter angle was probably due to principal stress orientation (assessed as sub-horizontal and from the north east, slightly shielding the southern cave faces) and weaker hangingwall conditions (allowing softer loading system around the bottom of the cave and more regular caving). During the last months of mine operation, the extended level separation created by the 20∞ mining front resulted in ground deterioration and seismicity on the 560 and 585 levels, therefore efforts were being made to steepen the angle. 7. CONCLUSIONS The principal factors controlling hangingwall cave performance include rockmass properties, mine geometry, and mine sequence. By understanding the caving mechanism in 1999, it could have been realised earlier that the more competent hangingwall promoted strain energy build up in the rockmass with the potential to generate larger rockbursts. The mining technique also allowed additional energy input into the mining environment when compared to non-hangingwall caving techniques. This is particularly Massmin 2004
important where the principal stress is sub horizontal and not parallel to the orebody. Improved management strategies may then have included; a slowing of the advance rate, changes to the mine development layout, and the implementation of the required heavy ground control scheme. Short term pain for a long term gain. These changes could have preserved additional levels and enabled a higher production rate than the 0.7Mtpa that was possible from 2001-2003, thereby improving the mining economics. Mining in a very high induced stress field, using longitudinal sublevel cave methods proved to be operationally and technically possible during 2002-2003.
ACKNOWLEDGEMENTS The author would like to thank Harmony Gold for permission to publish this paper, and my colleges and operators at the mine and within the mining industry for the effort involved in understanding and working with damaging seismic activity. REFERENCES • Barrett, D. and Player, J. 2002. Big Bell, High Stress at Shallow Depth. International Seminar on Deep and High Stress Mining, Perth, Australia. Australian Centre for Geomechanics. • Brady, B and Brown, E. 1994, Rock Mechanics for Underground Mining, Second Edition, Kluwer Academic Publishers. ISBN 0412475502 • Beck, D. 2000. Big Bell Gold Mine – Quantification of Seismic Mechanisms. • AMC200099 :August 2000. • Beck, D., Brady, B. and Grant, D. Induced Stress and Microseismicity in the 3000 Orebody, Mt Isa. Geotechincal and Geological Engineering,1997, V15, pp221-233 • Beck, D and Sandy, M. 2002. Mine Sequencing for High Recovery in Western Australian Mines. International Seminar on Deep and High Stress Mining, Perth Australia. Australian Centre for Geomechanics. • Brink, A. Hagan, T. Spottiswooode, S. Malan, D. Glazer, S. and Lasocki, S. 2000. Survey and Assessment of
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Techniques used to Quantify the Potential for Rock Mass Instability. GAP608. SIMRAC. South Africa. Grinceri, M. 2002. Structural and Spatial Controls to Mine Seismicity at the Big Bell Mine. Unpublished Undergraduate Honours Thesis, University of Western Australia, 81 pages + appendices. Morrison, D, Beauchamp, K. 2002. Stope Design and Sequencing Under Deep Mining Conditions. International Seminar on Deep and High Stress Mining, Perth, Australia. Australian Centre for Geomechanics. Player, J., 2000. Longitudinal Sublevel Caving, Big Bell Mine. Underground Mining Methods Engineering Fundamentals and International Case Studies. Ed W. A. Hustrulid and R. L. Bullock. SME Player, J., 2004a. Field Performance of Cone Bolts at Big Bell Mine. Ground Support 2004. Perth. Australia. Balkema, Eds, Villescusa and Thompson. Player, J., 2004b. Reinforcement and Support Investigation for Static and Dynamic Loading Conditions at the Big Bell Sublevel Caving Operation. Unpublished. Masters of Engineering Science Geomechanics Thesis. Curtin University, West Australia. 256pages Spottiswoode, S. Napier, J. Milev, A. and Vieira, F. 2000. The Relationship between ERR and Seismic Energy
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Release for different Geotechincal Areas. GAP612c. SIMRAC. South Africa. Spottiswoode, S and Milev, A. 2002. A methodology and computer program for applying improved, inelastic ERR for the design of mine layouts on planar reefs. GAP722. SIMRAC. South Africa. Sandy, M., and Player, J. 1999, Reinforcement Design Investigations at Big Bell, Rock Support and Reinforcement Practice in Mining. International Symposium Kalgoorlie 1999.Balkema, Eds Villescusa, Thompson, Windsor. Turner, M., and Player, J. 2000. Seismic Reinforcement at Big Bell Mine. Proceedings of MassMin 2000 – Brisbane Australia. Whittaker, B.N., Singh, R.N., Sun, G., 1992, Rock Fracture Mechanics; Principles, Design, and Applications, Elsevier, 570 pages Wiles, T. 2002. Loading System Stiffness a Parameter to Evaluate Rockburst Potential. . International Seminar on Deep and High Stress Mining, Perth Australia. Australian Centre for Geomechanics. Wiles, T. 2004. Map3D Course Notes. Mine Modelling Pty Ltd.
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A methodology for seismic risk analysis of mining facilities Jorge E. Crempien, Laborie Dean, School of Engineering, Universidad de los Andes
Abstract This work presents a methodology to estimate the seismic risk and design spectrum determination for mining installations in seismic areas such us the northern part of Chile. The methodology developed is applicable to any area. First the seismicity of the area is studied, looking at the frequency of events and their geographical distribution in order to establish a model of temporal and geographical occurrence for events of different magnitudes. The temporal occurrence model is based on the Gutemberg and Richter law which is determined using maximum likelihood. Also a law of attenuation for earthquake ground peak accelerations is determined for Chile; the attenuation law and the Gutemberg and Richter law are used to determine the probability of occurrence of accelerations for different design levels.
1. INTRODUCTION
2. ANALYSIS OF SEISMIC DATA BASE
In general, Chile is well known as an earthquake country, and since the arrival of the Spanish conquerors there exists the records of historic earthquakes that produced abundant loss of lives and great economic damages in the past. It is not strange to have earthquakes with Richter magnitudes over 7.2 in any place of Chile, as an example of this some earthquakes with magnitude over 7.2 that have been reported in the northern town of Arica are consigned in table 1.
In any seismic risk study the first thing to do is to review the data available. The typical data needed for the analysis are: the date of occurrence, the latitude and longitude of the epicenter, the depth of the hypocenter and magnitude. This means that if any of this data is missing in any record it should be discarded from the data base. In the case of the present work the methodology developed has been applied to a zone nearby the town of Arica, establishing for this zone an influence area of 500 km radius with center in Arica, the relevant data of earthquake occurrences in this area was obtained from the USGS catalog. From this catalog complete records and with magnitudes over 4.5 were selected only. Therefore, from an initial number of 4065 records of earthquakes from 1570 up to December 2002, only 2311 reported events were selected to perform the study.
Table 1. Earthquakes with Ms>7.2 in the last 150 years. Year
Lat.
Lon.
Prof. Km.
Ms
Effect
1868 1877 1950 1995
18.5 19.6 23.5 23.3
70.3 70.2 67.5 70.3
¿? ¿? 100 37
8.5 8.3 8.0 7.5
Tsunami Tsunami ---Marejada
Also, Fig. 1 shows a sample of the heavy seismic activity in the northern part of the country. In this figure the epicenters of the reported earthquakes since 1960 are plotted.
2. Model of tTemporal Occurrence Earthquake occurrence can be modeled as a Poisson random process model, where the arrival of earthquakes are events independent of each other. The model of occurrence is given by the law of Gutemberg and Richter that relates the number of earthquakes in a year within the influence area that have magnitudes Ms greater that a minimum Magnitude Mmin. This law is given by: Log10 (N) = a – bMs
(1)
In this last expression and are constants that are determined from the data gathered for the specific area for a certain time. In the case of Arica, the occurrence in time of earthquakes can be observed in Fig. 2
Fig. 1: Epicenters for different magnitudes in Arica. Massmin 2004
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From figure 2 is easy to see that the data is homogeneus starting from the year 1960, and that the data obtained before that year is a small fraction of the total data and can be neglected. The distribution of earthquakes by magnitudes can be observed in Fig. 3 for the same area
Fig. 3: Distribution of Earthquakes by Magnitude 3. MAXIMUM LIKELIHOOD ANALYSIS OF SEISMIC DATA
Fig. 4: Comparison between observed data and Gutemberg and Richter law
Several authors have proposed methods to determine the parameter b using maximum likelihood methods. The first work is due to Aki1 , in latter works Bender2 and Weichart3 improved Aki’s work by adding the possibility to have data in different time spans. In this work we adopt the method proposed by Aki, because it is simpler and also because the size of the data base. The maximum likelihood estimation of after Aki is given by: log10 e b=
4. SPACE DISTRIBUTION OF HYPOCEN-TERS. The space distribution of hypocenters is studied to establish a space model of occurrence The distribution of epicenters in the earth surface can be seen in figure 5 where the shore line is also shown. In this plot no difference is made for magnitude in contrast to Fig 1.
(2) M – MMin
Where M is the mean magnitude of the observed samples. Parameter a is obtained from the condition that for the lowest magnitude in the observed period of time, the computed number of earthquakes N should be consistent with the total number of events in the data base. This is: a = log10 N – log10 T + b • MMin
(3)
Where T is the time length of the data base in years and N is the total number of earthquakes in the data base. For the observed zone in Arica the values obtained for a and b are: a = 5.92400 b = 0.92966 In figure En la figure 4 it is shown the fit of the predicted curve obtained by the Gutemberg and Richter law and the one obtained from the observed data for the region. Fig. 5: Distribution of epicenters in the studied region.
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In figures 6 to 8 it is possible to see the distribution of hypocenters in a view from east to west for different latitudes. These figures clearly show the Benioff plane, where the Nazca Plate takes contact with the Continental Plate. The Nazca Plate is subducting beneath the Continental Plate, for this reason all the hypocenters appears clearly grouped close to the Benioff plane.
5. ATENUATION LAWS FOR MAXIMUM ACCELERATION The important thing in seismic risk procedures for earthquake resistant design is that the maximum acceleration that can happen in a place con be related with the occurrence of earthquakes of expected magnitudes, in the zone where a structure is being designed. What is needed is to obtain the probability of occurrence of an acceleration greater than a certain minim acceleration Amin. To obtain this relation it is necessary to know how varies the maximum acceleration in a point with the distance of the point to the hypocenter of the earthquake and also with the magnitude. These are the so called attenuation relations for a given region or zone. Many of these relations have been proposed in the past, and some of them have been plotted and can be observed in figure 9. In this work, the attenuation relation initially proposed by Arias and Crempien4 and latter modified by Crempien5, for Chilean data was adopted. In general, attenuation relations have a general form given by: c1 ≥ e
c2M
amax =
(4) (D + C3)c4
Fig. 6: East-West Distribution of Hypocenters.
Where c1, c2, c3, y c4 are constants. In this case, the values of the constants obtained by Crempien in the corrected model are: c1 c2 c3 c4
= = = =
422.0 0.79 60. 1.42
In eq. 4 M is Richter magnitude and D the epicentral distance in kilometers (km).
6. SEISMIC RISK MODEL AND DESIGN EARTHQUAKE.
Fig. 7: East-West Distribution of Hypocenters
To compute the seismic risk of a zone, the procedure suggested by Cornell6, and modified by Algermissen and Perkins7, is used. The attenuation law is jointly used with the all the Gutemberg and Richter law of temporal occurrence determined for the different seismic sources within the influence area. In the case of the zone being evaluated near the of Arica, we can observe in Fig. 1 that epicenters are well distributed almost all over the influence area, so the model of space occurrence adopted was a continuous source distributed over the influence area. This area was divided into square elements that form a mesh, and for every element i in the influence mesh the ratio of earthquakes happened in them to the total number of earthquake in the influence zone was computed.
αi =
Ni (5) NT
Fig. 8: East-West Distribution of Hypocenters
Massmin 2004
this ratio gives the earthquake productivity of that particular element in the mesh relative to the total area. This procedure was suggested by Der kiuregian and Ang8 for linear sources and extended to area sources by Arias and Crempien4. Then, for a certain level of acceleration that can happen in the point where the earthquake risk is being evaluated, the necessary magnitude to produce that level of acceleration is computed for all the squares defined in the mesh, and the total number of earthquakes that produce magnitudes equal or larger is computed. The probability of occurrence of the acceleration A grater than a minimum acceleration Amin is defined as the ratio between favorable Santiago Chile, 22-25 August 2004
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The maximum credible magnitude for this study was set to Ms = 8.75 based on observed data. The probability excedence curves obtained for Arica are shown are shown in figure 10. This figure shows the probability of excedence of four localities near Arica for small mining facilities. Because these localities are so close together the probability curves obtained for each one of them differ very little from each other.
Fig 10. Excedence probability for accelerations In Fig. 11 the period of return of accelerations in years is given for the same localities close to Arica.
Fig. 9: Atenuation Relations for Subduction Zones.
outcomes and total outcomes, the probability of excedence of a is defined as the total number of earthquakes that produce the acceleration divided by the total number of earthquakes in the influence area. In this computation the actual number in each square that produce acceleration greater that Amin is multiplied by the factor to take into account the productivity factor of each element in the mesh. This is m
∑ a xE {N ( A < a / M ≥ M )} F (a) = E { N ( M ≥ M )} i
Min
i =l
T
(6)
Min
where E {%} indicate expected value. The return period of this acceleration is computed as Fig 11: Return period for accelerations 1
1
F (a) =
(7) 1 – F (a) E {N (M ≥ Mmin) per year }
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7. SELECTION OF DESIGN EARTHQUAKE 8. CONCLUSSIONS According to Standard design practice it is recommended to use to levels of design earthquake. The first level is the operation earthquake, this is an earthquake that is likely to happen during the life span of the facility and that the facility have to withstand with no loss of functionality. For example if it is decided to use an earthquake with a return period of 250 years, then from Fig. 11 the maximum horizontal acceleration is Amax = 0.18g. The second level is the shut down earthquake. This is an earthquake that has low probability of happening during the life span of the facility and it can cause its loss of functionality. For example, if it is decided to use an earthquake with a return period of 600 years, the maximum horizontal acceleration obtained from Fig. 11 is Amax =0 .25g. The next step is to obtain the design response spectra. This can be accomplished using the design response spectrum given in the design code, or using the response spectra of representative earthquakes that have happened in the region. In this case, the spectra of the acceleration records obtained during the 1985 earthquake are used. These spectra are usually normalized so that he maximum acceleration of the ground be 1.0g. Therefore it is necessary to scale them to be usable, for this reason they are multiplied by the maximum acceleration obtained for the operation and shutdown earthquake respectively, the outcome is shown in figure 12 where the average design response spectra for the earthquake records obtained in the 1985 central valley earthquake is plotted for both levels design and shutdown.
A method for estimation of the seismic risk and for defining design levels was developed and applied to the zone northern zone taking the town of Arica as center for the study. This methodology is portable, and can be used in any location or zone for both studying seismic risk and defining the design level of mining facilities. 9. ACNOWLEDGEMENTS Support for this research was granted by FAI project CING 001-99, this support is deeply acknowledged. 10. RERENCES 1. Aki, K., "Maximum Likelihood Estimate of b in the Formula and its Confidence Limits.", Bulletin of The Earthquake Research Institute, University of Tokio, Vol. 43, pp. 237-239, 1965. 2. Bender, B., "Maximum Laikelihood Estimation of b Values for Magnitude Grouped Data.", Bulletin of the Seismological Society of America, Vol. 73, pp. 831-851, June 1983. 3. Weichert, D., "Estimation of the Earthquake Recurrence Parameters for Unequal Observation Periods for Different Magnitudes.", Bulletin of the Seismological Society of America, Vol. 70, pp 1337-1346, 1980. 4. Arias, A. y Crempien J., "Aislación sísmica de equipos eléctricos de la sub-estación de Endesa en Alto Jahuel", V Jornadas Chilenas de Sismología e Ingeniería Antisísmica, Vol. 1., pp. 1075-1083, 1989. 5. Crempien, J., "Un modelo de Atenuación para Aceleracioneas Máximas de Terremotos en la Zona Central de Chile", Documento de Trabajo No. 44, Universidad de los Andes, 2001. 6. Cornell, C.,A., "Engineering Seismic Risk Analysis," Bulletin of the Seismological Society of America, Vol. 58, pp. 1583-1606, 1968. 7. Algermissen, S.T., and Perkins, David M., "A probabilistic estimate of maximum acceleration in rock in the contiguous United States," U.S. Geological Survey Open-File Report OF 76-416, 45 p. 1976. 8. Der Kiureghian, A., Ang. H.S., "A Fault-Rupture Model for Seismic Risk Analysis", Bulletin of the Seismological Society of America, Vol. 67, pp. 1173-1194, 1977.
Fig. 12: Design spectra showing shutdown and operation level.
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A methodology for seismic risk analysis of mining facilities Jorge E. Crempien, Laborie Dean, School of Engineering, Universidad de los Andes
Abstract This work presents a methodology to estimate the seismic risk and design spectrum determination for mining installations in seismic areas such us the northern part of Chile. The methodology developed is applicable to any area. First the seismicity of the area is studied, looking at the frequency of events and their geographical distribution in order to establish a model of temporal and geographical occurrence for events of different magnitudes. The temporal occurrence model is based on the Gutemberg and Richter law which is determined using maximum likelihood. Also a law of attenuation for earthquake ground peak accelerations is determined for Chile; the attenuation law and the Gutemberg and Richter law are used to determine the probability of occurrence of accelerations for different design levels.
1. INTRODUCTION
2. ANALYSIS OF SEISMIC DATA BASE
In general, Chile is well known as an earthquake country, and since the arrival of the Spanish conquerors there exists the records of historic earthquakes that produced abundant loss of lives and great economic damages in the past. It is not strange to have earthquakes with Richter magnitudes over 7.2 in any place of Chile, as an example of this some earthquakes with magnitude over 7.2 that have been reported in the northern town of Arica are consigned in table 1.
In any seismic risk study the first thing to do is to review the data available. The typical data needed for the analysis are: the date of occurrence, the latitude and longitude of the epicenter, the depth of the hypocenter and magnitude. This means that if any of this data is missing in any record it should be discarded from the data base. In the case of the present work the methodology developed has been applied to a zone nearby the town of Arica, establishing for this zone an influence area of 500 km radius with center in Arica, the relevant data of earthquake occurrences in this area was obtained from the USGS catalog. From this catalog complete records and with magnitudes over 4.5 were selected only. Therefore, from an initial number of 4065 records of earthquakes from 1570 up to December 2002, only 2311 reported events were selected to perform the study.
Table 1. Earthquakes with Ms>7.2 in the last 150 years. Year
Lat.
Lon.
Prof. Km.
Ms
Effect
1868 1877 1950 1995
18.5 19.6 23.5 23.3
70.3 70.2 67.5 70.3
¿? ¿? 100 37
8.5 8.3 8.0 7.5
Tsunami Tsunami ---Marejada
Also, Fig. 1 shows a sample of the heavy seismic activity in the northern part of the country. In this figure the epicenters of the reported earthquakes since 1960 are plotted.
2. Model of tTemporal Occurrence Earthquake occurrence can be modeled as a Poisson random process model, where the arrival of earthquakes are events independent of each other. The model of occurrence is given by the law of Gutemberg and Richter that relates the number of earthquakes in a year within the influence area that have magnitudes Ms greater that a minimum Magnitude Mmin. This law is given by: Log10 (N) = a – bMs
(1)
In this last expression and are constants that are determined from the data gathered for the specific area for a certain time. In the case of Arica, the occurrence in time of earthquakes can be observed in Fig. 2
Fig. 1: Epicenters for different magnitudes in Arica. Massmin 2004
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From figure 2 is easy to see that the data is homogeneus starting from the year 1960, and that the data obtained before that year is a small fraction of the total data and can be neglected. The distribution of earthquakes by magnitudes can be observed in Fig. 3 for the same area
Fig. 3: Distribution of Earthquakes by Magnitude 3. MAXIMUM LIKELIHOOD ANALYSIS OF SEISMIC DATA
Fig. 4: Comparison between observed data and Gutemberg and Richter law
Several authors have proposed methods to determine the parameter b using maximum likelihood methods. The first work is due to Aki1 , in latter works Bender2 and Weichart3 improved Aki’s work by adding the possibility to have data in different time spans. In this work we adopt the method proposed by Aki, because it is simpler and also because the size of the data base. The maximum likelihood estimation of after Aki is given by: log10 e b=
4. SPACE DISTRIBUTION OF HYPOCEN-TERS. The space distribution of hypocenters is studied to establish a space model of occurrence The distribution of epicenters in the earth surface can be seen in figure 5 where the shore line is also shown. In this plot no difference is made for magnitude in contrast to Fig 1.
(2) M – MMin
Where M is the mean magnitude of the observed samples. Parameter a is obtained from the condition that for the lowest magnitude in the observed period of time, the computed number of earthquakes N should be consistent with the total number of events in the data base. This is: a = log10 N – log10 T + b • MMin
(3)
Where T is the time length of the data base in years and N is the total number of earthquakes in the data base. For the observed zone in Arica the values obtained for a and b are: a = 5.92400 b = 0.92966 In figure En la figure 4 it is shown the fit of the predicted curve obtained by the Gutemberg and Richter law and the one obtained from the observed data for the region. Fig. 5: Distribution of epicenters in the studied region.
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In figures 6 to 8 it is possible to see the distribution of hypocenters in a view from east to west for different latitudes. These figures clearly show the Benioff plane, where the Nazca Plate takes contact with the Continental Plate. The Nazca Plate is subducting beneath the Continental Plate, for this reason all the hypocenters appears clearly grouped close to the Benioff plane.
5. ATENUATION LAWS FOR MAXIMUM ACCELERATION The important thing in seismic risk procedures for earthquake resistant design is that the maximum acceleration that can happen in a place con be related with the occurrence of earthquakes of expected magnitudes, in the zone where a structure is being designed. What is needed is to obtain the probability of occurrence of an acceleration greater than a certain minim acceleration Amin. To obtain this relation it is necessary to know how varies the maximum acceleration in a point with the distance of the point to the hypocenter of the earthquake and also with the magnitude. These are the so called attenuation relations for a given region or zone. Many of these relations have been proposed in the past, and some of them have been plotted and can be observed in figure 9. In this work, the attenuation relation initially proposed by Arias and Crempien4 and latter modified by Crempien5, for Chilean data was adopted. In general, attenuation relations have a general form given by: c1 ≥ e
c2M
amax =
(4) (D + C3)c4
Fig. 6: East-West Distribution of Hypocenters.
Where c1, c2, c3, y c4 are constants. In this case, the values of the constants obtained by Crempien in the corrected model are: c1 c2 c3 c4
= = = =
422.0 0.79 60. 1.42
In eq. 4 M is Richter magnitude and D the epicentral distance in kilometers (km).
6. SEISMIC RISK MODEL AND DESIGN EARTHQUAKE.
Fig. 7: East-West Distribution of Hypocenters
To compute the seismic risk of a zone, the procedure suggested by Cornell6, and modified by Algermissen and Perkins7, is used. The attenuation law is jointly used with the all the Gutemberg and Richter law of temporal occurrence determined for the different seismic sources within the influence area. In the case of the zone being evaluated near the of Arica, we can observe in Fig. 1 that epicenters are well distributed almost all over the influence area, so the model of space occurrence adopted was a continuous source distributed over the influence area. This area was divided into square elements that form a mesh, and for every element i in the influence mesh the ratio of earthquakes happened in them to the total number of earthquake in the influence zone was computed.
αi =
Ni (5) NT
Fig. 8: East-West Distribution of Hypocenters
Massmin 2004
this ratio gives the earthquake productivity of that particular element in the mesh relative to the total area. This procedure was suggested by Der kiuregian and Ang8 for linear sources and extended to area sources by Arias and Crempien4. Then, for a certain level of acceleration that can happen in the point where the earthquake risk is being evaluated, the necessary magnitude to produce that level of acceleration is computed for all the squares defined in the mesh, and the total number of earthquakes that produce magnitudes equal or larger is computed. The probability of occurrence of the acceleration A grater than a minimum acceleration Amin is defined as the ratio between favorable Santiago Chile, 22-25 August 2004
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The maximum credible magnitude for this study was set to Ms = 8.75 based on observed data. The probability excedence curves obtained for Arica are shown are shown in figure 10. This figure shows the probability of excedence of four localities near Arica for small mining facilities. Because these localities are so close together the probability curves obtained for each one of them differ very little from each other.
Fig 10. Excedence probability for accelerations In Fig. 11 the period of return of accelerations in years is given for the same localities close to Arica.
Fig. 9: Atenuation Relations for Subduction Zones.
outcomes and total outcomes, the probability of excedence of a is defined as the total number of earthquakes that produce the acceleration divided by the total number of earthquakes in the influence area. In this computation the actual number in each square that produce acceleration greater that Amin is multiplied by the factor to take into account the productivity factor of each element in the mesh. This is m
∑ a xE {N ( A < a / M ≥ M )} F (a) = E { N ( M ≥ M )} i
Min
i =l
T
(6)
Min
where E {%} indicate expected value. The return period of this acceleration is computed as Fig 11: Return period for accelerations 1
1
F (a) =
(7) 1 – F (a) E {N (M ≥ Mmin) per year }
580
Santiago Chile, 22-25 August 2004
Massmin 2004
7. SELECTION OF DESIGN EARTHQUAKE 8. CONCLUSSIONS According to Standard design practice it is recommended to use to levels of design earthquake. The first level is the operation earthquake, this is an earthquake that is likely to happen during the life span of the facility and that the facility have to withstand with no loss of functionality. For example if it is decided to use an earthquake with a return period of 250 years, then from Fig. 11 the maximum horizontal acceleration is Amax = 0.18g. The second level is the shut down earthquake. This is an earthquake that has low probability of happening during the life span of the facility and it can cause its loss of functionality. For example, if it is decided to use an earthquake with a return period of 600 years, the maximum horizontal acceleration obtained from Fig. 11 is Amax =0 .25g. The next step is to obtain the design response spectra. This can be accomplished using the design response spectrum given in the design code, or using the response spectra of representative earthquakes that have happened in the region. In this case, the spectra of the acceleration records obtained during the 1985 earthquake are used. These spectra are usually normalized so that he maximum acceleration of the ground be 1.0g. Therefore it is necessary to scale them to be usable, for this reason they are multiplied by the maximum acceleration obtained for the operation and shutdown earthquake respectively, the outcome is shown in figure 12 where the average design response spectra for the earthquake records obtained in the 1985 central valley earthquake is plotted for both levels design and shutdown.
A method for estimation of the seismic risk and for defining design levels was developed and applied to the zone northern zone taking the town of Arica as center for the study. This methodology is portable, and can be used in any location or zone for both studying seismic risk and defining the design level of mining facilities. 9. ACNOWLEDGEMENTS Support for this research was granted by FAI project CING 001-99, this support is deeply acknowledged. 10. RERENCES 1. Aki, K., "Maximum Likelihood Estimate of b in the Formula and its Confidence Limits.", Bulletin of The Earthquake Research Institute, University of Tokio, Vol. 43, pp. 237-239, 1965. 2. Bender, B., "Maximum Laikelihood Estimation of b Values for Magnitude Grouped Data.", Bulletin of the Seismological Society of America, Vol. 73, pp. 831-851, June 1983. 3. Weichert, D., "Estimation of the Earthquake Recurrence Parameters for Unequal Observation Periods for Different Magnitudes.", Bulletin of the Seismological Society of America, Vol. 70, pp 1337-1346, 1980. 4. Arias, A. y Crempien J., "Aislación sísmica de equipos eléctricos de la sub-estación de Endesa en Alto Jahuel", V Jornadas Chilenas de Sismología e Ingeniería Antisísmica, Vol. 1., pp. 1075-1083, 1989. 5. Crempien, J., "Un modelo de Atenuación para Aceleracioneas Máximas de Terremotos en la Zona Central de Chile", Documento de Trabajo No. 44, Universidad de los Andes, 2001. 6. Cornell, C.,A., "Engineering Seismic Risk Analysis," Bulletin of the Seismological Society of America, Vol. 58, pp. 1583-1606, 1968. 7. Algermissen, S.T., and Perkins, David M., "A probabilistic estimate of maximum acceleration in rock in the contiguous United States," U.S. Geological Survey Open-File Report OF 76-416, 45 p. 1976. 8. Der Kiureghian, A., Ang. H.S., "A Fault-Rupture Model for Seismic Risk Analysis", Bulletin of the Seismological Society of America, Vol. 67, pp. 1173-1194, 1977.
Fig. 12: Design spectra showing shutdown and operation level.
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Chapter 16
Transition from Open Pit to Underground Mining
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Caving and fragmentation at Palabora: Prediction to Production Allan Moss, Principal Consultant Geotechnical Engineering, Frank Russell, Principal Consultant Underground Mining, Rio Tinto Technical Services Colin Jones, Consultant Mining Engineer
Abstract Palabora Mining Company has successfully transitioned from an open pit operation to an underground block cave mine. The transition faced a number of technical challenges, not the least caving and fragmentation of a competent rock mass. Initial predictions made during the feasibility study were that, due to the overall quality of the rock mass, oversize and drawpoint blockages would exert a major constraint on production build-up and cave progression. The actual fragmentation, while still coarse relative to other block cave operations, has been finer than predictions. There a number of reasons for this, the most important being the impact of cave induced stresses. Because of problems in dealing with hang-ups, initial draw rates were low. It is believed that this provided time for caving induced stresses to pre-condition the rock mass by extension of existing joints and the creation of new ones. The results of numerical modeling and microseismic monitoring are used to substantiate this view.
1 INTRODUCTION Palabora Mining Company is bringing into production a 30,000 t/d block cave operation below the now closed open pit. The rock mass being caved represents some of the most competent ground in which cave mining has been carried out, as is evident from the pit that has one of the deepest and steeply sloping excavations in the world. The caving process therefore relies very much on the generation of stresses of sufficient magnitude to induce fracture in the competent rock mass. This process provides the impetus for cave propagation. Equally important, it reduces the naturally occurring joint bounded blocks down to a size that can be managed on the extraction level. Cave breakthrough has occurred above the central section of the production footprint where the draw height is
greatest and the cave column the least. Sustained caving, once regarded as a major project risk, is now occurring. Fragmentation, which was also identified as a major risk, is generally finer than originally predicted. Because of problems dealing with large number of drawpoint blockages the initial draw rate has been low, allowing, it is conjectured, caving induced stress to pre-condition the rock mass. 2 OVERVIEW 2.1 Geology The Palabora copper ore body is an elliptical shaped vertically dipping volcanic pipe. The pipe measures 1400 m and 800 m in plan with resources identified to 1800 m below surface. Transgressive and banded carbonatites form the central core of the ore body with the banded carbonatites
Figure 1 Plan of Extraction Level Showing Dykes and Faults Massmin 2004
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and transgressive carbonatites dominant in the western sector and eastern sectors of the orebody, respectively. Barren dolerite dykes dipping steeply to the northeast are present as are a number of northwest and northeast trending faults, refer Figure 1. 2.2 Geotechnical Properties The uniaxial strength of the carbonatites ranges from 60 MPa to 180 MPa with an average value for the transgressive carbonatites of 120 MPa and 95 MPa for the banded carbonatites. The dolerites are very strong with uniaxial compressive strengths in excess of 300 MPa. Stress measurements indicate that, away from the influence of the open pit, the stress regime is hydrostatic and equivalent to the overburden load of approximately 38 MPa at the extraction level. There is substantial structural variability within the orebody. The central section of the production footprint is extensively faulted and is intersected by a number of closely jointed dolerite dykes. The west section has only a few minor dolerites dykes while the east section has a single major dyke. Two major moderately spaced relatively continuous subvertical joint sets are evident throughout the orebody. These trend north-south and east-west, approximately parallel to the major mine development. Other sub-vertical sets occur but are not as dominant. A widely spaced subhorizontal discontinuous set exists. Joint spacing ranges between 1.2 and 12 m. During the geotechnical drilling programme long lengths of unbroken core were obtained underlining the relatively un-fractured competent nature of the ore. The orebody was classified using Laubscher’s MRMR system. Values vary between 48 and 76 for the various rock types with an "average" of 61. For reference the footprint has been divided into West, Central and East zones that approximate to structural and geotechnical domains (note these zones are also coincident with zones of different seismic signature). 2.3 Mining The production level (see Figure 1) is located about 1200 m below the surface and 400 m below the final pit bottom. The undercut level is located 18 m above the production level. Crosscut drives in the undercut are excavated at 4 m wide by 4m high at 17 m centres above the 34 m spacing of the extraction drives (Calder and Russell, 2000). Undercut mining was based on the extraction of a diamond shape undercut with four fronts advanced simultaneously. Mining was initiated from a central slot drive and retreated to the north and south rim drives in each crosscut. North and south access drives are located 20 m from the orebody. Average crosscut length is 200 m, with undercut east-west width of 700 m. Drawbells were developed below the horizontal drill rings, with the major apex pillar and extraction drive beneath inclined drill rings. The undercut was advanced ahead of drawbell development. Originally 21 production crosscuts were planned. However, due to poor ground conditions associated with a major fault in the western extremity, only 20 crosscuts have been developed at present. In all 166 drawbells will be available for production. There are four crusher stations located along the northern side of the production footprint. Because of the anticipated coarse fragmentation LHDs dump directly into these crushers thus obviating passes. 3 CAVE DEVELOPMENT The Palabora orebody is one of the most competent rock masses in which caving has been attempted. Figure 2 586
shows the relative position of various caving operations in terms of rock quality, intact rock strength, induced stress and a simple assessment of cave performance. Palabora is positioned at the upper bound of experience.
Figure 2 Relative Cave Performance The principal method used to predict the size of footprint required to initiate and sustain caving was the MRMR approach developed by Laubscher (Laubscher, 1994) The hydraulic radius (HR) required to initiate caving was predicted to be 35 m and was based on the results of core logging and mapping in an exploration drive and in the pit. Because of natural variability of geotechnical properties, it was realised that this number was not absolute but represented the central estimate of a range of outcomes. Numerical modeling (FLAC3D and 3DEC) was also undertaken to provide further insight to the caving process and to allow an assessment of likely ground conditions as the cave developed. Undercutting and drawbell development commenced approximately in the centre of the footprint where the ground was weaker due to the presence of a number of major faults and more intense jointing. Though not ideal from a construction or operating perspective, it was considered that this strategy provided the lowest overall project risk with regard to caving and the impact of caving induced stresses on mine infrastructure. Cave development was monitored in a number of ways, including changes in ground conditions, seismic activity (Glazer and Hepworth, 2004) and observations of wall behavior in the overlying pit. Key milestones are shown in Figure 3. Five stages of cave development can be identified: Stage 1 Undercutting: Prior to production mining when undercutting was the major extraction activity, the amount of seismic activity was reasonably consistent with the elevation of this activity increasing with time as undercutting proceeded. This is believed to reflect the impact of stress re-distribution due to the undercutting process. The overlying ground, while experiencing changes in stress, did not cave during this stage. Stage 2 Initial Pull: Once production mining commenced, there was an increase in the number of seismic events and a decrease in elevation of this activity. This removal of significant volumes of rock results in stress change adjacent to where the ground is
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Figure 3 Seismic monthly activity and average depth of seismicity
removed, hence the lower elevation of seismic activity in Q4 2001. Stage 3 Cave Initiation: Caving was initiated upon the production footprint reaching a critical dimension (HR=45 m). This was marked by a substantial increase in seismic activity and increasing elevation of this activity (Q1 2002).
footprint. Local anomalies have occurred related to structure. All of the foregoing experience underlines the need for further investigation into caving mechanics. 4 FRAGMENTATION
Stage 4 Initial Cave Breakthrough: The pillar between the cave and the pit floor failed marking the onset of cave breakthrough and a major re-distribution of stress (HR~ 60 m). This was accompanied by an increase in seismic activity together with a trend of decreasing elevation of seismic events (Q1-2 2003). In addition, survey monitors in the pit showed an increase in rate of movement.
The mine is in the initial years of production and though cave breakthrough has occurred, the maximum height of cave draw is only some 60 m (less than 15% of column height). During Q2 2004 production rates approaching 30,000 t/d have been achieved. The main constraint to production has been secondary breaking of drawpoint blockages. Drawpoint hang-ups, blockages and oversize are common with, on average, a yield of some 350 t per hang-up event. Secondary breaking is therefore a critical component in achieving full and sustainable production.
Stage 5 Breakthrough: Evidence of cave breakthrough can be seen in the pit with the advent of wedge failures associated with general caving induced rock mass dilation. This demonstrates that the production footprint has reached the critical size to allow sustained caving to occur. This probably occurred during stage 4 at an HR of around 60 m. Comparison of the predicted hydraulic radius to induce caving with the actual value shows an under-estimate of some 30%. Though it is possible to adjust the prediction process to more accurately reflect observed conditions, it is considered that the error is a measure of uncertainty in empirical approaches to what is a complex engineering problem. Results of numerical modeling predictions were close to actual cave initiation. Conversely, ground conditions during cave development have been surprisingly benign in contrast to the predictions made with numerical modeling. Ground control problems in the undercut were only encountered after about 80% of the total undercut area was developed. Similarly, on the extraction level ground problems have generally only occurred once cave draw took place over about 80% of the
4.1 Predicted Fragmentation During the feasibility study it was recognised that the competence of the rockmass would result in a high proportion of oversize and the need for secondary breaking. At the time of the study, however, there was little experience with caving of competent rock masses and no tools were available for predicting cave fragmentation. A model for predicting block cave fragmentation (BCF) was therefore developed in conjunction with Premier Mine (Esterhuizen, 1994) BCF is a mixed analytical model and expert system that considers the effects of both primary and secondary fragmentation to predict the fragmentation range reporting to the drawpoints. Input data to the BCF model includes a measure of rock mass quality (RMR), cave geometry, rock jointing and in-situ stress regime. Blocks are generated from the joint set data and the effects of stresses and orientation are considered in determining the manner in which blocks separate from the cave back. This information is used to determine primary fragmentation These blocks form input to the secondary fragmentation module, in addition to geometry, height and specific gravity
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of ore columns. Accounted for in this module is the probability of blocks splitting according to their aspect ratios or from temporary arching. The comminution of corners and the effect of cushioning fines are also considered. Output from the second module is the size distribution of material reporting to the drawpoint. Any rocks that are greater than 2 m_ are considered oversize. This cut-off volume is based on the largest rocks that can be efficiently handled by the 6 m_ LHDs. Figure 4 shows the percentage oversize predicted during feasibility over the life of the cave, while Figure 5 shows the predicted numbers of high and low hang-ups over mine life.
oversize were identified as a major risk to the project, particularly high hang-ups where the technology was limited. Much work was carried out during the feasibility study, even to the extent of fabricating a prototype, to have a capability of dealing with hang-ups up to 16 m in height. Despite this effort and preparedness it is perhaps fortuitous that no high hang-ups have occurred! The current secondary breaking fleet (see Table 1) consists of a mixture of medium reach rigs (MRR) capable of reaching up into the drawbell a distance of 12 m, oversize rigs that use non-explosive technology to break oversize rocks on the floor of the drawpoint and water cannon and percussion rigs used to ensure hang-ups are stable before attempting to remove the hang-up using the medium reach rig. Table 1 Secondary Breaking Fleet Equipment
Purpose
No
Hang-ups and blockages up to 12 m
7
Commando rigs
General purpose
2
Robust rigs (RR)
Breaking oversize in drawpoint
7
Stabilise hang-ups prior to use of MMR
2
Bombing of blockages prior to use of MMR
2
Medium reach rigs (MRR)
Water cannon Concussion rigs Figure 4 Predicted Percentage Oversize by Year
Figure 5 Predicted Hang-ups by Year A second model was also developed to provide an estimate of the frequency and location of hang-ups and oversize. This model compared the fragmentation range to the cross sectional area of the drawbell to determine if a hang up would occur. High hang-ups were defined as occurring at more than 9 m from the floor of the drawpoint and low hang-ups were defined as those occurring up to 9 m from the floor. Oversize refers to rock greater than 2 m3 that has to be broken before loading. Information generated from the fragmentation and hang-up models were then input into a discrete event simulation in order for production rates and equipment requirements to be assessed. 4.2 Observed Secondary Breaking One of the critical issues with respect to predictions of fragmentation is matching the secondary breaking requirements. Methods of dealing with all hang-ups and 588
The type and number of secondary breaking equipment is different to that originally envisaged. The changes are part of an ongoing exercise to resolve secondary breaking issues of suitability, reliability and organisation. • Suitability is not just related to change in fragmentation but also to the fact that hang-ups now occur in a different location to that expected. For example, many hang-ups occur as "jumbles" of rock in the throat of the drawpoint and original equipment fleet was only capable of drilling either a 100 mm diameter hole inside the drawbell or oversize lying on the floor of the drawpoint. • Reliability was a problem with the original fleet, particularly in the demanding environment of production ramp-up. In addition, despite the use of water cannons and concussion rigs to stabilise hang-ups, boom damage has been high resulting in lower than expected availability. • Organisation of secondary breaking has been continuously improved by the use of multi-discipline process management teams. Current practice is to carry out sequential passes with the water cannon, concussion rig, medium reach rig followed by the oversize rig. The foregoing demonstrates the importance of good estimates of fragmentation and hang-ups together with reliable estimates of hang-up clearance times. With this information the appropriate type and number of equipment can be selected prior to the start of production. In terms of operations management, each unit maintains a record what work is done and where. Thus, a comprehensive database exists of secondary breaking events across the entire production footprint. This information demonstrates the impact of fragmentation in production terms and provides an indirect measure of fragmentation. The tonnage yield per hang-up event or the frequency of hang-ups is directly proportional to the size of fragments reporting to the drawpoint. The results show that geology, location within footprint,
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and draw height are key factors in the number of hang-ups and oversize that has to be dealt with and, by inference, the distribution of fragmentation. The following summarises secondary breaking achievements to the end of Q1 2004: • Overall average yield = 409 t • Average yield in the Eastern area = 248 t • Average yield in the Central area = 736 t • Average yield in the Western area = 433 t
subsequent rendition for image analysis in the course category. Over the six-month period since fragmentation sampling commenced at Palabora a total of 30 sampling sessions have been completed using the flip chart method, sampling a total of 6031 drawpoints.
In the Central and West mining zones, yield begins increasing from a height of approximately 45 m draw height . The East zone has not reached this height and is showing a relatively unchanging yield in the range 20-45 m in situ cave height. The numbers of hang-ups varies by mining area and, more importantly, by draw height as summarised in Table 2 and illustrated in Figure 6 Table 2 Tonnes/Hang-up and Draw Height Zone 15 East Central West
238 6861 440
In Situ Draw height (m) 30 40 193 457 318
170 448 374
60 621 833
The overall average oversize blast rate is 80 t/oversize blast though, like hang-ups, this varies by draw height and mining area. The rate of oversize drops beneath the average from heights of 40 m and above. Again, however, this trend may change as the East zone begins caving in this height range. Oversize is more frequent in the East than in the Central and West zones, but the difference is not as great as for hang-ups.
Figure 7 Fragmentation Analysis – Photo and resultant image Category C "Coarse" ore The general fragmentation tends to mirror the secondary breaking pattern. The following summarises findings to date: • The coarsest material is in the East followed by the West and Central zones. • The percentage of larger material (>2 m3 in size) is remaining steady with height. • The amount of fines is increasing with height, with a change being observed in the East and West zones at greater than 35 m draw height. This is much less than the 100 m predicted by BCF and is possibly due to enhanced secondary fragmentation and preferential movement of fines. On a monthly basis there has been an increase in the amount of coarse material underground. This fits well with the production data, which describes a reduction in yield over Q4 2003 to Q1 2004. This increase in the amount of coarse material underground is a general trend explained by the reduction in material coming from the undercut as a percentage of total material mined from underground. There is as yet insufficient material coming from a high draw height to counteract this reduction in undercut tonnage.
Figure 6 Average Tonnes/Hang-up Relative to Position in the Footprint (shown West to East) 4.3 Observed Fragmentation A site specific classification system ("flip-chart" method) is used to measure the distribution of fragment size across the production footprint. This system is based on an approach developed as part of the International Caving Study where image analysis is used to derive the distribution of the particle size reporting to the drawpoint. The method involves the determination of like fragmentation size ranges and then, by using a typical image, classifying the drawpoint into one of five size categories. This process enables data to be collected both simply and rapidly. With routine drawpoint inspections this data is easily merged with draw data. Figure 7 shows a photograph of a drawpoint and the Massmin 2004
4.4 Discussion Caving was initiated in the more faulted weaker centre section of the footprint. Stronger transgressive carbonatites predominate in the eastern sector and weaker banded carbonatites in the western sector. In addition, the Mica Fault, located at the western extremity of the footprint effectively concentrated stress within the western sector while by contrast a large dyke at the eastern extremity may have acted as a stress riser. The concentration of stress in the west induced additional fractures in the rock mass creating smaller blocks and thus finer fragmentation. Fewer fractures were generated in the east due to stress sheltering and a stronger rock mass. This resulted in larger blocks and an increased incidence of hang-ups. There is support for this hypothesis from the results of seismic monitoring that indicates relaxation (softening of the rock mass due to fracturing) in the West and increasing stress in the East (the less fractured rock mass able to attract more stress) together with the results from monitoring of secondary breaking and fragmentation. A number of fragmentation zones can be identified. These are: • Zone of undercut influence (0 to 20 m draw height): A rapid increase in fragment size as ground influenced by the undercutting process is pulled.
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• Zone of primary fragmentation (20 to 40 m draw height): Coarsest fragmentation, lowest yield per hang-up. Zone least impacted by caving induced stresses. • Zone of initial fining (40 to ? m draw height): Initial reduction in fragment size due to factors such as stress conditioning (as indicated by micro-seismic activity), and comminution in the draw column. Ongoing monitoring of secondary breaking performance and of drawpoint fragmentation will allow better definition of these and subsequent zones as the cave progresses 5 CONCLUSIONS Block caving at Palabora is at an early stage. Breakthrough into the overlying pit has occurred clearly demonstrating that competent rock masses can be successfully caved. However, there was a 30% under estimate in the size of footprint required to induce caving. The magnitude of this estimating error is large (particularly when the demands of feasibility studies are to estimate capital and operating costs to within +/- 15%) underlining the need for further work on cave prediction. Though the production ramp-up to 30,000 t/d is well underway it has been constrained by the secondary breaking process. This has resulted in the draw rate to be in balance or less than the caving rate. It is considered that in terms of cave performance this has been beneficial in allowing stresses to build and pre-condition the rock mass resulting in the creation of new joints and the extension of existing joints. The prediction of fragmentation allowed the critical nature of secondary breaking to be identified.
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Though fragmentation is substantially finer than predicted, the ability to identify the magnitude of the problem was remarkably good given the limited knowledge at the time of the feasibility study. Experience has shown that substantial demands are placed on secondary breaking equipment and that maintenance and task organisation are key factors in achieving breaking targets. Nevertheless, better methods are required for predicting fragmentation and the associated hang-up rates. ACKNOWLEDGEMENTS The paper presented here is the result of all those who participated in the Palabora Underground Mine Project and the role of the authors has simply been to report this work to a wider audience. The permission of Palabora Mining Company and Rio Tinto Technical Services to publish this paper is gratefully acknowledged. REFERENCES • Laubscher, DH, 1994. Cave Mining - the State of the Art. J S Afrr Inst Min Metall, 94(10);279-293. • Calder, K, Townsend, P, and Russell, F, 2000. The Palabora Underground Mine Project. Proceedings MassMin 2000. Brisbane (Ed: G Chitombo) 219-225. • Esterhuizen, G S, 1994. A Program to Predict Block Cave Fragmentation, Technical Reference and Users Guide. • Glazer, S and Hepworth, N, 2004, Seismic Monitoring of Block Cave Crown Pillar. To be published Proceedings MassMin 2004. Santiago.
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Geotechnical challenges of the transition from open pit to underground mining at Chuquicamata Mine Germán Flores, Codelco Norte Division, Codelco Chile
Abstract Chuquicamata mine is an open pit operation located in the northern part of Chile. This mine went into production in 1915, mining 1,000 tpd of oxide ore. Today it is a large scale operation mining 186,000 tpd of mainly sulphide ore. The present mine plan is for open pit operations to cease in year 2013 at a depth of 1100 m. Although the orebody continues below the bottom of the final pit shell, the cost increments associated with a deep pit operation do not allow for further open pit mining, therefore it become necessary to initiate a transition from open pit to underground mining. The underground operation will be implemented at depth, in a hard and massive rock mass and in high stress environments. To make this project economically viable requires application of a large scale and low cost underground mass mining method in order to achieve the required high production rates. The only methods that can achieve these requirements are block and panel caving. The transition from a large scale and deep open pit to underground cave mining at Chuquicamata will face with several geotechnical challenges. These include the presence of the large and deep open pit which will produce zones of stress concentrations and zones of low confinement, the magnitude of induced stresses due to the pit depth, cave propagation, simultaneous open pit and underground operations, the presence of the West fault and the shear zone, subsidence and water inflows. This paper discusses these geotechnical challenges identified at the scoping engineering stage of Chuquicamata’s project for a transition from open pit to underground mine by caving and describes technical strategies to reduce and manage associated risk at all stages of project development.
1 INTRODUCTION Chuquicamata mine lies at approximately 3,000 m elevation in the Atacama Desert of northern Chile, some 16 km from Calama city, in the Province of El Loa, and some 250 km north-east of Antofagasta city, II Region, as shown
in Figure 1. This mine is part of Codelco Norte Division which has three open pit operations with Chuquicamata being the biggest, as shown in Figure 2. Currently, Chuquicamata is one of the largest open pit mines in the world, as shown in Figure 3, with a strike length of 4.5 km in NS direction, a width of 2.7 km in EW direction and a depth
Figure 1: Map showing the location of Chuquicamata Mine Massmin 2004
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Figure 2: Aerial view of Codelco Norte Division
Figure 3: Aerial view of Chuquicamata mine looking to the North 592
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Figure 4: Reserves and geological resources at Chuquicamata orebody
Figure 5: Conventional panel caving method Massmin 2004
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2 GEOLOGICAL AND GEOTECHNICAL SETTING AT CHUQUICAMATA The Chuquicamata porphyry copper orebody is rectangular in plan, and dips vertically. The mineralization was controlled by the West Fault which is located at the toe of the West wall. From the fault to the West is waste and from the fault to the East is ore, as illustrated in Figure 4. About 2,470 Mt of ore, averaging 1.54% Cu, have being mined out from the Chuquicamata ore body since 1915, and 870 Mt will be mined out from 2004 to 2013 (final pit). However, the ore body is open at depth, with geological resources estimated to be 1,500 Mt at an average grade of 0.65% of Cu for the underground mining, as shown in Figure 4. At the Chuquicamata mine the predominant rock types are granodiorites and porphyries, whose western contact is defined by the West fault, a large regional fault with a NS trend, 4 to 6 m thick, and defining a 150 to 200 m wide shear zone on its western side. This shear zone has a poor to very poor geotechnical quality, and is located in the lower third part of the West Wall’s slopes. In the upper part of these slopes the rock is Fortuna granodiorite. On the eastern side of the West fault appears a massive quartz-sericitic rock, and beyond that porphyries with different types of alteration. Hence, from West to East the main rock mass types at Chuquicamata are: The engineering geology at Chuquicamata is such that twelve geotechnical units have been defined (Torres et al 2003), as shown in the plan view of Figure 6 and the EW and NS cross sec-tions shown in Figures 7 and 8, which also shown the main geological structures, as also the current and final pits. The main geotechnical units in the sector of interest to the transition project have the characteristics summarized in Table 1 (Flores et al 2004c). 594
Granodiorites
Waste
Fortuna granodiorite Moderately sheared zone Highly sheared zone
-------------- West fault ----------------------------------------------Quartz-sericitic rock / Highly sericitic rock Porphyries
Increasing grade
of 850 m. This mine started in 1915 and currently mines 88 Mt of ore and 115 Mt of waste. The current mine plan is to reach a depth of 1,100 m in year 2013. Although the ore body continues below of the final pit bottom the open pit, the operational cost at that depth will not allow for continued mining by the open pit method. Therefore, it has become necessary to initiate a transition from open pit to underground mining. This transition phase will require successfully overcoming a number of technical and economic issues (Arancibia and Flores 2004). The technical issues include the geotechnical challenges which must take into account the re-gional West Fault and its shear zone as shown in Figure 4, the presence of a large and deep open pit which will produce zones of stress concen-trations and zones of low confinement, a hard and massive rock mass and the depth of the un-derground excavations. It should be evident, therefore, that the decision on transition from open pit to underground mining should take into account the number of geotechnical factors which control the rock mass response during this phase. This is particularly crucial when simultaneous surface and underground operations are considered. The scoping study of this project has indicated that the most suitable underground mining method is panel caving based on the characteris-tics of the Chuquicamata deposit and the eco-nomic and business requirements of the project. Panel caving, illustrated in Figure 5, is considered to be the only method that could achieve high production rates and low operational costs. This paper presents and discusses the geotechni-cal challenges which may have a significant im-pact on the economic of this transition project from a large and deep open pit to underground cave mining at Chuquicamata mine
East porphyry with sericitic alteration East porphyry with chloritic alteration East porphyry with potassic alteration
Table 1: Geotechnical Units Geotechnical Unit Quartz-sericitic rock Highly sericitic rock East porphyry with sericitic alteration East porphyry with chloritic alteration East porphyry with potassic alteration UCS FF RMRL GSI
UCS (MPa)
FF (fract./m)
RMRL
GSI
20
1 to 5
55 to 65
70 to 85
10
> 10
35 to 45
25 to 40
31
1 to 5
60 to 70
55 to 70
84
1 to 10 55 to 65
55 to 65
85
1 to 10 55 to 70
55 to 75
Uniaxial compressive strength of the intact rock Fracture frequency (including weak veinlets) Laubscher’s rock mass rating Geological strength index
The stress field at Chuquicamata has been measured using a hydro-fracturing technique in deep vertical down holes. The in situ stress field is defined by a vertical stress proportional to the depth, with a magnitude in the range of 35 to 40 MPa at the elevation of a future UCL. The horizontal stresses are defined by minimum and maximum stress ratios, KMIN and KMAX, respec-tively. KMIN ranges from 0.5 to 1.0, with a direc-tion of N20ºE and KMAX varies from 1.0 to 1.7, with a direction of N70ºW (Torres et al 2003). These values will be verified using the CSIRO hollow inclusion technique to perform stress measurements from the exploration tunnels which will be available below the final open pit shell at the end of the year 2004. 3 THE TRANSITION PROCESS There are many near surface deposits that have considerable vertical extent. Although they are initially exploited by open pit mining, there is often a point where decisions have to be made to either continue deepening the pit or mining the same deposits by underground methods. At pre-sent several open pit mines are planning, or are in the process of implementing, a transition to underground mining. They include Bingham Canyon in USA, Chuquicamata and Mansa Mina in Chile, Grasberg in Indonesia, Palabora and Venetia in South Africa, Argyle, Mount Keith and Telfer in Australia. The decision to make the transition from open pit to an underground operation is often based on a simple determination of the NPV of the next feasible open pit
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pushback. Underground mining is only contemplated when a further pushback is shown to be uneconomic. However, any decision to go underground also requires consideration of a wide range of technical factors, and careful planning, which means a significant amount of time for achieving underground mining (up to 20 years has been suggested by Stacey and Ter-brugge 2000). This is in addition to the thorough assessment of the risks associated with the rock mass failure that accompanies underground cave mining, and its interaction with the open pit and the surrounding infrastructure. Perhaps one of the most important decisions, in the initial stages of a project for a transition from open pit to underground mining, is the definition of the most suitable underground mining method based on the characteristics of the deposit and, at the same time, the economic and business re-quirements of the mining company. If the business requires high production rates and low op-erational costs, then underground cave mining methods, such as block or panel caving, are the only methods through which these main objec-tives can be achieved. In such cases, it is desir-able that the open pit continues its operation during the first stages of underground mining, and that the underground mine gets to a high productivity quickly and before closure of the open pit operation. This means that there will be a period of simultaneous open pit and underground mining operations. This simultaneity implies an interaction between the open pit and the underground mining which makes the problem more complex than the typi-cal open pit or underground mine designs, be-cause the presence of the deep open pit will af-fect the stress field in which the underground mine will be developed and, conversely, the propagation of the caving will affect the stability of the surface crown pillar that defines the bot-tom of the open pit.
Figure 6: Geological units present in Chuquicamata Mine including lithology and alteration (Torres et al 2003)
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Figure 7: Section 3900 N illustrating the geotechnical units and the major structures (Torres et al 2003)
Additionally, there are many other factors or po-tential hazards that could make the problem even more difficult if they are not identified prior to making the transition from surface to under-ground mining. Some of the major hazards that could be expe-rienced in a transition from open pit to under-ground cave mining are: 1. mining by caving in relatively massive rock masses will induce seismicity which could trigger rockbursts; 2. if the caving propagation is arrested by the formation of a metastable cavity, a sudden failure of the cave back may occur which could trigger an air blast (de Nicola and Fishwick 2000); 3. an early break down or failure of the surface crown pillar, in simultaneous surface and un-derground operations, may cause an unex-pected end of the open pit operations, affecting the production plan and, if sudden, could put at risk the safety of the entire mine operations; 4. subsidence, which begins once the caving reaches the pit bottom, could affect not only the surface infrastructure close to the pit, but also the location of the main underground ac-cesses and infrastructure; and 5. the open pit could act as a catch-basin for heavy rainfall, increasing the risk of sudden water inflows and/or mudrushes into the un-derground mine. Even if these potential hazards are identified, it is difficult to assign a probability of occurrence to each one and to determine the potential safety risk/cost associated with it. This makes the issue of transition even more complex and challenging. It should be evident, therefore, that the decision on transition should not only be based on eco-nomic indicators, but must also take into account the number of technical factors which control the rock mass response during transition. This is par-ticularly crucial when simultaneous surface and underground operations are anticipated. The behaviour of a rock mass in a transition from open pit to underground mining by caving has been shown to be a subject of interest to today’s mining industry, because of the number of mines planning to start implementing the first engi-neering stages of such a transition. In addition, there are still a number of unknowns to be ad-dressed in order to solve key geotechnical issues associated with
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Figure 8: Section 3300E illustrating the geotechnical units and the major structures (Torres et al 2003)
such a transition, and the follow-ing questions must be addressed: 1. What is the optimum height of the ore column that can be mined safely from a economi-cal/geotechnical/ operational perspective? 2. Will the cave propagate upwards through the entire block height? 3. What is the minimum thickness of the surface crown pillar required to allow simultaneous surface and underground operations? 4. When is it no longer safe to be mining in the open pit while caving is occurring? How long could both mines operate simultaneously? 5. Will the subsidence generated by the under-ground mining affect the surface infrastructure surrounding the pit ? When? 6. What are the main geotechnical hazards, and how should they be dealt with? The potential consequences of an ill-defined transition project can be large, not only economi-cally but also environmentally, and even politi-cally. In addition, many aspects of the transition prob-lem are beyond the ranges of applicability of known solutions. For example, the simultaneous operation of the open pit and underground mines by caving methods requires a stable surface crown pillar between the cave back and the pit bottom. However, at the same time, cave propagation requires the failure of this pillar to connect to ground surface, so the definition of crown pillar failure is not the usual. Furthermore, the span of this surface crown pillar is much larger than the maximum span of surface crown pillar used in open stope mining. The answer to this and other questions requires an improved understanding of the behaviour of the rock mass, the mechanics of caving propagation, and the effects of a simultaneous surface and underground mining by caving methods. The quality and reliability of the geotechnical data is of paramount importance for the engineering of a transition project, and factors such as the strength, cavability and fragmentation of the rock mass could have a large impact on the project. 596
For example, if the layout design was based on a certain fragmentation finer than the actual one, the occurrence of hangouts would become a serious operational problem, and the need for secondary blasting and draw point repairs will be larger than expected. As the evaluation of the cavability of the rock mass is commonly based on Laubscher’s chart, the MRMR estimates must be as reliable as possi-ble, as illustrated by the following example: If the data available indicates that Laubscher’s RMR could vary from 55 to 65, a Monte Carlo simulation indicates that a value of 57 for MRMR have a 15% probability of exceedance, and caving initiation would require a hydraulic radius, HR, equals to 38, which corresponds to a 23000 m2 square area. If the data indicates that RMR varies from 45 to 65, the same analysis would indicate that a 22000 m2 square area is required (HR = 37). On the other hand, if the data indicates that RMR varies from 55 to 75, the same analysis would indicate that a 37000 m2 square area is required (HR = 48). Hence an overestimation of the lower bound for RMR has a minor effect on the project (-4%), but a underestimation of the upper bound for RMR could have a major impact on the project (+61%). Therefore, any additional cost incurred improving the reliability of the geotechnical data must be considered a very good investment and, at the same time, an insurance against changes from the expected geotechnical setting. The key geotechnical issues that are considered relevant in a project for a transition from open pit to underground mining by caving are: 1. The selection of the undercut level, which defines the block height. Proper selection of the block height is particularly important when there is the potential for simultaneous open pit and underground operations. 2. The cave initiation and propagation through the rock column to be caved. This is important to determine if the rock mass will cave or stall. 3. The minimum crown pillar thickness required for simultaneous open pit and underground operations. 4. Subsidence due to the failure of the pit slopes after the connection of the cave back with the pit bottom, with a
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Figure 9: Diagram illustrating the different stages of caving propagation in a transition from open pit to under-ground mining by caving Massmin 2004
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zone of influence with an important width. Knowledge of the extent of the subsidence zone is crucial for the location of the main accesses to the underground mine and the underground infrastructure. This is in addition to assessing the likely impact of the subsidence zone to the existing surface infra-structures, which eventually would have to be relocated. 5. The presence of a large open pit above the underground mine increases the likelihood of water inflows/mudrushes because the open pit could act as a catchment for heavy rainfalls, and the water could seep into the underground mine through fractures induced by the caving and the broken rock above the UCL. This may cause disruptions in production and, worst, safety risks for the underground operation. Figure 9 illustrates the different stages of the caving process in a transition for open pit to un-derground mining. The detail and mechanisms associated with these issues have been discussed in an accompanying paper (Flores et al, 2004b).
4 MAIN GEOTECHNICAL CHALLENGES AT CHUQUICAMATA The planned transition from a large scale and deep open pit to underground cave mining at Chuquicamata is expected to face a number of unique geotechnical challenges, as illustrated in Figure 10. Given their potential impact to the transition project these need to be addressed dur-ing the early and subsequent design stages of the project. These main geotechnical challenges are: 1. The presence of the large and deep open pit, which will produce zones of stress concentra-tions and zones of low confinement. These in-duced stresses will likely affect the propaga-tion of caving and must therefore be consid-ered in evaluating the likelihood of caving propagation through the whole ore column to be cave, as illustrated in Figure 11. 2. The level of induced stresses due to the pit depth in addition to the height of the block to be caved. When the final pit is reached in 2013 with a depth of 1,100 m, the undercut level will be located at a depth of around 1,500 m from surface. Therefore, the induced stresses are likely to be high and induced seismicity will be expected during under-ground mining, which eventually could gen-erate rockbursts. Figure 12 illustrates the con-sequences of a rockburst phenomenon in a caving operation. Due to this a seismic monitoring system is considered an absolute
Figure 10: Geotechnical challenges associated with a transition project at Chuquicamata 598
need, and it should be implemented when the initial developments of the underground mine begin. 3. Cave initiation and propagation. The initial stage of the underground mining will be in a hard and massive rock mass, where cave ini-tiation and propagation may be difficult. As the cave propagation approaches the pit bot-tom the rock mass above the cave back would be affected by the higher stresses associated with the presence of the open pit, which may affect the rate of the caving propagation by either accelerating or arresting the process. The main factors affecting cave propagation in a transition from open pit to underground cave mining are shown in Figure 11. Also, it becomes important to define and implement an instrumentation system to monitor the de-velopment of the cave. Considering the expe-rience at El Teniente (Rojas et al 2000) and Palabora mines (Glazer and Hepworth 2004), this system would include seismic instru-mentation, TDR´s and borehole camera ob-servations. 4. Simultaneous open pit and underground mine operations. The economic and business re-quirements of Chuquicamata are such that a period of simultaneous open pit and under-ground mining would be required. Hence, at least for a certain period, a stable crown pillar must be maintained between the cave back and the pit bottom. This period must be defined considering the stability of the crown pillar and the fact that its thickness is reducing due to the ore draw from the underground mine, as illustrated in Figure 13. A longer period of simultaneity requires a larger block height. A low block height would lead to a very short period of simultaneous operation, which could be non practical. Once the period of simultaneity has been established it is pos-sible to define when the underground mining should begin. 5. The presence of the West fault and the shear zone. The West fault forms an abrupt contact between the ore and the waste. The waste is a soft, weak and highly fractured rock mass which potentially could become a source of early dilution of the ore if the cavity reaches the West fault before connecting with the pit bottom. Hence, a rib pillar is required between the West fault and the undercut area as illustrated in Figure 10. If this rib pillar is too thin it could fail and early dilution may occur. On the other hand, if it is too wide some high grade ore would not be mined. Figure 14 il-lustrates the rib pillar at Chuquicamata. 6. Subsidence. Once the caving connects to the pit bottom the pit will become a subsidence crater with a zone of influence extending be-yond the pit perimeter. Of course, this condi-tion will evolve with time, and due to slope failures and the extension of the undercut area this crater will grow. The geometry of a subsidence crater at Chuquicamata will be defined by the crater depth, H, and the angle of break, a, which is the angle between the edge of the undercut level and the start of discontinuous deformations (large tension cracks). The influence zone adjacent to the crater perimeter is defined in terms of the in-fluence width, dIZ. The a and dIZ terms depend on the rock mass quality and the pres-ence of major geological structures. The morphology of the subsidence phenomenon at Chuquicamata is illustrated in Figure 15. The importance to define the a and dIZ is due to the requirement to determine the location of the main accesses to the underground mining and the location of the underground infrastructure which must be outside of the influence zone. In addition, it is necessary to know if the subsidence due to the under-ground mining will affect the current surface infrastructure related to the open pit opera-tions. 7. Ground water. Due to the presence of ground water in the slopes of Chuquicamata’s open pit and some rains during the Bolivian winter (January and February), there is a non zero probability of inrushes of water or mud into the underground mine. These inflows/mud-rushes could be facilitated by the presence of major geological structures
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Figure 11: Induced stresses affecting the caving propagation
Figure 12: Consequences of rockburst phenomenon in a panel caving operation Massmin 2004
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Figure 13: Diagram illustrating the surface crown pillar developed in a transition from open pit to underground cave mining and/or the frac-tures and the cave generated by the caving. Water inflows and/or mudrushes can cause damage to underground mines by caving methods due to the sudden inflow of wa-ter/mud from drawpoints, ore passes or other underground openings. Figure 16 illustrates the consequences of a mud rush in a mine by caving.
5 CONCLUSIONS The geotechnical challenges associated with the planned transition from open pit to underground cave mining at Chuquicamata have been identified at the early stage of the project (scoping study), and Codelco Norte Division is developing appropriate technical strategies to reduce and manage the potential risks associated with the geotechnical challenges identified. These strat-egies are now being incorporated into the ongoing engineering studies of the transition project. Hence, and as an integral part of the overall tran-sition project at Chuquicamata, a worldwide benchmark study and literature review on transi-tion from open pit to underground mining by caving was undertaken (Flores et al 2004a). This was carried out through the International Caving Study Stage II (ICS-II), managed by the Julius Kruttschnitt Mineral Research Centre, Brisbane, Australia, of which CODELCO is one of the sponsors. The benchmark concluded that there is currently neither sufficient experience in transition for deep pits nor available design methodologies in spite of the topic’s importance to mining in-dustry. The only documented transition involv-ing a large open pit and underground mining by caving is Palabora mine, South Africa (Glazer and Hepworth 2004). In 600
the near future a number of large open pit operations which include Bingham Canyon and Grasberg will be undergoing similar transition. As a result, the subject of transition was included as one of the major research topics in the ICS-II. This research was focussed on developing guidelines on rock mass charac-terization, caving propagation, surface crown pillar, subsidence and water inflows, all of which are important geotechnical issues for consider-ation in a transition project. The outcomes of this research will be used in the Chuquicamata transition project as part of the overall Codelco Norte strategy to ensure the suc-cessful transition from open pit to underground cave mining given the geotechnical challenges identified. ACKNOWLEDGEMENTS The author acknowledges the Division Codelco Norte for the permission to publish this paper. He wishes to also thank the Geotechnical Group of Codelco Norte Division for having provided material used in the paper. Special thanks are given to Professor E T Brown AC and Drs. Antonio Karzulovic and Gideon Chitombo for their encouragement and technical discus-sions. REFERENCES • Arancibia, E and Flores, G, 2004. Design for underground mining at Chuquicamata orebody. Scoping engineering stage. Proceedings MassMin 2004, Santiago, (Ed: A Karzulovic and M Alfaro). • de Nicola, R and Fishwick, M, 2000. An under-ground air blast - Codelco Chile - Division Salva-dor. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 279-288.
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Figure 14: Rib pillar between the caving cavity and the west fault
Figure 15: Subsidence phenomenon Massmin 2004
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Figure 16: Effects of a mud rush in an underground mine by caving.
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Australasian Institute of Mining and Metallurgy: Melbourne. Flores, G, Karzulovic, A and Brown, E T, 2004a. Current practices and trends in cave mining. Pro-ceedings MassMin 2004, Santiago, (Ed: A Karzu-lovic and M Alfaro). Flores, G, Karzulovic, A and Brown, E T, 2004b. Evaluation of the likelihood of cave propagation in mining engineering practice. Proceedings MassMin 2004, Santiago, (Ed: A Karzulovic and M Alfaro). Flores, G, Karzulovic, A and Gonzalez, G, 2004c. Geotechnical considerations for the scoping engi-neering stage of the transition project from open pit to underground mining at Chuquicamata mine (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile. Glazer, S and Hepworth, N, 2004. Seismic monitor-ing of block cave crown pillar – Palabora Mining Company,
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RSA. Proceedings MassMin 2004, Santiago, (Ed: A Karzulovic and M Alfaro). • Rojas, E, Cavieres, P, Dunlop, R and Gaete, S, 2000. Control of induced at El Teniente Mine, Codelco – Chile. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 775-784. Australasian Institute of mining and Metallurgy: Melbourne. • Stacey, T R and Terbrugge, P J, 2000. Open pit to underground – transition and interaction. Pro-ceedings MassMin 2000, Brisbane, (Ed: G Chi-tombo), 97-104. Australasian Institute of Mining and Metallurgy: Melbourne. • Torres, R, Araya, E, Córdoba, S y Domínguez, O, 2003. Geotechnical characterisation for the scop-ing engineering stage of the transition from open pit to underground mining at Chuquicamata mine (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile.
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Design for Underground Mining at Chuquicamata Orebody. Scoping Engineering Stage Ernesto Arancibia, Germán Flores, Codelco Norte Division, Codelco Chile
Abstract The current Chuquicamata mine’s long term mine plan has established that the final pit will be reached in year 2013 when the pit reaches a depth of approximately 1,100m. However, since there are still mineable geological resources underneath the final pit, the mine is currently assessing the feasibility of a transition to underground mining using panel caving method. This is the preferred caving method as it potentially should allow for highly mechanized operations with high production rates at low production costs. This paper presents and discusses the panel caving parameters and associated cost estimations that have been proposed at the scoping engineering stage including a number of both technical and human resources issues likely to be faced and how they are to be addressed. This scoping study is based on a combination of a series of preliminary analyses, other Codelco panel cave experiences supplemented by a world benchmark on transition. These may be refined as more data is obtained and made available followed by more detailed analyses. A key development of the transition from the Chuquicamata large open pit to the planned underground panel cave to date has been the identification of the potential major technical issues. These will now be taken into account in all subsequent stages of the transition and underground design process.
1 INTRODUCTION The Chuquicamata open pit mine is located in the II Region of Chile, El Loa Province, 16 km north from the city of Calama and 1,600 km from Santiago, as shown in Figure 1.
Figure 2: Aerial view of Chuquicamata open pit The operation at Chuquicamata began in 1915 and currently is removing about 600,000 tpd with a waste-ore at a stripping ratio of 2.1:1. Based on the current long term mine plan, the open pit mine will reach the final pit at a depth of 1,100 m in the year 2013, as illustrated in Figure 3. However, the continuity of mineraliza-tion has been proven at depth, and plans are now under way to carry out a transition from open pit to underground mining by a panel caving method. This paper presents the mine design param-eters for this transition project which have been determined from the scoping study carried out recently by Codelco Norte Division. The paper briefly discusses a number of technical challenges that this project has to overcome to ensure a successful transition project at Chuquicamata mine. Figure 1: Chuquicamata mine location
2 MINE DESIGN OF THE TRANSITION PROJECT
Currently (2004), the Chuquicamata pit is 4.5 km long, 2.7 km wide and around 850 m deep, as shown in Figure 2.
The scoping engineering study carried out recently by Codelco Norte Division indicates that it is feasible to exploit
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Figure 3: Overall schematic view showing geologic reserves
the ore below the final open pit envelope using panel caving as shown in Figure 4. Based on a combination of a series of preliminary analyses, other Codelco panel cave experiences supplemented by a world benchmark on transition (Flores et al 2004a), the following initial design parameters are proposed for the Chuquicamata panel cave. These may be refined as more data is obtained and made available followed by more detailed analyses. The undercut level, UCL, is planned to be located at a depth of 1,500 m below surface or approximately 400 m below the final pit bottom. The extraction level, EXL, would be located 18 m below the UCL. The extraction geometry would be the El Teniente layout with 15 ¥ 17.32 m (drawpoint x production spacing). The planned caving initiation area will be located just below the pit bottom where the block height is expected to be minimum and also where the rock mass is considered favourable for initiating the initial caving process. Given the characteristics of the local geotechnical environment (Torres et al 2003), the minimum area required to initiate the caving is estimated to be 15,000 m2 with a square or rectangular shape. A slot is considered a necessary measure to facilitate the initial cave initiation. The mining sequence will start from the centre of the footprint and then extended in two fronts, one to the East and the other to the West and after that towards the North and the South. The undercut rate would be of the order of 3,000 m2/month based on other Codelco panel cave experiences and the cave front would have an orientation of N85ºE. The average draw rate is estimated to be of the order of 0.25 m/day (0.65 t/m2/day). The maximum production rate is expected to be 40,000 tpd for each extraction sector or panel and 125,000 tpd from the whole underground mine when the caving reaches steady state or continuous caving. The production ramp-up period is expected to be 8 years. 604
The angle of break is estimated to be 60º for the East wall and 50º for the West wall. The influence zones are estimated to be125m for the East wall and 250m for the West wall. The rib pillar designed to reduce possible early dilution from the West fault is estimated to be 80 m thick at its base. The minimum surface crown pillar thickness required for the simultaneous open pit and underground operations is estimated to be 200 m. This will allow for approximately to 2 to 3 years of simultaneous mining of the final open pit and initial underground panel cave. Based on the experience of the other Codelco Chile operations (El Teniente, Andina and Salvador mines) the transition project has considered that the mine production plan would be affected by the occurrence of 3 collapses of up to 4,000 m2 each one during the first 5 years of operation. An analysis carried out also predicts a probability of a metastable cavity forming before continuous caving is achieved possibly affecting an area of 12,000 to 15,000 m2. As a result, the induced seismicity is expected to be high and as such a seismic monitoring system has been included during the early development stages. The support system is expected to be similar to the support used at El Teniente mine (Rojas et al 2000). The main accesses to the underground mine are to be located in the East wall and outside the estimated influence zone and would consist of two declines. One is to be used for production with two conveyor belts and the other for maintenance. In addition to these declines, there will be a service shaft and ventilation shafts. Figure 5 shows the location of the planned declines, and shafts. The total investment of the transition project has been estimated to be around US$ 500 millions and the operational cost is expected to be as low as 3.4 US$/t.
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Table 1: Panel caving parameters and costs estimation at scoping engineering stage (Arancibia 2002a and b and, Flores et al 2004b) Parameter
Recommendation
Geological resources
1,500 millions of tonnes with an average grade of 0.65% of copper and 0.03% of molybdenum
Underground mining method
Panel caving
Main accesses to the underground mine
2 declines, 1 service shaft and 4 ventilation shafts
UCL depth from surface
1,500 m
Block height
400 m
Crown pillar between UCL and EXL
18 m
Estimated initial fragmentation
35 to 50% > 2 m3
Initial area to initiate the caving process
15,000 m2 with an square or rectangular shape
Cave initiation sector
Below the pit bottom and where the rock mass quality is favourable to propagate the cave
Measure to facilitate the cave initiation
Slot
Mining sequence
Initiate the undercutting just below and centred to the pit bottom projection then to expand the area to the East and to the West. When the undercut has reached the footprint width then the sequence will go to the North and South at the same time.
Undercutting rate
3,000 m2/month
Cave front orientation
N85ºE
Extraction layout
El Teniente 15 m × 17.32 m
Tonnes per drawpoint
240,000 to 280,000 t
Draw rate
0.25 m/day average (0.65 t/m2/day)
Maximum production rate per sector
40,000 tpd
Total production rate from underground mine
125,000 tpd
Ramp-up period
8 years
Angle of break (subsidence)
60º for the East wall and 50º for the West wall
Influence zones (subsidence)
125 m for the East wall and 250 m for the West wall
Rib pillar width
80 m at its base
Minimum surface crown pillar thickness
200 m
Maximum time for simultaneous open pit and underground operations
2 to 3 years
Collapses
Expected the occurrence of 3 collapses affecting about 4,000 m2 each one during first 5 years of operation.
Metastable cavity
May occur at the beginning of the underground operation affecting an area of 12,000 to 15,000 m2
Rockburst
A seismic monitoring system at the early stage of the development phase.
Capital investment
US$ 500 millions
Operational costs
3.4 US$/t
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Table 1 is a summary of the expected main panel caving design parameters and associated costs of the transition project at Chuquicamata mine. 3 TECHNICAL CHALLENGES The technical challenges anticipated for the Chuquicamata panel cave are expected to be relatively significant given the planned high pro-duction rates and the requirement for low cost. The challenges will be exacerbated by the tran-sition from a large open pit to underground mining at great depths (1,100m), where both the geological and geotechnical environments are considerably more aggressive than currently experienced. The key technical challenges are related to the following (Arancibia 2002 a and b): 1. the depth of the underground operations which will be around 1,500 m below surface; 2. the location and type of the main accesses to the underground mining; 3. the construction period for a large scale underground caving project; 4. the block height, cavability and fragmentation of the anticipated hard and massive rock mass; 5. the number of mining fronts and sequencing of production sectors to achieve the produc-tion target of 125,000 tpd; 6. the mine layout and mining sequence; 7. draw control constraints, eg differential draw to control the West fault; 8. the ramp-up period required to reach the design production; 9. the interaction between the open pit and underground operations; 10. the ventilation system;
11. the material handling system required to achieve the production from a depth of around 1,500 m; 12. the location of the underground infrastruc-ture; and 13. the logistics system to mobilize the workers, equipments and materials to different places of the underground mine. In addition to these technical issues, there are a number of geotechnical challenges associated with the transition project at Chuquicamata mine. They are (Flores 2004): 1. the presence of the large and deep open pit, which will induce zones of stress concentra-tions and zones of low confinement; 2. the level of induced stresses due to the pit depth in addition to the height of the block to be caved; 3. cave initiation and propagation; 4. simultaneous open pit and underground mine operations; 5. the presence of the West fault and the shear zone; 6. subsidence; and 7. ground water These geotechnical challenges are summarized in Figure 6 and they have been discussed in detail in an accompanying paper (Flores 2004). In addition to the anticipated technical and geo-technical issues there are a number of social and human resources aspects that need to be con-sidered when converting form a large scale and deep open pit to underground mining by caving method. These equally need to be resolved with the same degree of urgency in order to ensure a successful transition from open pit to under-ground mining by caving method. In the case of Chuquicamata, the social issues include: 1. The workforce. The number of employees required in underground mining will be much lesser than the current workforce working in the open pit operations.
Figure 4: Panel caving underground mining method 606
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Figure 5: View showing the open pit and main accesses to the underground mine
Figure 6: Geotechnical challenges involved in the transition Massmin 2004
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2. The worker’s skills. The technology used in underground mining is completely different from that currently used in open pit oper-ations. 3. The contractors companies. The skills of the contractor’s employees and their companies should be different from that required in open pit mining. 4. The suppliers. The equipments and materials required at the underground operation will be different to those required at the open pit. 4 JUSTIFICATION FOR THE TRANSITION The technical and geotechnical challenges an-ticipated during and after the transition from a large open pit to a large panel cave may seem overwhelming to the point where undertaking such a project may be questionable. There are however a number of positive attributes associated with overall Chuquicamata underground project that have justified its initiation. These are: 1. the availability of large geologic resources below the final pit amounting to 1.5 billion tonnes of ore with average grades of 0.65% copper and 0.003% molybdenum ; 2. the availability of modern surface ore pro-cessing infrastructure, currently used to con-centrate, smelt and refine the ore from the Chuquicamata open pit; 3. the availability of the other geologic resources in the same Codelco Norte mining district which should allow for the careful planning of the transition from open pit to underground mining at Chuquicamata deposit without being subjected to much production pressures; 4. as a group, Codelco has considerable experi-ence in mining ore deposits by block/panel caving methods in relatively hard, massive rock mass and high stresses achieving large production rates and low operating costs. These should benefit the new Chuquicamata project; 5. the availability of the financial resources required to undertake the investment for this transition project, expected to be around US$ 500 million; and 6. the improvement of the NPV of Codelco Norte Division and Codelco Chile. 5. MINING STRATEGIES To date, the engineering objective has been to demonstrate to Codelco Norte that underground mining by panel caving beneath the final Chuquicamata open pit was feasible while meeting the economic expectations or criterion set by Codelco Chile (Arancibia 2002a and b). In order to ensure a successful transition phase, the ramup period has been estimated to be around eight years. This is considered feasible also given other geologic resources that Codelco Norte Division has available in the same mining district which have still not been included in the current mine plans. Hence, the strategy devel-oped for the transition project will need to con-sider bringing these geologic resources into the mine plan. The gradual increase of production from the underground mine will allow proper management of the geotechnical challenges identified at the early stage of the transition project thereby reducing the associated risks. From an economic perspective, this transition phase will require the simultaneous open pit and underground operations. As production from the open pit declines, underground mining will need to increase until it reaches the designed produc-tion rate. The strategy is also to design the underground mine with a low initial investment, which will mean beginning the underground production in one sector and then expanding into several production areas in different stages. This strategy offers the advantage of allowing the investments to be made over a much longer time frame. 608
Human resources are crucial for a state-owned company such as Codelco Chile. Therefore, con-sideration needs to be given to the utilization of the manpower the Corporation already has. The human resources strategy must be based on the retraining of site personnel to work in under-ground mining, bringing personnel with experi-ence in caving operations from the others Codelco’s Divisions and the use of contractors to develop and construct the underground mine. However, the above strategy should not necessa-rily prevent introduction of new technologies needed for the effective and efficient underground mining. The following technologies are therefore being considered: 1. Automation. This technology should be fo-cused on development, construction and pro-duction activities in order to reduce the time and costs associated with these tasks. Auto-mation should also allow for the improvement of the underground environment thus creating better quality environmental condi-tions for the workers. 2. New ore handling systems. Most of the underground cave mines have relied on the use of conventional LHD equipment. In the case of Chuquicamata underground project the plan is to look for alternative equipment with lower operating cost, fewer maintenance requirements, and better adaptability to new ore extraction design and handling concepts. 3. Development and construction of under-ground excavations. One of the key param-eters that potentially has a large impact on the economic indicators of the project, is the time required to develop and construct the transition project at Chuquicamata. New technologies are therefore required to reduce this time. 6. CONCLUSIONS Based on the assumptions made to date, the scoping engineering stage has shown that the transition from open pit to underground mining by panel caving at the Chuquicamata deposit is feasible. Aspects of the transition that may have a signifi-cant impact on economic indicators have been identified and engineering solutions are now being proposed and sought in order to better reduce their impact on the NPV. The geotechnical challenges identified need to be addressed at the earliest stages of engineering so as to reduce the potential consequences of undesirable events that could occur during the operation of this project. The human resources issues need to be solved early in order for the workforce to be prepared with the skills needed to achieve the production rate and the low costs required by the Codelco Norte Division. The challenges anticipated with this transition now require that next engineering stages focus on the following aspects: 1. rock mass characterization; 2. cavability and fragmentation assessment; 3. block height assessment and potential dilution control; 4. variants of the proposed panel caving method; 5. main accesses to the underground mine; 6. development and construction of the under-ground excavations; 7. material handling systems; 8. production ram-up to achieve the production rate of 125,000 tpd in 8 years 9. production sectors required to achieve the production rate; 10. simultaneity of open pit and underground mining; 11. geotechnical hazards such as air blasts, col-lapses, hang ups, rockbursts, subsidence and water inflows; 12. new mining technologies; 13. human resources; and 14. costs
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The mine design is currently based on existing technology, but the overriding philosophy is to use proven developments in technology and automation for the benefit of the operation in terms of safety and productivity.
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ACKNOWLEDGEMENTS • The authors are grateful to all their colleagues of the Gerencia de Recursos Mineros y Desar-rollo that helped them during the development of this work. Also, the authors want to acknow-ledge the permission given by Division Codelco Norte to publish this technical paper.
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REFERENCES • Arancibia, E, 2002a. Pre-scoping engineering study I for the underground mining at Chuquicamata ore body (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile. • Arancibia, E, 2002b. Pre-scoping engineering study II for the underground mining at Chuquicamata (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile. • Flores, G, 2004. Geotechnical challenges of the transition from open pit to underground mining at Chuquicamata
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mine. Proceedings MassMin 2004, Santiago, (Ed: A Karzulovic and M Alfaro). Flores, G, Karzulovic, A and Brown, E T, 2004a. Current practices and trends in cave mining. Pro-ceedings MassMin 2004, Santiago, (Ed: A Karzu-lovic and M Alfaro). Flores, G, Karzulovic, A and González, G, 2004b. Geotechnical considerations for the scoping engi-neering stage of the transition project from open pit to underground mining at Chuquicamata mine (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile. Rojas, E, Molina, R, Bonani, A and Constanzo, H, 2004. The pre-undercut caving method at the El Teniente mine, Codelco Chile. Proceedings MassMin 2000, Brisbane, (Ed: Gideon Chitombo), 261-266. Australasian Institute of Mining and Metallurgy: Melbourne. Torres, R, Araya, E, Córdoba, S y Domínguez, O, 2003. Geotechnical characterisation for the scop-ing engineering stage of the transition from open pit to underground mining at Chuquicamata mine (in Spanish). Technical Report, Codelco Norte Division, Codelco Chile.
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Mine design of The Argyle underground project Damien Hersant, Senior Underground Mining Engineer – Feasibility Study
Abstract Argyle Diamonds operate a 10Mtpa open pit diamond mining operation in northern Western Australia. Current schedules show the open pit finishing production by the end of 2007, having reached its economic depth. Some 60Mt of resource is below the ultimate pit bottom and consequently Argyle Diamonds is undertaking a Feasibility Study on the underground mining of this resource. The ‘Vision’ for the project is based on setting benchmarks for safety, productivity and sustainability. This will be achieved through the application of best practice designs and techniques, high levels of automation and remote operation and a commitment to the vision by all those involved. In essence, the Argyle Underground Mine will be designed and operated as a safe and predictable ‘rock factory’. This paper presents various aspects of the mine design philosophy, underground conditions and their management and the incorporation of the Argyle vision into the design.
1. INTRODUCTION Argyle Diamond Mine, a 100% owned subsidiary of Rio Tinto Limited, operates the Argyle mine located in the eastern Kimberley region of Western Australia (Figure 1). The operation comprises a large open pit mine, centred on the lamproite AK1 orebody, feeding a processing and recovery plant. Diamond product is shipped to Perth for sorting, further processing and marketing.
January 2003, examining various underground mining strategies, approvals processes and organisational options. In February 2003 approval granted by the Rio Tinto Investment Committee for Argyle Diamonds to undertake a two year Feasibility Study into underground mining of the AK1 orebody. Approval was also given for the development of an Exploratory Decline to further improve knowledge of the underground environment. 2. THE VISION In embarking on the Feasibility Study, all aspects related to the project were to be undertaken with the following requirements in mind: • Argyle will set the benchmarks for safety and productivity in underground mining through: - Commitment to safe design and risk assessment throughout the design stage; - Application of best practice designs and techniques; - High levels of automation and remote operation, thereby removing people from hazardous work areas; and - Commitment to this vision. • Argyle will be designed and operated as a safe and predictable ‘rock factory’. • The project will provide sustainable outcomes for Argyle and for the people of the East Kimberley region.
Figure 1 – Location Map Currently, the AK1 operation moves over 80 million tonnes of rock to treat 10 million tonnes of ore annually. The current schedule, sees completion of open pit mining of the AK1 orebody by 2007. The AK1 and associated mineralistion is known to extend some 600m below the planned pit bottom. Underground mining of this resource has been investigated a number of times between 1995 and 2000. In July 2000, an Order of Magnitude Study was conducted to provide a fresh look at underground mining given the alternative of a deeper open pit plan and the confirmation of further resources at depth. This study indicated good potential for underground extraction. In light of this, a PreFeasibility Study was undertaken between July 2001 and 610
In keeping with this Vision: • frequent risk assessments and updates have and will continue to be undertaken; • safety experts/personnel have been employed from the early stages of the project to ensure sufficient resources are available to proactively deal with safety issues; • communication with equipment suppliers is frequent to ensure that technological developments are known and incorporated into the design as appropriate; • Argyle is represented at the International Caving Study which continues to pursue improvements in the understanding and processes used in the design, planning and control of cave mining; • a ‘Lessons Learned’ methodology has been incorporated into the Argyle project which draws on learning’s and experience from other Rio Tinto operations and projects;
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• Argyle is undertaking the transition to a ‘Kimberley Centric’ workforce, with increased local employment; and • where possible, local Kimberley based companies will be used for on-going work at the mine. 3. REGIONAL GEOLOGY The AK1 deposit is a volcanic intrusion of Lamproitic tuff and magmatic Lamproite intruded into a stratigraphic sequence of interbedded quartzites, siltstones and mudstones that overlie dolerite and basalt units The deposit extends below the base of the final pit, plunging steeply towards the south and dipping to the west at approximately 70o. The main area of the AK1 orebody, where a majority of the mine production will come from, thins and bifurcates with depth. Figure 2 shows a Long Section of the orebody and LOM Pit.
Figure 2 – Long Section of the Argyle Deposit
4. GEOTECHNICAL CONDITIONS There are numerous regional faults within the area however the dominant structural feature expected to be a major controlling factor for any mining is the NNW trending Gap Fault, which is located at the northern end of the AK1 orebody. Infrastructure development is likely to intersect at least three main geological structures, namely the Gap Fault, Lamboo Thrust Fault and Eastern fault. The proposed development will also pass through the contact margins of the AK1 pipe, which in many places is a wide zone of healed brecciated lamproite and host rock. Table 1 below indicates the rock types and associated indicative UCS and IRMR results. TABLE 1 – Indicative UCS and IRMR Values Rock Unit
Rock Type
UCS (Mpa)
IRMR
RCFM (Pvb)
Basalt
50
46 - 49
RCFM (Pvd)
Dolerite
50
46 - 49
LCG (Pb)
Granite
110
Interbedded
35
40 - 45
AK1 (orebody)
Lamproite Sandy Tuff
80
51 - 57
RCFS
Mudstone
20
35 -40
RCFS
Quartzite
120
47 - 50
RCFS
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It can be seen that, in general, the orebody is more competent than the surrounding stratigraphic units, especially the mudstones. This is a major consideration in determining optimum locations for permanent infrastructure such as crusher chambers. Acoustic emission and Hydraulic Fracture stress measurements were taken from boreholes drilled within the pit. These methods gave conflicting results however and consequently design flexibility is considered essential. Additional stress measurements will be undertaken from the Exploratory Decline. 5. MINING METHOD In keeping with ‘The Vision’ defined for the Argyle Underground Mine, selection preferences of mining methods tended towards non-entry methods such as block caving and sublevel caving. These methods make it possible to implement high levels of automation, control and monitoring, thereby minimising the number of people required to work in potentially hazardous conditions. They also create an underground ‘rock factory’ environment, which can be operated in a safe and predictable manner. Earlier mining studies identified that open stoping options led to poor recoveries (low percentage extraction) as well as having the risks associated with entry methods, the stope and fill methods were not economically viable and it was concluded that only the caving methods, more particularly block caving, could be seriously considered for mining of the AK1 ore body. The grade, size, geometry and rockmass conditions of the AK1 deposit are also such that it is most appropriate to mine using bulk extraction mass mining methods. During the Pre-Feasibility Study, a number of mining scenarios where assessed. From these, two options were taken forward. Further work undertaken as part of the Feasibility Study identified that the option of an extended block cave, whose footprint extends out into the less competent footwall sedimentary units, holds the greatest value. Below the cave extraction level, a sublevel cave mining method will be used. 6. BLOCK CAVE The block cave area is some 230m in depth, below the base of the final pit, 450m along strike and 190m wide, giving a mining block of approximately 50Mt. The extraction level itself is 466m below the surface. Undercut Strategy An advanced undercut strategy is proposed. In this strategy a limited amount of extraction level development is done prior to the start of undercutting. At Argyle, this development will consist of the establishment of the footwall and hangingwall perimeter drives, extraction galleries and drawpoint stubs. Undercutting on the level above will then begin and the extraction level drawpoints and drawbells will be developed once de-stressed conditions have been established. De-stressed conditions are assumed when the drawpoint development lags behind the undercut advance by 45o (15m). Profile and Footprint The undercut footprint for Argyle is an irregular rectangular shape with dimensions 450m x 190m. The hydraulic radius considered necessary for caving at Argyle is ~40m. The proposed footprint has a hydraulic radius of 67m, and as such cave propagation is not considered to be an issue.
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The undercut profile, as shown in Figure 3, is described as a narrow inclined undercut and is the profile adopted by both Northparkes and Palabora.
allowing each gallery to be isolated from the perimeter drives and minimizing interactions between automated and manual activities. • ventilation and drainage raises connect directly into the mine exhaust system on the Transfer level such that all heat generated on the Extraction level is removed quickly and efficiently. Ground support design and implementation will be a major consideration in the construction phase of the Extraction level with emphasis being placed on ‘getting it right the first time’ in order to avoid costly rehabilitation and delays. A workshop/cribroom complex will be located within the orebody at the southern end of the Extraction level. Regular servicing and minor maintenance will be carried out at this facility with all major maintenance work being undertaken at the surface workshop. The Extraction Level Layout is shown in Figure 4.
Figure 3 – Undercut Profile Direction An analysis was undertaken using FLAC3D to determine the best direction in which to advance the undercut. Six different undercut directions were assessed with the preferential direction for undercutting being from the northwest corner to the south-east. This recommendation was based on: • the undercut progressing from weaker to stronger rock giving improved cavability. • the orientation to major structure is such that major wedges should not be formed. Figure 4 – Extraction Level Layout Extraction Level In defining the design layout of the Extraction Level, a number of considerations were taken into account, the main ones being automation, ground support, ventilation and production flexibility. In order to comply with the Argyle Vision, a high level of automation is needed. The single extraction horizon makes it relatively easy to setup and maintain the communication infrastructure required for the automated systems currently being developed. Rio Tinto, through its Northparkes operation, have already been trialling the Autotram system and given positive developments of this technology into the future it is intended that the Argyle underground mine will utilize a fully automated loader fleet. Secondary breaking remains an issue given the hazardous conditions often encountered. It is unlikely that much of this work will ever be fully automated but wherever possible teleremote operation will be used to remove personnel from the underground environment. A central control room will be located on surface from which all major activities will be controlled and monitored. Given hot and humid surface conditions, high ambient rock temperatures and the introduction of heat through ground and surface water, ventilation is of prime importance. As such, in order to minimize the ventilation requirement for equipment and in line with positive experience at Northparkes, a fleet of electric loaders will be used for ore production. Based on these requirements, it was determined that the Extraction level would comprise the following: • a transverse layout to give production flexibility given a relatively elongated orebody. • an offset herringbone drawpoint layout to facilitate the use of electric loaders. • extraction galleries with an orepass at the hangingwall end and a ventilation/drainage raise at the other thus 612
Transfer Level The lower level of the block cave has the multipurpose use of ore transfer to the crusher and transfer of exhaust air and drainage away from the mining area. It is proposed to use large capacity loaders for ore transfer. Each of the Extraction Level orepasses report to the western side of the Transfer level and are accessed much as a drawpoint. Twin transfer drives are proposed facilitating the use of multiple loaders operating on concrete roadways. The transfer drives report back to a centrally located gyratory crusher. The eastern footwall side of the level is quite separate from the ore transfer facility. Each of the ventilation/drainage passes from the extraction level terminates on the footwall perimeter drive of the Transfer level. This drive collects drainage water and return air, passing them south to the main ventilation exhaust and primary pumping infrastructure. All loading, crushing, conveying and pumping activities will be monitored and controlled from the central control room. Figure 5 shows the layout of the Transfer level.
Figure 5 – Transfer Level Layout
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7. ORE-HANDLING Requirement/Duty Mining studies have identified a sustainable underground production rate of 7.5 million tonnes per year during Stage 1 – Block Cave operation. Given the need to design for some waste handling capacity and providing some room for mining productivity improvements, a design capacity of 8.0 million tonnes per year (wet) has been selected for ore handling components. Crusher Numerous crushing options were investigated as part of the Pre-Feasibility Study. These ranged not only from underground versus surface primary crushing but also jaw versus gyratory versus hybrid crushers. Through this assessment, it was established that an underground primary Gyratory Crusher with surface secondary crushing was most appropriate for the Argyle Underground operation. A gyratory crusher in the order of 54-74 size, fed with ROM ore supplied through 900 millimetre to1200 millimetre orepass grizzlies and orepasses will be used. The crusher has been selected to suit the feed size and provide the required throughputs. The crusher itself will be located centrally relative to the block cave footprint. It is currently proposed to be within the orebody, given the unfavourable ground conditions anticipated in the hangingwall sediments. Locating this infrastructure within the orebody does however complicate the transition in production from the Stage 1 block cave to the Stage 2 sublevel cave. Conveyors A variety of ore haulage systems were evaluated, including conveyor, shaft and truck haulage. Ultimately however the relatively shallow nature of the underground
mine, together with the staged nature of orebody development and the requirement for significant lateral haulage away from the orebody, has pointed towards a conveyor solution for ore handling to surface (Figure 6). Recent advances in underground conveyor design, installation and operation has seen the planning and commissioning of long, steep underground conveyors at Palabora, Ridgeway and Northparkes. The ‘lessons learned’ from each of these sites will be incorporated and referenced in the Argyle design and operation. 8. DEWATERING Expected Inflows The monsoonal nature of the climate in the Argyle region results in the potential for severe tropical rainfall events. Combining with this is the location of the open pit directly above the proposed underground mine, which comprises a substantial catchment area ultimately funnelling surface water into the underground workings. Dewatering Strategy Given the wide range of potential inflows and uncertainty of the expected inflow quantities, a three part dewatering strategy has been developed. After cave breakthrough into the pit, three pumping conditions have been identified: - normal conditions, representing typical day to day pumping requirements; - seasonal conditions, where summer rainfall may increase pumping requirements substantially for a period of time; and - flood conditions, where unexpected intense inflows occur. Normal water inflow to the mine, including average rainfall, ground water and collection of raw water used for mining processes amounts to flows of between 40 to 120
Figure 6 – Schematic of Underground Orehandling System Massmin 2004
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litres per second. This will be handled by two Geho TZPM 800 positive displacement pumps, each of 60 litres per second capacity, with over 750 metre head lift capability. For seasonal/abnormal conditions, when inflows into the pump stations exceed the capacity of the two Geho positive displacement pumps, a series of 3 to 4 Warman 16/14 TYFC APH slurry centrifugal pumps will be used to dispose of the extra water, up to a total rate of 920 litres per second. Flood conditions prevail when inflows exceed the designed pumping system capacity. In this circumstance the mine will go into a shut-down mode. A series of water-tight doors installed in bulkheads will be used to protect key infrastructure such as pumping stations, crusher stations, workshops and conveyors. These doors will separate this infrastructure from level drives and declines, which would be allowed to flood and subsequently pumped out. 9. VENTILATION The ventilation design for the Argyle Underground Project has been undertaken with the following considerations in mind: • Air quality must comply with Occupational Exposure Limits of Australia’s NOHSC 1003. • As specified by the WA Mines Safety and Inspection Regulations 1995: - Re-circulation of air is minimized and air is drawn from the purest source available (single pass ventilation) - Where diesel units are used for production, a general criteria of 0.04 to 0.06 cubic metres per second per kilowatt of rated engine output will be applied to each ventilation district. - A minimum air velocity of 0.25 metres per second is maintained where electric vehicles are used. - If wet bulb temperature exceeds 25oC, a minimum air velocity of 0.5m/s. Block Cave The greatest ventilation requirement for the block cave is when both production and construction activities are occurring at the same time. The block cave production rate at its peak is expected to be 7.5 million tonnes per year. Even though electric loaders are expected to be used, diesel loaders have been assumed to ensure sufficient capacity for flexibility. Table 2 summarises the ventilation requirements. Table 2: Block Cave Ventilation Requirements 9715 BLOCK CAVE Production Drives Transfer Drives Crusher Conveyors Lower Infrastructure (Pump stations, etc.) Level Workshops Leakage Level Development/ Undercutting
150 m3/s 60 m3/s 35 m3/s 40 m3/s
TOTAL EARLY PRODUCTION
395 m3/s
TOTAL PRODUCTION
375 m3/s
15 m3/s 30 m3/s 45 m3/s 100 m3/s
Ventilation Circuit For all stages of underground operation, the primary ventilation system comprises fresh air being drawn down the Main Decline and the Fresh Air Decline, through the 614
various mine workings to the Main Return Airway. Schematic of the ventilation circuit is illustrated in Figure 7. The ventilation system has been designed as a push-pull circuit. In doing so the extraction level is maintained under positive pressure such that contaminants and air will not be drawn through the cave into the mine workings. The disadvantages of the system are the additional heat generated by the location of fans underground and additional capital and operating costs. It is considered however that the benefits of this system outweigh the disadvantages. It is proposed that exhaust fans be installed at the collar of the exhaust shaft. It should be noted that due to the length of the intake and exhaust declines, the fan would operate at high ventilating pressures. Establishing the primary ventilation system is on the critical path of the project and as such it is essential that the Main Return Airway is developed as soon as possible. 10. REFRIGERATION Design of the refrigeration system was undertaken with the objective of providing acceptable working conditions at major workplaces throughout the mine during fluctuations in climatic conditions. Expectation is that a majority of the underground workforce will operate within the confines of air-conditioned mobile equipment cabs. There will of course be occasions when work outside of such equipment will be necessary. Heat Load Heat is generated from a variety of sources including: • Climatic conditions • Heat from development walls and broken rock • Heat from machinery • Auto-compression of air • Heat from ground water Ground water is a major contributor to the mine’s heat load and as such it is important that water is removed from the underground environment as quickly as possible. Initial indications are that the mine will require between 9MW to 10MW of cooling capacity for operation of the block cave. Refrigeration & Cooling Plant Given the relatively shallow nature of the Argyle underground operation, a surface refrigeration plant will be used as it gives the advantages of ease of access, unlimited capacity for expansion and the ability to use refrigerants such as ammonia. Through the use of bulk air cooler serviced by a surface refrigeration plant, cooling of the mine can be accomplished in the most economical manner possible using a relatively simple system. The bulk air cooler should be positioned at the top of the intake ventilation shaft, with the refrigeration plant a further 200m away. The use of a number of refrigeration modules allows: • A build-up of refrigeration capacity consistent with the underground requirements at the time. • Flexibility in adjusting the amount of refrigeration given seasonal and diurnal variations in ambient temperature. • Capacity to maintain cooling during the maintenance of refrigeration units. • Easily expandable system. • Minimization of early capital expenditure.
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Figure 7- Block Cave Ventilation Circuit
11. CONCLUSION The major technical issues associated with the project are: • ventilation, given hot and humid surface and underground conditions, • dewatering, given a monsoonal environment and mining under an open pit; and • ground conditions associated with numerous structural features and relatively weak host rock units. These technical issues, whilst still posing a risk, are being addressed within the mine and infrastructure design
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processes of the Feasibility Study. In all aspects of the study, ‘The Vision’ for the Argyle underground project has been referenced and used as a point of focus. 12. ACKNOWLEDGEMENTS The author wishes to thank Argyle Diamonds for the opportunity to present and publish this paper and to acknowledge all of those contributing to the successful completion of the Argyle Underground Feasibility Study. Special thanks goes to Stephen Brennan, Brian Morris and David Nicholls for their help in preparing this paper.
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The alternate design considered for the Argyle underground mine Geoffrey Bull, Gary MacSporran, Campbell Baird, SRK Consulting, Perth, Australia
Abstract The Argyle Diamond Mine open pit reaches its economic limit in 2007. As mineralisation extends to beyond 600m below the final pit base, studies have considered potential underground mining options, with the two most viable combinations being, either: ß A large block cave for the upper part of the resource and a sub level cave for the lower parts, or ß A smaller block cave to extract the main portion of the ore body, a core-and–shell method to extract the footwall wedge (ore between the block cave and the dipping footwall contact) and a sub level cave beneath the block cave. The final feasibility study is progressing the former option, which is described by Hersant (2004), whereas this paper discusses the latter option, focusing on the challenging geotechnical and practical mining issues leading to the development of this combination of methods in the search for a safe and reliable underground mine design. 1 INTRODUCTION Argyle Diamonds Limited, a wholly owned subsidiary of Rio Tinto Limited, operates the Argyle Mine, an open pit operation located in the eastern Kimberley region of Western Australia. Pit production from the lamproite AK1 orebody feeds a processing and recovery plant at a rate of 10Mtpa. By 2007 the pit will have reached its economic limits in the AK1 ore body. The remaining substantial resource, extending some 600m below the final pit base, will only be amenable to extraction by underground methods. Since 1995 a number of underground mining studies have been undertaken, investigating the potential of a range of bulk mining methods. In the process, non-caving methods were gradually eliminated due to excessive costs, insufficiently high production rates and safety issues. The dipping and plunging nature of the AK1 ore body, it’s narrowing and bifurcation deeper down and the relatively weak inter-bedded sediments in the hangingwall and footwall introduced some technical challenges to the mine design and mining method selection process. Finally, the methods considered in the Pre Feasibility study, undertaken from July 2001 to January 2003, were narrowed down to two options: • A large block cave, with an extraction horizon covering the lamproite ore body and extending out into the footwall sediments, extracting the upper part of the resource, followed by a sub level cave (SLC) for the lower parts. • A smaller block cave, with an extraction horizon confined to within the stronger lamproite body and located at a higher elevation than that for large block cave, to extract the main portion of the ore body, plus a core-and–shell method to extract the "footwall wedge" (ore between the block cave and the dipping footwall contact) and, finally, a sub level cave beneath the block cave. Assessment of technical risk indicated that the former option (large block cave and SLC) provided a lower risk profile. The final feasibility study, currently under way, is focusing on this preferred option. To assist in final design and decision making a decline is also currently being developed to provide further information, including: • Additional information regarding the strength and competence of the footwall sediments • Measurement of in-situ stresses • Characteristics of the major structural features that will be intersected by planned production and infrastructure development 616
• Confirmation of achievable development rates in the different rock types for more accurate scheduling • Assessment of support requirements in the different rock types • More accurate development cost information The fact of an inclined ore body and the questions surrounding the competence of the footwall sediments and major structural features led to the inclusion of the second mining option in the pre feasibility study, in an attempt to optimise the mining of the AK1 ore body. This paper discusses this "alternate" option. 2 GEOLOGY AND GEOMECHANICS The AK1 deposit is a volcanic intrusion of Lamproitic tuff and magmatic Lamproite intruded into a Proterozoic stratigraphic sequence of inter-bedded quartzites, siltstones and mudstones that overlie dolerite and basalt units. The AK1 deposit consists of three north-south oriented pipe structures (Northern Bowl, Central AK1 and Southern Tail) that plunge steeply to the south (Figure 1). The central pipe of the AK1 ore body dips at approximately 70o to the west, plunges at 70o to the south and narrows and bifurcates at depth. The narrowing and bifurcation of the AK1 with depth is shown in Figure 2.
Figure 1 – Long Section through the AK1 deposit and Argyle final pit
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Figure 2 – Plan of AK1 ore body at different elevations showing narrowing and bifurcation.
The AK1 and surrounding country rock was domained according to a structural interpretation and the different rockmass characteristics and is shown in Figure 3.
stratigraphy here and all are probably components of the Gap Fault Zone. Investigation of pre-mining aerial photography for this region reveals that NW-SE trending faults extending from the North Bowl Region swing into more northerly orientations as they pass through the "Neck Zone". The major structures in the area can be summarised as follows: • The dominant active systems responsible for the tectonic deformation observed at Argyle are the Halls Creek Fault System (HCFS) and the Bow River Fault System (BRFS). • In the regional tectonic setting, the Gap Fault formation is currently interpreted as a second order feature and results from the activity of the HCFS and the BRFS. In terms of the AK1 deposit it is considered the major structural feature. • A few hundred metres to the north of the AK1 ore body is the Eastern fault that runs sub-parallel to the Gap fault. Cutting across the ore body is the Razor fault. The Gap and Razor faults are shown in Figure 2. Within the AK1 lamproite the Razor fault is generally a discrete, narrow planar feature, often difficult to see in the core, and not a wide feature like the Gap Fault. Figure 4 shows the main structural features in and around the AK1 ore body. The black lines are major faults, red lines are domain boundaries and grey lines are pit contours.
Figure 3 – Geotechnical domains
Domain 2, the AK1 lamproite consists of 60-80% quartz and is a relatively competent lithology. Irregular zones of more competent, fine grained, grey, almost aphanitic rock occur throughout the lamproite. Domain 3, the Western Contact Zone consists of an irregular and complex zone of breccia developed along the western edge of the AK1 pipe. It involves both the host sediments (quartzites and mudstone/shales) and the lamproite, and occupies a zone varying in width from 50m. Domain 4, in contrast to the western margin of the pipe, intrusion breccias appear much less well developed on the eastern margin. Breccias zones up to 5m thick are present along this contact but major changes in bedding plane dip and large blocks of country rock within the lamproite have not been observed. The lack of a complex contact zone suggests that either the intrusion processes were different on the eastern margin of the pipe or else the Gap Fault Zone in this region has faulted out much of the breccia. Projecting of faults mapped in the northern part of the mine indicates that the NW to NNW striking, moderate W to SW dipping Gap Fault Zone extends along the eastern margin of the AK1 pipe into this domain. The Gap Fault diverges into a number of strands at least one of which cuts into the lamproite body in Domain 2. On the SE wall of the pit, displacements in sedimentary units suggests that at least three N-S trending faults cut the Massmin 2004
Figure 4 – Domains 2, 3, 4, and 8 and structures The different rock types in each domain were classified into IRMR values as shown in Table 1. The strongest of the rock types are the lamproites and the weakest are the mudstones and interbedded mudstone/siltstone units. Attempts were made to measure the stress conditions using acoustic emission and hydraulic fracturing techniques, however, the results were conflicting and inconclusive. Analyses and modeling dependent on stress inputs used an assumed range of values. The resource model indicated that the upper, eastern side of the AK1 contains high diamond values, making it particularly important to recover this part of the ore body. 3 EARLY STUDIES In earlier mining studies conducted under the direction of the previous owners, Mineco Pty Limited, mining method selection for the AK1 ore body considered a range of bulk mining options, including: open stoping with and without fill, sub level caving and block caving (single and dual lifts). The open stoping options led to poor recoveries (low percentage extraction) as well as having the risks associated with entry
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TABLE 1 – Rock Properties Domain
Rock Type
IRMR (Avge)
IRMR (Min)
IRMR (Min)
2
Lmp (all) Lmp(p2000)
53 57.7
25.3 25.8
83.2 83.2
3
All (all) All (p2000) Mudstn (all) Qtzite (all)
46.4 52.5 36.1 47
25.2 25.2 27.3 25.2
77.2 77.2 57.6 77.2
4
All (all) All (p2000)
47.4 50.6
25.8 25.2
74.8 74.8
8
All (all) All (p2000) Mudstn (all) Qtzite (all) Dolerite (all)
42.4 49.5 31.7 46 46
11 11 24.8 23.1 11
70.4 70.4 61.1 69.4 70.4
(all) – denotes all geotechnical data, (p2000) – denotes only data collected after 2000 (less data was available but it is considered more reliable than pre 2000 data). methods, the stope and fill methods were not economically viable and it was concluded that only the caving methods, more particularly block caving, could be seriously considered for mining of the AK1 ore body. The early studies included a block cave that was situated within the lamproites, where it was considered that the cave would propagate up the perceived weak footwall contact enabling the "footwall wedge" of ore to be drawn on an incline from the east-most row of draw points.
Figure 5 – Cross-section and Long-Section showing the different mining areas
4 THE "ALTERNATE" MINING OPTION The components of the "alternate" mining option are shown in Figures 5 and 6. The essential difference between this option and the large block cave option (Hersant, 2004) is the selective mining of the "footwall wedge". The issues that prompted the development of this "alternate" option included: • Concerns that the footwall sediments, particularly the weak mudstones, may cause draw point stability issues, even if substantial support is included. This footwall area may be further weakened by the presence of the Gap fault and its splays. • In a large block cave option, with draw points extending out into the footwall sediments, a large wedge of waste occurs between the draw points and the inclined footwall of the lamproites. This waste has to be drawn off in order to draw down the valuable lamproites above it. Handling this waste separately from the ore generated from other draw points in a single material handling system will be very difficult to schedule. • Concerns that the weaker footwall sediments may cave and draw more readily than the lamproites, drawing in excessive dilution laterally from the footwall. • Disagreement with the postulation from early studies that an "inclined draw" would occur as the cave followed up the weaker contact zone, enabling the lamproites in the footwall wedge to be drawn from eastern drawpoints within the lamproites. This "inclined draw"assumption is considered very risky as it is quite likely that, while the cave may propagate readily into the weaker footwall 618
Figure 6 – Plan of Block Cave and Footwall Wedge mining areas
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rocks, large amounts of the weaker footwall waste material would be drawn preferentially in the attempt to draw the ore, diluting this ore excessively. As the grade of the footwall wedge of ore is relatively high it is important that it is extracted at minimal risk. Therefore, an alternate method of extracting this footwall wedge of ore was proposed. • The northern part of the AK1 flattens out towards the "neck" between the AK1 and Northern Bowl (Figure 1) making it uneconomic to include this in the block cave footprint. This area too had to be mined selectively by an alternate method. The advantages and disadvantages of selectively mining the "footwall wedge" (FWW) include: Advantages: • The FWW can be accessed and developed early so that production from this source can provide positive cashflow while the Block Cave undercut, extraction and ore transfer levels are being accesses and developed. • Early extraction of the FWW ensures minimal risk of dilution or loss of this valuable part of the AK1 resource. • Removal of the FWW will create a vertical free face along the caves eastern boundary, which, together with the undercut will assist caving of the main part of the ore body. • The FWW north area can continue to produce when the block cave starts to produce. Production from the FWW north can be used to supplement block cave production during the block cave ramp up to full production. • The FWW development on the lowest level can be used as drainage tunnels for dewatering and drawing down of the water table prior to commencement of production. • FWW development can be used for locating Block Cave propagation monitoring and seismic system equipment. • FWW production costs will, however, cost more per tonne than if this ore was extracted by Block Caving, but the mill head grades will be better due to less dilution. • Development of the FWW can commence while mining the open pit, however, the risks associated with underground interaction with the Open Pit must be identified and addressed. • The FWW production trucks will not add to the traffic congestion (Block Cave development and construction vehicles and equipment) as they will travel in the intake airway decline (exploratory decline). They only travel a short distance in the top section of the main decline. • The use of the intake airway decline for trucking will allow for production from the FWW of up to 2.0 Mtpa during peak block cave construction. Disadvantages: • There will be more up front capital expenditure incurred by the additional development necessary to effect the FWW production. • With the ore removed from FWW and having drawn down waste to fill this void, there will be the risk of this waste being drawn into the first row of block cave draw-points, undercutting the ore above. This risk can be minimized by good draw planning and control. • Potential stope stability issues given the variable geotechnical conditions in the ore body and undercutting of pit walls. • Scheduling risks associated with the production transition from FWW to Block Cave. • Reduced economic value due to lower production rates associated with the FWW and smaller BC. • Caveability risks associated with the small Block Cave footprint. Latest resource modelling (completed after pre Massmin 2004
feasibility study) has reshaped and reduced the ore body footprint at the extraction level of the small block cave. 5 MINING METHODS Block Cave Economic modelling indicated that the optimum elevation for the extraction horizon is around the 9775mRL elevation. A transverse Henderson offset herringbone extraction horizon is planned to make it possible to implement high levels of automation. The Hydraulic Radius (HR) of the economic footprint at this elevation is ~42m. This is just in excess of the required HR=40m for caving of the lamproites. The width of the extraction horizon is between 110m – 120m which is marginally greater than the minimum critical span (~107m) for caving for the lamproites (Figure 12). The design of the extraction level, undercut level and the transfer/drainage level are similar to the design described by Hersant (2004) for the large block cave option. Core and Shell Studies have examined various methods of mining the FWW and have concluded that the most appropriate is a Core and Shell method, which in recent studies has been referred to as the "Trough and Ring Retreat" method. The Trough and Ring Retreat mining method consists of modules of individual components which allow flexibility in its final layout and firing sequence (Figure 7, 8 and 9). The modules consist of Slot and Swell (Primary Stope), Ring Retreat (Secondary Stope) and Trough Undercut, which is blasted prior to the former modules. The dimensions of the components can be modified to suit ground conditions and orebody geometry. In areas where two modules are situated side by side, a central pillar is designed in between them to isolate the Slot and Swell components of the modules. These pillars are recovered in the final blasting of the two modules. Furthermore, to ensure good
Figure 7 – Modular stope layout for FWW
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fragmentation and to assist in initiating pit wall failure, cleaner rings can be employed on the eastern and western sidewalls of the modules and fired along with the Central Pillars. The regional sequencing of the modules is flexible with respect to blasting sequence. These modules can be produced either in a Continuous Retreat fashion (recommended for Footwall Wedge, Figure 10), or Mass Firing the Ring Retreat component of the modules (recommended for Footwall Wedge "North", Figure 11). Either of these sequences can be implemented in each of the Footwall Wedge zones.
• Establishment of Trough Undercuts provides early production and Mass Firing turns it into a "rock factory" yielding high tonnages. Schedules indicate that production rates of 1.8Mtpa will be achievable form the FWW. • The FWW development and production schedule does not hinder the infrastructure and production schedules of the main block cave operation. • The FWW "North" can produce concurrently with the block cave operation. Production from the FWW "North" area that is in contact with the Block Cave area must be completed prior to commencing Block Cave production. The mining sequence of the FWW "North" is south to the north in order to limit the mining interaction between the Block Cave and the Footwall Wedge "North". This south to north extraction is possible due to the establishment of an alternative access decline in the north connecting the Block Cave extraction, undercut and transfer levels with the FWW 9860mRL. The main risks of the method are: • Exposure to open voids for the period until mass firing takes place • Uncertainty concerning pit wall behaviour. Sudden large failures may affect stability of underground workings. • The ventilation requirements (volume and cooling capacity) for concurrent development of the FWW and block cave infrastructure may exceed design capacities.
Figure 8 – Modular stope layout FWW "North"
The main advantages of this method are: • Only two development levels required. • Low operating costs. • Flexibility in final design and blasting sequence through modular approach.
Sub Level Cave Below the Block Cave extraction horizon the ore body narrows down fairly rapidly, particularly below 9700mRL (Figure 2), and then splits up into two limbs. It would not be possible to consider a second, lower lift to the block cave as the ore body narrows to below the minimum critical span for caving of the lamproites. The minimum critical span can be determined from the HR for caving using the chart shown in Figure 12. For the lamproites, where the HRcaving = 40m locus intersects the 3:1 aspect ratio line the dimensions of the rectangular footprint having a width equal to 1/3 of the its length (107m x 320m) can be read off the chart. For the lamproites, this width of 107m is considered to be the minimum critical span, below which dimension the cave is likely to form a stable arch and stall. In some cases this ratio can be less than 1:3 (width to length) and the cave will continue to propagate, however it
Figure 9 – Typical Long Section FWW layout
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Figure 10 – Continuous Retreat sequence for FWW
Figure 11 – Mass Fire sequence for FWW "North"
has generally been found that this 1:3 ratio rule is a good guide for determining the minimum critical span.
Figure 12 – Chart for determining Minimum Critical Span Massmin 2004
SLC is a top down caving method and, if started below a large mass of caved ground, is not dependent on the width of the ore body or cavability of the rockmass, as the caved ground will rill down into narrow zones extracted by the SLC provided the dip and plunge of the ore body does not flatten out to angles less than the angle of repose of the rock within the cave. SLC is therefore the most suitable and economic method to use to mine the lamproites below the block cave horizon. There is one matter of concern that will need careful consideration in the ongoing design of the Argyle underground mine, which is, the possibility of a drop in production for a period during the transition from block cave to SLC. However, careful sequencing of the block cave to mine from north-west to south-east will allow the northern parts of the block cave to be completed in time for SLC to commence production in the north-west on the 9750mRL while the block cave is still being depleted in the south. Below the block cave there are only about four or five levels that are sufficiently large to maintain production at the
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high rates obtained in the block cave. Further down, where the ore body narrows and where it eventually splits into two and the tonnes per vertical metre decrease significantly, it will be impossible to maintain the same rate of production as obtained in the block cave. 6 SEVICES AND MATERIAL HANDLING FACILITIES For this "alternate" mining option, the main access decline, ventilation decline, return airways and conveyer declines and all crushing and dewatering facilities will be common with those being designed for the large block cave option that is being progressed to final feasibility study stage. 7 CONCLUSIONS The "alternate" mining method option for Argyle underground mine design combines three bulk mining methods (Block Cave, Sub Level Cave and Core and Shell) to allow for a high recovery of the ore resource and a rapid build up in production.
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However, along with the many potential advantages this option presents, there is also some uncertainty surrounding the caveability of the small block cave footprint, the risk of unstable stopes in the FWW, the potential scheduling risks associated with the production transition from FWW to Block Cave and the less economically attractive result this option produces. For these reason this alternate option will no longer be considered for the Argyle Mine Underground design. ACKNOWLEDGEMENTS The authors are grateful to Argyle Diamond Mines Limited, to Stephen Brennan and Damien Hersant in particular, for the opportunity to have been involved in this interesting project and for their permission to publish this technical paper. REFERENCES • Hersant, D, 2004. Mine Design of the Argyle Underground Project, Massmin 2004, Santiago Chile.
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Design of the Grasberg block cave mine Charles Brannon, Manager, Underground Planning, Timothy Casten, Senior Manager of Underground Planning, Mark Johnson, Senior Vice President and Chief Operating Officer, Freeport-McMoRan Copper & Gold Inc.
Abstract The Grasberg Block Cave (GRSBC) Mine will be the main source of mill feed after the Grasberg Open Pit has been depleted in 2014. When the pit is finished in 2015, the district-wide underground ore reserves will be on the order of 1.7 billion tonnes, of which the GRSBC mine comprises 874 million tonnes at a grade of 1.0% Cu and 0.8 g/t Au. Development of the access adits was initiated in 2004 in order to access and develop the GRSBC mine in time for the open pit completion. This paper summarizes a study undertaken on the GRSBC to confirm the viability of the mine project and to gain approval to commence with the long lead-time access development required. Block cave mining was determined to be the most applicable method for mining the deposit, with production rates of 115,000 tpd being considered. The method by which the large footprint (1km by 1km) will be developed and caved is discussed in the paper. The geotechnical issues, ventilation infrastructure and ore handling systems required to support the proposed tonnage rates are also described.
1 INTRODUCTION The world-class Grasberg copper-gold deposit is located in the Ertsberg Mining District, in the province of Papua, Indonesia (4o-6'S, 137o-7'E; see Figure 1). The district is in rugged mountainous terrain from 2,900 to 4,200 meters in elevation, with annual rainfall exceeding 300cm. The district has reserves exceeding 2.6 billion tonnes in eight surface and underground deposits that contain recoverable metal of about 54 billion pounds of copper and 66 million ounces of gold.
in 1970. The Grasberg deposit was discovered in 1988, and the district has seen almost continuous reserve increases and production expansion through to today’s production rate of about 240,000 tonnes per day. Current operations in the district include the Grasberg open pit (200 ktpd ore) and the DOZ block cave mine (40 ktpd). The Grasberg open pit is the flagship operation in the district, and when the pit is concluded in 2014, the Grasberg block cave mine (GRSBC) will be the primary source of mill feed. Figure 2 is a snapshot of the projected final Grasberg pit and surrounding topography, showing the planned underground workings and estimated cave shape of the block cave mine.
Figure 1. Location of PTFI’s mining operations. The Ertsberg District was discovered in 1936 during a mountain climbing expedition led by Dutch petroleum geologist Jean Jacques Dozy (Mealey, 1996). His detailed report of the surface geology and descriptions of outcropping high grade copper skarn mineralization led to investigations by Freeport Minerals Corporation in the early 1960s. Development of the initial Ertsberg open pit began Massmin 2004
Figure 2. Snapshot of Grasberg final pit and underlying block cave mine. This paper is a brief description of the results of the prefeasibility study completed on the Grasberg block cave mine. The study concluded that the project was viable; the
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feasibility study for the GRSBC was initiated in 2004. A project of this magnitude presents many challenges that need to be fully explored before development begins in 2008, and many of those challenges are discussed in this paper. 2 GEOLOGY, HYDROLOGY, & RESERVES 2.1 Geology Grasberg is a porphyry copper-gold deposit with many similarities to other porphyry deposits of the region (MacDonald and Arnold, 1993). The geology is typical of subduction-related arc systems formed at plate tectonic collisional boundaries. Calc-alkalic intrusive bodies are preferentially emplaced at intersections of major cross structures into tightly folded Tertiary-Cretaceous sediments. The Grasberg Igneous Complex (GIC) is a multi-stage dioritic intrusion about three million years old. The intrusions are emplaced into the center of a volcanic breccia complex about one kilometer in diameter that flares near the surface. A central intrusive phase forms a +/- 100 meter-wide septum of poorly mineralized rock that cuts through the otherwise roughly cylindrical GIC, resulting in a horseshoeshaped ore footprint in plan view. The mineralization is dominated by chalcopyrite although bornite becomes more significant with depth. The mineralization extends more than 1,600 meters vertically, from the 4,300 meter elevation at the original surface to at least the depth of 2,700 meters. Width of the mineralization ranges from about 200 meters to over one kilometer. Copper and gold grades are highest near the center of the deposit, adjacent to the low-grade core. On the margins of the GIC is an irregular zone of massive pyrite with minor magnetite and chalcopyrite, termed the Heavy Sulfide Zone (HSZ). This zone is up to one hundred meters thick and is related to a mineralizing event that postdates the main GIC. Portions of the HSZ are included in the block cave reserve. 2.2 Ore Reserves The diluted mineable reserve established for the prefeasibility study is 874M tonnes @ 1.0% copper, 0.8 g/t gold, and 2.9 g/t silver (as of January, 2004). The reserve was established using the PC-BC© block cave planning software. PC-BC simulates the block caving process and produces a predicted mined tonnes and grade, including estimation of "toppling" of pit wall material into the cave (Diering, 2000). The cut-off grade was roughly 0.90% copper equivalent (value of gold and silver calculated as copper percent). 2.3 Hydrology The passive inflow to the cave is predicted to peak at about 17,000 gpm, at the time of ultimate cave and cracklines, and in particular when the cave has broken through the HSZ into surrounding karstic limestone. Direct precipitation into the cave comprises approximately 7,000 gpm of the total; the remainder is primarily from major structures or certain limestone units susceptible to karst formation.
RMR has been utilized to establish mine design parameters (hydraulic radius and ground support) for the GRSBC . The rock mass rating, at the 70th percentile of the populations, is "good" for the Development Levels, the lower fifty meters of the mineable zone, and for the central Kali intrusive. In general, primary fragmentation of the GRSBC will be coarse, with more than 76 percent of the tonnes being in blocks greater than two cubic meters. The secondary fragmentation curves indicate that at greater than 250-meter column heights, 23 percent of the tonnes will have a volume greater than two cubic meters. The estimated hydraulic radius required for sustained caving is 33 meters. Rock burst potential does exist and will be evaluated in more detail during the feasibility study. Assumed draw rates range from 0.13m/day during the initial 100 meters of column height, up to 0.30m/day at greater than 200m of column height. The overall technical risk due to wet muck is thought to be relatively low, particularly for the early years of mining. The predicted coarse in-situ grain size for Grasberg should inhibit significant wet muck, despite the relatively high volumes of predicted passive groundwater inflow. However, communition of the HSZ, fines from adjacent open pit overburden stockpiles reporting to the cave, or secondary fragmentation of limestone derived from pit walls may eventually report to the drawpoints and present a wet muck issue. Immediately adjacent to the GRSBC is another 500million tonne block cave reserve, Kucing Liar (KL). The KL deposit has an extraction level some 220 meters below GRSBC. One of the major issues for Grasberg in the next phase of study is how to best manage the interaction between these two massive caves.
4 ACCESS TUNNEL DESIGN The extraction level is designed at the 2,820 meter elevation. The primary access to the mine will be via twin adits, each eight kilometers in length, that will be developed from surface at the 2,600 meter level (Figure 3). The tunnels are named the Ali Boediardjo (A.B.) Tunnel system. Due to the long lead-time for development of the A.B. adits, development was initiated in early 2004 so that the adits could reach the base of the mine by 2008 when block cave mine development is scheduled to begin. The adits will be fitted with an electrically powered rail haulage system that will deliver manpower and materials, and haul waste for the planned mining operation. The adits will also provide reliable gravity drainage of groundwater, ventilation, and service and maintenance lines. 5 MINE DESIGN The mine is scheduled to begin development in 2008 and will be accessed via the A.B. adits. The mine will be a mechanized block caving operation with a planned peak production rate of 115,000 tonnes per day. Undercutting is initiated in 2014, peak production is forecast by 2019, and closure estimated in about 2037.
3 GEOTECHNICAL STUDIES The average uniaxial compressive strength of the ore is about 110 MPa, with variation in values depending on mineralogy and alteration between 70 and 140 MPa (Srikant and Nicholas, 2003). The principal in-situ stress is in the northeast direction and corresponds to the regional structural geology (about 50 MPa). The principal horizontal stress is twice the overburden stress. 624
5.1 Level Description The extraction level (2820m) is based on an "El Teniente" style of layout, used previously in the Ertsberg District at the now-exhausted IOZ mine. Spacing of the panels is 35 meters with drawpoint spacing of 20 meters (Figure 4). The current mine layout is a very large footprint with a diameter of about one kilometer (694,000 m2) and contains nearly two thousand drawpoints.
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In order to develop and mine this size of footprint, the panels and drawpoints are broken down into five manageable sections about 260 meters wide. Each panel section contains an orepass and a vent raise, located in adjacent pillars. Extending the drawpoint through the drawbell to the next panel allows the creation of an open access and can be used as a short-term fringe drift. The sections allow for an LHD to access both orepasses located North and South of the section. This improves tramming efficiencies, with the longest tram being 140m. It also adds in the flexibility of a back-up orepass in case one is down for repair or maintenance. A flat, room and pillar undercut system is planned (2840m); headings will be directly above and parallel to the extraction level panels. Crosscuts are parallel to the drawpoint drifts but offset by 10 meters so that they lie on top of the drawbell minor apex (Figure 4). The service level (2790m) provides intake and exhaust to the undercut and extraction levels and is the primary ore handling level. The rail haulage system has four haulage lines that run under the orebody in a northwest orientation that match the section divisions on the extraction level. The lines gather into a northeasttrending fringe drift that runs over the coarse ore bin dumps. Two intake and two exhaust drifts parallel the haulage lines (Figure 5). 5.2 Caving The mine will utilize an advanced undercutting system. Expected draw column heights average 460 meters, and every method must be employed to protect the integrity of the extraction level drawpoints. A room and pillar advanced undercut is proposed for the mine. Lateral crosscuts are driven over the top of every minor apex. The narrow undercut removes only a narrow opening (4.0 m) between undercut level and cave back (Figure 6).
The caving concept basically sections the undercut horizon into a series of identical pillars that are blasted or "wrecked" using the surrounding development drift as expansion void, rather than the open cave. Pillar size has been set so that two pillars are equivalent in area to a single drawbell. The room and pillar undercut removes the swell mucking requirements, as the undercut drifts themselves are used as the expansion void for the blast. The major and minor apex are the same height with this plan, which helps ensure connection between drawbells. This geometry aids sequencing and scheduling because the undercut is blasted in drawbell-shaped sections rather than in incremental rings. Furthermore, the blasting process is simplified. A series of 32 near-horizontal holes are drilled into the pillar from the crossover, and then charged and blasted laterally to the open undercut drift. An expansion void in the cave is not required to be mucked open and an initiation raise is not needed. Some cleanup of blasted muck is required but an ongoing campaign of swell mucking is not required. The drawbell will be mined out once the undercut has passed by a minimum of a single drawbell ahead and the bell is clear of abutment stress. The excavation method consists of a portable Alimak nest and climber that is used to drive a 2.0m by 2.0m vertical raise 13.5m long. Once the raise is in place and secure the drawbell will be drilled and a mobile emulsion charging unit will be employed to load the bell and fire the round. Variable density emulsion and electronic detonators will be used to fire the bell in a single shot. Undercutting is initiated in 2014; active drawpoint production is initiated in 2015 after the open pit has ceased operations. Maximum drawbell opening rates are eight drawbells per month; maximum draw rates are 0.20m per day. The plan assumes that the cave will be developed in two separate areas simultaneously. The sequence has
Figure 3. Schematic perspective view of Grasberg block cave and AB access adits. Massmin 2004
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Figure 4. Undercut and extraction level layout, showing vertical position of undercut relative to extraction drifts.
Figure 5. Schematic layout of GRSBC extraction level and service level.
Figure 6. "Room and Pillar" advanced undercut concept. 626
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Figure 7. Generalized undercutting concept, split into twoyear increments as indicated.
Figure 8. Cross Section; extraction and service levels & the orepass/vent raise configuration.
been designed to maximize grade and minimize dilution effects from the toppling of the pit material. Figure 7 shows the generalized undercutting sequence, by two-year periods, utilized for this study.
The loaded trains will dump the ore into one of three coarse ore bins near the main fixed facility area. Separate feeders will draw from the three ore bins and deliver ore to one of three 60" x 89" gyratory crushers. Crushed ore is pulled from the crusher discharge bin over an apron feeder, onto a transfer belt, and delivered to the first leg of the main conveyor system. Ore is conveyed approximately three kilometers to the mill stockpiles.
5.3 Ventilation The GRSBC area will be supplied and exhausted with air via primary intake and exhaust drifts. These drifts will portal close to adits that access existing development at elevations of about 3,000 meters. Four main intake and four exhaust drifts, each sized at 6.8m square, are required to facilitate the total estimated airflow of 1,755 m3/s required for the mature mine. Details of the proposed ventilation system are in the MassMin 2004 proceedings paper by Duckworth (2004). Airflow is delivered and exhausted via connecting raises from ventilation service drifts on the 2790m service level. This allows the full width of the footprint to be segmented into five separately ventilated zones. The undercut level will be discretely ventilated via two ventilation drifts to be driven across the undercut level prior to caving. 5.4 Ore Handling Ore is delivered to the haulage level by a series of 4.0m diameter bored raises; the material will pass through the grizzly, with a 1.0m opening, to the haulage level. The raises can serve either as ore passes or as intake and exhaust during their life. The concept is to be able to swap raise functions during the mine life of the panel, allowing a worn ore pass to be quickly converted into a vent raise and the vent raise to an ore pass. Figure 8 illustrates the concept. The ore passes are monitored over their life, and when wearing becomes excessive the orepass will be switched over to the vent raise. Rail haulage was the system selected for ore handling of the operation. The system will operate on a two-way shuttle scheme with empty trains running into the end of rail line under the orebody and then reversing back towards the crusher station. With the side-loading chute configuration the trains can be loaded in either running direction. Rail cars are bottom-dump ASEA-type, 20-tonne capacity units. Split dump points allow multiple cars to load into one of three coarse orebins. Massmin 2004
5.4 Production/Development Simulation Simulations utilizing the Arena© software program were performed to confirm production capacity, production and secondary breakage equipment requirements, ore haulage, and requirements for the AB adits to handle all development activities. The results to achieve a daily production rate of 115,000 tonnes per day produced the following requirements: • 45 operating LHDs • 864 active drawpoints • 16 low hang-up drills • 26 non-explosive breaking units (Commando/Boulder Buster) • 4 haulage trains consisting of 24 cars of 20 tonnes per car • Three 60" x 89" crushers The simulation runs incorporate the fragmentation predictions for the mine, to include in the production simulations drawpoint hang-ups as well as drawpoint oversize. Estimated frequencies of hang-up or oversize events simulate secondary breakage activities. Maximum numbers of equipment are allowed in a panel drift at a given time, and interference with the LHDs is taken into account. Similarly, the rail haulage was simulated to confirm production capacity. Although there are additional investigations required to further evaluate and optimize design of the rail haulage system, the model shows that the required tonnes per day production capacity can be achieved. 6 SUMMARY The mine plan as currently developed is based on current technology; many of the design parameters are based upon experiences gained during successful development of the
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DOZ block cave mine. Grasberg clearly has the footprint size and the column heights to sustain production rates of 115,000 tpd or higher. However, it is recognized that the actual construction, operation, and cave management of a single mine with those dimensions presents numerous challenges that must be thoroughly explored during the feasibility phase. The Freeport operation will enter the underground-only era with the completion of the Grasberg open pit in 2014. The development that has begun in 2004 on the A.B. adits will provide the gateway to the major underground operations. The Grasberg block cave will be the flagship of the underground era. Freeport will continue to strive to improve the mine design and district-wide development plan so that the future of the district carries on the tradition of mining excellence and engineering achievements that have been trademarks of the Ertsberg district in the past.
7 ACKNOWLEDGEMENTS This paper is a brief summary of the work carried out by many who have participated in the prefeasibility study as well as in earlier precursor studies. In addition to contributions from many of the underground technical staff at PT Freeport Indonesia, significant input was provided from Crescent Technology Inc., Call and Nicholas Inc.,
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McIntosh Engineering Inc., Hydrologic Consultants Inc., and others. The permission of Freeport-McMoRan Inc. to present this paper is gratefully acknowledged by the authors. 8 REFERENCES • Diering, T, 2000. PC-BC: A block cave design and draw control system. Proceedings MassMin 2000, Brisbane, (Ed: G Chitombo), 469-484. Australasian Institute of Mining and Metallurgy: Melbourne. • Duckworth, I, 2004. Preliminary ventilation design for the Grasberg block cave mine. Proceedings MassMin 2004, Santiago, Chile. • MacDonald, G D, and Arnold, L C, 1994. Geological and geochemical zoning of the Grasberg igneous complex, Irian Jaya, Indonesia. Journal of Geochemical Exploration, 50 (1994): 143-178. • Mealey, G H, 1996. Grasberg; Mining the Richest and Most Remote Deposit of Copper and gold in the World, in the Mountains of Irian Jaya, Indonesia, 384 p. Tien Wah Press (Pte.) Ltd., Singapore. • Srikant, A, and Nicholas, D E, 2003. Geotechnical Design Parameters for the Grasberg Block Cave. Unpublished PT Freeport Indonesia document produced by Call and Nicholas, Inc.
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Combined open pit – underground operation at El Teniente: facing a new challenge Octavio Araneda O., Resources and Development Manager, Patricio Yañez U., Mine Engineering Superintendent, Patricio Vergara L., Chief Long Term Mine Planning, El Teniente Division, Codelco Chile
Abstract El Teniente’s Business Plan, a new strategic and long term view of resources development was defined. The work was based on a scenario analysis having as input Codelco’s mission and vision and some key strategic aspects regarding the mineral resource base, the current reality of the mine, and future possibilities of development. The process, that started analyzing several expansion alternatives, ended selecting an option combining both open pit and underground caving operations as the best long term option to capture value. A 100 years old full underground mining complex has now to face the challenge of raising an Open Pit mine. The paper presents the overall process, the final plan and discuss some challenging issues.
1. INTRODUCTION Codelco’s mission states the search for maximizing the economic value of the company in a long-term perspective. Additionally, one of the most demanding strategic objectives imposed by the current Administration is its decision to double the value of the Company by year 2006 in comparison with the results from year 2000. To achieve this particular objective of capturing value, the identified resources presented the most attractive potential. Clearly then, El Teniente pretends to raise a plan that maximizes, under accepted risk assessments, the profitability of its main asset, the deposit, known as probably one of the biggest copper deposit in the world. A process named the "Business and Development Plan" was then established, formally an intense strategic exercise to find out the best route to capture most of the deposit’s value.
Today, the ore is processed both in Colón, concentrator located 10 km from the Mine at elevation 1980m and with a capacity of 104,000 tpd, and in Sewell, the old milling plant situated 1 km from the mine, at elevation 2200m, and with a capacity of 23,000 tpd. Copper concentrate is delivered to Caletones smelter, where copper anodes and fire refined copper are produced.
As a primary goal, this means to set up a plan allowing the Division to increase mine output, and to reach it in accordance with the current development plan under construction (PDT) but also with the production commitments engaged in the current Base Case. Figure 1: Location of El Teniente. 2. GENERAL OVERVIEW El Teniente, one of the 5 Divisions of Codelco Chile -the Chilean state-owned company-, is situated 80km south of Santiago and 44 km (Figure 1) up to the Andes, and it comprises mining, processing and smelting facilities. At El Teniente over 1,100 million tons of ore have been mined out during almost 100 years of mining. Today the mine has begun an expansion program driving the production from 100 Ktpd to 130 Ktpd of mineral. This plan, the El Teniente Development Plan (PDT), precises that during 2004 the mine is to produce 127 Ktpd of mineral, i.e. around 450 Kt Cu per year, becoming one of the biggest underground copper complexes of the world. Massmin 2004
3. GEOLOGY AND RESOURCES El Teniente is one of the largest known deposits of porphyry copper in the world. The main rock types of the deposit are: Andesite (3 per cent), Diorite (12 per cent), Dacite (nine per cent) and Breccia (six per cent). Copper mineralization is set around a subvertical intrusive pipe of breccia named "Braden" that appeared once the mineralization process was completed. This intrusion moved apart some of the original volume with copper, creating a sort of backbone to access the mine.
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As a result from this, mining at El Teniente has historically followed a mining pattern around the pipe, where the infrastructure is built in (offices, maintenance garages, lifts, etc). See Figure 2.
The "planned" column correspond to the result from the former planning instrument in a 25 years exercise: an instrument designed for budgeting and control more than value creation. Because of its limitations, it could only capture 18% of the economic potential and didn’t provide a correct assessment on the company’s value. From previous figures it becomes clear that El Teniente’s deposit is an invitation to growth in mine production. In this sense, the expected result of this process is to declare a new strategic road map, to direct exploration needs, engineering decisions, technological innovations and expansion investments while transferring its opportunity cost to current operations. BDP’s first step consists in finding mutually excluding scenarios focused in creating value at different risk levels. Preliminary economic evaluations are computed with almost no capital restrictions and with special flexibilities regarding the technological basis and the engineering status of projects. A risk assessment is then also required in order to provide Head Office with a value at risk estimation. The overall BDP process from the scenario generation to the final decision is schematically shown in Figure 3.
Figure 2: Underground mining sectors and projects Identified in situ resources at El Teniente account for almost 80 million tons of copper plus some 13 million additional in tailing dams and remaining broken ore from the caved zone. Tables 1 and 2 present their distribution according to JORC’s. Table 1: In Situ Resources Measured
Indicated
Inferred
Total
Mtons Ore
1,463
3,097
8,344
12,903
% CuT
0.89
0.71
0.54
0.62
13
22
45
80
Mtons Cu
Table 2: Tailing Dams and Caved Zone Measured
Indicated
Inferred
Total
17
32
1,823
1,878
% CuT
0.54
0.63
0.71
0.71
Mtons Cu
0.1
0.2
13
13
Mtons Ore
4. BUSINESS AND DEVELOPMENT PLAN PROCESS The Business and Development Plan (BDP) is a new planning instrument aiming to estimate Codelco’s value under a long term perspective. It is basically a new, less restrictive, view on the resources available and focused on getting out most of its potential. The following table (Table 3) shows El Teniente’s in situ resources vs. planned reserves. Table 3: El Teniente’s In Situ potential
Mtons Ore % CuT Mtons Cu 630
Identified
Economic
Planned
12,903
7,500
1,100
0.62
0.83
1.03
80
62
11
Figure 3: The BDP Process Selection from the different scenarios is realized by a Head Office’s team of experts, in a Corporate perspective, and finally proposed to Codelco’s Board of Directors who finally takes the decision on which scenario to concentrate efforts. The selected scenario is then refined and some restrictions are applied related to the engineering basis and reserves category requisites. Finally, a plan and a definite economic evaluation are provided including: - Ore reserves and mineral processed - Ore grades, capacities, and metallurgical recoveries of the concentrator, and smelter - Total production by product - Investment profile over time, depreciation and amortization
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- Costs by business, process, activity and major equipment fleets - Operation cost by element: HHRR, materials, fuels, electric power and services - HHRR profile, payroll personnel and third parties - Incomes - Productivities and Performance indexes - Prices and Tariffs - Others 5. SCENARIOS GENERATED Five main scenarios were generated as part of El Teniente’s 2004 BDP process. Figure 4 summarizes how they were conceptualized.
6. EL TENIENTE’S 2004 BUSINESS AND DEVELOPMENT PLAN The finally selected scenario for El Teniente corresponded to a combined open pit – underground operation pushed at its almost maximum capacity with current planning parameters. Figure 6, shows a graph with sectors grouped as in: AT8 BO NL OP DL
: : : : :
Sectors Above level Teniente 8 Broken ore New Level, below Teniente 8 Open Pit Deep Level, conceptual planning continuity of the deposit
Figure 4: BDP Scenarios Figure 6: Plan selected as El Teniente’s BDP Scenario 1 is the Base Case scenario, the former long term plan available for El Teniente. It is conservative by definition, since it keeps the current expanded capacity of 130,000 tpd over time. Scenario 2 considers a marginal expansion using only underground mining and reaching the limits of the current infrastructure capacity. Scenario 3 (and its variants) was based on the possibilities given by an underground expansion through a brand new panel, reaching its maximum capacity. Scenario 4 (and its variants) explores the potential of a combined underground and open pit mining strategy. Scenario 5 is similar to scenario 4, but in a context of maximum capacity for both mining methods. Variants in scenarios 3 and 4 differ basically in the kind of reserves included, technological basis and operational and geomechanical risks. The relative economic evaluation of scenarios is shown in Figure 5:
It is a long term plan with a focus in the current expansion in progress (PDT) in the first years. Intermediate expansions area then based in the use of broken resources (remaining of past mined out sectors), investments in two additional mining sectors plus marginal expansions coping current infrastructure capacity (railroad, Colon concentrator). Finally, expansions that allow boosting the overall mine capacity are a whole new panel (2014) in addition to an open pit operation (2016). Selection was based on the following criteria: • Plan according to the huge amount of resources available • Mine capacity offering economic returns capable to afford investments • The diversification of the mining methods provides the opportunity to have flexibilities • Scenario lets the Division in a good position facing future expansions • The choice of having an open pit is still a reversible decision: resources can be mined underground • Value creation according to Corporate goal to double the value of the Company. The scenario is also attractive because: • The open pit project is planned only starting in year 2016, being very conservative regarding the current engineering phase of the project and its geological information • Without compromising the short term of the plan, the scenario sets a clear guideline in terms of long term strategy • Finally, it is almost a fact (according to external and internal engineering judgment), that the open pit project will improve during the next engineering phases
Figure 5: Economic Evaluation of Scenarios Massmin 2004
The open pit project is expected to be in the north-west side of the deposit, letting the east and south side too underground mining. Santiago Chile, 22-25 August 2004
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Figure 7 shows a plan view with the underground sequence and the open pit. It is clear the degree of interaction between both, also that an adequate sequencing becomes a critical issue in planning.
Interaction in certainly another of the most challenging aspects of the plan: first among underground sectors, but also between the open pit and underground operations. Sequencing was treated initially under a heuristic approach, however the complexity of the problem, and the fine-tuning required, forced the need of having it treated in a more detailed way. A four dimensions model was set up based in the 3D subsidence angles defining active caving zones and its evolution in time (fourth dimension). The model, works based on the output from the underground planning and allows to easily constructing actives caving volumes using any market design and planning software. This allows checking the validity of the overall sequence and eventually introducing changes whether in the underground or open pit operation. See Figure 8.
Figure 8: Dealing with Open pit and Underground interaction at a specific year
Figure 7: Open Pit + Underground sectors 7. CHALLENGING ISSUES Challenging issues regarding the open pit project may be grouped into three main families: information, interaction and planning. The first one involves general requirements of information needed for raising a mining project. The second one is related on how to keep a combined operation with underground panel caving sectors. Finally, and as a third but more general aspect, the expected benefits of having such a project make it attractive to have it planned before year 2016. The quality of the resources involved in the open pit is indeed a matter of challenge. Especially the first years of the plan: broken ore from past mining, reserves in special "mining difficulty" zones (due to accesses, material from the Braden pipe, loose material from the alluvial zone...etc.). In comparison with the rest of the deposit, the area involved in the open pit lacks -in quantity- of important information. An aggressive exploration campaign, was proposed in order to improve the knowledge in geology, geomechanics, geotechnics, water management, waste dump alternatives, infrastructure position and also environmental issues (to establish a base line). Among these, one special issue is related to the quality (grades, recoveries) of the depleted caved zone that is included in the plan. This broken ore is the remaining left from past underground mining and, up to now, has always been treated in a very conservative manner: just as waste in the case of the open pit. A special drilling campaign is expected to give detailed information. Meanwhile, planning exercises based on a better assessment on its copper content and aiming to see the impact on the open pit design will be conducted. 632
On the overall planning aspect, it becomes clear that by giving a better estimate on the Division’s value, the BDP should allow better decisions in terms of the opportunity costs involved. In this case, it appears indeed very attractive to have the open pit starting before year 2016 trying to capture its value as soon as possible. If this is to be planned then we would see an increase in interaction which will probably cause to plan a smaller overall plan capacity, but nevertheless returning a still higher value. The challenge being to have it around year 2010, exercises will be started this year, for the 2005 Business and Development Plan process. ACKNOWLEDGEMENTS The authors are grateful to all their colleagues of the Superintendence of Mine and Metallurgy that helped them during the development of this process. Also, the authors want to acknowledge the permission given by Division El Teniente to publish this technical paper. REFERENCES • Crorkan, Araneda, et al. Caso Base 2003, Codelco Chile División El Teniente. 2002. • Codelco, Corporate guides for Mine Planning: PEX Process. Internal document CODELCO, 2003. • Codelco, Procedimiento de Estimación de Riesgo PEX, Codelco internal document, 2003. • Skewes, A et al (2002), The Giant El Teniente Breccia Deposit: Hypogene Copper Distribution and Emplacement in Society of Economic Geologists, Special Publication 9,2002, p.299-332. • Metálica Asoc, 2003.Estudio Material Quebrado Mina El Teniente. • NCL, 2003.Proyecto Rajo Teniente, Ingeniería Exploratoria.
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Going to an underground (UG) mining method Sergio Fuentes S., Director, Metálica Consultores S.A., Chile
Abstract Recently, underground mining methods have been analyzed by some open pit (OP) staff people, mainly because they are anticipating the end of the economic life of those operations in the near future. Block caving is one of the lowest cost underground mining methods and can compete with some open pits. It is a natural substitute mining method for open pit because of the high production rates, levels of mechanization and of course the cost level that can be achieved. Underground mining presents more technical risks that open pit method, with the possibility of events such as air blasts, rock bursts, hang-ups etc. These risks can be quantified and managed in a rational, technical and reasonable way. This paper briefly describes some key issues regarding block caving, some basic information requirements, cost trends, potential production capacity, management issues and the expected evolution of some techniques that could improve or solve some of the main technical constraints of the method.
1. INTRODUCTION If we have an open pit operation and we do not visualize the final pit in the medium term, e.g. 5 to 15 years, some basic question appears over the planning area, such as: • How many years in advance we have to consider an underground mining configuration? or, • Why we have to evaluate an underground mining method in an early stage of the open pit planning or production? Or, • Is a crazy idea to analyse UG mining 10 years in advance of the final pit achievement? These are some of the most common questions feelings or statements that we could "feel" the first time when we discuss the point about open pit to underground transition. Indeed, people reactions are commonly quite magnificent and with some kind of aggressiveness about this idea. In the following section it is briefly described our experience and approach managing those issues. Figure 1 Open Pit & Underground Resources Configuration 2. FIRST EMOTIONAL SHOCK Traditionally, the decision making process in the open pit planning, does not take into account the opportunity cost associated to the underground exploitation of the remainder resources left by an open pit design. Standard methodology considers sequential push back evaluation, identifying the expansion that maximizes the net present value (NPV) of the design. In most of the cases, an asymptotic trend of the NPV is demonstrated in the final expansions. As an example, this could be because the economic weight of the waste removal is not considered by undiscounted outputs from moving cone valuations. If we take into account the economical potential associated to underground resources (see Figure 1), the final pit can be significantly reduced if an additional expansion of the pit can not generate large enough NPV than the financial effect of an underground exploitation displacement trough the time (see Figure 2 ). The first time this kind of analyses is carried out, the economic impact and mine planning consequences can be impressive and shocking. The life reduction could imply that there is no time to start with an underground project. Massmin 2004
Figure 2 Open Pit versus Underground, Break Even Analysis
This effect can be named "First Emotional Shock", and the entire organization involved in the decision process, begins to look for problems or good excuses to delay further analyses.
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3. STRETCHING THE TIME Usually, until this time a break even analysis (OP versus UG) has been carried out using a primary approach for the underground exploitation with big uncertainties within the UG project basis. We called it "Diagnosis Study", where the mine planning team begins to optimize the OP designs, modifying production schedules, investment profiles, etc., searching better options for open pit planning and consequently improving the economic performance of the expansions. On the other hand, the mine planning team and the entire organization start configuring a strong navigation chart for the UG project, where some main topics are: • Ore reserves estimations. • Geotechnical characterization for UG purposes. • General UG mining method definition. • UG production rate searching. • Material handling questions. • Preproduction time required estimations. • UG production ramp up approach. • Subsidence and UG mining sequence. • OP & UG interaction. • Mechanization and automation. • Working time effectiveness. • Human resources. • Capital expenditure constrains. • Environmental. For sure, there are many more not mentioned topics in this brief list. However, the issue here is to introduce the importance and sense of urgency at the business administration level, not avoid right and proper managerial decisions. Don’t worry, we have time! It is a common statement when we see some years in front to start an UG project under these conditions. But usually "some years" are not enough to develop proper and certain analyses in this matter, mainly when it is necessary to collect and define base information for UG purposes. Stretching time is a common attitude hoping to find a better option for OP mine planning, and it is a valid and possible situation, but this searching does not justify to stop important decisions about further UG related studies. 4. BLOCK CAVING AS FIRST OPTION Some of the biggest open pit mines worldwide will be achieving their final pit limits during the next 10 to 15 years (Metalica, 2002). Block/panel caving will likely enable these operations to continue achieving a high production rate and low costs. A review of the literature indicates that the first block caving mine was the Pewabic in Michigan USA at the beginning of the 1900’s (Peele, 1927). This mining method has evolved from total cost close to 7 US$/t down to 4 US$/t during the last two decades (see Figure 3). One of the most important technical aspects in cost improvement has been the evolution from an ore grade selective approach to a massive exploitation, increasing cave draw control while avoiding as far as possible any increased dilution in the material produced. i.e. improving regularity of the drawing surface instead of selective production from individual draw points to keep ore grade as high as possible. . Those major planning changes allows mining layouts modification and changes in management criteria, such as outsourcing of preparation and development operations due to variable preparation requirements trough the time . The 634
Figure 3 Chilean Block Caving Mines, Total Cost Trend costs of these outsourced operations represented almost half of the total mining cost. If there were proper geometric and geotechnical conditions, this cost trend suggests block/panel caving is the best choice as mining method. 5. BLOCK CAVING CONSTRAINS AND VULNERABILITIES Today, we can remark that main constrains and weaknesses of this mining method are related with geotechnical considerations, such as: • Caving induction and progression. • Fragmentation and block size distribution. • Induced stresses over mine infrastructure. ç Rock mass strength capabilities. • Dilution behaviour. If we consider high preproduction capital and time required to start with production, we can find from economical point of view some additional risks associates with block caving. Even though, massive exploitation implies some reduction in the ore grade fed to processing plants, this situation permits to mine out low ore grades because of ‘unit’ mining cost reductions achieving high productions rates . Planning and design for those marginal mineral resources (low grades ore) implicitly involved greater economic risk, especially when the prices of metals mined were depressed. Finally, considering the mining method change the work force skills adaptation must be carefully analysed, planning the training requirements and calculating the productivity projections associated with the new system. Indeed, these two factor wrongly combined could produce catastrophic results in cost, productivity and finally, profitability. All that constrains and vulnerabilities generate a risky heaven sense compared to open pit mining, usually inducing reactive planners behaviour about changes of the current mining method. 6. MINIMIZING THECNICAL CONSTRAINS Getting good enough knowledge of the rock mass characteristics, such as geology, geotechnical, environment, ground water, etc., could make the difference minimizing risk and uncertainty of the engineering assumptions to manageable levels. Additionally, new techniques such hydraulic preconditioning, blasting preconditioning or a mix of both, are gaining knowledge and expertise within block/panel cave miners, transforming and adapting rock masses to our purposes (Chacón, 2002).
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Rock Mass Preconditioning research field, is one of the most impacting research areas for block caving purposes because of improvements on material handling and caving induction conditions. Indeed, some of the modification achieved with those techniques, almost avoid most of the geotechnical constrains and vulnerabilities above mentioned. Some, industrial test shows quite good results, favourable changing rock mass conditions over the material to be mined. This mix, using hydraulic and blasting preconditioning techniques, it seems as the most successful combination, significantly reducing hang-ups and consequently, improving productivity of the active area. Also, cutting infrastructure reparation cost, due to reduction of stability problems, shall be some other advantages associated with this preconditioning implementation as standard in hard rock mining. An important consequence of successfully rock mass preconditioning shall be continuous material handling as much as possible. Indeed, this aspect could sustain important mining cost reductions increasing profit in low ore grade exploitations. 7. ADDITIONAL IMPORTANT TOPICS Other important topics, previously mentioned, cannot be underestimated. Those could introduce higher levels of uncertainty than the technical variables. When we are going from open pit mining to an underground method, it is necessary to develop and implementing a reengineering process over the entire structure of the mining business, taking into account new expertises, skills, risks, process controls, etc., generally speaking, appear new border conditions to introduce in the organization. For example, if there is no correction or cleaning of environmental liabilities associated with open pit operation, it makes possible the introduction of high no (under) estimated costs conditions into the projected results over the future underground operation estimations. The analysis and definition of the real choice to make a proper personnel searching and selection, and then the training programme associated, etc., could be an important risk factor that could imply success or fail of the project projections. Also, proper configuration, programming and management of the preproduction development (preparation), while a high level of simultaneous requirements are needed to achieve the goal in a short period of time, is an issue and could imply a huge uncertainty of the profitability of the project.
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Finally the change of the strategic position in the market is an important additional factor, while the most probable production scenario is that a large open pit with large fine copper production will achieve smaller output scale in terms of production. That situation introduces drastically changes in the relative strategic positions in the market, consequently we could expect some cost impacts because of this already changed strategic position. 8. CONCLUSIONS • The economic final pit is usually closer than we think. • Reactions against break-even analyses are usually found in an early stage of OP-UG studies. • Open mind of the entire organization is the best way to find a proper and profitable transition from OP to UG mining. • Underground mining using block/panel caving appears as first option in the transition of actual large open pit mining to underground. • Rock mass preconditioning appears a good guideline of research and development, contributing significant improvements in block/panel caving mining. • A successful rock mass preconditioning, suggests important improvements in material handling in block/panel caving designs. • Several non technical aspects, such as work force training, environmental constrains, etc., must be deeply analysed in primary stages of the transition studies. • If we consider the long time needed to start up with an UG project, OP limit taking into account UG opportunity cost, should be frequently analysed along the mine life. REFERENCES • Chacón, E., (2002). Pre-Acondicionamiento de la Roca Primaria en Codelco Chile, Seminario de Innovación Tecnológica en Minería, Mayo 2002, Santiago, Chile. • Fuentes, S., (2003). Planning Block Caving Operations with Metal Price Uncertainty, MSc(Eng) Thesis, Queen’s University, Kingston, Canada. • Fuentes, S., and J. Cáceres, (2003). How Improvements in Block Caving Method and Costs Have Impacted in Final Pit Definition, 2003 CIM Annual Mining Industry Conference & Exhibition, Montreal Canada. • Metálica S.A., (2000). Análisis de Límite Rajo Subterráneo, Internal Report. • Peele, R., (1927). Mining Engineers’ Hand Book, John Wiley & Sons , 2nd Edition, New York, USA, Vol. 1, pp 721 -727.
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Chapter 17
Mine Operations
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The Application of new underground mining technology and sound systems engineering principles to develop a cost-effective solution for the Finsch mine block 4 ore management system Danie Burger, Mining Systems Manager, De Beers Finsch Mine James Oosthuizen, Account Manager, Sandvik Mining and Construction RSA (Pty) Ltd. Brett Cook, Project Manager, Mine Automation Department, Sandvik Tamrock Corporation Japie Visagie, Divisional Project Engineer, De Beers Finsch Mine
Abstract The capability to excel in the acquisition or development of products, services, systems and processes has become vital in the global, competitive environment. This is true not only for the young, breakthrough industries but also for the traditional, mature industries. To be world-class, the critical business function of system/product and process development needs to be given high priority in order to ensure optimum performance at reduced operating and support cost. This includes, amongst others, the use of a formalized system product development and acquisition process. This paper describes the systems acquisition engineering process followed during the development of the Finsch Mine - Block 4 Project, Ore Transportation System (OTS). Furthermore, it addresses the business and production management requirements of the block cave operation in terms of the effective and sustainable management of the ore body, utilising the OTS as a part of the overall Ore Management System (OMS). The paper also describes the conceptual trade-off study methodology and selection model utilised to define the preferred OTS concept together with a description of the selected system configuration for the Block 4 Project. The selected system implementation and commissioning strategy together with the system’s operating and support concepts are also briefly highlighted. The paper concludes with a generic systems overview and description of associated benefits of the key technologies that make up the Integrated Ore Management System..
1 DE BEERS FINSCH MINE OVERVIEW Finsch Mine is situated approximately 165 kilometers west of Kimberley in the Northern Cape Province of South Africa. (Figure 1). The Finsch kimberlite pipe is a near vertically sided intrusion into the country rock, the latter consisting of dolomite, dolomitic limestone with chert bands and almost pure lenses of limestone (Figure 2). Exploitation of the kimberlite pipe commenced in 1964 utilising an open pit mining method. Approximately 120 million tons of kimberlite ore was mined from the open pit. Underground operations commenced during the latter part of 1990. The underground ore body is divided into mining Blocks 1 to 5. (See Figure 3). Each block terminates at varying depths below surface as 350m, 430m, 510m, 630m and 830m respectively. Blocks 1 through to 3 are being exploited via a modified blasthole open stoping technique. Blocks 1 and 2 were mined out while Block 3 is nearing depletion in 2007. Blocks 4 and 5 are planned to be extracted utilising mechanised block caving mining methods. Block 4 underground infrastructure development is well underway with production startup planning to commence during the latter part of 2004, while exploration drilling of the Block 5 Massmin 2004
Figure 1: Geographical Location of Finsch Mine resource is in progress already – aimed at improved qualification and quantification of the remaining kimberlite pipe at depth.
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Figure 4: 61 Level – Undercut level layout
Figure 2: Geological model of the Finsch kimberlite pipe
Figure 5: 63 Level – Extraction level layout
Figure 3: Cross section of mining blocks
Block 4 will be mined by means of a mechanised panel retreat block caving method, utilising an advance undercutting methodology. The undercut level of the block is situated on 61 level (610m below surface). The main extraction level, trackless workshop infrastructure, truck haulage tramming loop and primary crusher excavation and associated infrastructure are located on 63 level (630m below surface). The slot cutting process required to initiate the undercutting process has already commenced in April 2004. The Hydraulic Radius necessary to allow the caving process to commence, ranges between 14 to 18 and this specific milestone activity is scheduled to be achieved during the first quarter of 2005. The undercut tunnels are spaced at 15 m intervals, and will allow for the mining of an inclined 12 meter high undercut profile. The undercutting process is planned to be completed in 2009. The layout geometry of the extraction level is designed on a staggered herringbone configuration with 30m tunnel spacing and 15m drawpoint spacing. Based on geotechnical constraints and associated tunnel stability risks, the mining sequence on the extraction level is such that all drawbell and associated infrastructure will require development in the "shadow" of the advancing undercut face – with only the extraction tunnels and 640
drawpoint development allowed to take place ahead of the undercutting process. Depletion of the Block Cave draw columns will commence once the Hydraulic Radius is achieved in early 2005. The production buildup rate will be dependant and to a large degree, be governed by the rate of undercutting as well as drawpoint maturity rules. A full production rate of 3.8 Mta is planned to be achieved in 2007 and this steady state production rate will be maintained until 2011. The production from this block will then reduce significantly until the block 4 resource is planned to be depleted in 2015. To maintain production from the underground mine, Block 5 will have to start full production during 2011 thereby ensuring sustained production of Finsch Mine beyond 2020. 2 INTRODUCTION Growth within the De Beers organisation is essential for sustainability and increased market share. New acquisitions as well as expansion of existing operations such as the Block 4 project at Finsch Mine, requires that De Beers follows a process that produces world class outputs in both quality and performance. For this purpose De Beers adopted a formal project management and systems acquisition approach. The main concept of the De Beers Total Project Management (TPM) and system acquisition process
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involves the effective management of time, cost and quality of a complex multi disciplinary system applied throughout the system life cycle (Figure 6). The whole life cycle of the system is taken into account, from determining the needs and requirements, system acquisition, right through to system utilisation whilst always considering the interfaces and interactions with its environment.
Figure 6: System life cycle model This paper aims to explain the system acquisition principles applied to the Ore Transportation System (OTS) as part of the Finsch Mine Block 4 project. 3 OTS SYSTEM ACQUISITION PROCESS 3.1. Needs identification and requirements determination The baseline document of the Requirement Determination is a User Requirement Statement (URS), in which Finsch Mine Block 4 end-user and mining client requirements and constraints, from context requirements through to operating, support, cost and schedule requirements for the specific Ore Transportation System are covered. The URS defines how the resulting system will perform the ore transportation functions. The OTS is required to collect ore from the various draw points and to convey the production ore to a bin at the underground crusher. This transportation system shall furthermore receive ore from the mining development system and convey development ore to the underground crusher bin . The URS provides a point of departure in the search for the optimal system solution without considering any specific, preconceived solutions. It does however, include high level cost ,time and quality constraints within which the ultimate solution must be developed and implemented . These constraints are then effectively utilized as baseline key performance indicators against which the continuous performance of the system can be evaluated over the complete system life cycle. 3.2. System Acquisition There are four phases within system acquisition, with the following main objectives: 3.2.1. Concept exploration phase (Pre-feasibility estimate) The primary objective of the concept exploration phase is to identify and explore all potential solutions to the defined URS. All potential solutions are then evaluated against defined key performance criteria as specified within the URS using a selected decision model. By applying suitable weighting measurement criteria, each solution can then be suitably rated and the best rated solution will be ultimately selected. Therefore resulting in a single system specification being a formal deliverable for this phase. Massmin 2004
3.2.2. Definition and validation phase (Feasibility estimate) In this phase the preliminary design of the selected system specification is done in order to firm up on the cost estimate of the solution to enhance the certainty that the selected solution is the optimal alternative. This preliminary design includes layouts, general arrangements, preliminary equipment selections, etc. During this phase, solutions are engineered to a level of confidence sufficient to request financial approval for the project. All equipment, technologies and processes will have to be specified and costed at a more detailed level. The Engineering process followed during this phase, should effectively utilize integrated risk and change management principles aimed at identifying, eliminating or adequately controlling any undue risks associated with the selected system design. 3.2.3. Design and development (Control Budget) Following the definition and validation phase, the preliminary design is further developed to a detailed design level suitable for construction, manufacturing and commissioning purposes. The deliverable from this phase is a detailed system breakdown structure complete with definite time, cost and quality estimation, as well as detailed system commissioning, manufacturing and execution plans. 3.2.4. Construction, manufacturing and commissioning (detail budget) During this phase the system is constructed as a product. Items and equipment are manufactured and the system product as a whole is assembled, commissioned and handed over to production for use. At the end of this phase the system is signed off as a fully operational system, and is normally associated with the handover from project team to operational team (end-users). Continuous stakeholder involvement and management are essential throughout the entire system acquisition process. This allows a smooth and successful transition from commissioning to operating and support of the system. 3.3. System utilisation During the system utilisation phase, the system is operated and supported by the end-user and supplier until the end of the system life cycle has been reached when the system is disposed of. 3.3.1. Operations and support Operation and support is defined as the mature phase of the system acquisition process during which the commissioned system is operated and maintained in its intended operational mode. System performance during this phase continuously needs to be measured against the initial URS to ensure optimum operational and end-user satisfaction. Sub optimum performance and end-user dissatisfaction could negatively impact on the system life cycle performance as a whole due to the risk of premature system disposal. 3.3.2. Disposal and restoration This entails the decommissioning and termination of the system from its operational service in the most cost effective manner. 4 PARTNERING FOR SUCCESS Sandvik Mining and Construction (SMC) was awarded the role of system integrator and joint partner with De Beers Finsch Mine in the development, design, manufacturing, commissioning and support of the Ore Transportation System solution for the Block 4 project.
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SMC applied a systems acquisition process defined by De Beers Finsch Mine, with specific reference to the following scope of work:• Execute all the system acquisition phases of the system life cycle, from needs identification, up to the final commissioning, operation and support of the system accepting overall system acquisition and performance responsibilities. • Perform all systems engineering necessary to address the entire User Requirement Statement (URS) from high level requirements down to detailed levels. • OTS engineering design, manufacturing and commissioning project management responsibilities. • Satisfy all engineering design and development requirements. • Procure, manufacture and integrate all system elements. • Commissioning of total system. • Participate in operational risk assessments ( Safety and production related) • Relevant stakeholder engagement • OTS system training responsibilities. • Support the eventual system during utilization and phaseout.
This solution proved to be the most cost–effective with the capability of delivering optimum revenue over the life of mine together with improved draw control and management of the Block 4 resource. The hybrid OTS will operate within specified operational and associated underground infrastructural constraints and will significantly reduce safety related risks. The Block 4 Ore Transportation System (OTS) provides for all the functionality required to operate and supervise an automated LHD and dump truck fleet from a central control room located on surface - removed from the production area. 5.1. Ore Transportation System Overview Ore is loaded by LHDs from the 302 production drawpoints located within 11 extraction tunnels as well as from 5 undercut passes. The LHDs then tram the loaded ore to 4 designated transfer points located on the perimeter of the orebody after which the ore is dumped into autonomous trucks (Figure 8) . The trucks then automatically haul the ore and dumps it into the primary crusher located at the shaft.
De Beers Finsch Mine’s partnership role pertained to the following client interface responsibilities: • Project sponsor; • Owner of the needs identification and requirements specification; • Active partner and system acquisition process stakeholder; • Operational end-users. • Integrated Block 4 project management responsibilities. • Finsch Mine and De Beers group stakeholder engagement and management responsibilities. • System operational responsibilities. 5 BLOCK 4 PROJECT: OTS SOLUTION The Ore Transportation System forms a major component of the Ore Management System (OMS) for Block 4. The OMS is an integrated system that will manage the planning, optimal extraction, reconciliation and effective management of the Block 4 ore resource. As a sub-system of the OMS, the OTS is responsible for the optimal transport of ore from production draw points to the underground primary crusher.(Figure 7)
Figure 8: 63 Level - OTS Layout The LHDs selected will be Toro 007’s fitted with a 5.4m3 bucket with a 10 ton tramming capacity. LHDs are operated either manually or semi-autonomously depending on the implementation stage. (Figure 9) The trucks selected will be Toro 50D’s equipped with a 24m3 rock box capable of hauling 50 tons which will be operated fully autonomously (Figure 9).
Figure 7: OMS system context Figure 9: Selected loading and hauling equipment The preferred ore transport system solution selected from the concept exploration phase was a hybrid system configuration consisting of a combination of manual and automated LHDs and automated trucks. 642
This combination of LHD bucket size and truck box size allows for 4 pass loading.(Figure 10)
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which the manual and automated machine fleet will be supervised and controlled. It will facilitate the viewing and assignment of statuses and objects in the production area.
Figure 10: LHD tipping configuration 5.2. OTS System interfaces The following sections briefly describe the sub-systems below in the operating segment of the OTS and the two external systems that are interfaced to the OTS (As shown in the figure below).
5.2.3. Mission Control System The Mission Control System (MCS) has the following main functions: • Control of production execution including: - issuing manual production orders to manual LHDs. (Manual production orders are received from the PCS.) - creating and issuing drive orders to autonomous LHDs and autonomous trucks. (Drive orders are created based on auto mission lists received from the PCS). - traffic control (autonomous machines only) - management of truck loading by LHDs • Supervision of the machines and production area resources. • Monitoring of the total fleet production. • Monitoring of the fleet condition and status. • User interface for supervision and control of autonomous LHDs and autonomous trucks 5.2.4. Access Control System The fundamental safety concept for the operation of a hybrid OTS system is to provide an isolated area for automated machines from manually operated machines and personnel at all times. The Access Control System (ACS) conforms to this safety concept by using both physical and electronic means with the following main functions: • Control access to autonomous areas including: - restricting entry; - detecting unauthorized entry; - preventing unplanned machine exit; - providing controlled access for autonomous machines; • controlling the boundaries of autonomous areas including: - adding/releasing extraction tunnels to/from the autonomous LHD area - adding/releasing haulage tunnels to/from the autonomous truck area
Figure 11: OMS interfaces
system architecture with OTS system
5.2.1. Cave Management System The Cave Management System (CMS) is an external system that, from an OTS perspective, has the main function of providing a daily draw order (i.e. daily production targets and priority for each drawpoint) to the Production Control System (PCS). This daily draw order takes into account the undercut face position as well as the drawbell installation schedules. Each of the drawpoints will receive a draw order that takes into account geotechnical and mining related rules. This system also receives information from the previous shifts production in order to recalculate the required tons per drawpoint for the next period. This will facilitate effective draw control. 5.2.2. Production Control System The Production Control System (PCS) provides main functions to schedule shift production and to manage the production execution by manual and autonomous LHDs and autonomous trucks. The PCS creates Manual Production Orders (MPO) for manual LHDs that are grouped into a Manual Production List (MPL) for each machine. The PCS creates missions for autonomous LHDs and autonomous trucks that are grouped into an Auto Mission List (AML) for each machine. The system is based on a client-server configuration and users interact with a PCS client connected to the De Beers network. The PCS is operated through a Windows® based dedicated graphical user interface with Massmin 2004
Each access or entry point to the autonomous areas will be equipped with warning signs, zone status indicator lights, control panels, a lockable gate to restrict entry, a photocell to detect unauthorized entry of personnel or machines and also to detect unplanned machine exit attempts. (Figure 12)
Figure 12: Access gate system configuration Access points consists of 2 components: Access Gates and Transit Locks. Access Gates are used to barricade off normal access points to the autonomous areas. Access
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Gates are opened and closed manually by mine personnel. Transit Locks are a combination of 2 access gates and are used to introduce and/or remove autonomous machines from the autonomous areas under controlled conditions whilst allowing other autonomous machines to continue operating. The inner gate of the transit lock is remotely controlled whilst the outer gates are opened and closed manually by authorized underground mine personnel. In addition to the photocells to detect unplanned machine exit attempts, a "hard-wired" safety antenna system will be fitted to the autonomous machines with a tripping device to be installed at each access gate. The safety antenna (Figure 13) will be a pole that is raised when the machine is introduced to the autonomous area. When tripped the antenna will mechanically break the emergency stop circuit and immediately stops the machine independent of any communication system. The safety antenna tripping device will be a steel cable that is suspended at a height sufficient to trip the safety antenna (i.e. knock it down).
Figure 13: Safety antenna tripping device 5.2.5. Operator Station The Operator Station is the location from where the autonomous fleet is supervised and controlled and is located in the surface central control room. One of the main functionalities performed by the operator station is to assist and facilitate tele-remote bucket loading of the autonomous LHDs.
Figure 14: Operator station
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5.2.6. Citect SCADA Citect is an external Supervisory Control and Data Acquisition (SCADA) system that provides data to the Mission Control System essential for the monitoring and control of the OTS. Data includes: • crusher bin level: ton capacity available in bin - (% level) • crusher feed rate (at apron feeder)- (tons/hr) • crusher operational status-(running/stopped) • u/cut pass levels – (tons in pass, high level, low level) • cumulative crusher discharge tons - (tons) The data from the Citect Scada system is integrated with the PCS which utilizes this data to optimise production throughput by monitoring upstream and downstream resource-related and other operational constraints. 6 BENEFITS OF OTS SYSTEM ACQUISITION The following list summarizes the benefits of applying a formal acquisition process: • The risks of failing to meet technical performance, schedule and cost requirements of the project are significantly reduced. • The step-by-step authorization of baselines protects the project team, the end-user as well as the stakeholders. • Clear management responsibilities and transfer thereof are identified during the different phases of the system life cycle. • Clear statement of requirements upfront. • Traceable and auditable decisions/processes • Formatted documentation and revision tracking (change control process) • Clear/forced objectives and performance measurement criteria over system life cycle. • Knowledge capturing and sharing of intellectual capital • Integrated client/supplier team synergy. 7 OTS BENEFITS The Finsch Mine Block 4 Ore Transportation System provides all the functionality required to operate and supervise an automated LHD and dump truck fleet from a control room located away from the production area. It provides the following benefits: • Effective block cave management and control through automation.(improved quality) • Improved machine fleet utilization: More efficient fleet utilization results in increased production. (increased revenue) • Lower operating costs: One system controller is able to manage several machines. • Lower maintenance costs: Smoother operation and reduced damage to equipment cut down maintenance expenses. • Optimized tramming speed : machines are able to maintain optimized speeds throughout the route in each LHD and dump truck cycle • Increased safety: The Access Control System protects the automated production area from both unauthorized access and machine escape. People are removed from dangerous underground environment. • Improved working conditions: System Controllers are located in a safe and comfortable control room environment which implies decreased occupational injuries. • Improved production control and monitoring (reducing stoppage times which results in improved recovery rates). • Real–time production follow–up and response: Using the MCS interface, the system controller can easily monitor the production in real time.
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By following a systematic and methodological systems engineering approach to the design and development of this transportation system, the benefits described in the previous two sections would be realized.
not only to provide the OTS system solution on time and within budget but that it also meets all customers requirements thereby ensuring full customer satisfaction. 9 ACKNOWLEDGEMENTS
8 CONCLUSIONS Applying the core principles of high level up-front thinking early on in the lifecycle of a project will ensure that the most suitable requirements and solutions are defined and successfully achieved. The correct concept needs effective development, design, manufacturing and implementation to ensure that the process required to produce the final system solution is guaranteed, reliable and that it meets the customers requirements and satisfaction. By utilising these principles captured under the De Beers total project management and system acquisition methodology, effective concept exploration and design have been assured. Optimistic expectations related to the commissioning, operating and support of the OTS by the end of the year will be successfully executed to the specified performance criteria. This current focus on time, cost and quality will ensure that Finsch Mine Block 4 project has the ideal opportunity
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The authors are grateful to all their colleagues/partners at Finsch Mine and Sandvik Mining and Construction (SMC) that have contributed to the development of this system solution. Also, the authors want to acknowledge the permission given by the partners, Finsch Mine and SMC to publish this technical paper. 10 REFERENCES • De Beers Total Project Management System Guidelines for Project Management. • User Requirement Specification for Block 4 Ore Transportation System, Revision F, 04-06-2003 • System Specification for Finsch Mine Block 4 Ore Transportation System, Edition C, 14-02-2003 • Block 4 OTS-Operating System Preliminary Design Report, Revision B, 24-03-2004
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Visual grade control techniques and sub-level cave draw optimisation – Perseverance Nickel Mine, Leinster, Western Australia Geoff Booth, Senior Mine Geologist, WMC Resources Ltd., Perseverance Nickel Mine, Leinster, W.A. Ernie Gaspar, Sam Vine, Leanne Noble, Mine Geologists, WMC Resources Ltd., Perseverance Nickel Mine, Leinster, W. A. Scott Dunham, Geology Manager, WMC Resources Ltd., Leinster Nickel Operations, Leinster, W. A. Glenn Sharrock, Senior Rock Mechanics Engineer, AMC Consultants, West Perth, W. A.
Abstract Mass mining methods such as sub-level caving (SLC) require rapid and flexible grade control procedures. With SLC operations often suffering from erratic waste ingress, intensive inspection regimes are now widespread. Throughout its entire underground development, WMC’s Perseverance Nickel mine at Leinster, Western Australia has relied exclusively on inexpensive visual techniques to quantify grade. With ring designs ranging from ~4000-5000 tonnes, SLC rill inspections are carried out initially at 300 tonne increments. However, as dilution increases such examinations may occur more frequently, sometimes as often as every 50 tonnes. Repeated practice has shown that during a 5 minute inspection, hanging-wall and footwall dilution can be estimated confidently to within 5%. Simultaneously, fragmentation sizing and brow wear (break-back) factors are also easily recorded. Subsequent dilution draw curves can be used to depict and ultimately predict waste and ore surging. Such data forms the basis for real-time 3 dimensional flow modelling and can aid in delineating remnant pillars and dilution pipes. In the absence of any structured rill sampling programs, cumulative (yearly) SLC metal reconciliations lie at or near 99% of published design estimates. Accordingly, the creation and application of rigorous visual grade estimation protocols at Perseverance have become key to optimal SLC performance.
INTRODUCTION WMC’s Leinster nickel deposits lie ~380 km north of Kalgoorlie, Western Australia and 15 km north of the Leinster townsite. They are confined to south-eastern sections of the celebrated Agnew-Wiluna greenstone belt which is traced locally over a 200 km length and from 5-50 km in width. Acknowledged as the largest komatiite-hosted Ni orebody in the world, the Perseverance Class 1 deposit comprises a main disseminated (~2%) Ni sulphide lode, together with lesser massive (1A and F2) sulphide lenses located at or near the base of a thick (>700 m) ultramafic flow. These are further enveloped by an extensive sheet of weaker, Class 2 (~0.6%) Ni mineralisation. The Perseverance mine operates a large sub-level cave (SLC) and lesser open stopes with annual production rates of ~1.9Mt Its cumulative pre-mined resource was estimated at ~50 MT grading at 2.2%Ni for 1MT of total metal. 1 SLC GRADE CONTROL Current production and development (face-mapping) levels at Perseverance require a 3 person geotechnical crew and coordinating geological assistant for 24 x7, 365 day per year coverage. Shifts are 12 hours long and are based on a 7 day : 7 night: 7 day off roster, for a 21 day cycle. Geotechnicians nominally undergo 6 months prior training at a advanced technical college, receiving a degree in mining geoscience, 646
Figure 1: Aerial view of Perseverance Pit
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Figure 3: Visual calibration estimates vs actual %Ni for a typical geotechnician (0.9% Ni cutoff)
Figure 2: Perspective view of Perseverance Nickel Mine looking northwest while the assistant role is filled by a junior geologist. To facilitate data collection, dedicated underground offices were built at Perseverance for the grade control team with three separate computer stations for geotechnician, geological assistant and a roving duty / senior geologist. Grade control training at Perseverance commences with shift geotechnicians calibrating their visual assessment of sulphide percent against analysed development face and / or rill samples. Repeated comparison of these results permits progressive refinement of visual acuity, ensuring minimal ore:waste misclassification (Figure 3: - 1 error from 42 estimates). Once qualified in visual estimation techniques, geotechnicians progress to rill inspections (Figure 4). Designated crosscuts are accessed via light vehicle with permission from a given production bogger operator. To avoid climbing the rill, a single ~13T ore bucket is left at its base. Prior to inspection, this separate ore parcel and rill are watered down. Visual inspections then follow, requiring a minimum of 10 rocks to be split for lithologic confirmation. Commencing at 300T mining intervals, rill checks increase in direct proportion to waste ingress or extracted tonnes, often to as low as 50T by the end of ring. During an idealised 5 minute inspection, a given drawpoint is examined for 1) Waste type, including hangingwall and footwall percentages 2) Ore and waste fragmentation for 2 m intervals. 3) Brow wear of active ring 4) Hang-up presence by type (ore, waste; high,low) 5) Noteworthy draw characteristics including rill profile. Ore : waste discrimination at Perseverance is facilitated by rock colour and angularity contrast. Lighter and harder hangingwall gneiss and rounded footwall ultramafic are readily differentiated against a darker and more often angular ultramafic ore. Once singled out, traditional Massmin 2004
Figure 4: Shift geotechnician estimating waste ingress during a rill mark-up. percentage estimation charts / diagrams are employed to assess respective waste concentrations.
3 DILUTION MONITORING AND QUALITY ASSURANCE Cross-check inspections using nominal double blind techniques enables verification within and between crews. Significant changes for a given drawpoint within a 24 hour period trigger automatic re-assessment by senior staff. On rare occasions, a bulk sample may be collected for analysis as a final arbiter, however significant differences with primary visual estimates are rare. Sizing by class frequently focuses rill inspections on bimodal (£0.4m and/or >2.0m) waste ingress. Coarser hangingwall material often results in significant hang-ups that exacerbate dilution entry. Distinctive draw characteristics, favouring specific sides of a given rill are symptomatic of such hang-ups. Careful monitoring of coarser size fractions remains a useful aid in both understanding and predicting metal loss (Figure 5). On-line data processing at Perseverance permits automatic grade and dilution draw curve generation for a given drawpoint and ring (Figure 6). With the unpredictability of SLC flow and dilution entry, careful deliberation by all staff, particularly towards the end of design tonnes, is required to reduce or ideally eliminate ore:waste misclassification.
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5 PILLAR RECOVERY AND FLOW MODELLING
Figure 5: Waste size performance bar chart with zone of interest indicated.
Periodic overbogging or recovery of remnant pillar material from multiple levels is a common feature throughout the history of Perseverance’s SLC. Consistent with independent as opposed to interactive draw, some of these tonnages have been extraordinarily large, peaking at >61kT for a single fired ring. While over-reliance on extraction of unclaimed pillars remains highly questionable, simple accounting techniques can be applied to estimate likely tonnages as well as probable locations. In general, ore is readily recovered from up to two overlying levels (~50 vertical metres) by careful bogging and grade control. Especially vigilant extraction has identified markers from greater than 250 vertical metres, retrieved on designated "clean-up levels". Block models of recovered Ni metal (tonnes x grade) for individual rings are also useful indicators of likely recovery positions. In particular, analysis of N-S long-sections across the Perseverance SLC (Figure 7) have proved invaluable when depicting overdraw cyclicity.
Figure 6: Dilution draw curve with zone of interest indicated. Figure 7: Recovered nickel block model in long section with class percentages of reconciled metal vs design. 4 SHUT-OFF GRADES In anticipation of episodic waste piping between rill inspections, particularly at elevated tonnages, increasingly higher shut-off grades are imposed to lessen the potential for significant hangingwall or footwall ingress (Table 1). The tonnage limits used are derived from nominal interactive draw zone (IDZ) geometries, together with the upper limits of blast designs. Additional precautions include lower mark-up tonnages between inspections at specified lower grades (e.g. £50t at or below 1.1%Ni). Such simple precautionary techniques have facilitated creation and preservation of a formal dilution blanket at Perseverance. They have replaced prior practises of attempting to bog through waste or bogging for prolonged periods at or near a lower economic cut-off grade.
6 RECONCILIATION On an annualised basis, metal reconciliation at Perseverance illustrates the benefits of comprehensive geostatistical interpolation, com-paratively homogeneous in-situ mineralisation and meticulous grade controllership. In 2003, the cumulative claim grade (1.93%Ni) was 97% of its mill actual equivalent (2.00%Ni). When combined with their respective tonnages, total claimed and reconciled nickel metal at Perseverance were within 0.01% . This level of accuracy greatly facilitates mining factor generation, which in turn allows for improved mine scheduling and reserve estimation. 7 FUTURE DIRECTIONS
Table 1: Shutoff grades vs tonnage extracted Tonnes
0%- Forecast tonnes
Shut-Off Grade
0.9%Ni
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Forecast>Design Design tonnes tonnes 1.2%Ni
1.4 %Ni
Ongoing automation of grade control techniques are planned for the Perseverance mine which include: 1) Hardware upgrades with the introduction of superior ruggedised tablet PC’s for data collection. 2) Software upgrades with improved links to corporate intranet reporting and site production tracking systems.
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3) Digital image capture of all rills with links to a formal photographic database for further cross validation of grade and fragmentation estimates. 4) Comparative flow modelling using a variety of available software packages (CAVE-SIM, REBOP).
have formed the basis for SLC draw optimisation, significantly impacting on the bottom-line economics of this world class orebody.
8 CONCLUSIONS
The authors are grateful to all their production colleagues at Perseverance and in particular Paul Edsall, Diane Whibley and Reuban Campbell for their extreme care when performing underground geotechnical duties, frequently under trying physical circumstances.
Mature grade control techniques developed at Perseverance embody the collective efforts of geological and mining teams over a period of several years. Visually based systems allow for rapid and flexible estimation, requiring minimal sampling intervention during the extraction process. Extreme care in dilution monitoring, supported by automated data analysis and the implementation of variable shut-off grades has greatly facilitated ore-waste classification. Straightforward bookkeeping together with block modelling of recovered Ni metal have been used in concert to help reclaim pillar remnants. When fully integrated, such rigorous protocols
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ACKNOWLEDGEMENTS
REFERENCES Booth, G.W. Cervoj, K, and Dunham, S.F., 2003. New Frontiers in Research on Magmatic NiS-PGE Mineralisation, 10-14 February, 2003, Centre for Global Metallogeny, School of Earth and Geographical Sciences, Crawley, W.A., pp. 69-72.
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A supplier’s aspect on drilling for mine preparation and production Gunnar Nord, Construction Adviser, Veikko Subanto, Business Line Manager, Atlas Copco Rock Drills AB, Orebro, Sweden
Abstract Being a supplier of a tool like a passenger car, a dishwasher or a drill rig for mining purposes you are always interested in how the customer is making the best use of it. In case the customer is not using the full capacity of the tool the manufacturer of it is losing some incentive for improvement of the tool. The tool shall be like a race horse, always maximising the output and care for its well being. In this paper a supplier’s aspects on how the mine owners are using their drilling equipment is given. The major focus is on utilisation and availability of drill rigs. A minor part is spent on the quality of the drilling with respect to drilling accuracy. A few cases will be described without revealing the identity of the mines.
INTRODUCTION
equipment up to reason-able high figures. The reasons and measures taken are analysed and discussed.)
Each mine is a unique entity. That means that no mine is equal to any other due to differences in a number of conditions. Those are for example the geological and geophysical properties of the ore and parent rock, the depth, the climate, the staffing, local regulations, salaries etc. Still there are some conditions that are similar for most mines and one of those is to get as much as possible out of the investments that have been made. By that is meant that the added capital and running cost for a purchased tool, which might be the investment, should be as low as possible expressed as expenditure per ton ore handled by the tool. The time for activity of the tool is expressed as [Total available time]* [Availability of the tool]* [Utilisation] In discussions with customers the term availability is extensively discussed, while the term utilisation is almost neglected. Whereas the availability is commonly recorded close to 0,9 (90%), the utilisation often reaches 0,3 to 0,6 (30 to 60%). Sometimes it can be expressed much higher. Certainly the utilisation is a matter of the customer and not the supplier. Being an outside observer aware of the great potential for savings by improving the production conditions, the paper conveys experience gained at various mines having reached variable success in squeezing the utilisation up to a figure well beyond 0,6. Utilisation is not the only factor to be considered when aiming at an optimised mine operation. The relationship between excavated waste and excavated ore is also of utmost importance. Not so much for the mining operation itself as for the concentration plant which might have to process excessive waste. The waste factor can be improved by use of the technology that traces the boundaries between ore and parent rock. The geology, the size of the ore body and the mining method coupled with the blasting technique rule the frequency of boulders. Boulders are a pain in the mining activities as they block extraction points and ore chutes. Dismissal of the boulders is therefore recommended as early as possible in the process. Drilling accuracy is discussed here. A better drilling accuracy will bring the amount of boulders down. (This paper presents in general terms a number of mine projects having different success in raising the utilisation of 650
Utilisation and Availability The term utilisation needs to be defined before the discussion can start. The most obvious definition is as follows. Here all hours available is considered and that means 24 * 365 = 8760 hours per year. Utilisation is calculated by the following Utilised Hours Utilisation % = ------------------------------------------- x 100% Available Hours - Downtime Hours Utilised Hours Available Hours Downtime Hours*
= Time the drill rig is in operation = Days in Month x 24 hours = Planned Service + Repairs + Breakdowns
* Not included is hours spent on repairing damage i.e. rock falls, misuse and abuse. In many mines the utilisation is calculated on the time the miners are really employed in the production process omitting the time for lunch breaks, transport to the rigs etc. . This means that the available hours get considerably less than 24 hrs per day. As mine production is best governed by a steady and continues flow of ore to the concentrator to minimise capital tied up in intermediate storages, all 168 hours per week have been considered as total available time. In the automotive industry this minimising of storages is called "just in time production". There is nothing new or unique in discussing the utilisation of the mining equipment, but the aspects will differ depending on who is considering the issue. Problems will arise when the utilisation is unsatisfactory. Investing in new equipment could be an option in order to increase the utilisation. It should be stated however that even a supplier of drilling equipment realises that the drilling equipment does not represent the most costly item in the total process. In most cases it is the investment and running costs of the
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concentrator and mine infrastructure that is the ruling factor. The mining activities have to follow the budgeted flow of ore. This means i.e. that for an established production only 2,4 drill rigs (with a rated utilisation of 40%) will be needed. Here there are two options: either to employ three rigs or to raise the utilisation to 48%. In the latter case the risk exposure is higher as a major failure of one rig may cause a severe reduction of the flow of ore for one or two days. This example is picked to clearly show that improvement of the utilisation has effect on the economy. In many cases it is not just the cost for the rig that is involved, as the crews may be allocated to the equipment. Consequently this also results in low utilisation of the drilling staff. Utilisation is discussed below for a number of actual mine cases. The mines have on purpose not been designated. Mine production with Boomers An example on the monthly fluctuation of the utilisation is shown in the figure below. In this case the mining company used storages to ensure a constant flow to the concentrator. The drilling tool in this case is a two boom Boomer used for stoping and face drilling.
By arranging service contracts with a supplier the availability as well as the utilisation is now rising for the drill rigs and is presently running at xx%. It should be obvious to the reader to realise that in most cases it is much easier to boost the output of a rig by better discipline in the mine and thus push the utilisation up instead of going after the few percent possible to gain on availability. Long hole drilling for mine production Production drilling by use of the so called long hole drilling technology is in principal facing the same problems as face drilling by struggling to maintain high utilisation and availability. The blast holes are normally much longer and larger compared to those for face drilling. There is also more drilling from the set up and moving between the set ups is often only a couple of meter. As the penetration rate is considerably lower it would here be possible to achieve a higher utilisation. But is this always the case? It has been pointed out above that the number of rigs will never exactly match the demanded output. If the demand for blast hole production just exceeds the capacity of one long hole drilling rig with availability 0,9 and an utilisation of the same magnitude (0,9) the installation of the second rig means that utilisation will drop to 0,45. If the demand for rigs increase from 2 to 3, the drop will be down to 0,60 and from 3 to 4 the drop will be down to 0,675. The message with this discussion is that no matter that you possess the skill to run the production rig at 90% utilisation, the actual figure will be lower as the demanded production only rarely matches the capacity of an even number of rigs. Figure 2 supports this discussion.
Figure 1 Utilisation of a two boom Boomer for mine production year 2002 to 2003.
For the mine production activity two Boomers were used and they had very similar utilisation on a month by month basis (see Figure 1 above). The average utilisation over 24 months for the two rigs was not more than 40 percent and with an availability of 94% there might be a chance to reduce the number of rigs in production from two to one. However it is unrealistic, as the mine might never expose itself to a situation relying one Boomer only. A major breakdown of a single rig would cause a complete stop of the ore flow. This mine is located in a high salary region and the miners are not engaged in the drill rigs only. They will participate in other mining activities and consequently no salary costs are tied to the idle time of the rigs. In this case there is not much to do to improve the situation for the utilisation. Another mine located in a low salary region differed very much from the above case with respect to utilisation and availability but had fairly similar orebody geometry. The availability of the drill rigs was low and instead of squeezing that percentage up, the mine management acquired more drilling units. The mine became owner to an excessive fleet of drill rigs, whereof many could not be used due to shortage of parts. This resulted in lack of priority in making each drill unit running. The product of availability and utilisation was consequently as low as 15 to 20%. Massmin 2004
Figure 2. Schematic evaluation of what can be reasonable utilisations depending on number of drill rigs employed In reality the actual situation is not as simple as these calculations show, as there is often not just one type of rig available on the market. Therefore the worst case as described is unlikely to occur. Furthermore, the mine might consist of a number of orebodies spread in such a way that transport between different sites will be very time consuming. Therefore it could prove profitable to leave the rig idle for some percentage of the available time. There are examples of mines that seem to be properly operated and their figures match the figure above. Two such mining operations are shown below. In both cases there are production rigs located in mines with large ore bodies which means that transport of the drill rigs constitute only a very small share of the time. In the second case the estimated transport time is added on top of the bars (Figure 4).
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In both cases the availability is typically 95%.
Figure 5 Case, utilisation of a Simba based on working all available time (8760 hrs/year) Figure 3 Utilisation of three Simbas for one year of production. Month 37 represents the average for all 36 months.
cases used in this presentation which are picked randomly indicate that availability is not a factor that will have a major impact on the production output. Occasionally, the availability can be low but over the year the figures rarely go below 90%. These are commonly in the range of 90 to 95%. In two of these mine cases having reliable figures, the utilisation has been plotted versus the availability. The result is exposed in figure 6 and 7 below.
Fig 4 Exceptional good utilisation performed by a highly automated Simba working in a large scale mine. The red part of the bars covers transition to a new position. The drops in July are due to vacation time. Month 25 represents the average for the 24 months. Other examples show less successful records and one is exposed below. It should be stated that the mine on purpose only uses some eighty percent of all available time. But even so taking this reduced input of working time the utilisation is extremely low. The mine management is aware of the fact that they have too many Simbas employed (see figure 5). Consequently the figures should be adjusted upwards on average by some 2 to 3 percent. The drill rigs are also employed for drilling cable bolting holes and that type of drilling is not included in the bar-chart. By including this activity the average utilisation will raise from 28 to 33 percent. To some extent there is a self-explanatory reason for the low figures and that is that there are long distances between the extraction sites. The mine is fairly spread out and fewer rigs would mean more transport of the drilling units. The availability of the rigs is in the magnitude of 95% on a yearly basis. The monthly fluctuation is only marginal. What conclusions can be drawn from these bar-chart illustrations? The utilisation can be very high especially when dealing with large orebodies where transports are short. For production drilling using Simbas, the utilisation can run as high as 85%. In mines with minor and scattered ore bodies the utilisation will certainly be much lower. Figures below 50% are rarely possible to explain by other reasons than lack of focus from the mine management. Being a supplier we are often faced with accusations that low production is caused by unavailability of the drill-rigs due to shortage of parts or poor maintenance. The mine 652
Figure 6 a, b Case 1 Availability and utilisation monthly results from mine production with Boomer plotted in the two figures above.
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the side rock as untouched as possible. Therefore it is important to locate the holes where they are meant to be. The first question is then: "do I know where I want the holes to be?" The answer to this question would in general be "yes". However, if we give the ques-tion a second thought the answer may instead be "not always".
Figure 7 a, b Case 2 Availability and utilisation monthly results from production drilling with Simba plotted in the two figures above. It is obvious that there is a very poor correlation between availability and utilisation at least when utilisation is as high as 90% or more. Possibly it can be said that the availability goes up when the utilisation goes down but the correlation is weak. The conclusion from the presentation of the results above is that generally the availability for drill rigs both for mine production and mine preparation is in the range 90 to 95%. These figures are valid also for older rigs being properly maintained all the time from start up. Between 5 and 10% is consumed by maintenance whereof roughly half is preventive maintenance. It is difficult to cut this figure as it represents only 8 hours. What is left is only 0 to 5% and this time will be consumed by randomly occurring faults during operation. Improvements can certainly be made, but there are no major savings of time to be made by claiming a higher availability. In the exceptionally well run mines shown above it is not possible to squeeze more production hours out of the rigs. However there are a lot of mine cases where a better planning discipline would bring more production hours to the mine owner. SUMMARY A summing up of the discussion above will be done. Analysis of a number randomly picked mine cases has been made with respect to availability and utilisation based on monthly recording. In almost all cases the availability was in the range 90 to95%. In all the cases so called preventive maintenance was practised. It consumes half of the 5 to 10% down time due to planned service and repair leaving only some 3 to 5% of total for unpredicted repair. Certainly the availability can be squeezed to a higher percentage but this is judged to be very costly. Considering the utilisation of the mine cases the issue is more complex. Large ore bodies where the production calls for multiple long hole drilling units makes it possible to reach high utilisation figures like typically 80%. For smaller and scattered orebodies it is much more difficult to reach high utilisation figures which seems to be more difficult in face drilling than in long hole drilling. The real low figures as presented above seems difficult to explain. Figures below 50% are hard to defend unless special cultural habits on working time make it difficult to maintain a 168 hour working time per week. It is obvious that mine management that is looking for better produc-tion output from his drilling units should start to look for a better utilisation Accuracy in drilling Drilling for mine production and mine preparation is not just a production issue. It is also a quality matter. The holes are meant for explosives, used to excavate ore and leave Massmin 2004
Measure While Drilling In many cases the boundaries of the ore body are not established in detail. The available ore geometry information is often limited and incomplete, based on sparse diamond core drilling and/or available geological mapping. It is a fact that the less information available, the simpler and smother the ore geometry and ore boundaries appear to be. From production and planning point-of-view, uncomplicated ore maps are often very appreciated, and therefore not often debated or questioned. However, the major problem with such maps is that it does not coincide with reality. This type of insufficient agree-ment between reality and the maps used for mine- and production planning is rarely identified or evaluated since production targets (tons) as well as mining layout geometries still can be met. Nevertheless, it still means that an undefined amount of money is spent in absent incomes and higher costs. One solution to this problem is of course better and more detailed geological data as basis for ore maps and production planning. This can be achieved only by denser sampling of data, but the traditio-nal way to extract data, by diamond core drilling, is expensive and will therefore hardly be used for massive sampling of ore bodies. Methods to extract information directly from production drilling would exclude the cost, for drilling the hole, and reduce the sample cost to a minimum. A technique to extract rock properties while drilling is called MWD and it stands for "Measure While Drilling". This technique is a method for collecting data during production drilling. All main parameters influencing the drilling process, such as penetration rate, feed force, percussive pressure, rotation speed, etc, is recorded at predetermined intervals along the bore hole. With an adequate analyse rock mass quality parameters, such as rock hardness and rock fracturing can be calculated, based on monitored raw data. It has to be empha-sised that a prerequisite for successful monitoring is that the waste rock and ore have different drilling characteristics. Major advantages using this technology are: • Very high data resolution since data is extracted with a few cm interval in all production holes. • Very low cost since monitoring of data is conduc-ted automatically during normal produc-tion drilling. • Very low data risk, since monitoring is performed during the drilling of the hole, and no instruments have to be inserted after drilling is completed. • Minimal disturbance of production. MWD will require very limited extra work for the operator. However, many may object that the lay-out of the mining area, stops, drifts etc. is based on other things such as ore flow, rock mechanics, transportations, machine dimensions etc. and cannot be changed or modified based on last minutes information, which is the case for MWD. This is of cause true to some extent, but we must not forget that this information, even if it arrives in the latest moment, provide information on the economical outcome of the stope or even a single round. The excavation cost must always be exceeded by the ore value in order to make a profit. If the ore value is low there is always an option not to blast a stope, a round or even some holes in a round. The drilling cost is always neglectible in relation to the total handling cost of a blasted round. Also other changes can be made to
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optimise the outcome of the mining operation, see figures below.
Figure 8 The ore expires a few meters beyond the face, and continues to the right. Would you still continue to reinforce the drift according to the initial mine lay-out?.
interpretation is not always a strait forward process when there is only a marginal geological deviation between ore and not ore. There are though many cases where the ore boundaries are more distinct and with relative good accuracy can be detected by use of the MWD system. It is surprising that so few mining companies have shown such a limited interest in this technology when investing in new drillrigs. No estimates have been made on what the savings might be if the excavation is done to plus minus 0,2 m instead plus minus 1 meter. However it is believed that the cost-savings should be in the range of 5% when considering concentration process as well. This is also money to go after. Collaring errors and hole deviation Above a discussion has been given on the correct holedepth in order to minimise losses due to over-break and under-break of the ore along its boundary with the parent rock. It is not just the length of the hole that is of interest but also accurate the hole is drilled. There is nothing new in this statement but it is worth repeating as there is loss of much money due to poor drilling accuracy. This statement is valid for both long hole production drilling and face drilling. It is somewhat surprising to notice the little interest that is paid by most mine management to upgrade their boomers when old rigs shall be replaced. There are mines that go for latest technology when they invest in new equipment and one example will be given. In the Garpenberg mine in the middle of Sweden two old face drilling rigs used in face drilling and stoping. They were of the type called DCS which means direct controlled system and the boom movement are controlled by manual opening and closing of the hydraulic valves. The new rigs that replaced the old one were of RCS type with CAN bus ruled valve activity giving a great advantages when guiding the drill steel to the correct collaring position as well as alignment of the feed. The achievement of the new rigs is exposed in the table below. • Drift size 5,5*5 m and drilled depth is 4,05 m • Pull used to be 3,4-3,5 m but has been increased to 3,9 m • Over-break reduced from 20% to 9% (which means 11 cm on the radius or 3 m3/meter of tunnel)
Figure 9 A narrow vein type, complex copper-zinc ore, defined by higher hardness than the side rock. Is the thickness of the ore enough to cover the excavation cost? Is it possible to reduce the size of the drift to reduce development- and side rock handling cost?
Another possibility with MWD technique would be to stop the drilling as soon as the boundaries of the ore-body have been passed by the drill bit, without prior information on the ore geometry. This, of course, means that the drill rig must be equipped for on-line evaluation of rock properties and immediate identification of the ore boundary when crossing it. Atlas Copco can offer MWD options for many rig types. Monitored raw data is handled, analysed and presented in one of our software packages; Tunnel Manager for Boomer rigs, ROC Manager for the ROC product line and Ore Manager for Simba rigs. In the software’s raw data can be analysed and rock mass properties such as rock hardness and rock fracturing can be calculated. It is also possible to generate customised defined analyses based on monitored raw data. In the figure above MWD data have been analysed and presented in the software Tunnel Manager Pro. The MWD system has been on the market for a number of years. The collection of drill data is easily done when using modern drillrigs. The information is almost free of charge once the investment in the rig is done. For sure the 654
Figure 10 Elements of hole deviation
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What is the explanation to this improvement? It is believed that the whole explanation is in a much more correct positioning of the holes. It is believed that most readers will realise the profit from the new rigs. In this mining case with short holes only 4 meters the faulty drilling is mainly because of poor collaring faulty alignment as already mentioned above. When holes are longer like in long hole drilling with Simbas the in hole deviation will play a much more important part for the total fault of the hole location. What is meant by in hole deviation is explained in the figure below, which actually refers to bench drilling. The principals for hole deviation in the different types of drilling are though the same.
Figure 11 In-hole deviations due geological foliation
Collaring misalignment and in-hole deviation for long production holes play a major role for the blasting success. By success is here meant proper fragmentation at a reasonable input of drilling and explosives. Certainly the drilling equipment can be improved but the owners and operators using the drilling equipment can also contribute to a better drilling accuracy. A number of reasons contributing to poor accuracy have been observed but they are difficult to quantify. A too high feed-force has clear tendency to make the bit deviate. Some drillers still believe that a high feed-force will give better penetration. If any difference it is only marginal and what can gained is lost several times due to the increased amount of drill meters needed to compensate for poor precision. Bonus for the drillers linked to drill meters per shift is a hazard from quality point of view. In down drilling the feed-pressure is far from always correctly compensated for the weight of the drill string due to incorrect hold back force. This mostly results in a too high pressure in the bit and consequently a large deviation of the hole. Geology will also contribute to the deviation. In picture 11 it is obvious that a systematic error has occurred as all the drilled holes deviate in the same manner. The explanation is here believed to be the foliation of the rock mass. It is known that bedding planes, foliation and joint-sets will affect the drilling direction as is shown in figure below. Summing up; "accuracy of drilling" As supplier of rock drilling equipment we are fully aware of that the long hole drill rigs need improvement so that the blast holes can be drilled more accurately. Every step forward in the development of the rigs will mean savings on total drill meters as well as savings of explosives. Dealing with boulder is another problem that will be less costly for the mine owner. Major savings can also be made from reduction of over break which mostly will be run through the concentrator and only generate costs. But there are means and methods available today that can help the mine owner to achieve better quality of the drilling. It is believed that it is worthwhile to take a closer look at what is available offering help.
It is obvious that for longer holes that the collaring offset is playing a much smaller role than in the Garpenberg face drilling case above.
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Chapter 18
Automation in Mass Mining
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Supplier as solution provider for the mining industry, Sandvik Mining and Construction vision of the future in mining Arto Metsänen, President, Sandvik Tamrock, Finland
Abstract The mining industry has been through several years of consolidation, which has lead to fewer and larger mining houses. The competition continues to be hard and the long-tern sustainability is weighed against skills to survive through the rapidly changing economic cycles. Several massive underground mines are in the pipeline for the coming years. Investments made are carefully viewed for their lifetime rate of return. In many case this has changed the role of a mining equipment and service supplier to that of a long-term partner with similar business views and working synergies with the mining company. Sandvik Mining and Construction, SMC, sees its role in adding value and enhancing the profitability of our customers business.
1 INTRODUCTION We are experiencing an up-cycle in the need of commodities. The demand of various metals and minerals, aggregates and building materials is high, by far thanks to the economical growth in the Asian and specifically Chinese markets. The mines expand their production and new projects are being opened and it is the high-life for all in the industry. The natural follow-up of the up-trend is down trend, the length and level of the growth in China remains to be seen. Even without China, an increasing need of coal, metals and minerals will continue in Asia and those new renewing economies in other parts of the world. The uncertainty of the political climate seems to be here to stay and it will keep on affecting the economics and how business is done. The commodity prices and world demand continue going through the rapid changes and sustainability with in-built flexibility continues to be important. One of the competencies in the game of survival is the level of rate of return of the capital invested, the efficiency in which the available resources are utilised. High resource utilisation is a starting platform when lower operational costs with higher rate of return are the target. The trends both in surface and underground mining are towards optimised operation, lower costs and maximised output. Surface mines will look for means of optimising further the ore to waste stripping ratio and we will see deeper pits in the future. Underground mines are established for deeper deposits and increasingly complex environment. The scale of underground massive mining has grown to proportions that a few decades ago would have been unthinkable. Managing risk has a whole new level of meaning in setting up massive caving operations and in ensuring that all operational targets are met. When the large-scale underground mining operations are set to compete with the scale and cost of open pit mining, there is a need to create an environment, which has a high level of controllability. There is less room for excess capacity built into the system. The capital expenditures are carefully reviewed and suppliers selected in long-term partnership in mind. It is hard to Massmin 2004
see a future large-scale block caving operation to manage without comprehensive monitoring and process control system and, at some level of automation. This sets also new demands towards the suppliers of the technology and services for the operations. Sandvik Mining and Construction business has lived by the industry upturn and downturns. Much of our success has been born out of the needs and shared visions with the mining industry. Sandvik Mining and Construction sales in the year 2003 was 2 billion USD, one third of the mother company Sandvik Corporation. Sandvik Tamrock, Sandvik Materials Handling and Sandvik Rock Processing form the operational base for SMC. The importance of service and support can be seen in the distribution of sales where 67 % is coming from after sales and services and 33% from equipment and project sales. 2 KNOW HOW -KNOW WHY We define ourselves through the value we add to the surface and underground mining industry. This signifies the strong focus on ability to combine the understanding of
Figure 1: Continuous competence building is part of the SMC culture.
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Figure 2: Technical and project management knowledge is needed for quality service and maintenance offering.
Figure 3: Surface mining at Freeport Grasberg mine in Indonesia.
mining processes with the technology know how. Learning and innovative thinking continue to be important factors in making the mission of improving our customers business real. The focus is in the care and development of human resources by attracting professionals and supporting learning. Sandvik invests 4% back into research and development. We need to continue to invest into research and development to maintain our position at the leading edge of the core business areas. Surface and underground mining understanding is the key competency in SMC. We have a large number of mining engineers in our company globally today and we hope to attract even more mining professionals in the future. There is a continuously growing emphasis on training, service and maintenance in the mines. The success in service business is built in mastering the operating system. The skills of the service provider need to be in technology, logistics, scheduling and project management. The outsourced service function continues to be a part of the daily operation. It is a real test for working together and showing the value added every day.
secondary breaking solutions which meet the challenging operational and process needs. Sandvik Mining and Construction offers a wide range of underground mining technology. Our knowledge in surface mining has much focused on drilling and loosening of rock. The surface mining trends suggest that more emphasis will be put on managing the mine to mill concept as a whole process. Communition is a high energy and high cost part of a mining process. On the other hand, the beginning part of loosening rock, drilling and blasting, represents a fairly low cost low energy need portion of the process. What is being produced and where are the critical issues bear the notion of optimisation of the full chain, but mastering the fragmentation distribution at the beginning of the process will save costs in the whole process. Sandvik Rock Processing is a part of SMC and provides solutions for comminution of rock to sizeable material for milling, and thus completes the process from loosening rock to crushing.
3 WORKING TOGETHER WITH THE INDUSTRY To work together with the mining companies in a partnership on a global level gives reassurance in matching the mining company needs with solutions. The need of creating sustainability in mining in the years ahead will see further development of methods and technology. The research done in the ICS group has demonstrated that joint efforts and close relationships within the industry do support quality solutions in difficult issues. Since the operation eventually will demand technology and support systems, the understanding of what the industry expectations are is of utmost importance. We have been proud participants in the ICS and HSBM projects and look forward to participating in the MMT, Mass Mining Technology projects as well. Technology is in a key role in maximising mine output and the solution need to support the desired outcome. Much of SMC offering has been initiated from in-depth relationships with global mining companies. The work in partnership is essential when supplying systems that are integrated to the mining process. Mine automation system AutoMine‘ could not have evolved without partners like LKAB, DeBeers and Codelco. Fragmentation management and the issues round hang-ups, especially in underground block cave mines are one example of partnershiping. The work done together with Rio Tinto Palabora mine and Freeport DOZ mine has set the scenes for developing 660
4 SAFETY AND ENVIRONMENT The mine safety and environmental focus will become even sharper in the coming years. Not only are the deeper deposits setting rules for more demanding operational conditions, there is a more social and public need to continue to demonstrate the safety and environmental security of mining. The rule both for mines and suppliers is that safety can not be compromised. The legislation on safety and ergonomics pushes the cost of operation higher and new technologies and ways or working are established. Added automation is one way of coping with increased demands for safer and more controlled operations. Ventilation and refrigeration is a high-cost element in deep underground mines. The solutions lay in automation and non-emission approach with fuel-cell technology. Mining machinery both in surface and underground mining are viewed as high safety risk areas. This is why suppliers must have a strong role in ensuring that the right working procedures are followed in operating and maintaining the equipment. Close relationship between the supplier and mining company also ensure the future development of equipment safety and ergonomy. 5 CONCLUSIONS SMC business lives strongly with the industry upturn and downturn cycles. Like the mining industry itself, we also are searching for means to master this cyclical business environment. The consolidation within the mining technology suppliers has not been as dramatical as within
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the mining companies, merely because there has been fever players to start with. The move has been from being a supplier towards being a partner in value adding and risk sharing. Here the process understanding within comprehensive process offering and support presence sets the rules for future survival. Our successes are born out of the needs and shared visions with the mining industry. SMC wants to be the value adding solution and technology provider for the surface and underground mining industry. In massive mining the need is to be the lowest cost operator by moving large amounts of material in a continuous manner. The mining profitability is set to apply in the changing world economics when, for example, the copper price moves from 2400 USD to 1300 USD per ton.
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There is a need to create an operational environment that is ruled by predictability and proactive way of working. Machine data collection systems are available already today, but the ability to utilise the collected data as information for future decision making are becoming available. More machines are manufactured with in-built intelligence to cope with the new demands. The decreasing human involvement in mines will eventually lead to developing mining methods and processes that will be more suitable for automation. The future layout can be better fitted to suit the automated mining needs. The discussions between mining companies and machine suppliers will deepen as the new technology penetrates deeper into the mining process. There is a need to line up further according to our customers business.
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Mobile equipment resource management and its optimisation in an underground process Tuula Puhakka, Senior Advisor, Mining Business and New Offering Development, Sandvik Tamrock, Finland
Abstract The repeatability of work tasks and processes in an underground mine is traditionally low. In a study interview, mining professionals felt that the repeatability of a mineral processing system is between 75 and 100% whereas the repeatability in an underground process was considered to be 50 to 75%. There is a common understanding that too little attention is paid to resource management. Mining companies are becoming aware of the improvement possibilities through improved resource management, which looks at the mining processes in a holistic manner. The distinction of utilisation is understood often in various ways, which easily leads to the misconception that all utilised time posses the same quality of the use. Continuous improvement projects direct the change from reactive to proactive focusing on gaining more data and understanding on what actually happens in the processes. The current trend on operational optimisation and improved resource management also works towards improved ways of working and higher rate of return. When the repeatability has been established, a process environment has been created and outputs of automation are more reliable. This paper discusses the current status of the mobile resource use and management and ways to obtain repeatability in their operational use.
1 INTRODUCTION The repeatability of work tasks and processes in an underground mine is traditionally low. The underground mines are the ‘rock-ore harvesting sites’ for the mineral processing plants. The large variations in a task outcome have been explained by the harsh and uncontrollable underground mine condition, challenging logistic arrangements and ‘desired individualism’. In a mining industry directed study interview, mining and supplier professionals we presented with questions on repeatability and output estimations of various mining processes and automation. The need of process understanding and reliable operational knowledge, to be used for streamlining the underground processes for automation purposes became imminent through several mine automation projects during 90’s. The Finnish IMI, Intelligent Mine Implementation project and MAP, Mine Automation Project, both discuss the need of added process visibility. It was understood that successful automation of underground processes needed well-described process descriptions, rethinking of current mining processes and advanced decision making and support systems. The survey was done to add knowledge on how the industry it self viewed the repeatability and out-put of underground mining process and the management of the resources. The survey also had questions on mine automation expectations and on KPI’s for performance measuring. 2 RESOURCE MANAGEMENT A resource management and mine work and process repeatability related questions were presented to 30 people in the industry. The target group consisted of people on management level on operations, mine design and maintenance sectors. The questions were presented as statements, which the person could either agree or disagree 662
with on a scale from 1 to 5, where 1 equals strongly disagree and 5 strongly agree. 20 returned answers (total of 24 returned) were used to generate these results. The mine professionals indicated that too little attention is paid to resource management (Figure 1, E and F).
Figure 1: One result of a survey directed to mining industry representatives on resource management. 1=strongly disagree, 2=disagree, 3=neither disagree nor agree, 4=agree and 5=strongly agree. The questions were as follows: A. If the production lacks behind the first option is to add resources B. Mines optimize their material flow poorly before commencing mining C. Mines pay too little attention to resource management D. Mines understand what is the optimised number of resources needed E. All the mine resources are well utilised
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F. Low resource utilisation is the result of insufficient management The questions did not go deeper into the details of the lower utilisation in resources. The practical knowledge of operations indicate that many of the reasons are related to mine logistics (preparation of the site and availability of services), availability and compatibility of mining resources and training to utilise them. In the question on acceptable KPI’s for operational efficiency in an underground mine, the total resource cost per utilised % or utilised hour, the resource utilisation % and tones produced per time unit received of the highest marks among the responses. The distinction of utilisation is understood often in various ways, which easily leads to the misconception that all utilised time posses the same quality of the use. For example in a mismatch situation of loader and truck operations neither of the resources are quality utilised even if the utilisation rates are high. The tones produced in given resource utilised time are different in match and mismatch situation and the rate of return for capital employed changes. Resource management dictates that concept of right place, right time and right work description are valid simultaneously for the given resource. The acceptable loader utilisation was considered as 50 to 75% whereas the acceptable crusher utilisation was 75 to 100%. By accepting lower machine utilisation rate the mine accepts the lower rate of return for invested resource capital. The large variation in some utilisation rate acceptance indicates that also the lower repeatability is imminent.
The information turned into knowledge-system generates plans for executing needed corrective actions in the process or with the resources and suggests possible improvement areas for optimised outcome. The current mine information of tends to flow differently in different parts of the underground process. The information flows within a task, less between the tasks and less with the total process flow. This supports better task optimisation than the total process optimisation. The interview of current system outputs stated that the repeatability of a mineral processing and hoisting systems was considered to be between 75% and 100% (Figure 3). The output of an underground process of maximum available was rated down to 50 to 75% with some answers coming back as low as 5%. Repeatability of one individual task, long-hole drilling, was estimated between 25 and 100%. The individual machine performance was considered to be 50 to 75% of their capability.
3 SYSTEM REPEATABILITY The development of equipment is ongoing with implanted intelligence and readiness for improved communication systems that withstand the underground environment are setting the basis for on-line process monitoring. Massive amounts of machine functional data can be collected already today. The challenge is to find the relevant data to be turned into information and further to knowledge, which then could be used for value-added purposes. There is a limited amount of software available to be used for holistic resource management in an underground mining operation. Such a system or systems software will have relevant data turned to information on all process relevant resources (people, machines, material, ore), the status of the resources and the system (performance, condition, location) the current and expected out-put of resources and the system (related to planned).
Process flow-material flow and information flow
Information flow
Tasks performed
Sub-processes D
Information flow
D
Total UG-process
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PL RR
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ore flow
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Figure 3: Process output estimation. The questions: A, The repeatability of mineral processing system is % B, The repeatability of hoisting system is % C, Underground mine out-put is today how many % of possible maximum D, The repeatability in total underground process is % E, The repeatability of long-hole drilling out put is % In a low repeatability-works or does not work situationsystem the system optimisation for best output becomes an impossible task without a deeper look at the reasons behind the non-repeatability. Typical challenges of this type are related to mine maintenance and service process and ability to tie them up with operations. The end result is a continuous work rescheduling. A resource comes with a cost and has more relevance when utilised to its fullest potential. The high marks KPI, which is measuring resource cost per utilised % or ton also indicates the desire to receive a higher return for investment made. 4 WORKING TOWARDS HIGHER RATE OF RETUN IN INVESTED CAPITAL The actual working procedures in mining have not significantly changed since the onset of mechanisation. Added cost deduction in mining has rather been a result of increased power and more material being moved than educated and calculated methods of improving the rate of return of resources invested. Detailed description of working procedures and practices are needed as well as linking them with one another in
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adjusting technologies together. Management level interest in controlling the mining processes has increased with improved information technology and developments in systems engineering. Mining companies are now more aware of the improvement possibilities through improved resource management. The objectives include improved quality, reduced cost, decreased cycle times, increased flexibility and improved efficiency together with better social image.
B, Mine automation needs a process environment to work C, Mine automation will process mining D, Automation allows for fewer mistakes in the system E, Mobile equipment are the weakest link in mine automation There is an understanding that higher system repeatability needs to be established to gain all of the expected the benefits of automation (Figure 5). The expectation is that underground mining will become a process through automation 5 CONCLUSIONS
Figure 4: Mining process control approach; IMI, Intelligent Mine Implementation project. If in deed the repeatability of underground process is between 50% and 75% the possibilities of improving output of any existing mine are huge. Any random approach will not meet the desired outcome of taking the remaining 25 to 50%. There is a need to collect data systematically from the resources and ways of working in the current process and turn that into information, which can be used for actual operational process description (Figure 4). The collected data when turned into information reveals the realism in current operation and helps to understand the reasons of low repeatability. A system that has no acceptable measurements can not be identified, controlled and further optimised.
A system that at any given time can give any given result has little possibilities for guidance and control. The underground system has not kept up with the technology change. The improvements in the output have been directly related to individual machine efficiency increase and less to improvements in processes. The technology has been inbuilt on top of existing systems inheriting all the previous methods and ways of working. Introduction of automation underground has been an eye opener for possibility of improving current out put through processing and systems engineering. The simplest systems are easiest to automate. The coming years will see more massive mining systems with large mobile resource fleets and sophisticated process solutions. A massive block cave operating with 50 loaders will be based on automation and well-established mine-wide process control and resource management system. The smaller and bigger mines with cyclical operations will gain major benefits through monitoring and improved resource management systems. The lower process repeatability in the underground system and unused resource potential clearly describes the opportunities for better rate of return. The current trend on operational optimisation and improved resource management works towards improved ways of working and higher rate of return. When the repeatability has been established, a process environment has been created and outputs of automation are more reliable. Increased level of repeatability has also a positive impact to mine safety. The change from reactive to proactive focuses on gaining more relevant data and creating a system understanding on what actually happens and what needs to happen in the processes. The resource management lack of reliable and timely information for intelligent decision making purposes, is seen as the main reason to inability to achieve higher repeatability.
REFERENCES
Figure 5. Inputs on automation in a mine. 1=strongly disagree, 2=disagree, 3=neither disagree nor agree, 4=agree and 5= strongly agree. The questions: A, Repeatability is needed in underground mining before automation can be applied fully 664
• Baiden GR., Strom RE., Preston CJ., 1997. Mining automation program, CIM Bulletin 90: 71-77 • Baiden GR., Future robotic mining at INCO limited1996.The next 25 years, CIM Bulletin 89:36-40 • Sarkka, P., Liimatainen, J., Pukkila, J., 2000. Intelligent mine implementation-realization of a vision, CIM Bulletin 93: 85-88 • Puhakka, T., Särkkä, P., 2003, 'Process and System Improvements in Underground Mining', World Mining Congress, New Delhi 1-5.11.2003 • Puhakka,T., 2000, "Optimising Mobile Equipment Resources in Massive Mining," MassMin 2000, Oct. 30Nov. 2 Brisbane • Raimonaho, J. 2001. Mine-wide communication system, Proceedings of the 4th Symposium on Computer Applications in Mineral Industries:181-189
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Rapid LHD advance using laser guidance and 3D vision systems for block-cave mining applications Roy Jakola, President, Automated Mining Systems, a Division of MD Robotics Robert Ward, PEng, Senior Research Engineer, Leif Bloomquist, Research Engineer, Ken Martin, Manager, Mining Applications, MD Robotics
Abstract Traditionally, LHD’s have been operated by miners who are caught in a restrictive cycle which depends on availability of work faces and on geological and survey guidance. This paper describes the capabilities that Automated Mining Systems (AMS), www.automatedmining.ca, has developed in the fields of underground vehicle auto-guidance and 3D imaging technologies. The use of laser guidance, and advanced driving and steering algorithms in a safety critical fault tolerant architecture has resulted in a safe high-performance automation and guidance system for underground vehicles. The evolution to autonomous systems utilizing a 3D vision system for real time control while simultaneously creating photo-realistic 3D models of the geology and surveying of the advancing mining faces is introduced. This evolution towards rapid advance and flexibility within the existing process will add a new level of value to the autonomous mucking cycle in block cave applications.
1 A BUSINESS CASE FOR CREATING AUTONOMOUS CYCLES IN BLOCK-CAVE MINING Underground mining brings with it unique challenges. It is a materials handling exercise interrupted by series of roadblocks and bottlenecks, in a dynamically changing environment. Block-cave mining epitomizes these challenges as loading patterns change often; there is frequent inability to remove material from the loading points to surface and often a lack of effective coordination of services to support the process. In essence, underground mining must succeed in an environment where challenges are frequent, and significant improvements to process are not easy to find. One area where process improvement is still possible is LHD utilization. Mine production reports continue to show that an 8-hour operating shift turns into 4 1/2 hours of effective LHD operation. There are many reasons for this, but despite many efforts being made to increase this number, it seems to defy significant improvement with conventional mining techniques. Mines have experimented with longer shifts, only to learn that the ratio of hours worked to length of shift is not significantly increased. One exception in favour of 10-12 hour shifts occurs when travel time to and from work faces is excessive. Even in this case, however, operators become fatigued and productivity tends to falls away over time. In a recent business case study of an underground block-cave application, creating an autonomous muck cycle for 3 LHD’s showed a 50% improvement in operating time per shift, and a corresponding increase in material moved. Autonomous LHD cycles allow machines to operate for longer periods of the work shift than traditional methods with operators in the work area. Recent advances in automation technology by AMS also provide the flexibility necessary to successfully automate LHD cycles within a block-cave mine. Vehicle routes and mucking patterns can now be changed quickly and as often as necessary to meet changing mucking patterns. This level of flexibility was not possible Massmin 2004
with the previous generation of automation systems relying on fixed guidance media. Further advances have been demonstrated using a new 3D imaging technology to record the geology and take survey volumes and direction while an LHD is at the work face. It will soon be possible to record geology and survey data without human intervention during an automated mucking cycle because of advances AMS’ parent company, M D Robotics, has made in the field of 3D imaging. When fully implemented, geological and survey data will be integrated with mine management databases to allow accurate data to be presented in real-time to technical and management personnel. This will represent an unprecedented degree of flexibility and allow important process decisions to be made faster and with better background data. The underground mining industry is therefore at the threshold of benefiting enormously from autonomous mining/geology/survey work cycles that will improve the safety of process, but more importantly, will allow a significant breakthrough in productivity that all mines need to remain successful. 2 INFRASTRUCTURELESS GUIDANCE (IGS) MEANS FLEXIBLE AUTONOMOUS OPERATION Autonomous tramming of production vehicles has been available for a number of years and has demonstrated the economic returns of increased production and efficiency. Unfortunately, earlier generations of autonomous vehicles have relied on installed infrastructure (light rope, reflective tape, bar codes etc.) to guide the vehicle along predetermined routes. This extra installed infrastructure represents additional cost for installation and maintenance. It is susceptible to damage from blasting activities and rock falls. It is also extremely inflexible and places restrictions on the routes an automated vehicle may follow because a continuous guide-path must be established for the entire vehicle route. This often means that work crews must move or install new guidance cues before a route can be traveled autonomously.
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IGS, as the name suggests, does not require the installation of a continuous guidance medium. Instead, it uses 2D laser scanners to be its "eyes" to follow an assigned route and avoid obstacles. IGS provides guidance and control within a strict safety regime utilizing SICK laser scanners (see Figure 1) mounted fore and aft on the vehicle as the primary sensors. The on-board vehicle controller, StrongBox, uses IGS to control the vehicle. IGS operates on two distinct levels, guidance and navigation. The RoboScan software module performs the guidance function. The Route Profiling software module performs the navigation function.
Figure 1: SICK Laser Scanner and AMS Equipment Mounted on LHD 3 IGS SYSTEM ARCHITECTURE IGS is comprised of a number of major software modules. This is illustrated by the system block diagram shown in Figure 2.
are discarded, as data beyond that distance is not reliable or useful enough to be used for guidance. RoboScan employs a novel fast algorithm that results from M D Robotics’ research into extra-terrestrial vehicle navigation. This algorithm quickly determines where viable paths exist in the direction of travel. Steering commands are sent directly to the vehicle steering controller to provide a fast response to changing path conditions. These steering commands keep the vehicle on a safe trajectory to avoid collisions with walls and obstacles. RoboScan keeps the vehicle straight and parallel to the drift walls when driving along straight drifts. Intersections are safely negotiated by plotting trajectories based on Bezier curves. RoboScan plots Bezier curves using these four control points: • Center of Scanner (or center of the front axle and center axis of the vehicle) • Immediately ahead of the vehicle, along the vehicle’s center axis, in the center of the desired crosscut. • The center of the desired crosscut. • A point projected to be in the center of the crosscut, 10 meters along the crosscut. In the case when the desired crosscut is the one straight ahead, the Bezier curve will essentially be a straight line. Figure 2 shows a Bezier curve around a right-angle intersection, but the algorithm handles more general branching cases as well. There is no requirement for there to be cross-cut openings on both sides of the drift, and the cross-cut drift can meet the current drift at any angle within ± 45º of the perpendicular. Drifts meeting at shallower angles are dealt with by wall following techniques. RoboScan sends steering corrections to the Steering Controller that interfaces directly to the vehicle steering. These steering corrections ensure that the vehicle correctly executes turns at intersections maintaining safe distances from the drift walls. When no paths are visible, RoboScan signals a controlled stop. A configurable distance threshold of 10 metres is used to make this determination. This is the normal mode of stopping the vehicle at the end of route.
Figure 1 - Infrastructureless Guidance System Block Diagram 4 ROBOSCAN SOFTWARE MODULE, THE GUIDANCE HEART OF IGS RoboScan receives laser scanner data from both the forward and reverse scanners. It analyzes the data in the direction of motion to recognise obstacles such as walls and other vehicles that could pose a collision risk. Scanner points that are more than 25 meters away from the vehicle 666
Figure 2 - Bezier Curve Path Planning
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5 COLLISION AVOIDANCE MINIMIZES ACCIDENT RISK RoboScan performs collision detection by counting the number of scanner points within an adjustable distance threshold. If this number exceeds a threshold value, the potential for collision is reported for higher-level software to take action. Because the drift roadbed is not flat nor the back uniform, it is possible for the scanner to pick up points from the roadbed or back, which could be incorrectly interpreted as potential collisions. However, as the vehicle moves down the drift these points quickly move off the scan plane. If a viable path exists around an obstacle then RoboScan avoids collision by following such a path. If no safe path exists around an obstacle, RoboScan stops the vehicle. It is important to recognise that RoboScan can only avoid obstacles it can see. It is possible for obstacles to exist that do not show up in the plane of the laser scanners or which present a small enough cross section that they could be ignored. For this reason, all operating areas for autonomous vehicles must be considered as total human exclusion zones unless other measures are taken (technical or procedural) to lower the risk of collision to an acceptable level. 6 STEERING AND DRIVE CONTROLLERS KEEP THE VEHICLE ON COURSE The steering controller implements the steering corrections it receives from RoboScan to follow the planned path as closely as possible. The controller has been designed to be parametric, so that different mining vehicles can be handled simply by changing the control parameters. The drive controller maintains the correct vehicle speed by manipulating throttle, gears and brakes. This controller is also parametric. The control parameters required by the controllers are determined from kinematic models (one for steering and
one for the drive train) of the particular vehicle; constructed within a MatLab/Simulink modeling environment. A model of any vehicle may be created based on the manufacturer’s specifications. The steering model details every major aspect of the steering mechanics, hydraulic valves, hydraulic cylinders, linkage geometry and dynamic effects due to friction. The drive model details the every major aspect of the drive train, engine, throttle, transmission and brakes. Once the models are created they may be validated against telemetry data captured during actual vehicle runs in a mine environment. The validated vehicle models are used to create suitable linear model approximations. This is done to simplify the final steering and drive controllers. The model approximations are then compared to the more accurate non-linear models and adjusted until the errors are less than a few percent over the expected working range of input parameters. The steering and drive controllers are based on the model approximations. This means that both controllers are optimized for the particular model of vehicle being controlled. 7 ROUTE PROFILING FOLLOWS THE CORRECT ROUTE The IGS Route Profiling module interfaces to RoboScan and the drive controller to determine the route followed by the vehicle. It uses a previously recorded file called a "route profile" to generate navigation information. The teleoperator teaches a route to IGS simply by driving the vehicle along that route. IGS generates a route profile as the teaching phase progresses. Route profiles are divided into a number of segments as illustrated in Figure 3. Each segment contains a single navigation instruction and is calibrated according to the distance traveled. Route profiles therefore contain a series of navigation instructions such as "drive straight for 150m
Figure 3 - Examples of Route Profile Segments Massmin 2004
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until a right intersection is detected, turn right and drive straight for 20m etc". The Route Profiling module reads and interprets a route profile and sends appropriate navigation instructions to the guidance module as the vehicle follows a route. When the end of a route is reached, the navigation module instructs the vehicle to stop. Route profiles also record the vehicle speed during the teaching phase. The Route Profiling module uses the recorded vehicle speed as a target speed for the drive controller. The drive controller manipulates the throttle, gears and brake to make the vehicle match the recorded speed at all points along the route. The distance the vehicle has traveled is determined by odometry and is subject to error as a result of the wheels slipping. The Route Profiling module keeps this type of error to a minimum by using landmark features along the route as datum points. The distance traveled is continually corrected as landmark features are encountered. Route planning is integrated into the IGS system such that an operator is able to select a route for the vehicle to follow. An IGS equipped vehicle will autonomously travel the entire route from the draw-point to the dump-point (and back), following a previously recorded route profile. On leaving the draw-point, the vehicle will automatically follow the selected route towards the dump-point. The vehicle will stop at a predetermined safe distance from the dump-point. When returning to the draw-point, the vehicle will follow its return route and stop a predetermined safe distance from the draw-point. From the above description, it can be seen that IGS differs from competing systems because of the way it defines and records routes. IGS does not rely on a global map and requires no survey data to operate. It will operate in any area that has suitable radio coverage. Routes can be taught quickly and are easily changed by the teleoperator, all without electronic access to maps and survey data. The overall flexibility of IGS results in a system that requires less processing and communications resources, is easier to install and operate. 8 TEACHING PHASE Route profiles are generated during the teaching phase of IGS operation. Each route profile must be named by the teleoperator before the teaching phase can begin. Route profile names are made up of two parts, a location name and a direction name. The location name is selected from options previously set-up on the teleoperator’s control station (see Figure 4). The direction name is selected from two options depending on the direction of travel. The direction is typically given as "dump" or "load". Route profiles are usually generated in pairs. Each one of the pair describes the route in one of the two possible directions of travel between the same two points. The Teleoperator teaches a new route to IGS by simply driving the vehicle along the required route. He drives at the appropriate speed for each part of the route; the Route Profile module will use the vehicle speed as a target during subsequent route profile playback. The Teleoperator stops the vehicle at the end of the route, by releasing the control joystick or applying the service brake; thereby leaving the vehicle in the approximate position it should be delivered to at the end of autonomous operation (route profile playback). After the teaching phase, the Route Profiling module parses the new route profile and appends a checksum before saving the profile to flash memory. During parsing, the following values are calculated: • Overall length of route • Individual segment lengths 668
Figure 4 - Operator Control Station
• Teleoperator action during each segment (whether a turn was made and whether that turn was to the left or to the right etc.) • Distance to next segment boundary • Distance to a Y-Switch (where the direction of the vehicle is reversed) 9 AUTONOMOUS OPERATION The teleoperator begins autonomous operation by positioning the vehicle correctly and selecting the correct route profile for playback. Selecting a route profile for playback consists of selecting a location and direction in an identical manner to the teaching phase. Playback begins when the teleoperator holds over the control joystick for 2 seconds to select the correct direction of travel. Intersections encountered during route profile playback are used to synchronize the playback odometry to the recorded route profile. This minimizes the effect of wheel slip. If the actual odometry reading differs greatly from that recorded in the route profile for the same intersection then route profile playback is aborted. In this case, the vehicle is safely stopped and an error message displayed to the teleoperator. The vehicle stops automatically at the end of the route after the last line of the route profile has been read. The teleoperator interface includes a video overlay inserted onto the normal vehicle camera view. This display provides appropriate feedback to the operator on: • Current status of the playback or record run • Segment ID • Anticipated features along the route (segment type) • Distance to next feature • Action at next feature (turn right, turn left, straight ahead). • Distance to walls. 10 SAFETY REGIME Safety was a major consideration during IGS development. Safety experts reviewed the entire system design ensuring that IGS is a robust system that is failsafe and single fault tolerant. Ensuring safety is based on these principles: • The exclusion of personnel from the autonomous operating area while vehicles are operating under remote or autonomous operation. Depending on a mine’s particular topography and procedures, this requirement may be met in a variety of ways:
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• Physical barriers with or without verification means (gates, doors, chains) • Signs, lights or other warning indicators • Procedural restrictions • Tagging systems (manual tagging, switch-board tagging) • Personnel detection systems (RF tags, light curtains, etc.) • Secondary safety shutdown devices (safety stop switches, etc.) • A combination of the above • The restriction of the area of operation of the machine to a known area of the mine. IGS will automatically stop the vehicle if the route conditions do match those encountered during the teaching phase. This means that the vehicle will stop if an intersection is encountered too early or too late (within the limits set for correction of odometry), if the drift width is incorrect or if an intersection is of the wrong type (left instead of right, for example). Absolute localization is achieved by the use of small, selfcontained radio ID tags placed at strategic points along a route. Because these tags have a unique ID number they serve to "mark" their location in a unique way. IGS records any tags encountered during the teaching phase and will stop the vehicle if the tags are not encountered at the same point along the route during autonomous operation. Missed or incorrect tags will also cause the vehicle to stop. Area limit RF-ID tags are a secondary method that will cause an immediate stoppage of the vehicle if a limit tag is detected. Two independent limit tags are placed at each entrance/exit to the operational area in such a way that they will not be detected if the vehicle follows its correct route. Used in conjunction with odometry-limited travel, this creates an outer limit to the operating area. The vehicle will stop should it travel beyond the range of reliable communications, or if the system fails to give periodic, regular permissive control packets to the vehicle. The vehicle will also stop due to failure or shutdown of the fixed communication system or due to failure or shutdown of the fixed automation components. • Designed-in features of the StrongBox hardware and software, such as dual-processor architecture that is internally self-checking, prevent unintended operation in the event of single point faults. 11 COMMUNICATIONS StrongBox requires a low speed Ethernet communications link to the control station. Commands pass from the operator to StrongBox and telemetry data passes from StrongBox to the diagnostic display on the control station. StrongBox will function well with a full duplex link of 19.2kb/s but faster telemetry updates will be achieved with higher data rates. StrongBox also requires a single video channel to convey teleoperation camera video coupled with a diagnostic overlay. The video stream is in analogue form and conforms to NTSC or PAL video standards. IGS places no additional demands on the communications infrastructure as IGS related commands and telemetry data are merged in with the conventional StrongBox data transmissions. IGS requires continual radio contact with the operator control station to allow autonomous operation. The vehicle will stop safely if data radio contact is lost. StrongBox and IGS will function well in conjunction with any communications system that meets the requirements for real-time teleoperation. 12 3D Photo-Realistic Imaging System by LHD for Instant Geology, Geotechnical and Surveying information 3D Imaging System (3D IS) is an exciting new technology that is now being demonstrated in mines by AMS. This Massmin 2004
technology is unique in its field and was developed from work for the Canadian Space Agency to support Mars Lander missions.
Figure 6: shows an LHD taking a 3D model of a mining face. The image can be transported by RF or hard wire link and integrated into mine management software to get a 3D asbuilt rendering of face geology, and survey coordinates. To be studied in real colour from any angle . 3D IS analyzes the signals from stereoscopic cameras to construct a three-dimensional model of the scene being surveyed. Details of the scene are built into the model by mapping precise textures taken from the individual camera images. This results in a truly photo-realistic three dimensional VRML model that may be manipulated (pan, zoom, rotate etc.) using conventional tools; including web browser plug-ins. 3D IS captures the spatial arrangement of the various components in the scene model so it is possible to measure three dimensional distances between any two such components. This allows the size and distance of rock features to be quickly measured. Presently, a geologist and other geotechnical or survey personnel visit each fresh workface to assess where the ore is, and as a result, make a decision about where to drill the next round. From visually inspecting the face, the geologist creates a hand sketch of the ore and features, and brings the sketches to surface and ultimately logs the data into a mine planning software. The geologist is able to extrapolate where the ore is going, from an accumulation of consecutive sketches. 3D IS has the potential to significantly improve the time to collect information, and also the visual information that is available at each work face. 3D IS scans the workface, collecting a photo realistic colour picture from a sequence of images. The system is portable and may operate either in hand-held, automatic or mobile vehicle mode. The 3D model is accurate and is created within a few minutes of acquisition. The images can also be integrated into a mine database to facilitate mine planning. The data can be interpreted from successive images to better determine the grade at the face, and where the ore is migrating, thereby allowing for a considerable opportunity for better controlling the amount of dilution. The collected images can be transported from underground to surface via an RF or hard-wired link to be viewed by the geologist in his office. A stitched image can also be sent via the Internet to other locations for assessment and to support decisions by the mining company’s most qualified personnel. Without actually visiting the working faces, corporate experts can easily visualize the features for a variety of
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purposes including mineral content, unstable ground, and many other re-occurring workplace issues. The autonomous LHD mucking cycle could be enhanced by incorporating 3D IS into the vehicle control system. Analysis of images from the stereoscopic cameras would provide estimates of the vehicle motion (visual odometry, which would overcome inaccuracies of wheel odometry due to wheel slippage). 3D IS would also allow for the localization of the vehicle using natural features observed in the environment. 3D data could be used to create 3D photo-realistic models. These models could be applied to autonomous bucket loading by giving the vehicle control system independent knowledge of the instantaneous muck pile shape and flow. Fragmentation and grade of the ore could also be determined, which could then be sent from the vehicle to a mine database. Eventually, 3D IS could be used as the vehicle guidance sensor itself, replacing the laser scanners in the current version of IGS. Three dimension versions of the current RoboScan algorithms would detect walls and other obstructions and plot safe vehicle trajectories into observable drifts and cross cuts. Currently, the processing time of 3D IS will not support real-time guidance applications; but with anticipated improvements in algorithms and computing platforms this prospect is not very far removed. 13 CONCLUSION Infrastructure-less guidance techniques and real time photo-realistic imaging of the advancing workface have the potential to significantly advance the speed and flexibility with which mining and drifting are performed. One worker can operate two or more vehicles from a safe surface location allowing for increased machine utilization, and more tons moved. Also, geological, geotechnical and survey data can be captured during the mucking process and transferred seamlessly to surface databases to further speed up the mining process, and also offer significantly
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better information to the technical and management personnel who can make decisions. The mining process is proving to be at the threshold of evolution to a safer and significantly more productive process, which will translate directly to quick rates of return, and increases to the financial bottom line of any mining company. REFERENCES
• Bloomquist, L. P.Eng., Law, J. And Arnoldi. C., 2004. The Evolution of Advance guidance Techniques for Autonomous Underground Vehicles, CIM Bulletin, January 2004 • Anderson, A et al, 1998. Mining Industry Roadmap for Crosscutting Technologies. U.S. Department of Energy, National Mining Association Technology Committee meeting, Denver, Colorado October 1998. • M. Born and E. Wolf, Principles of Optics, Sixth Ed. (Pergamon Press, Toronto, 1980) pp.38-41. • Baiden, G. and Henderson E. 1994. LHD Teleoperation and Guidance – Proven Productivity Improvement Tools. CIM Toronto Conference 1994. • Bloomquist, L. and Hinton E., 2001. Towards an infrastructureless guidance system - A Proposed Guidance System for Autonomous Underground Vehicles. CIM Bulletin January 2001. • M. Piotte, C. Coache, A. Drouin, P. Gaultier, and L. Geoffroy, 1996. Automatic control of underground vehicles at high speed. CIM Edmonton Conference 1996. • Wber, Koller, D, Luong, Q. and Malik, J. 1995. An integrated stereo-based approach to automatic vehicle guidance. International Conference on Computer Vision, Boston June 1995. • Vagenas, Sjoberg, and Wikstrom, 1991. Application of Remote-Controlled/ Automatic Load-Haul-Dump System in Zinkgruvan, Sweden. Proceedings of 1st International Symposium on Mine Mechanization and Automation, I p. 6/21-6/30.
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Telerobotic experiments for mining Dr. Greg Baiden, School of Engineering, Laurentian University Sudbury, ON, Canada
Abstract Telerobotics focused on mining is currently being introduced into production systems around the world. Mining companies in Canada, Sweden, South Africa and Australia have tended to lead the international charge to this form of technology for the mining of underground and open pit operations. While this introduction is taking place a few basic questions have yet to be answered. How many machines can a single operator run? How many types of machines can a single operator run? A new Canadian Research Chair in Robotics and Mine Automation has been established at Laurentian University. This chair will investigate these questions and many more through a series of experiments in a newly established telerobotics laboratory that connects Laurentian teleoperation workstations to model mining machines at Cambrian college’s eDome. The laboratory will support many experiments allowing researchers to physically run multiple robot scenarios differing the numbers and types of machines at one/quarter scale. Further experiments are being designed to investigate the potential for managing time delays in telerobotics. This paper reviews the state-of-art in the field and describes the laboratory, experiments and some preliminary results achieved to date.
INTRODUCTION Telerobotics systems used for mining are currently being introduced into production systems around the world. Mining companies in Canada1,2, Sweden3, Finland4, South Africa5 and Australia6 have tended to lead the international charge to this technology for mining. NASA and the Canadian Space Agency have begun to discuss the use of telerobotics for rock drilling on other planets and large scale space construction8. While organizations are looking toward telerobotics for applications, a few basic questions have yet to be answered. How many machines can a single operator run? How many types of machines can a single operator run? A new Canadian Research Chair in Robotics and Mine Automation9 has been established at Laurentian University around this subject. This chair will investigate these questions and many more through a series of experiments in a newly established telerobotics laboratory that connects Laurentian teleoperation workstations to model mining and full scale machines at Cambrian college’s eDome. The laboratory will support many experiments allowing the researchers to physically run multiple robot scenarios differing the numbers and types of machines at one/quarter scale. The paper describes the laboratory, experiments and some preliminary results. Robotics as field of study is the creation of machines and control systems that can perform the tasks of a human operator. Within the robotics field is the subset of telerobotics. Telerobotics allows the human operator to remain in the control of a semi-robotic machine that incorporates limited automation to provide the opportunity to free time for the human operator. The free time allows the operator to focus on other more productive work. For example, one operator can run multiple semi-robotic machines achieving a significant potential increase to the output of a process. Mining is a process where mineral is broken and treated to extract the valuable portions for use by society. Mining consists of the discovery, delineation, development, extraction and closure of an ore body. Each process is currently mechanized. In the future, teleoperation of robots will become the norm for these processes altering the performance of the mining industry physically and Massmin 2004
economically. Some expected results include enhancements in: safety, productivity, cycle time, quality, value and costs. REVIEW OF TELEROBOTICS Telerobotics in the mining industry started to be used in production in the early 1970’s with the implementation of line-of-sight radio remote controls underground and truck dispatch systems in open pit mines. As the ideas have progressed, line-of-sight remotes have been extended to longer and longer distances underground. Open pit mines leaned towards autonomous operation of truck due to a lack of communications bandwidth. Today many underground mining companies and suppliers are developing true teleoperation systems where operators sit as far as hundreds of kilometers from the units they run. New research into a high bandwidth communication system for open pit mining is underway at Laurentian University. Line-of-sight radio remote controls have been integrated into mining methods over the last 30 years. In fact, some methods such as Vertical Crater Retreat mining depend on this technology on a day-to-day basis. For example, companies such as Inco and Falconbridge could not produce safely without the technology onboard LHDs and some drills. Long distance teleoperation while starting from line-ofsight techniques has seen surface teleoperators run multiple LHDs and drills from comfortable control rooms. The safety benefits of this style of operation are obvious as teleoperators are not exposed to the mine environment. Moreover, the support staff for the teleoperator only enters the mining area when no mining work is happening. Therefore, safety is maximized for the operation. While safety is an important benefit, teleoperation offers many more. For example: • The ability for labour productivity gains as a result of instantaneous connection to many machines at the flip of a switch • Maintenance improvements through the operation of machines within their capability • More intimate control of the mining process through direct measurement and communication of real time information
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• The combination of these benefits alters the traditional economics of mining a deposit making it more profitable.
TELEOPERATION CHAIR TEST-BED LAURENTIAN LABORATORY
Given that these benefits have been achieved and mining companies will gradually embrace this new method of mining it is important to be building the human and technological resource base to support the implementation.
A teleoperation chair test-bed is located at Laurentian University (figure 2). The chair consists of an ergonomically designed seat that has joysticks and displays to allow an operator to teleoperate a machine. The main idea behind the teleoperation chair is to study the interface between mining machines. As the research progresses it is working toward the development of a universal interface to allow a configurable computer interface for all types of machines without the need to physically alter the chair. The connection between Laurentian University and Cambrian College is through a "dark fibre". The fibre is connected to a head-end located in the laboratory and networked to the computers on-board the teleoperation chair.
MINING TELEROBOTICS LABORATORY A Mining Telerobotics Laboratory is now commissioned and operating at Laurentian University and Cambrian College in Sudbury. This facility is unique in the world today. The laboratory is a collaborative research site between the two Sudbury educational institutions. It has been put in place to build on the strengths of the two institutions and with Penguin ASI fulfills the innovation cycle of research, development and commercialization. The Laurentian portion of the laboratory consists of a teleoperation chair and a telecommunications control room for research into teleoperation. At Cambrian College a series of laboratories that includes the eDome, supports the ability to design, build and teleoperate from 1/4th scale model to full size mining machines and full mine simulations. As this work progresses new control systems, machines and processes will be developed building on the idea of telerobotics. Since telerobotics will become an emerging industry that has and will grow from the mining industry into other fields such as manufacturing, underwater, space and military, the building of this laboratory is and will be important to the development of telerobotic applications.
Figure 2: Teleoperation Chair at Laurentian University
Figure 1: Teleoperation Laboratory Layout
TELEROBOTICS TESTBEDS CAMBRIAN COLLEGE LABORATORY
The laboratory is really the combination of two labs in different geographical locations in Sudbury. Figure 1 shows a map of the lab locations and the equipment in each. Jointing the two locations is a dedicated "dark fibre" to facilitate teleoperation experimentation. This laboratory connection allows the study of a number of telerobotic issues that include: • How many machines can a single operator teleoperate given different levels of automation? • Is it possible to teleoperate the task of development using cooperative robotics? • When the telerobotic systems are scaled up, what are the implications for the safety, timing and economics of multiple mine operation scenarios?
The dark fibre is brought into Cambrian College at the eDome (figure 3). It is connected to a second head-end switch that supports mobile data communications and mobile video. The wireless network established between Laurentian University and Cambrian College will allow the study of mobile machine operation and multi-machine operation. As the research progresses full scale machines developed for mining and various other industries will be tested at the Cambrian facility. Cambrian College through the instrumentation and machine design classes are developing a 1/4th scale model Load-Haul-Dump machine. The first unit is one of the most sophisticated mining machine models that has been built to date. A general schematic of the structure of the model LHD is shown in figure 4.
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the electronics being an exact duplicate this allows experimentation with new systems on scale models both at the teleoperation lab and in real mine facilities. The model LHD has the potential to operate at any mine in the world with the correct telecommunications infrastructure. Following the building of the first 1/4th scale model of an LHD and the connection of the laboratory a fleet of models will be built to answer the research questions outlined at the beginning of the paper. First, five more telerobotic units will be built to aid in chair and control system software experimentation. Some of these new units will have fuel cell power plants instead of batteries. The fuel cell offers the same capability as batteries while increasing the range of the unit. Eventually, a Honeywell Ore Retrieval and Tunneling Aid (HORTA) will be incorporated onto new units. The HORTA is a Ring Laser Gyroscope (RLG) combined with accelerometers. This unit outputs azimuth and x, y, z coordinates that a pair of laser scanners can be linked to allowing the collection of tunnel or drift data and experimentation with alternate guidance systems. This makes the machine an experimental test-bed for teleoperation and data collection.
Figure 3: Cambrian College eDome
TELEROBOTIC OPEN PIT OPERATION While the bulk of the telerobotics work to date has been in underground mining, open pit mines are in need of similar opportunities. Open pit mining equipment suppliers have tended to work towards autonomous mining equipment because of limitations in communications bandwidth. Given past experience in the implementation of underground automation it is important to note that remote control followed by telerobotic control leading to autonomy is the logic path of implementation. Mine operators need to be comfortable with the operation of their fleet of equipment. Therefore, telerobotic operation of open pit mining equipment is more than likely going to be the next logical step as opposed to autonomous operation.
Figure 4: 1/4th LHD Scale Model Desing
The base machine is a CANBUS compliant unit that is electrically driven. Steering and bucket control are done with electric screw actuators. The unit has an onboard AMS Strongbox controller for machine control and telecommunications connection. The first of these units is complete and operational as shown in Figure 5.
WHAT ABOUT TELECOMMUNICATIONS FOR OPEN PIT? High bandwidth communication for open pit mining is a huge challenge if telerobotics is to be employed. Radio frequency communications are inherently low bandwidth in comparison to the need. Therefore how can we deal with this? One solution is using optical communication systems. As part of the expansion of the Telerobotic Laboratory at Laurentian University we will begin experimenting with high bandwidth optical communication systems. Our first experiments will take place this summer using an underwater optical system. Underwater was chosen as this is the most difficult environment for high bandwidth communication. Figure 6 is an artist concept of the underwater optical communication system working for telerobot control. RESULTS
Figure 5: Actual 1/4th Scale Model
The unit is designed with exactly the electronic and telecommunication systems that are on current production robotic machines working in the field at Inco Limited. With Massmin 2004
The Mining Telerobotics Laboratory is now completed between Laurentian University and Cambrian College in Sudbury. On May 25, 2003 we ran our first telerobotic machine between Laurentian University and Cambrian College. The completion of this marks the beginning of a test facility to support the development of telrobotics for mining. The initial experiments are beginning to be laid out to investigate the implications of telerobotics for underground mining. At this time, a new agent based algorithm for telerobotic dispatching of machines is being developed. This will be initially targeted for block cave mines.
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Development work is now underway on an optical communication system for open pit mining. This development will initially start off in an underwater environment although testing for surface applications will happen early in the research program.
Figure 6: Underwater Telerobotics
SUMMARY A great deal of work is yet to be done to hone the idea of telerobotic applications for mining. This paper has discussed a new Telerobotics Laboratory created through the Canadian Foundation for Innovation, Ontario Innovation Trust Fund, Materials and Manufacturing Ontario, Inco Limited and Penguin Automated Systems. The laboratory has begun experimenting with telerobotics for mining, cooperative telerobotics and multi-mine teleoperation. The laboratory was completed in the Spring of 2003. This laboratory was designed for studying telerobotics for mining will have implications for manufacturing, space, underwater and military applications. In early 2004 a research program was brought together to begin to experiment with high bandwidth telecommunications for underwater and surface applications. This work is just underway but is yielding promising results already.
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REFERENCES [1] Baiden, G.R., "MAP - Mining Automation Programme Results", Final Presentation, October 25, 2000, Copper Cliff, Ontario, Canada. [2] Piche, A., "LHD Automation at Noranda, 33rd Annual Robotics Congress, Montreal, Quebec, Canada, April, 1999. [3] Samskog,, P. and Wigden, I., "Process Automation in Ore Mining", MineTime 99, Dusseldorf, June 8, 1999. [4] Schweikart, V.S.,"Control and monitoring systems at the Palabora Underground Mine", Proceedings from Telemin 1, June 1999, Sudbury, Ontario, Canada. [5] Pukkila, J. and Kosonen, K., "The implementation of Intelligent Mine Technology at Kemi Mine of Outokumpu Crome OY", Proceedings of Telemin 1, June 1999, Sudbury, Ontario, Canada. [6] Penswick, D. and Gilliland, K., "The missing link: The gap in mining automation hierarchy", Proceedings of the 2001 – 6th Symposium on Mine Mechanization and Automation, pp 129-134. [7] Dudley, j. and Ferrier, S.,"Tele-remote LHD implementation and the progression to LHD automation at Northparkes E26 Block Cave Mine", Proceeding of Telemin 1, June 1999, Sudbury, Ontario, Canada. [8] Mankins, J., "An Overview of the Space Solar Power Exploratory Research and Technology Program" Advanced Projects Office – Office of Space Flight, Spring 2000. [9] Canadian Research Chairs Website, Government of Canada, 2001. ACKNOWLEDGEMENTS The author would like to acknowledge the support of the Laurentian University, Canadian Research Chairs (CRC) program, Canadian Foundation for Innovation (CFI), Ontario Innovation Trust Fund (OIT) and Materials, Manufacturing Ontario (MMO), Communication and Information Technology Ontario and FEDNOR. Further the generous support of Penguin Automated Systems Inc. and Inco Limited that made this work possible. The City of Sudbury and Sudbury Hydro have made a significant contribution by supplying the dark fibre link between the two facilities.
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Bringing underground mining into the Wi-Fi revolution Sergio Blacutt, CEO, Jigsaw Technologies Inc
Abstract This paper addresses the advantages of reliable, high-speed data and voice networks in underground mining operations. Jigsaw Technologies is combining proven technologies used extensively on aboveground applications and transforming them into state of the art tools for underground operations. At the heart of these mobile applications are sophisticated networking techniques with efficient software designed from the ground up for underground mining operations. Custom applications are deployed on all mobile units in order to optimize their performance and reduce network traffic. This technology allows engineers and other underground personnel to query and/or update critical information directly from the field. A key advantage of implementing this system is its ability to carry voice over the same data network (VoIP). Jigsaw Technologies wireless network utilizes modified 802.11 access points that run custom routing software, providing a robust and secure VoIP communications platform. Access points are interconnected via a fiber-optic or standard CAT-5 backbone that ultimately connects to the mine network allowing bi-directional data/voice communications between authorized users. In conclusion, the combination of powerful custom software for underground operations, plus the ability to carry crystal clear full-duplex voice communications on one single mobile device, will enable a company’s plan to lower costs and achieve higher profits. 1 INTRODUCTION Utilizing the latest wireless technological advances and data-gathering devices along with standard and user friendly information displays, Jigsaw Technologies is developing easy to maintain reliable data and voice networks for underground mining operations.
only predefined alarm conditions or unscheduled events. A major advantage of all Jigsaw’s applications is their ability to integrate and support off the shelf hardware components. Different operating systems are also supported and field units can either run under embedded Microsoft Windows or Linux. 2 CORE NETWORK TECHNOLOGY To ensure reliable and continuous operation, the system communications network is entirely based on mesh network technology. Unlike traditional tree or star network architectures, mesh networks utilize intelligent access points that can communicate with each other without the need of having to go through a central switch.
Fig 1. Overall application diagram This application will allow mining operations to coordinate and monitor all field operation activities by providing mine personnel with powerful wireless rugged portable computers connected to a central application. Included in the central application is a standard "Session Initiation Protocol" (SIP) local service that allows for clear full duplex voice communications over the standard 802.11 wireless network. All mine equipment can also be fitted with field computers that allow for real-time communications, and data gathering providing significant on board intelligence to continuously analyze and store data locally. Although field units can continuously send information to central, they can also be configured to analyze trends locally and report to central Massmin 2004
Fig 2. Tree type network. Failure at a top branch will render all units under it unusable. Mesh networks incorporate a grid type of network with one of the major advantages being the elimination of a single point of failure.
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The Jigsaw Technologies physical network infrastructure is deployed using embedded Linux based mesh access points powered via Power Over Ethernet (POE) technology. These rugged access points are capable of operating within a wide range of temperatures and are able to withstand high transient voltages.
operate uninterrupted even without communications with the central server. Continuous mining operations data is collected and analyzed locally on each device with only relevant information sent in real-time to the central server. This approach maximizes bandwidth utilization and ensures high network availability for voice and video communications. 4 DATA ANALYSIS AND SYSTEM INTEGRATION The Jigsaw application cleanly separates data access, business logic, and user interface. User interfaces are designed as XML files that instantiate interfaces on mobile or stationary devices on request providing unique and custom interaction screens for each different user. Maintenance personnel for example can log into a mobile unit and be greeted with the equipment engine parameters and equipment-troubleshooting screen. All system interaction and reports generation is done through a web-based user-friendly interface that requires minimum or no user training.
Fig 3. Mesh type network. Without a central switch, mesh networks provide self-healing and self-organization characteristics. Power Over Ethernet eliminates the need to deploy additional cables between mesh access points requiring only the use of power injectors at certain points. As new mobile or stationary network devices are deployed into the network, they are set to automatically detect all neighboring nodes, configure themselves, and based on factors such as signal strength, transmission errors, and latency determine the best path to use when communicating peer to peer or with the central computer. Any changes or sudden faults on the network are immediately detected by affected field devices, which automatically recalculate the best communications path to the rest of the network. 3 DATA AND VOICE COMMUNICATIONS Having a robust and reliable network permits the Jigsaw application to include voice over the same data network. A SIP server application is deployed on the central computer and all mesh access points, with all mobile devices equipped with VoIP capabilities. This capability allows all field devices on the network to talk to each other as well as any user connected to the network. All field units are deployed with custom software applications, and have the ability to process data locally and
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Real-time field data is automatically stored in powerful SQL databases and can utilize a number of commercial SQL database servers, such as Microsoft SQL, MySQL, OpenBase, Oracle, Sybase ASE, and others. This allows for easy data integration with most ERP applications. In addition the main application can be deployed on various operating systems including Mac OS X, Linux, and Microsoft Windows 2000 or XP. 5 CONCLUSION As companies increasingly focus their efforts on their core business, it is becoming clear that state of the art open technologies will play a major role in their competitiveness and successful business execution. Systems like the one described above, that utilize advanced and standard hardware with powerful embedded software applications designed for the underground mining industry will help companies achieve their goals and increase productivities. Strong communication channels have always been the backbone of any successful company. Adding reliable voice capabilities to monitoring and control applications results in higher productivity and most importantly safer underground mining operations. Mine personnel will be able to communicate freely with other underground personnel or personnel on the surface without the need of dedicated restrictive cables.
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Automation of mineral extraction and handling Fredy M. Varas, Civil Mining Engineer MBA, Project Management, El Teniente Division, Codelco Chile
Abstract Underground Mining, in order to maintain its competitiveness, need constant innovation thus it must make use of the technological advance which allow it to face not only challenge of cost diminution and productivity increase, but that also to give better work’s conditions to his workers which is translated in diminution on exposition to the risk and better environment quality. Codelco and especially El Teniente Division, conscientious of these considerations, have implemented strategic impulses to incorporate technology as an essential competitiveness requirement. For that reason when Teniente Development Plan was formulated, these concepts were incorporated in his mining projects and they are been materialized. So, we have that projects like Pipa Norte are pioneers incorporating semi-automated LHD equipment and complementing this technology with an innovating material handling that get closer the size diminution issue toward mineral extraction points, establishing a production line so that after reaching the scoop’s dumping point it allows to reduce mineral, through a jaw crusher, to 7 inches size very appropriate for a belt conduction and improving the filling factors for the railroad cars to its later hauling to the surface. In order to materialize these concepts, it has been necessary to consider a series of edge conditions, some of specific technical nature, being notorious by its importance the subject of communication systems and production control, and in the same way the production level designs and layout and the incorporation of material handling elements not used before at El Teniente Mine. They are also notorious the incorporation of complementary technologies like hydraulic fracturing and the utilization of plate feeders for material handling. Finally as case of study, this work shows the main technical elements that has been considerate in the Pipa Norte Project, being notorious the considerations of productive nature, security and the standards that have been established for project operation.
INTRODUCTION The present tendency in underground mining is to give a product of homogenous granule size, initiated by means of a natural fracturing process, for then to incorporate technology of downsizing in places near to mineral extraction. Thus the mineral is transported to surface by means of the best automated systems. Associated to this the underground mining constantly needs to improve his productivity, for which the new projects must respond to needs of improvements in production costs and so life and environmental quality of workers. An effective way to obtain these results is making use of the technology available on the modern world. Thus it is as the new mining projects of the El Teniente Mine have looked for to obtain important advances in the use of ultimate generation technology. With this it looks for to assume the strategic leadership of technological innovation, which is being materialized in two relevant aspects in the operation of productive sectors. First one is related to using of semiautomatic equipment of greater size for mineral extraction, because the greater synchronization of their acceleration movements contribute to wait for a lower maintenance cost per extracted ton. The second aspect of this technological innovation talks about to an approach of the downsizing and handling of materials towards mineral extraction points. Massmin 2004
Applicability of the Technological Strategy To define applicability of available technology of mineral extraction automation in underground mines is a decision that don’t should loose the orientation of give continuity to the mining business, for which is very important to evaluate its real utilization, therefore antecedent compiling related to automation in the world on material extraction processes in underground mining must response to the convenience of operating a system that be compatible with strategic objectives that delineate enterprise orientation in its long term projection. For the studied case this decision aims to fulfil with one of the strategic objectives of División el Teniente, that shows the orientation and endorse decision in terms of incorporate technological advances to mining activity, in addition to increase productivity levels, improve security indicators for the personnel and ensure a quality level that allows to obtain the certification, in accord with standards fixed for ISO norms that regulate quality in its widest concept, finally aims to minimize the operational interference and guaranties a greater continuity of productive activity. Tendency of Materials Handling In a traditional scheme, materials handling in underground mining consist in mineral extraction by some explode technique and, from there, guided it down through a flux, which can be transported horizontally in intermediate levels, and then, going up or down, without any processing or selectivity on this path. Today this situation begins to change, going to maintain automated transport circuits and, therefore, material that is handled and transported must be homologated to certain granule size characteristics, allowing the minimum number of stops and make easier the autonomous control systems, that are operated from the surface. Description of Pipa Norte Project This is one of the three mining projects of División El Teniente and it is in construction from ends of 2002, and will finish to ends of 2004, this project involve reserves for 27.1 Mtons, with average Cu law of 1,00%, and with an infrastructure and with an infrastructure that will have a nominal capacity of 10,000 tpd. and it will be started up with the best technology, considers a high degree of automation.
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Table Nº 1.- Comparative Table in classic format and current tendency on Material Handling Characteristic
Traditional
Current Tendency
Caving Altitude
High one taller than 10 meters
Down one equivalent to caving gallery altitude
Type of Caving
Simultaneously overture of collector galleries
Pre-caving or Advanced Caving
Verticals Mining
Great quantity of meters
Minimum Verticals Mining
Mechanicals
Automated
Material Handling
Transportation to surface in big granule size
In situ Size Reduction
Communications
Wired
Wireless
Explode Method
Panel Caving or Sub Level extraction nets reduced
Big extraction nets or macro galleries
Of production Onevariable Sensibility
Of Multi-variable business
Expertise
Competency
Equipment
Orientation of the Administration Human Resources
The project will integrate the management by competition, gathering the best standards of competitive companies, the best practices of work and a harmonic development of the human resource during the life utility of the project. The location of the sector in the deposit El Teniente can be observed in the following figure.. The geologic and geotechnical antecedents, demonstrate that more of 90% of the existing rock it corresponds to primary and secondary andesite. Certain important intrusions of hydrothermal Breach and Braden Breach are verified mainly towards the Southern of the sector, adjacent with Pipa Braden, which present predictions of heavy fragmentation. Exploitation Method. The exploitation method used in this sector is a panel caving with a variant of low undermining of 4 m. and advanced to the limit of abuttmen stress, this undermining is
obtained by means of the pillar blasting with a design of parallel shots of 3.5"- 4.0" of diameter. This design has considerable operational advantages in security and for removal of the residual material, thus its application, with the restrictions stipulated with respect to the out of phase condition between the advance in undermining and the advance in extraction, is confirmed. The mining design of the project emphasizes the following general characteristics: • Extraction Net of Teniente type, 300 m2, (30 m x 20 m), 30 m separation between streets and 20 m ditches. • Advanced Caving, by blasting of solid pillars of 13 meters wide, 23 meters in length and 4 meters of high (4 m x 4 m section). • Height of the Crown Pillar, 19 m between production and caving levels floors and between extraction points and rolling carpet.
Figure Nº 1: Location Pipa Norte Mine 678
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Figure 2: Diagram of Material Handling in Pipa Norte ProjectFigure 2: Diagram of Material Handling in Pipa Norte Project
Material Handling The mineral handling system has concluded with the consolidation of designs already established in the Conceptual Engineering, that is to say, with a pre-crushed hopper system of 400 tons with a 70" x 70" grill, fed directly by LHD equipment, and a plate feeder that feeds a jaw crusher with a maximum gap opening of 84"x 66". This unit unloads on a conveyor belt of 60", 20 m on length, the one that as well feeds another conveyor belt of 48" and 330 m on length, which takes the mineral up to two main resentments, that allow to transfer the mineral until Teniente 8 Level, to be transported by the railroad towards Colón Concentration Plant. Description of LHD Semiautomatic System The system is formed by a system operator located in Control Room located in Colón Alto, approximately to 15 km of the production area of Pipa Norte Mine. This operator is the person in charge to control and to operate the SCL (LHD Control System) and the handling of personnel and equipment access to the Autonomous Operation Area that is the area of work in production area around where LHD equipment circulates. The SCL is structured by: • Mission Control System (MCS): It provides supervision and control for operation, also planning for production and functionality control. The MCS also provides interphases to the external systems like the SCGP (Production Control & Management System), SCGM – (Maintenance Control & Management System) and SCADA, that is a system oriented to supervising control and data acquisition. • Operator Station: It is the physical place located in Colón Alto, where System Operator is located. It provides Massmin 2004
control and user interphases to the system operators. • Access Control System (ACS): It controls the access for personnel and mobile equipment to the Autonomous Operation Area. Its main function is to maintain the confinement for production tasks of semiautonomous LHD with respect to the personnel and equipment in the production area. • Wireless Underground Communication System (WUCS): It provides video, audio and data communication, like the vital signs of semiautonomous LHD, between Control Room and these equipment. • Semiautomatic LHD Equipment: They have a capacity of 13 yd3, and their production cycle begins with the load in Extraction Point on tele-commanded mode, continuing with the material hauling, material that is dumped directly on grill, and finally LHD returns to the same point or another one, according to production plan or availability of extraction points at that moment, on autonomous mode. • Operation: There are two people directly located in the Operator Station, controlling the SCL. To one of them corresponds the function of control of LHD equipment and the tele-commanded operation of these equipment in his loading task; the other person is responsible for control and supervision of the operation, work that is mainly focused to the planning and coordination of the activities during the turn. Operation of the System The LHD will load mineral at the extraction points on telecommanded mode, to go then on automatic mode to the Dumping Station, point in which they will unload the transported mineral. This point has three flanks for mineral
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unloading. After this, the LHD will be destined, according to the production program given by the Production Control & Management System (SCGP) to the Mission Control System (MCS), to the same extraction point or another one where the system operator, from Control Room, will come to load it again, completing the cycle. An Access Control System (ACS) will maintain confined the operational area of LHD equipment, in such way to make production tasks a safe operation, maintaining an adequate access control and monitoring of Autonomous Operation Area. It must be mentioned that certain production support tasks will have to be made in Pipa Norte Mine, such as: • Secondary reduction in Extraction Points • Off-hook in Height in Extraction Points • Sampling in Extraction Points • Production Infrastructure Maintenance and Material Handling.
The system, as result of to have so advanced level of technology never before applied in an integral form, presents weaknesses that are fundamentally focused to: • Inexperience in production capacity for systems with demanding production programs • Preparation of an equipment of maintenance for the automation system that be efficient and a market that supplies spare parts for the system. • The automation takes associated a change of practice in complementary elements of the extraction operations like, for example, secondary reduction. • Material handling through production line and a planning of the made mining preparation in such a way that avoid interferences to the zone in production, are necessaries.
Considering this, the entrance of mobile equipment operated in manual mode and personnel to the production area while the different operational works are being executed, will mean that the Autonomous Operation Area will be modified, through doors located in the production galleries and in the accesses to the area, isolating, of this form, the Autonomous Operation Area from the sector where the support tasks are developed.
Additionally with this system, the own difficulties of heavy material handling originating of the primary rock exploitation improve the effective operation time in their handling. The automation produces an increase in the utilization and an increase of the yield, because of operations can be made at constant speeds and with equipment of a greater capacity; also maintenance associated to the equipment is diminished in costs fundamentally because the automatic operation diminishes the premature wearing down of many elements due to the better synchronization of component functioning and also improves the information for preventive maintenance of mobile units.
CONCLUSIONS
REFERENCES
Of the exposed, it can be concluded that the Pipa Norte Project incorporates the last standards in material handling for Underground Mining, incorporating LHD equipment with semiautomatic operation, secondary reduction with hydrofracturing system, material handling system with plate feeders, primary crusher near to production levels, and conveyor belts, all that are translated in simplification and automation of the whole process. Mineral Handling in systems like the observed one allows to define strategies that assure with high level of satisfaction for the client (quality, amount and opportunity), the accomplishment of mining plans, under standards of first cuartil of the underground mining sector, that position to companies in an efficient and competitive way. Application of these modern operation technologies, must actually be reflected with high productivity standards, low operation costs, and security, life quality, and environmental conditions that assures the technical and economic viability waited for the project.
1.- Pozo Rodrigo.- Operational Procedure for 13 yd3 automatic LHD, Thesis Work, University of Chile, Santiago, Chile, 2003. 2.- Varas Fredy & Crorkan Paul.- Seminary of Automation, Technical Visit, Rovaniemi, Finland, April 2001. 3.- Varas Fredy & Pozo Rodrigo.- Operation & Security Procedure for 13 yd3 automatic LHD. Sernageomin. Santiago, December 2003. 4.- Varas Fredy.- Automation of LHD Equipment, Internal Presentation, CODELCO Chile Division El Teniente, Rancagua, Chile, December 2002. 5.- Varas Fredy.- Material Handling Tendency in Underground Mining. 21st Congress of Geologists and Mining Engineers, Acapulco, Mexico, October 2001. 6.- Varas Fredy.- Mineral extraction with Semiautomatic LHD Equipment in Block Caving Method. MININ 2004, International Conference-Mining Innovation, Chile, Santiago 2004.
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Mine technology and its implementation and control – Reservas Norte – Sub 6, El Teniente Mauricio Barraza, Juan Francisco San Martín, Manuel Montecino, El Teniente Division, Codelco, Chile
Abstract Mining is commonly seen as a typical basic industry with rough and even dangerous working conditions, heavy environmental load, and a low level of high-technology with much manual work and operations. An underground mining operation is a very complex one consisting of many manual, physical, mechanical and logistical operations with different human interfaces and decisions. For this reason it is a demanding and potential area for all applications of information technology: controlling difficult non-linear, time varying multivariable processes and machines; solving of ergonomic safety and environmental problems; automation of logistical systems, and information handling, etc. In addition to that, big economic sums are involved with these operations. On the other hand, technology has developed towards intelligent and adaptive systems including some from the above mentioned fields. Connecting them with modern information and communication technologies makes the idea of an intelligent and unmanned mine more and more feasible in the near future. This requires intensive co-operation and coordination between the mining industry and manufacturers of mining machinery. This has been successfully accomplished in the few-year Intelligent Mine research and development program as well as the last years of subsequent implementation of its results in El Teniente Mine. The basic elements of an Intelligent Mine are: • Mine-wide information and data acquisition system. • High-speed two-directional mine-wide communication network for real-time monitoring and control. • Computerized information management, mine planning, control and maintenance systems. • Machinery and equipment which are connected to the information system. The degree of high-technology in mine depends on many technical and economical factors. The basic precondition for such an approach is that it will improve the total economy of the mine.
1. INTRODUCTION
was done in a real mining environment with the assistance of the mine personnel.
The productivity of mines can be significantly improved and the costs of the final product decreased by applying the advanced mining technology and automation. Today emphasis is given to automation of production machinery and mine-wide information and communication including maintenance and production control systems. The objectives of the Implementation Technology Programs were to increase the productivity and profitability as well as improve the working conditions of the miners, equipment, communication and data transfer systems. A mining process controlled and managed in real-time most economically according to internal and outside conditions. All machines and activities are integrated by bidirectional high-speed mine-wide communication and data networks to enable the real-time communication, monitoring and equipment control. Mine planning, production planning and equipment maintenance planning systems are integrated by this network allowing all control and decisionmaking of the mine to be centralized.
The program was as follows (Table 1):
2. MINE TECHNOLOGY PROGRAM The development program was formed in co-operation between the mining and Modular Mining Systems (DISPATCH - SISPAC) the emphasis is on the requirements of the end user, the mines. The practical development work Massmin 2004
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The technology development program is a beginning for future development which will lead to a concept of the Technology Mine _- an automated processes and teleoperated machinery (Pick Hammer) which are controlled in real-time to provide the best possible economical production according to the internal and external conditions. The basic elements of an Intelligent Mine for Teneinte Sub 6 / Reservas Norte are: • Mine-Wide information and data acquisition systems. • A high-speed, two-directional, mine-wide communication and information systems network for real-time monitoring and control. • Computerized information management, mine planning, control and maintenance systems. • Tele-operated machinery and equipment connected to the mine-wide communication networks. • Communication and monitoring systems to other sector in the mine.
Figure 1: Development steps for automation and controlled process in mine.
All these elements could not be implemented in the same time. Therefore the development was done gradually according to the priority requirements of the mines and step by step to avoid ‘bottlenecks’ in the total production process (Figure 1). The research and development work was divided into main areas: • The real-time management of production. • Machine Tele-operated (Pick Hammer) • Production maintenance. • Safety, training and motivation. Under the real-time management of resources and production constituted information acquisition, communication and information transfer as well as the processing and utilization of this information for management purposes. It also contained the major part of the application of computers in mine-wide production planning and control (Figure 2). The projects included control of production machines in such a way that it resulted in increased machine-working hours, dependability and higher productivity. Included were fault diagnostics and data acquisition systems incorporated within the machines. Navigation systems, modular type of production monitoring and fault diagnostics were developed in mines. The production and condition control, fault diagnostics and monitoring systems were tested in the mine-wide intelligent condition control system (Figure 3).
Figure 2: Real - Time Management
Machines have become more reliable in operation. The main goal is to increase effective time. Features of technology machines and mine will include: • Remote operation (tele-operation). • Data communication. • Maintenance support. ç Task planning and reporting. The technology of production methods and production maintenance covered a large area in the program. It constituted such fields as machine maintenance, new hoisting methods and mine profitability analysis. New machine constructions such as a mobile underground crusher, an automatic LHD machine, an automatic haulage truck and charging equipment were also included in different area of development in the mine. Figure 3: Intelligent Production and Control System. 682
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Technology will bring new type hazards of accidents and have an effect on work motivation. Although safety has been taken into account within each development project, the human role in the environment, especially in a mine, has not yet been extensively studied. In this part of the program, the methods and tools for introducing automation successfully into the mine environment. Research and Results The period to develop the systems needed for try organize the Intelligent Mine was made in steps: first to investigate what was available and then develop the most urgently needed systems and machines and make them functional. During that time, the testing was done in Sub 6 / Reservas Norte an Esmeralda, the program was a success and a good start for implementing the technology.
3. MINE IMPLEMENTATION The Implementation Technology was development in the major part of the machinery and systems with MODULAR and Codelco – El Teniente, the Technology Program were being developed and tested to the stage in which they can be implemented in the planned underground mine of Teniente Sub 6 and Esmeralda (Figure 4).
• Systems integrated to the networks. • Machines and equipment integrated to the networks. • Mine infrastructure and personnel training. The program was realized by the management of Codelco EL Teniente, how contract Modular Mining System, and the program consisted: • Further development and testing of the machinery • Developing and testing necessary additional systems and computer programs: • Further developing and testing the compatibility of systems in order to form one integrated system suitable for the underground mine of Teniente Sub 6 / Reservas Norte and Esmeralda. • Planning and carrying out a training program for the operating personnel of the mine. During the implementation, the mine was made ready to apply the concept. The Teniente Sub 6 / Reserves Norte Mine acted as the testing ground for integration of all partial components into one functioning system and will be, in the future, the first mine to fully utilize the concept. 4. TENIENTE SUB6 / RESERVAS NORTE All of the projects were started in the beginning of 2000 (Table 2). Table 2: Mine Technology Program.
Figure 4: Integrate Service Communication Network
The results of this program were: readiness to realize the Intelligent Mine. The vision, have the machinery and systems developed ready for application, and show the know-how of old sector in the El Teniente called Teniente 4 SUR. The technology program lasted three years (2000 - 2002) and its estimated budget was KUSD 500. It contained 4 development projects, which were divided into four main areas. • Implementation of advanced technology. • Data utilization. • Training and adaptation of mine personnel to the new technology, systems and environment. In addition the program contains two supporting projects: • Co-ordination of the program. • Project concerning with layout of advanced mining. The objectives of the program were set according to the needs and requirements of the Teniente Sub 6 / Reservas Norte and Esmeralda mine. The contents comprises of the following: • Fast mine-wide communication and data networks (Figure 3). Massmin 2004
During this implementation stage the new technology was tested. Possible unknown faults or new requirements in systems and applications were recognized and corrected. The tasks, character of work, safety matters and organization will change. This might cause opposition and if not properly dealt with, can make the implementation of new technology difficult The mine’s organization has also been planned to suit the applied new technology and best utilization of the systems created (Figure 6). The organization structure of the mine as described cannot be very complex. The information is available to everybody and the decision-making of day-to-day production decisions is given to lower levels of organization. The management of the mine will have more of a supporting
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Figure 6: Layout and Equipment distribution and control in real time. and advising role. The lower management will deal more in the safety issues and be ready to give advice whenever necessary. The operators of the different sections work as teams and take responsibility of their own section. Different sections assist each other and support their services when needed. For example, the maintenance teams give services to the production teams (Figure 7). This way the responsibility will be shared. The final results of the implementation are not yet seen, but during 2004 many of the systems will be in practice. The Teniente Sub 6 / Reservas Norte Mine will most probably apply the technology that has been developed and has developed it further for its own use. Some problems are yet to be solved but as the mine development continues the solutions for these problems will be found. It is too early to
say that all applications of the systems, machinery and organizational changes will be a success but so far there have been no impossible situations. 5. FUTURE The future development at Reservas Norte mine will be in the consolidation of the Handheld PC used by the shift boss. PC that allows to have on line the information of the production and machines running on the field. This Handheld device makes possible the access to the same information that is available at the UNIX or main PC’s. To make this possible it was necessary to install a wireless communication system (access point) so that information can be captured and showed in real time for the
Figure 7: The Maintenance Products 684
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users of this device, the configuration of the wireless system is showing in figure 8.
Besides that efforts must be set in the integration of the total production area (LHD, secondary reduction and pick hammer) and haulage (Trucks) activities. 6. CONCLUSION • As a conclusion, it could be said that both of these Technology Programs have given much to the Teniente Sub 6 / reserves Norte mining and mining machinery manufacturing industries. • The mining industry has had a chance to develop its mine management systems, methods of increasing productivity and safety. • The mining machinery manufacturers have had a good opportunity to develop their machinery, technology and systems integrated to the machines and so to increase their competitiveness in the market. • The Company has had a good opportunity to increase and develop its know-how and so to be able to produce more confident mining engineers and workers.
Figure 8: Access communication for the real time information
The signal makes possible to move around the room offices at the Sub-6 level and through the main access, at the production area. The figure 9 shows in green where this wireless net is available.
7. REFERENCE • Barraza, M., and Zamora, A. (2001) "Automatización de la Extracción en Pannel Caving" IM2, Codelco Internal report, Brisbane, Australia. • Barraza, M., San Martin, J., Bonani, A., and Alvial, J. (2000) "Grupo de tarea planificación mina Sub 6", PL-I200/2000, Internal report, Codelco Chile División El Teniente. • Baiden, G.R. (2001) "TeleminingTM System Applied to Hard Rock Metal at Inco Limited" Underground Mining Methods, ed. SME 2001, pp 671 – 712, USA. • Brunner, D.T. (2001) "Simulation of Underground Mining Operations" Underground Mining Methods, ed. SME 2001, pp 705 – 679, USA. • San Martin, J., and Alvial, J. (2000) "Contrato con Modular Mining System’, Internal report, Superintendencia Mina 2000, Codelco Chile División El Teniente. • Sturgul, J.R.(2000) " Using Animation of Mining Operations as Presentation Models" Mine Planning and Equipment Selection, Panagiotou & Michalakopoulos (eds) 2000 Balkema, Rotterdam .
Figure 9: Net layout in the production level for wireless
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Codelco El Teniente - Loading automation in panel caving using AutoMine™ Vic Schweikart, Prof. Eng., M.Sc., MBA, Sandvik Chile Timo Soikkeli, Lic.Sc., Sandvik Tamrock Corp., Finland
Abstract In the world of commodities, copper is a highly competitive market that forces mining houses to continuously improve and optimise production. El Teniente’s objective is to use the best technologies to automate the production processes. Following an extensive review of equipment and system suppliers available worldwide, Codelco selected Sandvik Tamrock of Finland, industry leader in this field, to provide Sandvik Tamrock’s AutoMine™ system, a proven commercially technology, to automate the LHD fleets at the Pipa Norte and Diablo Regimiento mines. Pipa Norte and Diablo Regimiento deposits will be mined with pre-undercut panel caving at 10 000 t/d and 28 000 t/d respectively. Pipa’s fleet, the three Toro 0010C, is currently being commissioned. Delivery of the Toros for Diablo Regimiento mine started in February 2004 and will continue until 2016. Fully automated Toro 0010C LHDs will be operated from a single control room outside the mines. A powerful communication system transmits control and operating data between each unit and the control room. This system connects to the mines’ production and maintenance systems enabling high fleet utilization and quality draw control. The autonomous LHD fleet sets revolutionary standards of performance cost cutting and safety for underground ore extraction and transportation processes.
1 INTRODUCTION Monitoring, instrumentation and automation have become attractive technological choices in massive mining applications. Mining methods like block caving, where large amounts of material are moved during steady state production support a well automated operation. The higher degree of automation in a mining application will be best applied to a smooth and streamlined mining process and operation. The challenge is then to master the effect of process disturbances or discontinuities like secondary breaking. The obvious expected result of an automated system is the increase in utilisation of equipment and decrease in operative unit cost. To achieve this objective, several optimisation steps must be followed. The actual performance of the selected mobile equipment technology will rely on the support functions, infrastructure, and user interface and user acceptance. Optimising the technology to match the circumstances is an increasingly important phase in a mine feasibility study. The automated system must be functional in the given mining process and the mine layout applicable to automated operation. Only when the full system with machinery, process control, communication and operator interface has been identified, the final economical justification can be made. Careful evaluation of both technical and economical aspects of different degrees of automation in given circumstances should precede any automation commitment. The approach that Sandvik Tamrock and El Teniente mine have taken in evaluating automation for the Pipa Norte panel block cave mine included a review of the given processes and conditions and the suitability for automation and the evaluation of risks and critical points in the systems. Then, the actual automation feasibility by using the AutoMine™ solution was reviewed against the overall economy and rate of return of the given options and in further detail analyses the proposed automated system. 686
Mines are moving towards real time process control and resource monitoring systems. They combine the use of simulation tools for production forecasting and evaluating changing situations. This in turn means that processes are well defined and the resources carefully thought of. The resources will be increasingly more dedicated in autonomous system where moving a unit from one area to the other may not be possible at all. An increasing desire for lower cost and more massive operations moving to deeper deposits and decreasing human involvement will eventually lead to developing mining methods and processes more suitable for automation. 2 EL TENIENTE PIPA NORTE MINE Pipa Norte and Diablo Regimiento deposits will be exploited by pre-undercut panel caving at 10 000 t/d and 28 000 t/d respectively. In the pre-undercut panel caving operation planned for Pipa Norte and Diablo Regimiento, fractured rock gravitates through to drawpoints on the production level. Drilling and blasting or other secondary breaking methods will break oversize boulders that block the drawpoints. Pipa Norte’s fleet uses three fully automated Toro 0010C LHDs, each with a nominal capacity of 17,5t currently being commissioned. Diablo Regimiento mine will operate with ten Toro 0010C that will be commissioned by June 2005. Obviously, these two mines have their own characteristics, but both of them will have the same AutoMine™ solution, and for simplicity, this paper refers to Pipa Norte mine only. The Pipa Norte AutoMine™ system transports rock directly to the one crusher. Crushed rock is conveyed to an ore pass and then transported by train outside the mine for processing. The secondary breaking and loading process takes place 24 hours per day and 7 days per week. Operational delays including shift change, lunch and tea breaks constitute a major percentage of total LHD downtime. These delays are reduced in both frequency and
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duration by removing the local operator from the LHD. Additionally, removal of personnel from operational areas reduces the potential for accidents to occur. The local operator on board of the LHD is replaced by tele-remote control from a central control room located some 10km outside the mine for functions, such as loading, which require decision making. On-board computers control routine functions, such as tramming and tipping. Pipa Norte will operate a panel cave comprising 188 drawpoints arranged on 15 production tunnels (Figure 1). Initially, the automatic LHDs operate in the seven most northern tunnels of the mine (top of Figure 1), an area that is both static and concentrated. As time goes pass, production is taken from southern tunnels and the northern tunnels are released once their drawpoints are exhausted. Pipa Norte production area will be developed progressively from north to south and so will be the extraction of caved ore from drawpoints. There will consequently be little interaction between development and production crews, so that extraction operations will be of low complexity and routine. All underground infrastructure is new and designed for purpose. The fifteen production tunnels and the main haulage ("zocavon") that comprises the production area will remain discreetly static for the 10-year life of the operation. This situation encourages the establishment of a rigid LHD operating concept supported by the AutoMine™ system in order to achieve higher than average availability and utilisation factors.
production area (refer to bottom of Figure 1) and scheduled through the Mission Control of the AutoMine™ system. A team of maintenance personnel will take the LHD out of the autonomous area through the Transit Lock and the shift and daily maintenance functions will be performed while the machine is being serviced. Fleet availability is based on planned downtime hours to perform scheduled maintenance services and unplanned downtime hours to perform unscheduled repairs or breakdown work. There is no experience to accurately predict unplanned downtime in an automatic operation. However, with real time condition monitoring it should be feasible to achieve a diagnostic capability so that at least 15% of potential failures of the manual operation are diagnosed and rectified during existing planned downtime. It is assumed that this improvement offsets any additional downtime associated with the AutoMine™ system and the downtime is therefore about the same as manually operated LHDs. As the operator is located in the surface Control Room, the operator line-up time will be used to accomplish a 'hotseat change' with no time lost. During the shift, the operators will change out for rest and lunch breaks as required. Only one operator per shift is scheduled for all three Pipa LHDs. This same operator is responsible for the crusher hammer. However, this situation will be reviewed during the operations phase. Time lost during changes of operators has been estimated at five minutes per shift. From the cost point of view, the main advantage of operating automatic LHDs is primarily related to labour and maintenance costs. For example, Pipa Norte AutoMine™ system requires (four) 4 operators; if these LHDs would have been operated manually, it would have required 16 operators on the four shifts. On the maintenance side, the number of full time fitters required to maintain a fleet of automatic LHDs is similar to the requirements of a conventional fleet however, the former require a full time electrical/instrumentation technician per shift. It is also correct to say that the autonomous fleet requires a higher level of technical skills. Some conclusions can be made about the maintenance and operating costs of an autonomous LHD fleet from a combination of first principal (Predictive Operating Cost Model) and experience at Kiruna. Service and maintenance spares costs will be less expensive for automatic LHDs due to smoother operation and less frequent contact with the walls of the haulage. Tyre life is significantly higher, about 33%, for automatically operated LHDs, as proved at Kiruna. Bucket costs, fuel consumption per tonne of ore produced, and oil and lubricant costs in an autonomous LHD are about the same as the one for manually operated machines. 3 AUTOMINE™ OPERATIONS
Figure 1. Pipa Norte Mine Layout To achieve this, the LHDs will not leave the production area, except for planned maintenance events or major breakdowns, when they will be removed to the Diablo Regimiento workshop, some 500 meters away. The process of refuelling, topping up oil and lubricating the machines will be carried out at the south-east end of the Massmin 2004
The AutoMine™ system of three autonomous LHDs being commissioned at Pipa Norte was developed by Sandvik Tamrock to achieve the economic and safety objectives of removing the operator from the machine. A schematic of the major subsystems and interconnections between the systems is presented in Figure 2. When the autonomous LHD approaches the allocated drawpoint, the Mission Control System (MCS) alerts the operator that an LHD is about to require the bucket filled. The operator takes command of the LHD. After the teleremote mode is selected, the control system switches data and video signals so that direct communication is established between the LHD and tele-remote operator station. The tele-remote operator fills the LHD bucket using joysticks and pedals at the operator chair located in Colon.
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The filled bucket is then trimmed to drop off excessive ore, thus avoiding spillage on the road. After loading is finished the operator gives the 'go to crusher' order, the specific crusher tipping point having determined by the system. When the bucket is emptied, the LHD alerts the MCS that it is available for a new drawpoint assignment. The MCS informs the system that the LHD has completed its dumping mission and waits for a mission to tram to a drawpoint. Upon receiving a new assignment, the LHD exits the crusher tip and trams autonomously to the assigned drawpoint. MCS also prepare the automatic LHDs for routine re-fuelling and shift services as well as The LHD trams autonomously from the drawpoint to the crusher and back under the navigation system. The LHD is in Automode. The operator can, at any stage switch the LHD to Remote Standby, Semi-Auto or Remote mode if required. The switching takes effect on-line without stopping the LHD. While tramming, the MCS controls the LHD path, performs traffic control and communicates the LHD status, location and heading to the system. When an LHD approaches the dumping position at the crusher tip, a water mist spray system is activated automatically. In the event of the tip becoming unavailable when a loader is on its way, the Mission Control System will alert the LHD to avoid tipping. The LHD will not tip if the hammer is in use or the crusher bin level is high. In the event of an LHD breakdown the system informs the workshop and the Control Room, then exclude the LHD from the assignment logic and prepare the LHD for manual control. Other LHDs will be excluded from the tunnel where the breakdown has occurred. The Access Control System, ACS, performs a shutdown procedure and a "Safe Path" so that maintenance personnel can gain access to the machine. The production area isolation is achieved with both physical gates and optical barriers that are part of the
Access Control System. The gates are constructed so that they physically prevent human access into the isolated production area. The optical barriers detect any entries into the production area and will stop the LHD fleet if one of them activates them. The Access Control System for the production area safety barriers is an independent system and operates as a supervisory system to the MCS. The production tunnels isolation is done per unit by using "Access Gates". An Access Gate includes a physical steel gate with an electric lock that is engaged when the gate is activated. The lock also detects that the gate is closed and locked with feedback to the Access Control user interface in the Control Room. Also, attached to the gate are zone status lights and a horn that indicate the status of the autonomous area and warns personnel, respectively. An Access Gate also includes one set of infrared barriers to allow entry for service and maintenance personnel without stopping the production in other tunnels. The main haulage drive, "main zocavon" is the backbone for the autonomous LHD traffic. It is therefore free of any isolation device. The Access Gates field cabinet and control box each have a fast stop button that stops all autonomous machines in the production zone when pressed. Pressing a fast stop button also disengages the electric lock to permit evacuation of the area in a mine emergency situation. The field cabinet and control box also have a key switch that is used to activate the gates. This ensures only authorised personnel with a key can activate the ACS. The whole production area isolation is achieved by means of the "Transit Lock". A transit lock is a double gate arrangement used to introduce and remove an autonomous machine from the autonomous area whilst allowing other autonomous machines to continue operating.
Figure 2. Semi-autonomous LHD Subsystems. 688
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The transit lock utilises two interlocked access gates. The outer gate, on the manual zone side, is opened and closed manually by authorised underground mine personnel. This gate is kept closed during the automatic operation. The inner gate, on the autonomous zone side, has electric motors to enable the gate blades to be opened and closed remotely from the field cabinets. The inner gate can also be opened from the ACS user interface in the control room. The Transit Lock uses the same components as an Access Gate with the addition of an ultrasonic sensor for detecting when a machine is in the transit lock. Furthermore, a transit lock has an identification system used to detect the ID of the machine so that the system knows which machine is being introduced or removed to/from the autonomous area. Violation of the isolated production area by opening a physical gate or crossing an infrared barrier stops the automatic LHDs operating in the area and also creates an alarm at the Control Room. The tele-operator at the Control Room also has the option of reaching the production area by radio or telephone (the latter at the Transit Lock only). When a production tunnel is allocated for automatic operation, it is isolated by opening the infrared barrier in the main haulage end. The optical barriers are also disabled at the main haulage end. In the event of fire, all automatic LHDs go into emergency stop status and all safety barriers are opened for free evacuation.
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4 CONCLUSIONS Sandvik Tamrock has developed the technology required to implement automatic loading, and it is commercially available under the name of AutoMine™. AutoMine™ is being commissioned at Pipa Norte mine at El Teniente in Chile at the time of writing this paper. In parallel, another AutoMine™ system is in its early installation phase at Diablo Regimiento mine, also at El Teniente in Chile. A major benefit is the removal of the operators from the underground to a safe controlled working environment on surface. A financial evaluation indicates significant potential cost savings for automatic operation. The additional costs of the system and the skilled labour to maintain it are offset by a reduction in the number of operators, greater LHD utilisation and a reduction in spares consumption. REFERENCES • Internal Report Sandvik-Tamrock, Document Number 62100-2, 2003. Automated Hauling System, Technical Description. • Schweikart, V., Control and Monitoring Systems at the Palabora Underground Mine, Telemine 1, June 1999, Sudbury, Ontario, Canada.
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PCRB-Surface control centre for mine automation Fredrik Kangas, Technical Development Sub-level Caving and Transportation Level, LKAB Carl-Erik Emmoth, Superintendent Sublevel Caving and Transportation Level, LKAB Pär Lindahl, Vice President of Ifa Production Development AB
Abstract The development and implementation during the last years of advanced technology such as remote control and automation within the Kiruna mine has changed the character of mining. Mine operation is now becoming more and more alike any advanced process industry. In a development project carried out jointly by LKAB and Ifa Production Development AB these new pre-requisites have been full. Exploited, A new concept of production control, process management and organisation has been developed where all production units will be linked together in a joint Production Centre Kiruna, placed at surface level. The corner stones of the concept are using the benefits of modern technique in combination with taking advantage of human skills for key tasks. In a production centre the personnel that control and co-ordinate the operations are placed. The Production Centre is in turn linked to Operation Stations at key areas in the mine for specific tasks such as loading and haulage. The concept gives room for a constant development off skills and offer opportunities for continuance improvements off the mining process The mining operators in the production centre will control and supervise the load to hoisting process by remote control and be in charge of planning and delivering of ton and quality from the ore factory on a weekly basis to the refining plant above surface. 1. INTRODUCTION Kirunavaara iron ore mine is the third biggest underground mine in the world and is located in the north of Sweden about 200 km north of the polar circle. The mine daily produce between 60 to 75 kton raw material iron ore in three qualities containing more or less Phosphor, Potassium and waste rock. The mine started to produce about 1902 in an open pit mine and went underground in 1962 and is today a highly automated mine. The production technique is sub-level cave mining with draw points mainly at the 820 meter level, loading to ore passes down to the 1045 meter main level for haulage by train to crushers and then hoisted in two steps before surface. Every year the automated machines drill 70% of our production drilling, load 30% of our raw material iron ore and remote controlled boulder breaker treats 30% of our boulder on grizzlies. We remote control all chutes when we load our trains on the fully automated main level and hoist it to our refining plants above surface. Kiruna iron ore mine is a few years from fully automated production and taking the step over from traditional mining to general process industry. We are in this very minute operating the automated process from above surface and preparing the organisation to work in a new way. The organisation is based on the concept of providing power of decision and skills night and day through process focused organisation. 2. PROJECT HISTORY At 1999 the first semi automated and remote con-trolled LHD was tried out in the Kiruna mine. About one year later we first observed how little influence the automation had on the operators. The only difference was that the operator did not sit in the machine. In every other way he was still working like he was sitting in the LHD far from the process information that now was available. 690
Figure 1: An ilustratation of the Kiruna mine ore body with infrastructure for transport of personnel, ore and air. From that observation the curiosity made us look over our total process from mine to harbour and we found out that the operator in most cases worked in a similar way. Every operator optimised his own process with little regard of the process before and the process after. Several examples were found that showed weaknesses in communication between the process steps with production losses as result. The investigation lead to a project with the goal to give an alternative solution to monitor and steer the process from mine to harbour. In 2001 the project delivered a case model of how the process could be controlled from a production centre. And the same year the project group was asked to realise the idea in a first step were the mine should be operated from a production centre above surface. The development of the practical solution took about two years and involved more than sixty people, mainly workers. To succeed with the idea we had to create a will of change in the organisation so we could take full benefits of motivated
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and high knowledge of its own area. The daily production relies on those teams. To overcome the distance between the teams we support communication with regular meetings and the same process information in real time solutions. The production team that has its home in the Production Centre is working with a strong focus to optimise the process. The Production Centre is the heart of the production assignment. The process is supported by high quality maintenance that keeps the facilities and machines in full production. That service is given by Operation Centres that perform mainly scheduled maintenance but has a central function in teams that perform supervision and emergency maintenance on the facilities. This Operation Team is a link between process and facilities and is in constant communication with both the Production Team and Supporting functions as scheduled maintenance, production system maintenance, technicians, etc. If you can take a decision, take it! The key is to give as many as possible the information and knowledge to take it. Figure 2: Over view picture that describes remote controlled loading system from navigation to control. employees who are qualified to take fast and accurate decisions. Employees who supervise and operate the process and maintain the facilities to achieve right quality in the raw material ore and high reliability in the facility. 3. PRODUCTION CENTRE CONCEPT The Production Centre Concept introduce enrapture to traditional mining technique where every operator or driver is working with a well-defined part of the process and is an owner of his specific facility or machine. The traditional solution creates constant sub optimisations and depends strongly on high accuracy in over all planing. Every day things happen for a single operation that makes it impossible to produce according to the original production plan. The overall process is thus often disturbed. The process becomes dependent on a strong staffs and special functions that supervise the process and takes the decision on how to operate every part of the process. The solution has long lead times and poor motivation, and does not support own initiatives from the mining personnel. In the same way the strong relation between op-erator and machine affects in a negative way the contact between operator and maintenance personnel. The Production Centre Concept in a way separates production and maintenance in teams that have a clear focus
4. CONCLUSIONS The conclusion of the project is that automation of a process is far more than expensive investments in sophisticated technique. Often companies have long time plans fore technical development and investments but ignore the fact that we need qualified staff to operate and maintain the production system. When the production speed increases and the quality expectations accelerate many companies stand without the personnel to match the expectations. The result is long payoff times on investments and higher risk to deliver low quality products during the rebuilding and adaptation of the organisation to the new production system. You will never regain the losses, and in most cases never reach the optimised solution you will get when you have an organisation that asks for automation and better technical solutions. Only a thoroughly designed organisation with high degrees of decision making power on the right level can give you that benefit.
Figure 4: Mining industry has the possibility ta take huge steps in automation level but it takes big effors to succeed. That hardest struggle will be organisation related. If you get the right organisation around the produc-tion assignment a good guess is that you can reduce the time to gain full production on a new facility with 30 to 50 percent and increase the full capacity on a new or old facility with 5 to 20 percent.
Figure 3: Production Centre Concept in Kiruma mine. New York places that are tools for the self-learning organisation described in the right corner. Massmin 2004
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Introduction of autonomous loaders to Olympic dam operations, Australia Charles McHugh, Group Manager Mining Technologies, WMC Resources Ltd.
Abstract WMC Resources Ltd has been involved the development and trialling of an automated load haul dump vehicle system since 1997. This paper describes the introduction of the system into the Olympic Dam Operation and its performance compared with existing manual loaders. In certain situations the autotram loader will outperform the manual operation, particularly on the longer distances. There were no personnel injuries associated with the operation of the system during the testing period. The major operational issues with the system are discussed and the direction for future improvements outlined.
1 INTRODUCTION The Olympic Dam Operation (ODO) is owned by a subsidury of WMC Resources Ltd (WMCR). ODO is Australia’s largest underground mine located some 550km north of Adelaide in South Australia. The orebody was discovered in 1975 and had undergone several expansions with the last one completed in 2002. The mine is polymetallic and produces 200,000 copper tonnes per year along with uranium, gold and silver. The mine produces approximately 10 million tonnes of run of mine ore per annum from open stoping methods and backfilling with cemented aggregate fill. The mine has highly automated transportation systems that includes an automated underground train and shaft. The mine has been involved in the development and testing of an automated load dump vehicle system for four years. 2 HISTORY OF PROTO-TYPE SYSTEM DEVELOPMENT WMCR had sponsored two automated loader projects from as early as 1997. One was with the Company, Lateral Dynamics (LD) and the other was the Australian Mineral Industry Research Association (AMIRA) Project P517. LD had tested a proto-type at the WMCR Perseverance Mine on an Elphinstone R1600 and it had proven to have greater productivity than tele-remote systems in retreat longhole stopes by at least 40%. LD formed a joint venture company with Caterpillar. The company was called Dynamic Automation Systems (DAS). DAS under licence combined parts of the AMIRA P517 into their system. The system is now marketed as MINEGEM™. The DAS system was taken to ODO for testing in 1999 and fitted to an Elphinstone R2900 loader. The control centre was set up underground near the production area. The loader had 18,000 hours of service and was very old. There were problems of control when speeds were increased over 5km/hour. In March 2001 WMCR decided take the DAS system to the Queensland Centre for Advanced Technologies (QCAT) to determine the cause of the speed problem. It was determined to be a combination of latency from an internal remote control modem and hydraulic pilot valve. By April 2001 the DAS system had demonstrated that a R2900 loader was capable of speeds over 15km/hr in third gear at the QCAT test track. 692
The DAS system was transported back to ODO for production testing. The control centre was set up on surface and communication to the production area was via optic fibre and microwave radio network. The cable transmission was 100 Mb/s between radio cells and 11Mb/s for the wireless ethernet between the autotram vehicle using 802.11b standard. The small latency in the on board cameras did not affect tele-remote bogging performance. One week after re-commissioning at ODO, the loader was involved in an uncontrolled movement. The loader drove 25 metres out of the production control area, through the safety barrier gate, colliding with a Volvo service loader and running over a light vehicle. The loader came to rest in a stockpile and shutdown before any analysis could be performed. All automation testing was immediately suspended and a thorough investigation was performed. Ironically this was one of the most computer monitored mining accidents at the mine but it remains unclear as to the exact cause. The investigation concluded the most likely cause of the accident was a severe shock in the lowering of the bucket during loading causing the on board computer to crash. The computer continue to output the final control signal of first gear reverse. The continual collision with the wall as the loader was moving prevented the computer from doing a complete reboot. When the loader stopped in the stockpile it was able to reboot and immediately shutdown as it had lost communications. The system was redesigned based on the investigation with the following important safety features. • A completely independent safety system (ISS) that shuts down the fuel solenoid. • The ISS is a Category 3 European rated and Australian Standard rated electrical safety system complying with EN 954-1 and AS4024.1 respectively. • The onboard computer is a PC104 industrial computer. • An independent onboard mission recorder similar to an aircraft "black box". • Software re-written using real time operating systems. The system was then tested on a new R2900 loader at the Caterpillar mine simulation facility in Burnie, Tasmania for 200 hours. It was dismantled and sent to ODO for further testing in July 2002. A new R2900 was hired for the duration of the first trial in the 42 Orange 20 Stope. The display of the unit from the ODO autotram control room is displayed in Figure 1.
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Figure 1: Autotram control room showing orepass grizzly, vehicle health, cameras, map route layout, independent safety system status and control stick.
3 PRODUCTION TRIAL AT THE 42 ORANGE 20 STOPE The loader was setup in the 42 Orange 20 stope with 24 hour coverage using 2 x 12 hour shifts. The route included a 100m section with a 1:8 gradient. It was necessary to make two, three point turns during a full cycle due to the position of the safety barrier gate. The stope had two active draw points. The results are displayed in Table 1. Initial results using a 15 tonne per bucket assumption were very encouraging. The average over 18 days or 36 shifts was just over 80 buckets or 1160 tonnes per shift. Table 1:Summary of Automated Production Trial 42 Orange 20 Stope Operating Hours Total Cycles Average Cycle Time (min) Tonnes One Way Distance (m)
Average of Cycle Time 266 2,786 3.75 41,790 220
The 42 Orange 20 stope contained a low amount of oversize rocks. The operators achieved very good productivity even during on the job training. Within four shifts operators were generally very comfortable with the system. The tramming and dumping were automated so the variation in terms of buckets per hour was very low. The tele-remote bogging took an average of 45 seconds. The system was influenced more by factors other than the skill of the operator. The Caterpillar Autodig™ was used in conjunction with the automated tramming but was discontinued after three shifts because of the following reasons. • The large variation in rock sizes which seemed to pulse between large and small made correct manual setting of the Autodig™ difficult. • The Autodig™ managed to fill buckets in one pass approximately 30% of the time. • The Autodig™ would try to straighten out the machine while bogging on left full lock causing the rear end to swing into adjacent wall during reversing. The use of Autodig™ was discontinued as tele-remote bogging is more efficient even until this time.
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The speed of the loader was tested up to 20km/hr in 3rd gear but this speed was limited by second gear during production due to concerns about road conditions. The speeds were determined by the operators based on manual operation experience. The maximum speed obtained in 2nd gear reverse was 14km/hr. It was observed that if the machine was programmed to attempt a corner too fast in reverse then it would have a "rear bucket swipe" collision. This would cause the bucket teeth to touch the wall. The major collision damage was caused during tele-remote operation with operators breaking off bucket teeth adaptor plates. This would take approximately five hours to repair. Collisions with the wall under automated mode were very minor. This indicates that the major collision damage occurs in the 5m approach to the muckpile during tele-remote operation. The autotram system was shown to operate between shift breaks and even during major stope firings. Shift handover took just a few minutes on surface. The system was able to work through major dust episodes in which normal manual operation would have ceased. Dust was generally caused when there was major rock pile movement of an open drawpoint. Re-fuelling was performed at mid shift at the mine central refuelling point by the underground maintenance crew. This allowed operation between shifts. The fuel capacity was for 14 hours continuous production and the operators adjusted re-fuelling schedule according the maintenance schedule. Road maintenance was scheduled during refuelling. The guidance lasers were cleaned during servicing with a clean rag. The maintenance personnel commented that in some places the road was so badly damaged it was not possible to drive manually faster than first gear at 5km/hr without causing injury to oneself, whereas the autotram loader had been travelling at 12km/hr. Towards the end of the testing the stope was almost empty. Normally this would require tele-remote bogging then manual tramming to the grizzly. In this case it was decided to continue using the autotram loader reducing the need for double handling. An early comparison was performed when road conditions were good between a manual loader and the automated loader for 15 cycles for instantaneous production rate and results are displayed in Table 2. Even with variations in rock sizes it can be said that manual loader was quicker however the following observations were made. • The operator could not sustain the rate of production due to operator fatigue. • The operator was allowed to use third gear. • The operator used a sightly quicker path because of no path restriction on the safety barricade gate. • Autotram dump was set at 30 seconds (this was reduced later to 12 seconds).
Table 2: Average Instantaneous Cycle Time 42 Orange 20 Stope
Average of Cycle Time
Automated Loader
3.75 minutes
Manual Loader
2.83 minutes
A high level comparison of the autotram at the 42 Orange 20 stope and the average of the manual loaders across the whole mine at that time was performed. It was found that the autotram loader was 67% productive in terms of tonnes per shift per loader but 175% in terms of tonnes per shift per loader per stope. The results are displayed in Table 3.
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Table 3: Comparison of Manual With Automated 42 Orange 20 Stope Tonnes/Shift/Loader Stopes Number of Drawpoints Tonnes/Shift/Loader/Stope One Way Distance (m)
Manual
Automated
1740 2.7 6.75 660 220
1160 1 2 1160 220
The automated loader was not able to move to other available areas when oversize rocks were encountered or roads degraded as there were no other communication network setups. All situations had to be dealt with immediately. Manual loaders were able to move to other stopes when they could not deal with the problems. The large amount of oversize rocks in all stopes required a mobile rockbreaker operator to work as a team with the manual loader operator. The automated loader operator had no rockbreaker because the rockbreaker at the time was not remote controlled. Large rocks were broken by repeat dumping on the grizzly. If this was unsuccessful then rocks would be transported to a popping bay. Estimation of loader productivity lost due to absence of rockbreaker by production personnel was approximately 400 tonnes per shift. The autotram has demonstrated by ensuring that the loader stays in one stope it is more likely that operators will follow the correct sequence of mining The loader was taken to a new stope to commence work. The loader was involved in a collision with the wall during some routine testing that resulted in severe damage to the unit including twisting of the chassis. An investigation revealed the unit was not in automated mode. The loader was removed from service and the MINEGEM™ system installed on an older production loader.
4 PRODUCTION TRIALS AT THE PURPLE STOPES
operator free time between tramming cycles allowed for very accurate recording of the causes of production downtime. There were no reported injuries associated with the autotram loader during the production trial. In the longer haul distances, operators reported of becoming bored and had taken to reading between bogging. A subtle but important change had occurred, the operators had become supervisors. 5 MULTIPLE AUTOTRAM TRIAL AT THE 56 AMBER 25 STOPE In August 2003 in the 56 Amber 24 stope, a second unit was fitted with MINEGEM™ and two loaders were demonstrated to be operating in the same area, controlled from surface by one operator. The map route is displayed in Figure 2. The traffic control system while about 50% reliable showed that operator intervention provided almost no lost time in productivity. Over 430m it was demonstrated that the productivity was approximately 1.8 times the estimated time for one autotram loader due to queing. Queing generally occurred if the cycle was delayed due to difficult bogging. The cycle time average was 9.5 minutes. The traffic control system calculated the most appropriate place for passing to optimise cycle times of both units. The barrier gate was replaced with a laser system and warning sign. This significantly reduced setup time. Manual tramming to the orepass was considered too fatiguing for the operators. During this trial the Co-operative Research Centre for Mining based in Brisbane, Australia, had developed an underground haul road duty meter with a road severity index. This unit was fitted to one of the loaders and was able to map the roughness of the road during each cycle. It could be seen there was greater roughness at the drawpoints and turning points as the road deteriorated. Future work will focus on productivity versus road quality and maintenance regimes. The trial was suspended after one week and 10,000 ore tonnes removal due to a major misfire in the stope.
A four month production trial was performed from 10th of October 2002 to 17th of January 2003. A summary of the key performance indicators are given below in Table 4 by Bryan (2003). Table 4: Summary of Automated Production Trial
Operating Hours Cycles Average Cycle Time (min) Tonnes One Way Distance (m)
Purple 21
Purple 71
363 5,723 3.8 71,556 178
134 748 10.7 11,489 349
After initial testing the total system demonstrated availabilities over 90% with the main cause of down time being the radio cell reliability. The biggest cause of utilisation downtime was breakdown maintenance and oversize rocks. A comparison between average manual stope loaders and single unit automation was made, allowing for factors such as large rocks, hang-ups and full passes. Over the four month trial it was shown that the average manual loader worked approximately 4.8 hours per shift and the autotram worked 6.7 hours per shift. The instantaneous production rate of the manual loaders was up to 30% higher but the autotram cycle time had significant room for improvement in the same order of magnitude. The computer monitoring and 694
Figure 2: 56 Amber 24 multiple autotram loaders with haul road duty meter fitted.
After this trial it became imperative that a form of remote rock breaking at the grizzly was required to ensure clearance of large rocks. A project was formed in November 2003 to remote control a Caterpillar 325 excavator with a rock breaker attachment.
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It was successfully controlled from the surface using existing communication infrastructure. It is now possible for one person to control two loaders and one rock breaker at the same time. Without automated digging it was felt three machines would be the maximum number to be operated by one person efficiently. 6 COST BENEFIT ANALYSIS A single MINEGEM™ system would cost approximately 40% of the capital for a new 20 tonne capacity load haul dump vehicle. This would reduce to 25% for multiple units. This assumes three radio cells per area. The training time for new operators with some computer experience is a few days compared with manual loader system of several months. When one operator begins to operate several units the advantages become very apparent. Tele-remote systems are cheaper in capital however the MINEGEM™ offers productivities at least 40% greater. The support for the system requires a set of modular hardwared spares kept at the mine site. DAS provide 24 hour coverage via the internet for system trouble shooting which reduces the need for a permanent on site technician. Important achievements during autotram loader trialling include: • The unit has performed up to 235 cycles in a 12 hour shift over a 70m haul distance. • The unit has performed up to 155 cycles in a 12 hours shift over a 220m haul distance. • The unit has performed up to 70 cycles in a 12 hour shift over 430m haul distance. • The system has been used on 8 different loaders of different sizes and manufacturers. • The system has been used in 7 different stoping areas. • The system can installed in one week if the optic fibre is available in the area. The integration of the autotram system into the mine must be considered carefully as initially there will be large production penalties as manual system interactions place restrictions on the automated systems.
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7 CONCLUSION The MINEGEM™ automated loader system at ODO can be considered very successful although not fully integrated into the mine. It cannot yet be said that the automated system will outperform manual system in every situation all the time. The anecdotal evidence suggests that the autotram will outperform manual operation in distances over 200m. There is no comparison with tele-remote systems in terms of productivity. The system has shown to be reliable and accepted by operators. The training time to learn the system and get to full productivity is less than 10% of manual systems. The additional 25% cost in capital for multiple automated systems must be clearly considered in terms of mine integration to ensure maximum productivity is achieved. The change to the automated system requires full management and operator support if it is to become the new business process. The following issues in order of importance must be addressed in the future to increase the productivity of the system. • Methods for remotely removing large rocks at drawpoints. • Road construction and maintenance. • Bigger fuel and grease capacity. • Smart automated digging. ACKNOWLEDGEMENTS The author would like to thank DAS, QCAT and ODO mine personnel, especially J.Lachmund, J.Lever, J.Kerr and D.Leonard for having the positive attitude to make the vision work. A special thanks to I. Bryan for mentoring everyone during the trials. REFERENCES • Bryan, I, 2003. Auto-tram Loader Transfer Project Close Out Report. STEM Partnership Consultant Report 25p.
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Use of the modular dispatch system to control production operations at the DOZ block cave mine Rudy Prasetyo, Superintendent Cave Management, David C. Flint, Technical Expert – Cave Management, Tarsisus B. Setyoko, Chief Dispatch EngineerP. T. Freeport Indonesia Eddy Samosir, Project Engineer, Strategic Planning, Freeport McMoRan Inc.
Abstract The Modular Mining Dispatch software and hardware systems have been installed at the DOZ Block Cave Mine to aid production operations. The system is being utilized to implement the daily detailed production plan, prioritize the location of production for each individual loader, collect actual drawpoint production data, provide real-time drawpoint status data and to measure loader productivity. Discussion is provided as to how use of the system benefits each of these points. This paper also describes the software and hardware configurations, as installed to control extraction and haulage operations of the mine. The Intelimine Dispatch software logic has been customized for DOZ conditions and production control requirements. The hardware system consists of a central computer, a micro-cell network and field computerswhich are installed on each mobile equipment piece. Communication between the mobile equipment and the network micro-cells is by UHF radio transmissions. Hubs manage the communications between the equipment and the system’s central computer, and monitor vehicle vital signs. Dispatch operations are monitored by Dispatch Engineers from a central underground location. Use of the Dispatch software has increased the reliability of production data and compliance to the production plan, and allows for a higher degree of management of the block cave.
1 INTRODUCTION The Deep Ore Zone (DOZ) block cave mine, operated by P. T. Freeport Indonesia (PTFI), is located in the Ertsberg District, West Papua Indonesia (Figure 1). Development of the mine was initiated in 1997, production commenced in 2000 (Barber et. al, 2000), and currently produces at 40,000 tonnes per day. DOZ is the third level of block caving to exploit the coppergold Ertsberg East Skarn System (EESS). The Gunung Bijih Timur mine (GBT) exploited upper levels of the deposit (Owen, 1992), from which 68.7 million tonnes at a grade of 1.93 % copper was extracted. The Intermediate Ore Zone (IOZ) mine produced from the deposit between the 3456 meter and 3706 elevations, with 43 million tonnes of ore on an average grade of 1.21 % copper and 0.43 g/t gold extracted. The Intellimine Dispatch system, of Modular Mining Inc., has been installed at DOZ to control block cave extraction and haulage operations.
2 THE DOZ BLOCK CAVE MINE The DOZ mine exploits the Ertsberg East Skarn System (EESS) deposit between the 3470 and 3120 meter (base of the IOZ mine) elevations. Draw column heights are 350 meters where below the IOZ, and a maximum of 500 meters elsewhere. The DOZ block cave mineable zone extends approximately 900 meters along strike and varies between 200 to 350 meters wide. The mine is an advanced undercut, mechanized block cave, utilizing truck haulage and a gyratory crusher. The extraction level has been developed at the 3120 meter elevation. Panel drifts are oriented perpendicular to the strike of the deposit and are developed on a spacing of 30 meters. Drawpoints, developed with the 696
Figure 1: Ertsberg Mining District Location Map.
herringbone layout, are spaced at 18 meters along the panel drifts (Barber, et. al. 2001) (Figure 2). Currently, DOZ produces from 240 drawpoints, accessed from 15 panel drifts. The undercut level is developed 20 meters above the extraction level. The advanced undercut method is utilized for caving. Caving has been completed for the eastern portion of the mine, and will be initiated for the western portion in the 4th quarter of 2004. Block cave ore is delivered from drawpoints to orepasses by load-haul-dump (LHD) equipment. A four (4)-meter diameter orepass is located in the center of each panel drift, as ground conditions allow (Barber et. al, 2001). In addition, three (3) orepasses, two (2) fitted with rock breakers, have been constructed along the extraction level fringe. All orepasses bottom at truck-loading chutes at the haulage level.
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Figure 2: Eastern DOZ Extraction Level Showing Dispatch Hardware Installations
Figure 3: DOZ Haulage Level Showing Dispatch Hardware Installations The haulage level is located at the 3076 meter elevation. Truck haulage is through a limited access, one-way traffic, racetrack-type loop system between the orepass loading chutes and the crusher (Figure 3). The haul trucks dump directly, at one (1) of three (3) dump positions, into the Fuller-Taylor 1,372 x 1,956mm gyratory crusher (Barber et. al., 2001). The production equipment fleet is comprised of Elphinstone LHDs, and Elphinstone and Toro haul trucks. Massmin 2004
Typically, 15 production loaders and six (6) haul trucks at active at any time. 3 DOZ DISPATCH SYSTEM DOZ employs the Intellimine Dispatch software system of Modular Mining. Intellimine is a large scale, computer based mine management system that allows PTFI to monitor and control production loaders, trucks and secondary sizing drills.
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Figure 4: Schematic of the DOZ Dispatch hardware network The hardware system consists of a central computer, a network of micro-cells, equipment field computer systems (FCS) and fixed radio frequency (RF) tags. Communication between the FCS and the micro-cell network and the RF tags is by UHF radio transmissions. Hubs manage communication between the equipment and the central computer system. Dispatch operations are monitored by Dispatchers, from a central underground location (Figure 4).
For data packages transmitted by the equipment, the process works in reverse. Micro-cells, with dipole antennae, have been installed at strategic locations throughout the extraction and haulage levels (Figures 2 and 3). The micro-cells are internally connected by token-ring cable, and networked to the master hub by fiber-optic cable.
3.1 Computer System The Intelimine (version 5.1) Dispatch software operates on two (2) Sun Ultra30 workstations, running on the Solaris 2.7 operating system. These servers, one as a backup, are situated six kilometers from the mine. Fiber-optic cable links the server to the mine Dispatch hardware network. Dispatchers monitor operations on two (2) emulator monitors located in the mine. The Intelimine software has been highly modified to suit the block cave environment and the DOZ mine configuration. It has also been customized to include communications in the Indonesian language.
3.4 Field Computer System A field computer system has been installed on each mobile production equipment piece. The FCS receives and displays instructions from the central computer system and also accepts input from the equipment operator. The FCS is comprised of a: • UHF data radio (inside the hub) and antenna, to transmit and receive signals. • Hub, which manages communication between the microcells and RF tags. • Touch-screen graphics console (GOIC). • Tag reader, integral to the hub, and antenna, to communicate with the RF tags.
3.2 Master Hub Two (2) master hubs manage and distribute communications between the mine equipment and the central computer. The master hub reads the data packages sent from the server and routes each to the correct microcell loop. Likewise, the hub organizes data generated from the field computers and forwards it to the central computer. The master hubs are connected to the micro-cell networks and the PTFI network by fiber-optic cable. 3.3 Micro-cell Radio Network The micro-cell radio network consists of a series of lowpower radio modems that run on a simplex radio channel operating at 451 MHz. As information is received from the system (via the master hub), the micro-cells read the data packages to determine the intended recipient of the communication. If for equipment with which a particular micro-cell is in radio contact, the data package is translated to radio frequency and relayed to the accompanying antenna. The antenna then transmits the radio message to the equipment FCS. 698
The function of the FCS hub, like that of the master hub, is to manage and distribute data between the equipment FCS components and the micro-cells. The hub is comprised of a processor, interface card, data radio and tagreader. The interface card translates data between the digital and radio format. The tag-reader, and associated antenna, transmits radio signals that interrogate fixed tags to determine location. The data radio communicates radio signals with the micro-cells. The GOIC is a 640 x 480, VGA compatible, touchsensitive screen that displays instructions from the central computer and allows for the operator to transmit data to the central computer and messages to the Dispatcher. The unit is sealed, has no moving parts, and is physically robust. It can be easily configured for truck LHD, or auxiliary equipment operation, and can display messages in either text or icon mode. 3.5 RF Tags Passive radio frequency tags have been installed at drawpoints, orepass collars, orepass chutes and at each of
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the three crusher dump positions. These tags have been programmed with a unique identification that is crossreferenced to the physical map location. The RF tags are read by tag readers, which are a component of the FCS, to determine location in the mine. Tags installed at drawpoints are covered by rubber to protect from secondary blast damage. The performance of the RF tags is fundamental to determining the specific location of production, so tag integrity is monitored by the Dispatch Engineers on a regular basis. 3.6 Extraction Level Network Three micro-cell loops, connected to the master hub by fiber-optic, have been installed at the DOZ extraction level. Micro-cells are located along the north and south perimeter of the level. A dipole antenna, connected to the micro-cells, has been installed at the north and south of each panel drift (Figure 2). Passive RF tags have been installed at each drawpoint and orepass collar. 3.7 Truck Haulage Level Network A network of twelve (12) strategically located micro-cells with antennae, also connected to the master hub by fiberoptic cable, has been installed though out the truck haulage level. Passive RF tags have been installed at each loading chute and crusher dump point (Figure 3). 4 DISPATCH CONTROL OF PRODUCTION The DOZ Dispatch system is utilized to communicate the production plan to the equipment operators and to collect data on the details of actual production. A production control system was chosen for DOZ due to challenges experienced at the GBT and IOZ mines, where production plans and actual performance data were transferred between engineering and operations by paper copy. The integrity of some of this information was suspect, and therefore resulted in inaccurate reporting of daily mine tonnage and grade, and calculation of remaining drawpoint reserves. The quality of long-term production planning and overall cave management suffered, accordingly. 4.1 LHD Dispatching At the beginning of a production shift, the Dispatcher uploads the detailed production plan generated by the Cave Management System (CMS) production planning tool (Samosir et. al., 2004). This plan includes the number of desired buckets of production for each drawpoint, plus the order of production priority for drawpoints within a panel. The LHD operator travels to the desired panel and logs-in on the GOIC to the Dispatch system. Based on the predetermined order of production for that particular panel, the central computer directs the LHD to produce from the initial drawpoint. The operator either accepts this order or actively logs-in to an alternate drawpoint. The operator then commences production from that drawpoint. When at the drawpoint, the tag-reader component of the FCS interrogates the RF tag installed at the drawpoint to determine the drawpoint identification. The operator also pushes the "load" icon on the GOIC. The LHD’s FCS then relays the RF tag and the LHD’s FCS identification numbers to the central computer via the micro-cell radio network. This informs the system that the LHD has produced a load from the specific drawpoint. The system then directs the loader to the assigned orepass for tipping. En route, the LHD passes beneath other RF tags (known as call points) strategically located in the tunnel which inform the system of the LHD’s progress. At the orepass, the FCS reads the RF tag to determine the tip location. Once the load is tipped, the operator Massmin 2004
pushes the "dump" icon on the GOIC to register the load to the orepass. The system records the dump and then automatically directs the LHD to the point of next production. If an LHD is late to complete the next expected action, such as "arriving at the assigned rock breaker," the system alerts the dispatcher by an exception message. The Intelimine software has been customized to calculate the height of material in the orepass, based on LHD bucket factors. When the "dump" button is pushed, the orepass material level is recalculated. The LHD operator then returns to the drawpoint and continues to produce according to the draw order. Once the planned number of buckets has been produced from the initial drawpoint, the LHD is automatically dispatched to the next priority drawpoint. As a drawpoint becomes unavailable for production, the LHD operator has the capability to change the status of the drawpoint within the database. The system will then dispatch the loader to the next priority drawpoint. If the operator desires to produce from a drawpoint out of sequence, he may do so by actively selecting the particular drawpoint off the GOIC icons list. The FCS tag reader continues to communicate with the RF tags installed at the drawpoints to confirm the production location. Should the operator key-in a drawpoint that is in conflict with the drawpoint RF tag, an exception report is automatically submitted to the operator and Dispatcher by the central computer. Once the available draw order for a panel has been completed, the LHD operator sends a message to the Dispatcher for further instructions. The customized system has codes for various classes of equipment delay. The operator actively selects the appropriate category off the icon list on the GOIC. The central computer and Dispatcher are thereby notified of the current operating condition for the equipment. At the end of shift, the Dispatcher makes whatever manual adjustments to the recorded data that are necessary. Once daily, Dispatch production information is downloaded to the central production-planning database. This information is then utilized within CMS to calculate daily production, draw order compliance, and remaining drawpoint reserves, and to analyze potential sources of induced stress in the production area. The production strategy for the subsequent day and the balance of the month is then re-established based on analysis of this daily information. On a monthly basis, the Dispatch data is utilized to analyze LHD equipment and manpower efficiency and productivity. 4.2 Truck Dispatching Trucks at the DOZ haulage level are generally automatically dispatched to orepass chutes according to rules established by the Dispatch Engineer. These rules consider the following: • Distance between the truck and an available loading point. • Dispatching of other trucks. • Orepass material inventory. • Time since the loading point has last produced. The Dispatcher is able to also actively direct trucks to particular loading points, as operational conditions dictate. At the beginning of a shift, the truck operator logs-in to the Dispatch system and the central computer automatically directs the truck to the initial chute for loading. When the truck arrives at the chute, the FCS reads the tag installed at the chute and the operator pushes the "arrive at LP" icon. Once the truck is loaded, the operator pushes the "load"
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icon on the GOIC. The system then registers the equipment as loaded, and directs the truck to travel to the crusher. The material inventory in the orepass is also adjusted for the load, based on the factor for that truck type. At the crusher, the FCS reads the tag for the particular crusher dump point and the operator pushes the "arrive at crusher" icon on the GOIC. Once dumping has been completed, the truck operator pushes the "dump" icon on the GOIC. At this point, the system records the load as dumped and adds the tonnage to the shift total. The central computer then dispatches the truck to a chute for the next load. Dispatchers have the capability to classify as a chute as "Down" or on "Standby", based on operational conditions, and thereby remove the unit from the dispatching logic. A utility screen has been created which provides orepass information to the dispatcher for: • Tonnes remaining, • Last time it was drawn from, • Dispatch priority level, • Truck identification number, if one is currently assigned to it; plus the expected arrival time, • Status of the loading chute. 6 DISPATCHER MONITOR AND CONTROL Dispatchers monitor and control DOZ production activities, on each shift, from emulator monitors in a control room located within the mine. The Intellimine software system has been customized to provide a number of realtime informational screens that allow the Dispatchers to determine critical operating parameters and to interact with equipment operators. The Transaction Screen displays real-time tonnage summaries (last hour, this hour, shift-to-date) for extraction, orepass inventory and haulage to the crusher. The Exception Screen displays messages generated from the central computer for operator actions that are unexpected (Figure 5). Examples include: 1) locking-in to a drawpoint that is different than the drawpoint tag identification, 2) producing from a drawpoint above the draw plan, 3) loss of radio contact with the micro-cell network, and 4) equipment operating without the operator actively logging-in to the system. The Exception Screen demands that the Dispatcher either "accept" or "reject" these operator actions. If "rejected", the central computer automatically sends a message to the equipment FCS to request remedial action from the operator.
Figure 6: Dispatch Utility Screen
and the current draw compliance. A daily summary report is automatically generated by Dispatch and emailed to DOZ Management. The email includes data, by shift, for production tonnage; equipment operating hours, delay hours, availability and utilization, and orepass material inventory. An hourly summary of the same information, plus cycle time data (Figure 7), may be accessed from the Report Utility Screen.
Figure 7: Truck Cycle Time Report
5 CONCLUSION
Figure 5: Exception Screen A range of equipment status information is displayed on the Dispatch Utility Screen. The utility also provides a means for the Dispatcher to send text messages to the equipment field computer system (Figure 6). The Draw Card Screen provides the Dispatcher real-time information on the drawpoint production plan, drawpoint status and buckets of ore produced shift-to-date. The screen also tabulates the percent drawn for each drawpoint 700
Use of the Dispatch system provides many benefits to controlling production activities at the DOZ Block Cave Mine. For extraction, the system provides an effective means: 1) to communicate the details of the production plan to the equipment operators, 2) to collect accurate actual production data, 3) to calculate compliance to the production plan, 4) to summarize daily actual production for transfer to other production planning tools, 5) to identify circumstances when equipment operators produce in an unexpected fashion, and 6) for the loader operator to effectively communicate with the Dispatcher. All these benefits allow for precise compliance of actual production to the long-term plan and aids PTFI to effectively manage the DOZ cave.
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The Intelimine system allows haulage operations to be conducted and optimized with only minor active input from the Dispatchers or supervisors. Custom features of the system also provide real-time information on the material inventory in each orepass. Use of the system, in general, provides a continuous monitor of all production activity. Information stored to the database provides the basis for calculating equipment productivity and efficiency. Since allocation of equipment to a particular task is automated, supervision is able to instead focus on resolving mine issues. ACKNOWLEDGEMENTS The authors are grateful to all their colleagues working at the DOZ Mine. Also, the authors acknowledge the permission given by P. T. Freeport Indonesia to publish this paper.
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REFERENCES • Barber, J., Thomas, L., and Casten, T., Freeport Indonesia’s Deep Ore Zone Mine, Proceedings from MassMin 2000, pp. 286-294. • Barber, J., Ganesia, B., and Casten, T., 2001, Developing the DOZ Mine at PT Freeport Indonesia, Mining Engineering, vol. 35, no. 11, pp. 19-24. • Owen, T., 1992, Ertsberg East Mine Freeport Indonesia Inc. Papua Indonesia, SME Mining Engineering Handbook, vol. 2, p. 1830-1835. • Samosir, E., Brannon, C., and Diering, T., Implementation of Cave Management System (CMS) Tools at the Freeport DOZ Mine, Proceedings from MassMin 2004, in press.
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Large diameter vertical raise drilling and shaft boring techniques as an alternative to conventional vertical shaft sinking techniques Pete H Ferreira, Manager Marketing and Business Development BSc. Eng (Mining) Pr.Eng , FSAIMM Murray and Roberts RUC Limited
Abstract Raise drilling in South Africa started in 1968 with machines capable of drilling 1.2 metre diameter raises up to a length of 90 metres. Today’s raise drilling machines are capable of drilling vertical shafts to a diameter of 6.1 metres to depths in excess of a 1 000 metres and 7.1 metres to 200 metres in depth. That is an immense improvement from its humble beginnings. In 1971 the first shaft boring machine was designed and manufactured in the Federal Republic of Germany and bored a 4.88 metre diameter, 231 metre deep shaft. In the quest for mechanical shaft sinking technology it is now possible to sink vertical shafts mechanically up to 8.5 metre in diameter to 2 000 metres in depth. Fewer personnel are required with mechanisation, and due to the methodology used, safety aspects are improved with overall risk drastically reduced. Conventional shaft sinking is briefly discussed and compared to large diameter raise drilling and shaft boring. The technological improvements in shaft boring machines (raise drilling and V-mole) have progressed at an accelerated rate. Under certain geological conditions, with the increase in diameter of raise drilled holes, however, comes greater potential for instability of the exposed sidewalls of the drilled hole. A systematic flowchart developed by Stacey & McCracken, is discussed to quantify the risks associated with raise drilling and shaft boring and thereby quantifying the risk attached to drilling a relatively large shaft prior to commencement of the excavation in order to assess the stability of the bored hole. Managing this risk now becomes the engineer’s challenge. The capabilities of these machines and associated risks are explained with reference to specific drilling projects.
1. INTRODUCTION Murray & Roberts RUC has been involved in raise drilling contracting since 1978. Murray & Roberts RUC has become the world’s largest raise drilling contractor and is considered a leader in the field of large diameter raise drilling. The Company operates a total of 23 raise drills, which includes four Wirth HG 330 SP and one Robbins 123R machines, these being some of the largest raise drills ever manufactured in the world. Since 1989 RUC has also gained operational experience in shaft boring using the V-mole shaft boring technique. To date, four major shaft projects, have been completed utilizing the V-mole technique, these being:• Oryx No 1B Ventilation Shaft in South Africa • Pasminco’s Broken Hill No. 5 Airway in Australia • Anglogold’s Western Deep Levels South Mine sub vertical ventilation shaft and • AlpTransit St Gotthard project in Sedrun in Switzerland. These projects were undertaken in joint venture with Thyssen Schachtbau GmbH of Germany using a Wirth SBVII rodless shaft boring machine, better known as a Vmole. 2. RAISE DRILLING TECHNIQUES AND
HISTORY
2.1 Modes of Operation Raise borers can be used in various modes of operation and the modes most often used are: 702
• • • • •
Conventional pilot drilling Conventional up reaming of vertical and inclined holes Down boring with a pre-drilled pilot hole Blind up boring Directional piloting and raise drilling used in conjunction with a shaft boring machine (V-mole), for the boring of a large diameter shaft.
2.1.1 Conventional Pilot Drilling A tri-cone pilot bit is normally used varying from 9 inches (229mm) to 15 inches (381mm). The 15 inch (381mm) bit is normally used on long holes with a 12 7/8 inches (327mm) integral drillsteel string with 10 1/8-inch DI 42 tool joints. During drilling, a fluid is pumped through the center of the drillstring to the cutting face, where the rock cuttings are flushed and raised from the bottom of the hole through the annulus around the drillstring to the collar of the hole. The drilling fluid is settled in a closed loop via a series of settling dams so that the drilling fluid can be re-used. 2.1.2 Conventional Up-reaming of Pilot Holes On completion of pilot drilling and at such time that the pilot hole breaks through into the lower excavation, a reaming head is attached to the end of the drillstring. The size of the reaming heads range between 1.2 metres and 7.1 metres in diameter. The head is rotated by the machine and is pulled back against the rock face at the same time. Cutters with tungsten carbide inserts are fitted to the head and these cut grooves in the rock in a rotary crushing mode. The ‘kerfs’ of rock in between the grooves ‘spall’ out and
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rock failure occurs in a tensile mode. The rock cuttings fall to the bottom of the hole where they are mucked out by a mechanical loader. It is a safe, efficient and cost-effective method of making holes through different geological formations with the use of powerful machines, high strength drillstring and reliable reamer heads. The maximum loading capacity of the drillstring limits the diameter as well as the length of the shaft. The loading is dynamic and only approximately calculateable as tensile, torsional and bending stresses are overlapping. (See Figure 1)
Figure 3 2.1.6 Recent Achievements (Breaking New Ground) Murray & Roberts RUC currently holds world records and has had the following achievements: • Largest diameter shaft raise drilled to 7.1 metres in diameter and 178 metres of vertical depth at Sasol Coal’s Secunda Collieries’ Bosjespruit Mine.(See Figure 4) Figure 1 2.1.3 Down Boring with a Pre-drilled Pilot Hole In this case an oversize pilot hole is drilled. The cutting head is installed at the top of the pilot hole and drilling takes place in the downward mode. Rock cuttings are flushed down the oversize pilot hole to the bottom of the hole where it is removed. In the case of smaller holes, the machine provides cutter thrust and in the case of large diameter shafts the cutter head is weighted through the addition of steel collars. The down boring method is not used often as the risk of blocking the pilot hole and creating mud rushes at the bottom of the hole is high. (See Figure 2) Figure 4
Figure 2
2.1.4 Blind Up Boring In this case the machine is placed at the bottom of the planned hole position and the cutting head drills upwards. Rock cuttings fall to the bottom of the hole where they are deflected into muck cars or an alternative mucking arrangement. This application is widely practiced in South African deep tabular orebody mines with hole lengths ranging from 30 to 90 metres under normal circumstances and up to 190 metres in special applications. (See Figure 3) 2.1.5 Directional Pilot Drilling Directional pilot drilling is costly and is therefore only used in applications where a high degree of accuracy is required. The accuracy of a vertical pilot hole can be guaranteed to depths within the capability of the raise drill machine and accuracies of 0,036% have recently regularly been achieved. Refer to paragraph 4.1. Massmin 2004
• Longest vertical hole reamed to 1.83 metres in diameter and 1260 metres deep at the Primsmulde Project, Germany. • At Kloof Gold Mine South Africa a 1100 metre deep 4.1 metre diameter hole was drilled through the hardest rock formation being Lava with an UCS between 600 and 750Mpa . • Longest inclined raise drilled hole to 3.5 metres in diameter and 755 metres deep at BCL in Botswana ( See Figure 5 ) • Deepest shaft V-mole bored in South Africa to 6.5 metres in diameter and 972 metres deep at Oryx Gold Mine • Deepest shaft V-mole bored at Prismulde Germany to 7.8 metres in diameter and 1260 metres deep by Thyssen Schachtbau of Germany • Largest diameter V-mole shaft bored in hard rock to 7.1 metres in diameter and 752 metres deep at Western Deep Levels’ South Mine, South Africa • At Impala Platinum Mine South Africa drilled a 1 050 metre long 5.1 metre diameter raise drill hole through norites with RVDS with 0.05 % accuracy • At Sedrun in Switzerland drilled a 785 metre long 1.83 metre diameter hole with RVDS with 0.035 % accuracy i.e. 280 mm deviation • At Moab Khotsong in South Africa drilled a 360 metre long 3.8 metre diameter hole with RVDS with 0.063 % accuracy i.e. 223 mm deviation • At Prismulde in Germany drilled the 1 260 metre long 1.83 metre diameter hole with down-the-hole motor with 0.04 % accuracy i.e. 450 mm deviation
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• At Impala Platinum Mine South Africa drilled a 1 090 metre long 5.1 metre diameter hole through norites
has been done before. The next challenge will be to drill a 1.83 metre diameter hole over a continuous length of more than 1 300 metres. The shaft depth is therefore unlimited as long as a center core pilot hole is available. A V-mole shaft construction is carried out in various stages: • the raise drilling of the pilot hole and center core to serve as a rock pass • the construction of a pre-sunk foreshaft to facilitate the installation of the V-mole machine • the installation of the necessary and required hoisting facilities • V-mole boring, shaft support and equipping • the final removal of the V-mole at shaft bottom • and commissioning of the permanent shaft system The raise boring of the pilot shaft: The pilot shaft is raise bored using a Wirth HG330 raise drill (See Figure 6). The verticality is ensured by using directional drilling tools, i.e. the Navi-Drill or the preferred RVDS.
Figure 5
3. SHAFT BORING (V-MOLE METHOD) 3.1 Background In the late sixties, following the successful application of tunnel boring machines in tunnels, thought was given to use this new excavation technique to underground coal mines with a view to fully mechanise tunneling and shaft sinking. In 1971 the first shaft boring machine was put into service in the coal mines in Germany by a consortium of specialist mining contractors namely Deilmann-Haniel GmbH (Dortmund) and Thyssen Schachtbau GmbH (Mulheim). The shaft boring machine used was a Wirth GSB-V-450/500 capable of reaming shafts with a diameter of up to 5m from a center core pilot hole. Numerous improvements have been made to this machine since 1971, which is reflected in the three machine generations with the latest model in use being the Wirth SBVII.
Figure 6
3.2 Mode of Operation The rodless shaft boring machines (V-mole) can be applied to sink deep vertical shafts with a diameter of up to 8.5 m. The requirements for this method are: • relatively competent rock as determined through the study techniques mentioned above (unsupported center core to stand up) • and a reamed pilot hole between shaft head and shaft bottom of approximately 1.83 to 2.4 metres in diameter with sufficient verticality to serve as a center core. During the boring operation this center core pilot hole is used to drop the reamed cuttings to the bottom of the new shaft and is also used for ventilation purposes. The shaft boring machine constructed similarly to a tunnel boring machine, (TBM), widens the center core pilot hole to the final shaft diameter by reaming downwards. Reaming, muck disposal, shaft support and permanent shaft equipping are performed continuously and concurrently. The steering system of the machine guarantees the verticality of the bored shaft with the aid of a laser beam through the centreline of the shaft. The boring diameter can be varied within a range of 5.0 to 8.5m. The depth to be bored is not restricted by the shaft boring machine parameters, but becomes a factor of the ability to drill and ream a pilot hole to 1.83 metre in diameter. We know that the drilling of holes of 1 300 metres in length is possible and
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Figure 6 (a)
Construction of the pre-sunk shaft (foreshaft) and the installation of the hoisting facilities On completion of the reaming of the center core pilot hole, the foreshaft is sliped and lined to a depth of ± 11 metres for the assembly of the V-mole. The foreshaft can be sunk before the pilot hole is drilled with the raise borer. The installation of the hoisting facilities are done concurrently with the pre-sink. The hoisting facilities are required to transport the men and material to the shaft borer (V-mole).
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V-mole boring and the installation of the permanent rock support The shaft borer then reams the center core pilot hole to the required size with the rock chips being loaded at the bottom. The shaft can be concrete lined or shotcreted by means of a robotic arm mounted on the stage. The "drilling" and "lining" is co-ordinated in an innovative construction unit. The shaft boring machine SB VII is mainly built with a stable frame – outer kelly – which is hydraulically clamped against the shaft wall by means of 12 gripper pads, arranged symmetrically in 2 levels of six pads. The rotating inner kelly includes the main drive shaft, bearings and gears. The upper non-rotating end is square shaped and accommodated in an articulated frame of the outer kelly. The lower end is carrying the rotating cutterhead, which is powered by six electrical motors of 132kW each. 3.3 Track record Some fifty-five vertical shafts around the world have been successfully bored with V-mole machines. Four shafts were constructed by the Joint Venture between Thyssen Schachtbau GmbH and Murray & Roberts RUC. The Wirth SB VII shaft borer in South Africa is jointly owned by these two companies. Joint Venture first The joint venture started its first project in 1989 in South Africa, constructing the 972 metres deep, No 1 B ventilation shaft at Oryx Gold Mine, to a diameter of 6.5 metres, with the shaft boring machine SB VII. Prior to this operation in South Africa the machine was thoroughly overhauled and prepared for the application in hard rock formations. Six electrical motors with 132kW each were installed. The penetration rates achieved on this project were satisfactory, considering hard quartzite rock formations with compressive strengths ranging from 220 to 280 Mpa were penetrated.
• the shaft will be used as a return air ventilation shaft and will pass through the stoped out reef horizon between 84 and 109 level • the exposed shaft sidewall is to be supported by means of "splitset" anchors, providing the primary support followed by steel fibre reinforced shotcrete (SFRS) using the shotcrete method in place of conventional concrete lining • the support of the sidewalls must occur concurrent with the sinking operation of the shaft • on various levels throughout the depth of the shaft, return air way (RAW) holings must be provided for into the ventilation shaft Competitive tenders were submitted for the above consisting of conventional blind sink, slipe & line operations and the shaft boring option. The decision was made in favour of the shaft boring technique after considering the cost and time advantages of this method. The shaft had to be drilled in extremely hard conditions. The geological profile showed that approximately 40% of the formations had a uni-axial compressive strength of more than 300Mpa. In the Alberton Lavas the compressive strength was as high as 550Mpa. The achieved drilling rates of 0,4 m/hour were very satisfactory. The experience gained from the two preceding projects "Oryx" and "Pasminco" and the knowledge of the geological profile caused the joint venture to design a new cutter head for the project (See Figure 7). This cutter head has a closed design in order to achieve better stabilization and to provide the head with back loaded cutters (See Figure 8).
Joint Venture second Upon completion of the Oryx project, the joint venture could start a second project namely the ventilation shaft No. 5 in the south field of Pasminco Mining in Broken Hill, Australia. In the course of the sinking of the 810 metre deep shaft with a diameter of 6.5 metres, steep rock formations of amphibolites with a compressive strength of 350Mpa and gneiss with approximately 150 Mpa were penetrated. After these two remarkable projects the following conclusions could be made: • both shafts were be completed on schedule • performances of up to 18 m/day (Oryx project) and 12 m/day (Pasminco) concrete lined shaft could be achieved with an average performance of 7.6 m/day • the concurrent support drilling and installation as well as concrete lining, could be done undisturbed • during the course of shaft boring, the necessary injections could be performed to seal off the fissure water inflows • a number of intermediate stations were excavated during the sinking process Joint Venture third V- mole at depth, the ultimate challenge Anglogold awarded the 752 metre, 7.1 metre diameter Sub Ventilation Shaft construction project at the then Western Deep levels South Mine to the joint partners Murray & Roberts RUC and Thyssen Schachtbau GmbH. The design brief read as follows:• the new SSV (Sub Shaft Ventilation) is to be sunk from 84 level to 109 level (3312 metres below datum) • the SSV shaft must have a 7.0 metre diameter and a depth of 752 metres Massmin 2004
Figure 7
Figure 8
The new cutter head of 7.1 metre diameter can be equipped with disc-cutters as well as with tungsten carbide cutters.
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The shaft sinking activities commenced in May 1995 with the drilling of the pilot hole and center core to a diameter of 3.5 metres. This was carried out with Murray & Roberts RUC’s Wirth HG 330 raise borer. The assembly of the shaft boring machine took approximately 6 weeks. The 752 metre deep ventilation shaft was completed under extremely difficult geological conditions. The highest performance attained under these conditions was a monthly advance rate of 137 metres in the Alberton Lava formation. After drilling through the lava formations, scaling occurred in the quartzite formations at 104 level (+-3 200 metres below surface), which made increased temporary support density necessary. The center core pilot hole "dog-eared" excessively from the original 3.5m diameter size due to stresses at this great depth underground. The shaft sidewall bolting with a rock bolt density of one bolt per m2 was done fully automatically and concurrently with boring by using two Atlas Copco drill rigs mounted on the rotating and telescoping top deck of the shaft boring machine. (See Figure 9)
Joint Venture fourth, the St Gotthard Base Tunnel Sedrun Ventilation Shaft The latest venture between the joint venture partners Murray & Roberts RUC and Thyssen Schachtbau is a 785 metre deep 7.1 metre diameter ventilation shaft in Sedrun, Switzerland. The Sedrun shaft is part of the AlpTransit Gotthard project which could well be the construction project of the new century. The 57km long base tunnel under the St Gotthard massif is the core project of a new railway link through the Alps, with future passenger trains travelling at speeds of up to 250km/h through the longest railway tunnel in the world. The shaft construction works commenced mid May 2002. The raise bore machine Wirth HG 330 SP was installed at the shaft collar for both directional pilot drilling to 381mm diameter and the subsequent reaming of the center core pilot hole to 1.83 metre in diameter. The vertical drilling system, RVDS, of Micon was applied for the directional drilling. Despite the unfavourable geology with numerous vertical joints, the deviation from the vertical was only 280mm (0,036% accuracy). The reaming of the pilot hole from 381mm to 1.83 metre diameter was conducted without problems and completed by mid November 2002. After the completion of the pilot shaft the installation of the shaft sinking equipment and the shaft boring machine commenced. The V-mole, type Wirth VSB VI with a 7.1m borehead was dressed with a combination of disc and tungsten carbide cutters and a total installed power of 520kW. The shaft boring operation started in February 2003 and was completed by the end of June 2003 with an average boring performance of approximately 7m/day. The continuous drilling and shotcrete lining system, which was perfected at Western Deep Levels Gold Mine, was used. (See Figure 11).
Figure 9 The application of the steel fibre micro silica wet shotcrete with 50Mpa strength was also fully automated by means of a robotic nozzle, which was situated on an independent operating sinking stage with three decks above the shaft boring machine. (See Figure 10).
Figure 11
4.0 RAISE DRILLING RISKS AND RISK MANAGEMENT
Figure 10 The project construction was remarkable for its successful penetration of one of the hardest rock formations with compressive strengths of up to 550Mpa. 706
4.1 Deviation of the pilot hole Accuracy of pilot drilled holes has been a concern almost since the invention of mechanized rotary drilling. This problem became much more apparent as operators required the drilling of longer holes with lengths of 1 100 metres, this not being uncommon anymore and competing with blind sunk vertical shafts.
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The deviation of a borehole from its intended path can be attributed to both geological and technical factors, which can be divided into three categories:Controllable factors • Set-up accuracy of a machine • Equipment condition • Machine • Starter pieces • Correct boring tool (bit) • Bit sub • Stabilizers • Bit contact pressure (force) • Rotation speed • Flushing rate • Starting procedure Semi-controllable factors • Build rate • Stiffness (design) of the drill string • Bit walk rate Non-controllable factors • Variations in rock hardness levels • Strata dip • Ground conditions – jointing, fractures, partings, etc. • Geological features – faults, dykes, bedding planes, etc The graph, refer figure 12, shows a typical deviation with a constant build rate of 0,25 degrees per 100 metres of drilling. The build rate represents the increase in inclination of the hole and is measured from the vertical axis. It can be seen that due to the compounding effect, the deviation becomes exponential and increases drastically with increased depth.
Drill is fitted with an adjustable bent-sub. The motor angle is set to a suitable angle and is lowered down the hole. The motor is then orientated, 180 degrees opposite to the direction of deviation, using the steering tool. The steering tool is fitted with a series of magnetometers and accelerometers that relay information via the wireline conductor to the equipment at the machine. All data is processed by computer at the collar of the hole and the operator can monitor the motor toolface, as well as the hole direction, on the drillers display unit. A high viscosity mud is then pumped through the drill string, which causes the mudmotor to rotate at a speed of roughly 120rpm. The drill string is now moved downward to provide sufficient thrust to the bit, no rotation of the drill string takes place during the correction-run, and rotation is provided through the mud-motor directly to the bit. On completion of the correction-run the directional drilling gear is removed from the hole and conventional pilot drilling is resumed. The biggest disadvantage of the Navi-Drill system is that it is used re-actively. To rectify the hole deflection the drillrods must be removed from the hole, the Navi-Drill attached and lowered to the bottom of the hole being a very tedious process. The Navi-Drill must be removed when the hole direction has been rectified. To remove the Navi-Drill, the drillrods must be removed again. The drill bit must again be attached and lowered before piloting can commence. To overcome this problem the self-steering drilling system was developed. (See Figure 13).
Figure 13
Figure 12 Highly sophisticated survey tools are used to monitor the inclination and direction of a hole. The instruments are capable of detecting movement off the vertical through angles as low as 0,05 degrees in inclination and 0,5 degrees in Azimuth. There are two ways in which the direction of the pilot hole can be steered and these being: • Using navigational drilling equipment The downhole motor such as the Navi-Drill is used reactively, i.e. when the hole deviates it is rectified. The NaviMassmin 2004
• Rotary Vertical Drilling System (RVDS) The ZBE3000 (DMT GmbH) self-steering directional drill has been in use since the mid 1980’s. High maintenance costs and occasional problems necessitated the further development of this type of equipment. New models were developed, namely the Well Director, the ZBE4000 and the ZBE5000. Micon developed a rotary vertical drilling system, ( RVDS ), and this has been available since the mid 1990’s. This equipment is particularly suitable for directional drilling in conjunction with raise drilling. This system uses a pair of incline sensors to measure the borehole inclination and transmit the data to an electronic unit. If the preprogrammed directional limits are exceeded, the steering function is initiated by the hydraulic steering system, which extends or retracts the four external, independently operated control ribs. The extendable stabilizer ribs generate radial forces and work against the angle build-up. The RVDS is supplied to the rig as a complete system consisting of the downhole tool and an independent PCbased management system. The downhole tool can be divided into two parts, each of approximately 1.5m in length, being the pulser sub & the steering sub. In order to monitor the self-steering drilling process, data signals are transmitted to the surface via positive water pulses and are received, decoded and visualised by an information unit. (See Figure 14).
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The upper part of the RVDS, called the tank-sub, rotates with the drill string. The outer steerable stabiliser is a part of the lower steering-sub. It is non-rotating and runs on bearings on the drive shaft. The drive shaft transmits the torque of the drill string to the bit. The non-rotating lower part contains the sensors, the data processing electronics and steering unit. This sub is fitted with radically extendable ribs. The required steering force is generated hydraulically by an oil pump inside the pulser-sub and is transmitted to the borehole wall by pistons and steerable ribs. The directional data signals are also transmitted to the mud pulser for further communication to the surface. Both the water pulser for data transmission and alternator for the electrical power supply as well as the pump for hydraulic power supply are housed within the pulser/generator sub. A water turbine drives the alternator and pump. A fully digital electronic unit located in the steering sub, supplies the two accelerometers with the required voltage. Inclination data is then compared to predetermined data and, if necessary, transformed into steering signals. Subsequently, one or two of the four control valves are being supplied with current. The valves control the cylinder oil pressure, which in turn generates the compensating forces necessary to achieve vertical drilling. The directional data signals are also transformed into a pulse pattern by the digital electronics and they are transmitted to the surface. The system for large pilot holes, between 15 – 17_", has been jointly developed by Murray & Roberts RUC and Micon. The improvements are reflected in the new design (See Figure 15). Various holes have been drilled with accuracies of 0,04% using the RVDS. Hoisting shafts can now be raise drilled by using the Rotary Vertical Drilling System (RVDS) to drill the pilot from there the final hole.
Figure 14 4.2 Geotechnical risks associated with large diameter raise drill shafts The two biggest geotechnical risks are boreability and stability. Prof. T R Stacey and A McCracken have written numerous papers on risk analysis and some aspects are briefly discussed in this section. 708
Figure 15
Boreability is determined by the hardness and abrasiveness of the rock material and the structure of the rock mass, that is, it’s jointing and then also by the raise drilling machine factors. Stability is determined by the rock mass structure, which defines the potential freedom of movement of the rock blocks and by the stresses acting, which provide confinement to the rock mass, but may also be of such a magnitude as to induce failure in the rock material and rock mass. A detailed geotechnical evaluation or raise bore rock quality assessment based on the Stacey and McCracken method is recommended in the case of deep and/or large diameter shafts and is briefly discussed. The risk attached to any raise bore project is dependant on the confidence with which the rock mass conditions are known. The level of confidence in, or reliability of, information depends on the amount of information available, the variation of individual parameters, the impact of this variability on the probable quality and the required minimum rock quality for compatibility with the proposed raise drilled shaft specifications. The important aspect is to assess the rock conditions with respect to the required minimum quality for stability. A flow chart, developed by Prof. Dick Stacey, that sets out the activities to be followed for a systematic assessment of the risk related to the geotechnical aspects of any raise drill project is presented in Figure 16 attached. • Initial risk assessment The preliminary geotechnical assessment should be aimed at determining average and lower bound conditions in terms of "raiseability" and "stability". The range and distribution of the rock quality, QR and the important parameters RQD/Jn and Jr/Ja, must be compared with the required minima for stability at the proposed shaft diameter. The Q value for the rockmass is obtained from the relation Q = RQD x Jr x Jw Jn Ja SRF
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where RQD is rock quality designation Jn is joint set number Jr is joint roughness number Ja is joint alteration number Jw is joint water reduction factor, and SRF is stress reduction factor. RQD/Jn gives an estimate of rock block size, Jr/Ja provides an indication of discontinuity shear strength and Jw/SRF indicates the conditions of active stress surrounding the excavation. To obtain the raise drilled hole quality index, QR from Q, the following adjustment factors, which are cumulative, must be applied. • Wall adjustment • Discontinuity orientation adjustment factor • Weathering adjustment At the preliminary evaluation stage the risk should only be deemed "acceptable" if the quality consistently exceeds the requirement throughout its entire length. This presupposes the availability of sufficient information for this conclusion to be drawn.
Figure 17 stability of different types of excavations are presented in Table 1. These provide guidelines for other raise drill shafts. Table 1: Suggested acceptability of risk for various raise drill shafts Excavation Type Unlined hoisting Shaft Ventilation Shaft Ore Pass Ore Pass
>15 10 >2 1
0,01 0,05 0,15 0,25
• Conclusion A method of quantifying the geotechnical risk associated with a raisedrill or shaft bore shaft, is presented above, and based on shaft diameter and a raise drill rock quality index, QR. The approach outlined provides an indication of overall geotechnical feasibility. All excavations must, however, be considered individually and the potential problems should be addressed on merit. The chart presented in Figure 16 does not replace classical analysis as a means of evaluating the incidence and stability of potential failure wedges, but does allow the probability of failure to be predicted in a simple manner. Comparison of the probability so obtained with the required reliability, permits assessment of the overall feasibility and the risk associated with a proposed raise or shaft. In many cases adverse ground conditions can be treated by cement grouting prior to raise drilling, or alternatively, advanced planning can be done to carry out support works directly after raise drilling.
Figure 16
• Final risk assessment The assessment of risk will depend ultimately on the acceptability of failures within the raise drilled shaft and on the incidence and volume of failures that can be tolerated. In general, an acceptable probability of failure of a raise drilled shaft, given its function, is considered to be 0,05 i.e. 5%. This is commensurate with an RSR value of 1,3. Given a proposed raise drill diameter and a rock mass of a certain range of QR values, the range of probability of failure can be obtained. If the length of the raise is known, the likely length of raise liable to be affected by failures can be calculated and the volumes of failure determined from stability analyses. A chart showing the probabilities of failure, P(f), or alternatively the reliability, R, of a shaft (where R = 1 – P(f) x 100%), for the range of raise bore diameters and rock mass qualities is presented as Figure 17. Suggested levels of reliability, R, and probabilities of failure P(f), that are considered acceptable for the raise wall Massmin 2004
Service Reliability Probability Life (Years) R (%) of Failure P(f)
4.3 Risks and the control thereof One of the bigger risks in the shaft boring method is the potential scaling and deterioration of the center core pilot hole and the other to ensure the center core pilot hole remains open during the shaft boring. To manage these factors the risks need to be managed and engineered to ensure success. 5. CONVENTIONAL VERTICAL BLIND SHAFT SINKING Conventional blind shaft sinking using drill and blast techniques has been practiced for as long as underground mining has taken place. Various shaft sinking methods are being used and these being: • Hand held drilling of the bottom with nominally 2.0 metre advance per blast • Jumbo drilling of the bottom with either pneumatic or electro-hydraulic drifters with advance per blast of up to 6.0 metres
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• • • •
Mucking with an Eimco 630 type loader into a kibble Mucking with a cryderman type clam system into a kibble Mucking with a cactus type lashing unit into a kibble Concurrent shaft concrete lining from the sinking stage above.
Rock is hoisted to surface or the bank elevation therefore not interfering with other rock hoisting operations. Bottom access is therefore not necessary and this is the method to follow where a shaft is sunk in green fields operations. Large winding facilities are generally required to hoist the rock from the shaft bottom as well as a relatively large stage winder installation as a large stage has to be supported from such a winder, especially with the cactus grab cleaning method. Shafts are generally equipped on completion of a sink to shaft bottom unless a cryderman type cleaning method is used, which supports concurrent equipping of a shaft with the sinking. A smaller stage winder is required with this cleaning method. Blind sink operations are generally done with shafts of diameters of 4.5 metres and more. Advance per blasts will vary and advances per day will generally average around 3.5 metres to 4.5 metres depending on the depth of shaft and diameter. The deepest one lift blind vertical shaft sunk to date has been the South Deep shaft in South Africa to a depth of some 2 963 metres below collar. Blind sink shafts can be sunk from very shallow to very deep depending on needs as well as to any diameters. 6. PROS AND CONS OF CONVENTIONAL VERTICAL BLIND SINK SHAFTS COMPARED TO SHAFT BORED SHAFTS Shaft boring becomes an economical option from depths of around 600 metres and deeper and at that point becomes cheaper and faster and can be bored to great depths. Comparison Items Conventional Blind Sink
Shaft Bore
1 Depth Restrictions
None
Competitively economical from 600 m
2 Diameter restrictions
None
4.5m to 8.5m in diameter
3 Speed of sink
Faster up to 600m
Faster from 600m onwards
4 Blasting
All sinking
None
5 Need bottom access
No
Yes
6 Lining thickness
Thicker
Thinner
7 Safety aspects
Poorer
Best
8 Stage requirement
Large
Small
9 Kibble hoist requirement
Large hoist Small hoist for men for rock & material & material only
10 Stage hoist
Large
Smaller
11 Sinking crew size
Larger
Smaller
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Raise drilling fills the gap between a very small shaft and a larger shaft and is the fastest means of sinking a shaft provided bottom access is available. Every technique has it’s place in the business and the pros and cons must be weighed up against each other before a final decision is made as to the required method of sinking. 7. CONCLUSION The mining industry’s requirement for safe, rapid and economical mine development is met by the mechanical large diameter raise drilling and shaft boring methods described. The technique has provided an economically sound solution for a large variety of different requirements, especially in those projects executed in recent years involving deep, large diameter holes. Raise drilling to depths exceeding 1 000 metres and at diameters of up to 6 metres, is no longer uncommon. The method continues to be developed to cover an increasingly wide range of circumstances. The improvement made in directional drilling now enables hoisting shafts to be raise drilled, either in one pass or in combination with the V-mole. By using the systematic risk assessment developed by A McCracken and TR Stacey, a quantitative assessment of the risk attached to any shaft prior to commencement can be done. The capabilities and effectiveness of the raise drilling and shaft boring techniques have been proven in the execution of more than 50 projects throughout the world, with an accumulated depth of 21 000 metres and in a wide variety of rock types. "Using alternative scenarios, the future literally becomes a matter of choice, not chance" - (Wolfgang Grukke) 8. REFERENCE • Geotechnical Risk Assessment for Large Diameter • Raisebored Shafts TR Stacey, a MC Cracken Rock Mechanics Considerations in Raiseboring TR Stacey, Steffen, Robertson and Kirsten • Fully Mechanished Sinking of Deep Shafts Dr Ing. B. Schmucker and Dr Ing. C. Cetindis Thyssen Schachtbau GMBH Germany • Directional Drilling in the Mining Industry Industry E. Berger, Thyssen Schachtbau GMBH
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Lower cost and higher productivity through deepened partnership but separated responsibilities Sverker Hartwig, Vice President Technology, Atlas Copco CMT Business Area
Abstract In 1986 Atlas Copco introduced its first fully automated drill jumbo at Las Vegas. Miners were talking about the automated mine but in reality such advanced equipment had limited sales success throughout the past century. Remotely controlled loaders and trucks were introduced but have largely been abandoned and replaced by manually operated vehicles. One of the major drawbacks related to the success of mine automation has been the safety issue. As high speed tramming vehicles operate in confined, dark places it is of utmost importance to incorporate fail-proof devices detecting all personnel being in close proximity to the equipment. The author discusses the reasons for recent trends in mine automation. As Atlas Copco at the Sudbury ISMMA conference in 1999 took the initiative to IREDES, a common communication platform now exists to go one step further to get unified systems for navigation and other safety features.
PROLOGUE This paper is not written by an old grumpy man who never believed in technological breakthrough and thinks still, nothing can beat pusher legs. This paper is presented by a man even if old and grumpy, who his whole life has worked with technological development and still thinks we have seen nothing yet. But this old man has with time realized that we have to give the technique its own chance based upon its own merits and not try to make it copies of mankind. He has also realized that we no longer can afford being proud miners doing everything by ourselves. We have so much to learn from other industries! Finally, he has also understood that the mining business is too small for a lot of parallel developments. We need to work together and think longterm profit before short-term greediness. If we do so, we will be greatly rewarded. BACKGROUND Atlas Copco’s first attempt to develop an automatic drifter drill rig was made in the early seventies, using technique from military anti air guns. For the computer freaks I can mention that we based the electronics upon ECL circuits, since the first microprocessor Intel 4004 was not yet born. Never before or later was there a faster positioning rig, it could almost chase a fly with the tip of the drill rod, but the amount of printed boards was scaring not to say horrifying and the power consumption for the electronics was close to consumption of the rock drills. Later in 1986 Atlas Copco introduced the first commercial fully automatic 2 booms drifter rig at Las Vegas The Robot Boomer. Think about it, this is already now very close to 20 years ago. Of course we thought at the time that in a couple of years almost all rigs would be automatic. Yes there were also others, a Norwegian contractor, Furuholmen, had developed a rig of their own, of which traces are found with AMW or Bever Control, if you wish. Also Montabert came with their Robofore. Nevertheless, I guess that from 1986 until 1998 fewer than 10-20 rigs all brands were actually sold, and maybe none really operated in the automatic mode. In practice none of the rigs could really finish a whole round, something Massmin 2004
always went wrong or there were holes which were impossible to drill etc. As a consequence the customers always kept an operator aboard to solve problems, and our main argument for the added cost of the automation in those days, saving on personnel, fell flat. Now today at least a double-digit percentage of Atlas Copo’s rigs are sold in the full ABC version, as we call them, but still today, they very seldom operate unmanned. However, two things have been changed since then. Firstly, the rock drills are much faster so it’s difficult, if not impossible, for one man to keep up with three booms. Secondly, we know today that electronics provide other important qualities than manpower savings such as hole quality and repeatability. These arguments were only vaguely understood or mentioned in 1986. I would like you to keep the following sentence in mind for the rest of my presentation. After more than 30 years of development we have not yet reached a level of automation were rigs could be run autonomous. But the automation added features that were not foreseen at all when we started. These features may very well give a better payback than the "possibility" to run the rig unmanned, and at least have they greatly improved the working conditions for the remaining operators. Productivity measured in brutal tons per hour, meters per day etc. are still difficult to match when you compare with an all out manual operation! Is this just because we are stupid miners? Some of you may have heard about the one million dollar NASA contest in the USA recently. Mission: Have a vehicle of your choice with whatever electronics sensors and whatever gadgets onboard, to go from A to B passing C, D, E. and F in dessert type terrain, but with no interference from man. The winner, sorry there was no winner, managed only a few percent of the track before it collapsed. Those rocket science guys cannot do it either! How far away aren’t we from the long reaching ambitions from various Auto(mated)mine etc concepts. I also remember one project that strived for an unventilated and maybe not even drained mine, where unmanned rigs with close to 100% availability, were doing an everlasting high productivity job.
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Nothing wrong with the ambition, but maybe by the time we have made this possible, we also have found new ways of excavation with bacteria or other means. So far I have only discussed drill rigs, or better drifter rigs. It is true that the level of automation has gone further in production drilling such as Atlas Copco’s BK Simbas at LKAB, every rig are now producing an impressive 100 000 meter per annum in partly remote and partly autonomous operation. How come? A part of the explanation is obviously that these rigs are close to standstill in their drifts, but maybe more important. LKAB has fine-tuned the operation baered upon the pros and cons of such rigs, not doing absolutely the same thing in the same way as in previous manned rigs. So, what about the other unit operations underground. I, the undersigned, Atlas Copco and our dear colleague machine suppliers, and dear miner customers have all burned our fingers here too! Not to mention anyone in particular, we all know of automated loading and/or hauling projects, initially promising, which have been abandoned due to lack of productivity in a wide sense. In cases of the operation itself and in other cases, due to the look up in a mine having critical areas blocked. I will not put my head into the fight between truck haulers and conveyor friends, but the safety measures around for instance an automated wheel haulage system sometimes makes it less flexible than a conveyor system. A sad part here is that at the same time we learn that in many a case did the machines perform better than an operator regarding tire wear, breakdowns etc. But still the whole thing did not perform! An other interesting observation is that in some projects the actual loading was made automatic but hauling manned and in some applications the opposite, loading was supervised but hauling left to automation. What conclusion can we draw from that? In a way these above applications represented an unnecessary exercise. In the manufacturing industry selfmoving pallet movers became out of fashion at same time as we started to move them underground and even worse, they became out of fashion due to the same limitations we now see underground. There are two problems; one is in our own mind and the other in the machine. Most systems of this type use magnetical guidance from a cable in the floor. The pallet mover can move along this wire at a low speed for safety reasons (much lower than a regular fork lift truck with driver) Whatever obstacle along route will stop the pallet mover including an empty paper box. How difficult it would be, to make this mover advanced enough to understand type of obstacle, move around if possible, or take another way to target, we all know from the NASA experience. And if, would we ever allow this vehicle to move around with forklift speed? And again horrible thought, even though we unfortunately face multiple accidents with manned vehicles inside factories, would above unmanned system survive one single lethal accident? The other problem with the pallet movers is in our minds. The concept was intended to give freedom and flexibility. Without effort we could move goods around in the shop irrespective where the different machines were placed. But we got almost the opposite and the transparency of the system was only visible for the few planners who really understood how the control computer behaved or misbehaved. For all others things just moved around, in and out of pallet storage in a magic way, and we had no chance to realize if we would get parts needed at our place, now or never. In a shop designed around the flow concept is really not a need for pallet movers, cruising around in the factory. 712
In the flow concept in the best of worlds, the output from one machine centre is the intake of the next. No material is in transit or even worse in storage. From the factory comes a continuous steady flow of products. Isn’t that what we always wanted from a mine? Or do we prefer mining equipment moving around by itself in a magic dance for minutes but standing for hours waiting for a human to sort out problems, obstacles or just the fact that somebody rattles the protective fence. Where are we today in the mining industry? What to do and what should not be done? COMMUNICATION UNDERGROUND The IREDES Initiative I would lie if I did not admit being proud of taking the IREDES Initiative at the Sudbury ISMMA conference in 1999, a common communication protocol platform, which now exists. But I am equally honest when stating that I am proud of our industry, fellow machine manufacturers and users joining in, unselfishly forgetting short term considerations for long term profit for us all. The progress of The IREDES has been presented elsewhere but please, if you are still not a member - sign up! Now I will repeat something I said in Sudbury, maybe it was wrong then, maybe also today, but it could very well be true. The stock value of Microsoft is higher than the market value of all publicly listed mining companies. My estimate is, that the global mining industry (machine suppliers and users) employ less than 1000 professional "computer" engineers, hard and software, working with automation and control. This can be compared to an estimated 170 000 ditto working with cellular phones and networks. Who gave us the idea that we should develop our own communication systems? Mind you IREDES deals so far only with the protocols "The language" not with what type of systems we use for transmitting it. It’s easy today to be ironic about the performance of "leaky feeder systems etc." It’s also unfair because in those days, it was not much that could be used. BUT now there is WLAN. What lovely automation network, or Wireless local area network as some old-timers still call WLAN. Yes, some of these components need to be mine ruggedized and some works remains on roaming, but it can still do a lot (and very much more than present mining systems in terms of video channels etc) and at unbeatable cost levels. Will it change over time? Yes! Will the old WLAN work in future WLAN? Yes to most parts! Will it continue to offer better communication at lower cost? Yes! I say, let’s go for WLAN (and TCP/IP) Using IREDES Protocol and I say, let’s together sort out the few remaining question marks about mining applications. Maybe we could use the IREDES organization tofacilitate it over company borders. Give me a call! I will happily host a kick off meeting. Think about it. All mining personnel and vehicles in South America or Russia of make A, B, or C are using communication system and same digital language. Plug and play as they say. And fixed installation in mines at a fraction of today’s cost. And finally, the system transfer capacity will grow with future needs.
NAVIGATION AND OBJECT (LIVING) DETECTION UNDERGROUND In view of the NASA competition I fear that this is an area where the mining industry is going to pump a lot of good money into a large black deep hole.
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We, LHD manufacturers, think that we have come a long way with our different scanning laser systems. But I am not sure that we are right. Will they ever work with drill rig accuracy? And what about living object detection? If we cannot solve that we are stuck with locked up areas in our mines and that we know how far it takes us! I would have loved to suggest an all out cooperation in this field but I fear that there may be too much pride about the scanning lasers systems. BUT what about living object detection? How much work has already been done by for instance the automotive industry? Is not this an area where we all need to work together; manufacturers, mines, unions and safety regulation authorities? Things can and will go wrong and we need to stand up as one man to fix the problems. Without a 99.999999% functionality here, I am sure that we will never reach commercial autonomous operation underground. On the other hand I am sure that we, with a combination of different systems such as detectors on vehicles, transponders on living objects and mine staff mapping systems, can reach there. Again I say, Give me a call. Atlas Copco is willing to host a starting session that would lead to an independent organization like IREDES with some money at least to find out what has already been done by others. MINE PLANNING This is an area, which I always treat with respect and prefer to stay out from. I am a designer of mine machines and not a designer of mines. Nevertheless, I have been involved in a few lucky projects where the mining company could start out all fresh in a new found deposit. On occasion these plans were made for different types of "mobile miners" i.e. mechanical rock cutting machines. Having worked a number of years with these "monsters" I still have the dream, but will it come true in my life? Close your eyes and dream with me! One machine with few, if any, operators are doing all the operations, pumping out ore in the rear end onto a conveyor or in a pipe. No logistic problems whatsoever. Nothing of all preparation works, minimum ventilation and so on and so on. The underground flow concept!
THE MINERS DREAM! As we all know few of these magnificent plans came true. Nobody dared the risk even though calculations looked fine. Only in coal potash and the like is ore excavated this way. But also for drill and blast a new deposit opens up possibilities not obtainable in an old operation. It is my understanding that the way to fortune and happiness here is in split-level operations. If all haulage could be done on one level, loading other one etc we have much bigger chances to design a mine where different machines running or at standstill do not interfere. Maybe we could even have areas which are more conveniently sealed off for people. I don’t hesitate saying that separate mini-drifts for people leading to the "few" points where man is inevitable should be considered. In communication between supplier and user we have a problem. With all rights, the users are chasing us suppliers for faster, cheaper, and more reliable machines. But are we chased in the right directions? Is there a common understanding - when we are discussing automation - what is state of the art - what is the future? Let me give a simple example. Mines of today are designed for manned machines. Therefore we have rigs Massmin 2004
with cabins, lots of lights, ventilation and so on and so forth. An unmanned rig does not necessarily need all that and some of it is both costly and cumbersome. A manned rig has all the possibilities in the world to solve or work around an unexpected problem. The unmanned has only little or nothing of that. We can cut out a lot of gadgets needed for manned operations yes, but instead we need to create an environment suitable for the autonomous rig where things are alike and stay that way. I am sorry to conclude that we started all wrong. We have tried to design computerized mine-machines that act like manned machines. Suppliers and users are to blame alike, because then at least, we know what we are talking about. I don’t like to admit it, but I think many of us deep inside understood, that it would never work. In all we have made great progress in designing systems that help and unload an operator and are doing many of the details equally well, or even better than the operator. But to take him away, forget it! We need a restart, starting from a theoretical viewpoint and we should stop making funny machines. We need an autonomous mine research think tank with advanced researchers who are not stuck in the old way. Many of these, certainly more than 50%, have to be miners. The approach should be to, theoretically quality- and quantitywise analyze what the unmanned machines can really do with reasonable availability and productivity, and to start designing the mine around it to help them out. Yes I understand that these were harsh words for some miners. We suppliers should design equipment that fit the mine not the other way around. Let me offer a compromise. We do our best along that way on the manned side and you give my idea at least a thought on the unmanned. Also here I say, Give me a call. Atlas Copco is both willing to help this process going and also to support it on its way inviting all makers and users alike! SUMMARY I have from my experience tried to understand why we are so far from autonomous mines today as we in fact are (there could be papers from Atlas Copco and others saying something different) The main reasons seem to be that we started all wrong trying to design human like "robot" operated machines assuming that these "robots" could think. To speed up the process we didn’t bother to learn from other industries nor did we cooperate. We followed instead the multiple path system and this in an industry with very limited resources. But there is light in the tunnel. There may very well be other ideas and concepts than those I presented here, which are equally important but those I would like to suggest are important to maintain momentum in our efforts Lets together develop WLAN and HUMAN DETECTORS so that we at least have a chance to talk to our machines in a global way and so that we protect the future of autonomous mines not being stopped by causing injuries to our workers. And finally let’s take a step back from the route we are following and sit down and think. What can we really expect from machines with their pros and cons and how do we design our mines to maximize the output thereof? In all these cases Atlas Copco is also willing to invite you to a starting session to form an independent approach. Atlas Copco is also willing to support the work and safeguard a positive outcome. Give me a call!
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• Sverker Hartwig is Vice President Technology in the Business Area Construction and Mining Technique at Atlas Copco AB in Stockholm, Sweden. In this position he coordinates all development and design in Atlas construction and mining divisions.
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• Sverker Hartwig joined Atlas Copco 1975 and has then had numerous technical and commercial positions within the Atlas Copco group working for Atlas Copco companies in Sweden, USA, and Switzerland. Sverker Hartwig graduated as an M.Sc.EE from the Royal Institute of Technology in Stockholm 1974.
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Chapter 19
Miscelaneous
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Air inrush risk assessment for caving mines Andrew Logan, Business Improvement Manager, Newcrest Mining Limited, Melbourne, Australia Duncan Tyler, Geotechnical Manager, WMC Resources Ltd, WA. Australia
Abstract Hazardous air inrush events occur in underground mines using caving methods or large roof spans. This paper outlines a simple, practical methodology to assess potential wind gust velocities in mine tunnels associated with collapse of roof material into a void and diffused through a broken rock pile. It also outlines tolerances based established meteorological wind classes. These guidelines and associated control measures were developed by Newcrest Mining Limited for safe sublevel caving at Ridgeway Gold Mine. INTRODUCTION The Ridgeway Gold Mine (Ridgeway) is located near Orange in Australia. Ridgeway and its neighbouring open pit mine, Cadia Hill, are part of Newcrest Mining Limited’s Cadia Valley Operations. The Ridgeway ore body is well suited to the sublevel caving (SLC) mining method adopted. Mining starts at the top of the ore body, 550 m below surface with ore extraction progressing downwards and cave propagation upwards. The porphyry gold-copper ore is broken by drilling and blasting and is extracted incrementally in 25 m horizontal slices over the ore body footprint. Ore is extracted at 5.5Mt per year from the SLC, tipped to an underground primary gyratory crusher and transported via a 4km inclined conveyor system to surface. There are several principal geological and geotechnical domains (Figure 1). The North Fault also represents a significant weakening structure considered important for caving initiation and propagation. The major principal stress is sub-horizontal, oriented east-west. The production layout has been designed to allow effective draw control by including transverse extraction; sublevel intervals of 25 m; 6 m wide x 4 m high ore drives; and 14 m ore drive centres. Caving of the overlying strata is induced as it is undercut by ore extraction. Broken rock created by caving of the overlying strata fills the void created by ore extraction. Ore dilution by waste rock from above is controlled by optimising the design of the mining layout, by disciplined production mining practices and draw management. At Ridgeway’s Feasibility stage, the recognized standard method for predicting caving initiation was the Laubscher (1999) method. It predicted Ridgeway caving initiation was likely. The effect of the North Fault on Ridgeway’s caving assessments was unknown. Caving occurs as a result of two mechanisms: gravity collapse of well structured rock masses and induced stress fracturing of less structured rock masses. With its good rock quality, induced stress fracturing was the primary mechanism anticipated at Ridgeway. The conceptual cave model presented by Duplancic and Brady (1999) describes the cave front behaviour in this case (Figure 2). As shown in Figure 2, the cave roof ‘necks’ as it propagates upwards. The better the rock quality, the more quickly the necking with occur. The larger the ratio between height (distance from cave initiation to free surface) and initial minimum caving width, the higher the risk this necking will cause the cave to stall. Confidence in cave initiation at Ridgeway was high. However the height to minimum width ratio of the cave Massmin 2004
Figure 1: Schematic cross section through Ridgeway deposit (2001) (550m:180m or approximately 3:1) was believed to be greater than any successfully attempted. The Ridgeway cave was therefore beyond previous precedent, leading to some uncertainty about cave propagation. INDUSTRY AIR INRUSH EVENTS An air inrush, wind gust, or air blast event can be defined as "a sudden movement of air, displaced by a collapse of rock into an underground void, which causes the air to move through the adjacent openings", Fowler et al (1996). A wind blast can be also described as an event resulting in air movement underground that causes injury and / or seriously disrupts ventilation. There needs to be a void for an air blast to occur. In caving operations an appropriate expansion void is an essential part of the mining process. Geometric factors such as volume; plan area; height; shape and hydraulic radius must also be considered.
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develop and maintain hazard management procedures for the management of: • The void above the muckpile; • The height of the muckpile above the extraction horizon, and • The air blast hazard and shall include all the appropriate controls for the air blast at all openings or potential openings into the cave zone Management of the major hazards in a block cave mine must include recognition of the facts that these issues are interrelated and cannot be managed as discrete elements".
Figure 2: Conceptual model of caving (After Duplancic & Brady, 1999) Rock or other material (water or fill) above a void provides a potential energy source if a void is allowed to grow beyond tolerable levels and the progressive roof collapse slows relative to void growth. This potential is converted to high air pressures if the roof becomes rapidly unstable and collapses into the void. Factors that impact on this instability include span, shape, mass, proximity to the surface, rock fabric, major structures, stress levels, mode of failure and trigger mechanisms. The high air pressures are converted to wind gusts when they find an escape pathway to areas of lower pressure (e.g. surface). In caving operations, the broken rock pile in the caved zone, provides a buffering effect to dissipate some of the high air pressures. Considerations include: broken rock pile thickness and permeability; the number of openings; location of openings size of openings; path of least resistance; ventilation controls; access to openings; and connection to surface. High wind gusts become hazardous when the paths of rapid air flow concide with working location of people and/or infrastructure. Considerations for assessment of potential hazard to people and infrastructure include: location of manned work areas with respect to the escape pathways; and the provision of pressure relief pathways and restricted access. Air inrush events have been recorded in both the underground metalliferous and coal mining industries. Both rely on caving and progressive roof collapse to enable safe production. Air displacement from falling rock is, therefore a common event in such mines, which require an appropriate expansion void or air gap to allow progressive collapse and/or cave propagation to occur (Figure 2). Most of these events remain unnoticed due to their very small magnitude. Although there are numerous accounts of air blasts in general and technical literature, very little detailed work has been done on understanding the phenomena and providing practical prediction tools. In November 1999, a catastrophic event occurred at Northparkes Mine where the cave back collapsed into the mine’s air gap. The cave back had formed a stable arch and an air gap of some 180m was allowed to develop, with nominal muck pile height of 60m. The force of the blast was such that roof bolts and mesh were bent, motor vehicles destroyed and 4 workers in the vicinity were killed. The following salient findings and recommendations have been extracted from the resulting coronial report, Bailey (2003): "Any mine operator intending to employ the process of block cave mining is to identify and analyse the elements of all the risks associated with its block cave operations and 718
Fowler and Hebblewhite (2003) also discuss in detail issues associated with Australian coal mine wind blasts. The most recent fatality to wind blast in the Australian coal industry was in 1976 at the Eastern Main Colliery, NSW. They report that, air blasts resulting in serious injuries have occurred in the following NSW coal mines Cooranbong (1983), Wallarah (1989), Myuna (1990), Newstan (1995-96), Newvale No 2 (1995), Gunnedah (1998) and Moonie (199899). Air blasts in Australian coal mines since 1895 have resulted in the deaths of nine workers. A potentially disruptive air inrush was thus considered to be one of Ridgeway’s major hazards that required systematic assessment and effective controls. This required research, development and innovation. RIDGEWAY INNOVATIONS New processes were developed by Newcrest to manage the air inrush risk at Ridgeway. These were incorporated into an inrush hazard management plan with preventative controls, monitoring, triggers and responses to proactively manage air inrush, NML (2002): 1. Blasted ore was left in the cave as a ‘blanket’, predominantly on 5330 & 5305 levels, and as a guard to dissipate potential air inrushes. 2. An integrated cave monitoring system was developed, using the data from a number of sources allowing cross validation of data, and the effective tracking of the cave through to the surface: This included successful use of 500m deep holes for • surface seismic system; • open-hole depth plumbing; and • extensometers. 3. A wind gust model was developed specifically for Ridgeway in order to reduce the technical risk associated with air blasts. This model predicts air inrush velocities resulting from a massive cave back failure, and is fed into triggers and responses built into the hazard management plan. 4. Air inrush triggers and responses were developed to match existing wind force standards using the Saffir Simpson Hurricane Scale as no guidelines were available in the mining industry. 5. Response plans were specifically developed and implemented (Table 2). 6. Access drives were stood further out than common practice to allow for cave footprint and width extension responses. 7. Hydro-fracturing to stimulate caving was researched and trialled as a response. From systematic measurement, new knowledge was acquired regarding initiation and propagation of caving in strong rock masses where mine scale discrete structures are predominant. Under these circumstances initiation and propagation of the cave appear to be controlled more by the discrete structures rather than the general rock mass rating.
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This suggests that major structures should be explicitly taken into account in predictive methods. The Ridgeway cave was designed and developed at a height to minimum width ratio outside current worldwide experience - greater than height to width of 2:1. Flores and Karzulovic (2002). This significant extension of existing mining practices was made achievable by the beneficial effect of North Fault as a weakening structure. WIND VELOCITY ASSESSMENT A credible peak wind velocity due to an air inrush event is a function of several key variables: Wind gust potential (m/s) = f (expansion void, muckpile thickness, permeability of the broken rock pile, nature of collapse, number of exit paths, measurement error) (1) Ridgeway first developed a simple, practical model to assess the instantaneous compression pressure generated in the broken rock pile due to cave roof collapse in a void, using Boyles Law. Conservative assumptions were made regarding the nature of the collapse as a piston. Wind velocities where then subsequently assessed from these overpressures by air flow principles. Buffering effects were considered and measurement accuracy and uncertainties were included in the analysis. The model can be used to assess the overpressure developed over a range of broken rock pile heights, air gap heights and broken rock pile fragmentation types following the instantaneous collapse of the cave back, using a leaky piston model. The potential compression pressure (P1) from atmospheric pressure (Pa) generated by the collapsing cave roof was estimated from the formula:
Different muck pile porosities and resistances were assumed for the different density broken rock piles. The assessment is sensitive to the permeability assumption. Permeability of the large scale muckpile is difficult to directly measure and requires interpretative assessment from qualitative muckpile fragmentation interpretation and back calculations. For example, an extrapolated broken rock resistance for a finely fragmented material could be in the order of 0.11 Ns2m-8. Significantly more work is required to establish working guidelines for fine, medium and coarse material. The theoretical peak velocity (TPV) is thus a combination of the previous formulae and site specific measurements and variable assessments. TPV = from (2), (3), (4), (5)
The Ridgeway formula and parameters are intentionally not shown here, as they are site specific. The intent here is rather to explain a method whereby individual locations might develop their own site specific formula. Further formal research along these lines may also lead to more widely applicable guidelines, which would be beneficial to the mining industry in general. Mine ventilation assessment tools such as VENTSIM, can also analyse multiple flow paths as they have many of these concepts embedded in their methodology. Fowler and Sharma’s (2000) measurements of wind velocities from coal mine goaf collapses showed that there were buffering effects that limit the TPV through relatively small underground development drives. The TPV is thus reduced by a buffer reduction factor (BRF) to give a practical peak velocity (PPV): PPV = BRF x TPV
Boyles Law
P1 = Pa [(Vag + Vrp) / Vrp ]γ
The principal unknowns regarding the prediction of pressures are the rate at which the back falls; the degree to which it breaks up (thus permitting leakage and pressure loss); vertical height of the fall (the rate of pressure build up must relate to the rate at which the back falls – up until terminal velocity is reached); plan area of the fall and the air flow resistance of the connected openings. The conservative assumptions that were made to permit the simplistic estimation of the void pressure which could be generated during a caving event were: i) pressure rise is very rapid (in reality peak pressure may take several seconds or more to be reached); ii) compression of air below the fall is uniform; iii) the complete cave back collapses as a single slug, with no friction on the sidewalls; iv) all of the air below the collapse is compressed and a vacuum is created above the fall (there is no leakage); and v) the air below the collapse is compressed into the voids in the broken rock pile which is in the base of the cave and there is no overall change in the volume of air contained in the system. The assumptions identified represent a "worst case’ air pressure model, and in reality the generated pressures will be lower than those predicted. Using the peak over pressures calculated above and applying them to different muck pile resistances it is possible to assess the corresponding flow rate through the broken rock using the ventilation equation:
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(7)
(2)
where γ is 1.4 for air, Vag is the volume of the air gap and Vrp is the volume of the broken rock pile.
Pressure = resistance x flow2 Velocity = Flow / Cross Sectional Area Volume Air = Porosity x Total Volume
(6)
(3) (4) (5)
When substituting the variables back into formula (7) developed from formula (1), the following points were considered. The expansion void value is assessed by direct and indirect measurements, and draw control records. The muckpile thickness is the thickness of broken rock overlying the first extraction horizon. This value is also assessed from direct and indirect measurements and draw control records. Measurement error needs to be considered and a conservative view taken. The large scale permeability of the broken rock pile is assessed from a site specific fragmentation interpretation and estimated from back calculation of inrush records from mines with similar characteristics. The nature of collapse was considered as a plug failure as previously discussed. This is the most efficient failure mechanism for driving the air out of the air gap and through the muck pile. This is a considered a conservative failure mechanism, but it has been interpreted to have occurred at both Northparkes and Salvador mines. The duration of the failure can be conservatively assumed to be 1 second for a first pass assessment. The number of exit points relates to the pathways available for any potential air blast to exit the uppermost extraction horizon and vent to surface, and Measurement error is related to errors associated with the above measurements, interpretations and estimations. This assessment can be entered into a spreadsheet coupled with a probability tool such as @Risk to aid the assessment of input variability. Wind velocity values can be assessed for a series of expansion gap and broken rock pile heights for different broken rock porosities. Upper quartile values can be used to increase the conservatism of the assessment. A series of charts can be developed into a nomogram (Figure 3) that links the expansion void height
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(first x axis) to variable broken rock heights to give a wind velocity assessment through a single exit tunnel (y axis). The nomogram also shows the reduced wind velocities for multiple exit paths (second x axis). Separate nomograms are required for different broken rock fragmentation and permeability. WIND TOLERANCES Limited international mining guidelines are available for definition of tolerable wind velocities. Meteorological classifications of tropical storm wind velocities are well established however. These scales provide estimates of the possible impact of winds of the same magnitude as those that might be generated by a significant cave failure. Using this information, estimates of potential damage over a range of possible wind gust magnitudes can be made. Wind velocities from these scales are linked to injury / damage. The following wind velocity models can be used: • The Beaufort Wind Force Scale ; • The Saffir-Simpson Hurricane Scale; • The Tropical Cyclone Intensity Scale; and • The Fujita Tornado Intensity Scale. It was interpreted that the most applicable meteorological scale to use is the Saffir-Simpson Hurricane scale; because wind gust velocity is measured at ground level in this scale. All of the other scales measure wind gust velocity 10m above ground level. Based on the above, five maximal wind gust potential classes were identified and are presented in Table 1. This classification is support by Australian Coal Association Research Program (ACARP) work which indicates that laceration injuries of uncovered skin would occur for wind velocities exceeding 15m/s for projectiles weighing 10 grams or less. CONTROLS The primary preventative control for major air inrush events is to inhibit the development of an inappropriate void or air gap above the broken rock pile. This can be achieved by adopting ‘strict’ draw control strategies in conjunction with cave back monitoring. Systematic measurement, ongoing assessment processes, well defined action ‘triggers’ and management agreed responses are important process controls. Responses target void reduction, dissipation of potential energy, securing of escape pathways and tightening of working controls. Actions may include increased monitoring intensity; cave inducement (hydrofracturing or blasting); deferred draw; level closure; extension of cave footprint; and surface access exclusion of the potential subsidence zone (Table 3). SUMMARY The Ridgeway sublevel cave was implemented with the understanding that air inrush events have historically occurred and needed to be proactively managed. Preventative, monitoring and responsive risk management strategies were implemented to effectively manage this air inrush risk. A new methodology to assess potential wind gust velocities associated with cave roof collapse was also developed, along with wind velocity tolerances from meteorological wind classes. These controls significantly contributed to successful cave propagation and continuing safe production at Ridgeway.
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Figure 3: Ridgeway wind blast nomogram
ACKNOWLEDGEMENTS The author is grateful to all of the people at Cadia Valley Operations who helped produce information and methodology presented here. In particular Professor Ted Brown for his review of the airblast hazard management systems reported here. The author also wishes to acknowledge the permission given by Newcrest Mining Limited for publication of this technical paper. REFERENCES • Australian Mining Consultants, 2000. Ridgeway Gold Mine Ventilation Bulkhead Design. Derrington A.S., AMC Job Number 300029. September 2000. • Bailey J. 2003. Findings and recommendations: Inquest into the deaths of R Bodkin; M House; S Osman and C Lloyd-Jones on the 24th November 1999 at the E29 Lift 1 Mine. NSW Coroner’s Court, 18 March 2003. • Duplancic P. and Brady B. H., 1999. Characterisation of caving mechanisms by analysis of seismicity and rock stress. In Proceedings 9th International Congress on Rock Mechanics, Paris, Eds Vouille G and Berest P, 2: 1049-1053. Balkema: Rotterdam. • Flores G. and Karzulovic A., 2002. Geotechnical Guidelines for the transition from open pit to underground mining, ICSII task 4. Report to the International Caving Study • Fowler J.C.W. and Hebblewhite B.K., 2003. Managing the hazard of wind blast/air blast in caving operations in Australian underground mines, In Ground Control in Mining: Technology & Practice. Ed Hebblewhite B.K., ISBN 0 7334 2085 0, UNSW School of Mine Engineering, Sydney. • Fowler J.C.W. and Sharma P., Jan 2000. The dynamics of wind blasts in underground coal mines, Final Project report (No 4), Project No C6030. ISBN 0 7334 0700 5, UNSW School of Mine Engineering, Sydney. • Fowler J., Torabi S. and Daly C., 1996. Field Investigation into Windblasts Resulting from Large Falls of Roof in Australian Underground Coal Mines, In Mining Science and Technology. Eds Guo and Golosinski. AA Balkema. • Laubscher D.H., 1999. Block and panel caving, Internal report to the International Caving Study. • Newcrest Mining Ltd., 2002. Ridgeway Gold Mine Inrush Major Hazard Management Plan, Internal report to Newcrest Mining Pty. Ltd.
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1: Wind gust damage classes Wind gust potentials
Classification & reference
Potential impact
Up to 15m/s, (55 kph)
Moderate gale (Beaufort scale no. 7). Wind gust class = green;
Difficulty walking into wind.
15 to 33m/s (55 to 125 kph)
Cyclone / hurricane Category 1, (F0) tornado. Wind gust class = yellow
Some damage to windows, signs, vent bag, air and water pipes. Unprepared person may be knocked over.
35 - 45m/s (125 - 170kph)
Cyclone / hurricane Category 2, Moderate (F1) tornado. Wind gust class = orange;
Significant damage to signs vent bag and pipes. Some automobiles overturned, small projectiles (sand).
45 - 60m/s (170 - 225kph)
Cyclone / hurricane Category 3, Significant (F2) tornado. Wind gust class = red
Major damage – some de-mountable offices pushed over or destroyed. Light object missiles generated.
>60m/s (225kph)
Cyclone Categories 4 and 5, Severe to Inconceivable (F3-F6) tornado. Wind gust class = red
Severe Damage – Heavy cars lifted off ground and thrown. Metal buildings collapsed or severely damaged
Table 2: Wind gust speed, trigger and response plan Measurements
Trigger or key decision point
Planned response
Estimated expansion void. Broken material including ore above uppermost active production level. Time since last cave movement seismicity
Condition Yellow Level 1. Potential velocity 15 – 35m/s; (55 – 125 kph) OR Cave back has not moved for 3 – 6 months & little/no recorded
Increase monitoring intensity; review draw records; prepare for level closure
Condition Yellow Level 2. Potential velocity 15 – 35m/s; (55 – 125 kph) AND Cave back has not moved for 3 – 6 months & little/no recorded seismicity
As above; implement underground hydro-fracturing cave inducement strategies
Condition Orange. Potential velocity 35 – 45m/s; (125 – 170 kph) OR Cave back has not moved for 6 – 9 months & little/no recorded seismicity
Deferred draw; full level closure; extend cave footprint; plan to evacuate mine if there is a marked increase in the seismicity in a 4 hour period; exclude surface access to subsidence zone if cave could collapse through to surface
Condition Red. Potential velocity >45m/s; (>170 kph) OR Cave back has not moved for over 9 months & little/no recorded seismicity
Immediate and full mine evacuation; remote hydro-fracturing if safe to do so; exclude surface access to subsidence zone if cave could collapse to surface
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Cavity monitoring system and stope analysis John D. Lupton, B.Sc., Arts, AscT Optech Incorporated, Toronto, Ontario, Canada
Abstract This paper will discuss the benefits and advantages of using 3D survey data in stope analysis. Obtaining detailed 3D survey data of underground mine progress and production is a challenge. It is almost impossible to "see" around corners, underneath overhangs, or deep into dark, dusty, humid underground voids. Thus, valuable ore can remain permanently encased in backfill. Without the benefit of 3D survey data, underground mines have the difficult task of meeting grade quotas without the benefit of knowing exactly what has been mined. Lack of this data usually results in unfavorable month-end reconciliation factors, where grades must be adjusted to match grades assessed by the Mill.
1 INTRODUCTION
to determine the over/under breaks. A simple formula is used to calculate dilutions:
Any new technology adopted by an underground mine requires a learning curve. However, laser survey technology is the quickest way to move a 2D mine into a 3D world. 3D survey data is much easier to comprehend and communicate. This data makes it easier to make and back up difficult decisions. Combined with one of the many powerful 3D mining software packages available, a mine can improve grades by building a CMS database, as seen in Figure 1.
Dilution (%) = CMS Survey Data Reserves Building a CMS database begins with surveys of primary stopes. Planning will then design secondary stopes based on actual mined data of primary stopes, thus maximizing recovery and minimizing dilution. This will improve monthend reconciliation and overall grade control, resulting in a more efficient mine.
Figure 1: CMS Data.
Figure 2: Blast Ring Comparison.
The value of 3D survey data is to provide a safe benchmark that will be used for production data. This actual accurate 3D survey production data will be compared to blast design and mine reserves. For example, a set of blast rings will be designed to maximize recovery and minimize dilution. Once these new rings have been blasted, excavated and surveyed, the results will be compared with ore reserves and the original blast designs. The results will be analyzed for tonnage of lost ore (underbreak), tonnage of waste rock (overbreak), and accurate grades sent to the mill (reconciliation). Ultimately the results will be analyzed for the success, or lack of success, of drilling and blasting.
For example in Figure 2 the secondary stope was designed based upon the survey results from the primary stope (which is now backfilled). The secondary stope is then blasted, mucked and surveyed with a CMS (dark line). CMS scan results are sliced at the same coordinates as the blast rings and draped over the ore body (dotted line) and blast rings. The first blast ring (left) shows unsuccessful blasting and the second blast ring (right) appears to be marginal. A close examination reveals a substantial amount of material left in the lower hanging wall. However, when the grade is draped over, the remaining material actually contains very low-grade ore (0.90). Strategic blasting in this stope, that left low-grade material in the stope, saved the mine almost $9,000 (CDN). Further analysis will reveal exact tonnage of ore/waste left in the stope, and ore/waste hauled, or not hauled to the mill. Plans to recover any remaining ore will be implemented when designing the tertiary stope.
2 STRATEGIC BLASTING IN 2D Working with 2D data, planning would simply slice the 3D survey data at the desired coordinates, azimuth and dip. This 2D "slice" would then be draped over the 2D ore body 722
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3 STRATEGIC BLASTING IN 3D With one of the many powerful mining software programs such as Gemcom, calculating overbreak/underbreak is painless. Users can overlay CMS polygons with the block model and a simple "clipping" routine provides the tonnage of overbreak/underbreak such as below. In the event of unsuccessful blasting, new bast rings must be designed to remove remaining pillars. In Figure 3, the image on the left illustrates unsuccessful blasting potentially leaving behind thousands of dollars in ore. Since the stopes on both sides have already been backfilled, a new set of rings must be designed, blasted and cleared quickly without disrupting production. The recovered pillars seen on the right contributed several thousand extra dollars in ore.
For example, an underground mine in Ontario was experiencing difficulty mounting a sloughing ore pass, most of which went undetected. With the use of a reflectorless automatic laser scanner, the mine engineer was able to obtain an accurate image of sloughing ore pass as seen in Figure 4. Immediately the chief engineer justified the closing of the drift above the pass. Almost two weeks later the closed drift completely failed into the ore pass below without injury to any employees or equipment. Sloughing ore passes and hanging walls are perilous areas to send miners. With a CMS it is possible to obtain detailed information without unnecessarily risking injury to mine employees.
Figure 5: Stope Comparison In Figure 5, the original stope design included mining 19% backfill, 22% air, and left 33% additional available ore in the stope. The grade would have been 0.277. The redesign increased the grade to 0.344, an increase of 20%. The redesign included blasting under part of a backfilled stope. After blasting, mucking and surveying, the final grade was 0.337, which were still 18% better than originally designed. Figure 3: Blast Success Comparison. 5 CONCLUSIONS Sloughing rock usually goes undetected. By monitoring rock behavior with a 3D laser survey instrument, specifically hangingwall stability, the 3D survey data can be used to predict unplanned dilution (overbreak) in future stopes. Rock mechanics use CMS data to determine parameters surrounding hangingwall stability. Using this data along with the known amount of undercut, dip, strike length, and mucking time, a model is developed to predict dilution in future stopes. This new model will provide the information to reduce future dilution, thus creating a more efficient mine. 4 MINE SAFETY Since sloughing rock usually goes undetected, monitoring unsafe areas prevents injury. By monitoring inaccessible areas potential problems, including injuries, can be avoided.
CMS data cannot be accepted blindly. Errors, including those caused by water in the stope, are always possible and should keep the processor awake. Laser beams reflect off water and can be detected immediately as unusually long distances. Also any hanging cable bolts will interfere with the laser contacting the rock. All these errors must be edited before processing. It is important to build a CMS database that can be consulted constantly when developing adjacent drifts or stopes. If a stope is backfilled without being surveyed, then your database is not complete. Ultimately a more efficient mine means better grades, increased production, and less month-end reconciliation. CMS continues to enhance efficiency and safety in over 100 underground mines around the world. ACKNOWLEDGEMENTS The author wishes to thank Optech Incorporated for permission to present this work and for the many CMS customers over the years without whose assistance this paper would not be possible. REFERENCES • Yao, X. (Mike); Allen, Gary; and Willett, Mike, Dilution Evaluation Using Cavity Monitoring System at HBMS Trout Lake Mine, 1999. • Knight, Adam, The Use of the Cavity Monitoring System and VULCAN at the Jerritt Canyon Joint Venture, Northeastern Nevada, 2001. • Duke, Cliff, Cavity Monitoring System as a tool for Reserve Reconciliation, 1999.
Figure 4: Sloughing Ore Pass.
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Preliminary ventilation design for the Grasberg block cave mine Ian Duckworth, Technical Expert Ventilation, Ketut Karmawan, Superintendent Ventilation, P.T. Freeport Indonesia, West Papua, Indonesia Timothy Casten, Senior Manager of Underground Planning, Freeport-McMoRan Copper & Gold Inc.
Abstract P.T. Freeport Indonesia operates a mining complex located in the highlands of Papua, Indonesia. This complex consists of both underground and surface operations. By approximately 2014 the existing Grasberg open pit will be exhausted, and a new underground mine will have been brought into production to cave the deposit below the pit. This paper describes pre-feasibility ventilation planning for the Grasberg Block Cave Mine. The mine is being designed based on a nominal production rate of 115,000 tonnes/day with a panel caving footprint of approximately 1 km by 1 km. The proposed ventilation design allows for the long production panels to be broken into five discrete ventilation zones, and ensure the economic delivery of large volumes of air to the working regions. Discussion is provided on ventilation criteria, network modeling, examination of shaft versus drifting options, proposed infrastructure requirements, and recommendations for future study.
1 BACKGROUND ON PTFI
2 GRASBERG BLOCK CAVE MINE
mine support services will utilize diesel-powered mobile equipment. Due to relatively cool virgin rock and ambient air temperatures, mechanical cooling or heating of the air will not be required. The main access for men and materials will be via the AB Tunnels and a Service Winze. The GRS BC operation will include the following components: • An 8 km long light rail passenger/freight system connecting a surface station with an underground terminus. This supply tunnel will also provide for three other proposed mines, called the MLZ Block Cave Mine, Kucing Liar Block Cave Mine and the Big Gossan Sublevel Stope Mine. It is likely that all these operations will be developed during the period of the GRS BC mine life. • An underground crushing system with twin conveyor drifts linking the crusher bins/feeders to the surface milling facilities. • A concrete-lined Service Winze with man cage providing access between the rail terminus and block cave operation. The length will be approximately 300 m. • Single ramp linking the GRS BC rail terminus with the orebins/conveyors and the production levels. • Block cave operation consisting of 2840 m Undercut, 2820 m Extraction and 2790 m Service Levels (elevations above sea level).
The Grasberg open pit is scheduled to cease production in 2014. Since the orebody continues at depth below the economic limits of surface mining, the reserve that is left behind will be extracted using block-caving methods. Development of the Grasberg Block Cave Mine (GRS BC) is scheduled to commence during 2008 and will be accessed via the Ali Budiardjo (AB) Tunnel system. The mine will be a mechanized block caving operation with a nominal production rate of 115,000 tpd. The option for an increased production rate of 160,000 tpd was also considered in the planning and design. The GRS BC will be a high-tonnage mine utilizing electric ore handling equipment on both the Extraction and Rail Haulage Levels. Caving, development, pre-production and
2.1 GRS BC Design Overview The mine design for the GRS BC utilized the VulcanTM mine planning package. Output from VulcanTM was imported to the VnetPCTM 2000 ventilation simulation program to develop a three dimensional network representation. Figure 1 shows a general isometric view of the proposed GRS BC. This network includes all components such as the tunnel access, ramp, Service Winze and conveyors. Figure 2 shows a plan view of the Extraction Level for the GRS BC. For 115,000 tpd there will be approximately 870 active drawpoints that will be brought online as the cave moves from the Northeast towards the Southwest. The ventilation and track drifts comprise the underlying Service Level. The Undercut Level overlies the
P.T. Freeport Indonesia’s (PTFI) project site is located in the Sudirman Mountain range of Papua (formerly Irian Jaya). West Papua is the Eastern-most province of Indonesia, which occupies the Western half of the island of New Guinea. PTFI acquired and have been developing the Ertsberg district since 1967. The presently known ore reserves are located approximately 100 km North of the Southwest coast of Papua, between elevations of 2,800 m and 4,000 m above sea level. Total ore tonnage has increased from 15,000 tonnes/day (tpd) in 1978, to 60,000 tpd in 1992, to the present tonnage of about 230,000 tpd. At the time of writing this paper approximately 45,000 tpd is being produced from the DOZ Block Cave Mine. Present reserves consist of six coppergold-silver orebodies and several other resources located within a 15-kilometer square area. The Grasberg porphyry copper-gold deposit possesses the world’s largest gold deposit and is among the five largest copper deposits in an operating mine. The present reserves for the Grasberg ore body are about 1.7 billion tons.
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Figure 1. Schematic showing the layout of the GRS BC. Extraction with the same general footprint. A schematic layout of the proposed ventilation system for the Extraction, Undercut and Service Levels is shown in Figure 3. The airflow is delivered and exhausted via the ventilation service drifts on the 2790 m Service Level. There will be four groups of three service drifts, with each group comprising two parallel ventilation drifts (intake or exhaust) and one-track drift. Each group of service drifts will be connected to every Extraction panel by two 4 m diameter smooth raises. One raise will serve as a chute, and the other for ventilation. The ventilation service drifts and connection raises allow the full width of the Extraction Level to be segmented into five separately ventilated zones. The Undercut Level will be discretely ventilated. To avoid the requirement for many ventilation raises interconnecting the Undercut and Extraction Levels, two ventilation drifts will be driven across the Undercut Level ahead of caving. These drifts will be connected down to the Service Level exhaust at the West end. As caving progresses, intake air will be brought around the cave using temporary fringe drifts and
advance Drill Drifts (DDs). Auxiliary ventilation will be used to supply air to the active DD headings from the perimeter. The air will be exhausted via the advance Undercut ventilation drifts as shown in the sketch. The Service Level track drifts will be ventilated in the opposite direction to the Undercut, which will help balance the air velocity through the intake and exhaust service drifts. The GRS BC area will be supplied and exhausted with air via primary intake and exhaust drifts. To facilitate the large quantity of air required, there will be four main intake drifts and four exhaust drifts as shown in Figure 1. 2.2 Ventilation Design Criteria PTFI adopt certain ventilation criteria to assist with underground mine planning. Tables 1 and 3 present some of the criteria that have been established based on Indonesian Mining Regulations (Decree of the Minister of Mines and Energy 1995), US Mine Safety and Health Administration
Figure 2. Plan view of 115,000 tpd Extraction Level with underlying rail loops.
Figure 3. Sketch showing the proposed GRS BC ventilation system.
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(MSHA), and good engineering practice. Although the design criteria do not specifically address diesel particulates, PTFI has a comprehensive program in place to monitor and minimize diesel particulates in all underground areas. This program integrates baseline monitoring of worker exposures with reduction initiatives, including fuel changes, upgrade of equipment to tier-rated engines, and increased ventilation. Table 2 provides a list of the Atkinson friction factors used for the ventilation design. These values have been obtained from detailed friction factor tests conducted in PTFI’s existing underground operations, and hence there is a good level of confidence in the data. 2.3 Airflow Requirements Table 4 provides an estimate of the airflow requirements for the GRS BC operating at 115,000 tpd. This includes a comprehensive list of all mobile equipment, with actual engine power and estimated operating factors. To provide for an economic ventilation design, every attempt has been made to reduce the amount of diesel equipment operating in the mine, and a significant component of the production operation will be run remotely using automation techniques. Major contaminant producing areas of the mine such as the crusher stations and shops will be established on separate air splits direct to exhaust. The rail haulage and chute loading systems will be fully automated from a remote control station. The nature of the Extraction Level lends itself to automation of the loader fleet although at this level of the study it has been assumed that all mobile equipment on the level will have an operator. The total airflow estimate for the mobile equipment is about 627 m3/s. The requirement for the active Extraction and Service Levels is 736 m3/s. Other fixed allowances include four shops, with a total airflow of 161 m3/s exhausting directly to return ventilation drifts. There is an additional 189 m3/s intaking the AB Tunnels which is exhausted via the GRS BC conveyors. This ventilation split requires an underground booster fan. Based on this list, the total airflow provision for the mine, excluding leakage and balancing, is about 1,755 m3/s. 2.4 Ventilation Modeling and Infrastructure Ventilation analyses were conducted which included modeling the GRS BC during two different phases in the proposed mine life. The analyses took into account all anticipated air leakage paths (doors, bulkheads, old panels, etc.) to provide a realistic representation of the mine. The modeling incorporated the AB Tunnel system and made allowance for interconnection with the proposed Big Gossan and Kucing Liar Mines. The analyses assume primary underground exhaust fan installations close to the main portals. Table 1. Air velocity design criteria. Airway
Air Velocity (m/s) Min Opt Max
Conveyor Drifts - Homotropal - Antitropal
0.8 0.8
2.0 1.0
4.0 2.0
Truck Haulage Drifts
0.8
4.1
6.1
Primary Ventilation Drifts
0.8
8.1
10.2
Rough Large Raises (+4 m)
-
14.2
19.8
Typical ALIMAK Raise
-
12.7
19.8
Drop Ventilation Raise
-
6.6
19.8
- Optimum values are used for design purposes. - Velocity criteria are based on an economic assessment factoring in power and development costs. 726
Table 2. Atkinson friction factors. Description
Friction Factor (kg/m3) Actual* Stnd
9 m2 to 15 m2 XSA - Drifting 15 m2 to 20 m2 - Drifting 20 m2 to 30 m2 - Drifting 30 m2 and Up - Drifting
0.0102 0.0093 0.0083 0.0074
0.0132 0.0120 0.0107 0.0095
Conveyor Drift (with 2.1 m Belt)
0.0093
0.0120
Primary Raise – 6.0 m f+ Typical ALIMAK Rse – 3.0 m f+ Drop Ventilation Rse+
0.0074 0.0111 0.0148
0.0095 0.0143 0.0191
* At average mine air density = 0.93 kg/m3 + Does not include shaft/raise entry and exit losses
Table 3. Other ventilation design criteria. Criterion
Value*
Min. Airflow per 100 Workers
7.1 m3/s
Design Airflow per 100 kW Diesel
7.9 m3/s
Shops and Facilities – Exhaust Diesel Equipment Shop Lube/Fueling Shop "Hot" Shop or Rock Breaker Stn Explosives Magazine
40.1 m3/s 28.3 m3/s 23.6 m3/s 9.4 m3/s
Common Gas TLVs Oxygen Carbon Dioxide Carbon Monoxide Hydrogen Sulfide Nitrogen Dioxide Sulfur Dioxide
> 19.5% TWA=5,000 ppm TWA=50 ppm STEL=400 ppm TWA=10 ppm STEL=5 ppm TWA = 2 ppm
*TWA = Time Weighted Average for 8-hr working shift STEL = Short Term Exposure Limit for 15 minutes The basic ventilation model was established as shown in Figure 1. The model was adapted to simulate two different scenarios: • Production from the first 12 panels, with a further 4 panels under development or construction. • Final (most Westerly) 4 panels are under development or construction, with the 12 adjacent panels in production. The center of mining has shifted to the West, and leakage in the system has increased (termed Mature Mine.) The predicted fan requirements for these two scenarios are given in Table 5. For the early mine the total airflow is predicted to be 1,813 m3/s. Based on a nominal daily tonnage of 115,000 tpd this equates to 0.0158 m3/s/tpd. The total installed fan motor power requirement is about 7,200 kW with a predicted annual power cost of US$2,836,000. These values assume 75% efficiency for the main fan and motor installation. For the mature mine the total airflow quantity increases to 1,888 m3/s, or about 0.0164 m3/s/tpd. This is higher than the value shown in Table 4 due to the system leakage. As a result of the efficient layout for the ventilation system, with few connections between intake and exhaust, the predicted leakage is less than 10%. The installed fan motor requirement for the mature mine increases to 8,660 kW, with a projected annual operating cost of about US$3,414,000. Both of these operating characteristics (Early and Mature cases) will be achievable with the same fan installations.
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Table 4. Airflow Required - 115,000 tpd. Nº
Mobile Equipment LHD 6.0 m3 (Dev.) - Dsl Truck 40 t (Dev.) - Dsl Drill Jumbo (Dev.) Production Drills Rock Breaker LHD 6.0 m3 - Electric Truck 30 t (Dev.) - Dsl Drill Jumbo (Sec. Breaking) Commando Sec. Breaking) Drill Jumbo (Rockboting) LHD 2.7 m3 (Clean-up) - Dsl Scissor Lift Explosives Truck U/G Road Grader Shotcrete Jumbo Shotcrete Truck U/G Lube Trucks U/G Boom Truck U/G Flatbed U/G Service Truck U/G Electrical Truck U/G Fork Lift u/G Personmel Carrier U/G Tractor (General) U6G Tractor (Engineering) Rail Haulage Locomotives Subtotal Mobil Equipment
Table 5. Predicted fan requirements for 115,000 tpd operation.
Op. Unit airflow Total Factor Pow. per Unit Airflow (%) (kW) (m’/s) (m’/s)
8 3 5 6 13 45 8 16 26 2 2 12 4 3 5 10 6 4 7 4 6 4 10 10 2 8
70% 80% 30% 30% 30% 80% 70% 60% 60% 30% 50% 30% 70% 30% 30% 30% 30% 75% 75% 30% 30% 15% 30% 30% 30% 80%
186 354 58 113 123 224 43 36 42 138 61 123 112 58 57 115 115 115 115 115 32 115 24 24 -
14.7 28.0 4.6 9.0 9.7 17.7 3.4 2.8 3.3 10.9 4.8 9.7 8.8 4.6 4.5 9.1 9.1 9.1 9.1 9.1 2.5 9.1 1.9 1.9 -
82.6 67.3 6.9 16.1 38.0 99.1 32.3 44.2 2.0 10.9 17.4 27.3 8.0 6.9 13.6 16.4 27.3 47.7 10.9 16.4 1.5 27.3 5.7 1.1 626.7
Fixed Allowances Development Panel Segment 45 Rail Haulage Drives 3 Personnel 600 Maintenance Shop 4 Crusher Bins and Belt Drifts 2
100% 100% 100% 100% 100%
-
14.2 33.0 0.1 40.1 94.4
637.2 99.1 42.5 160.5 188.8
Subtotal Total Airflow Required
1.128.1 1.755
The following drift sizes have been determined for the 115,000 tpd design (all ventilation drifts will be fully shotcreted): • Primary Ventilation Drifts from Portals to Service Level = 6.8 ¥ 6.8 m • Ventilation Service Drifts below Extraction = 5.5 ¥ 5.5 m • Ramp and Access Drifts = 4.5 ¥ 5.0 m • Conveyor Drifts = 4.5 ¥ 5.0 m • Extraction panels = 4.5 ¥ 4.5 m • Undercut DDs = 3.6 ¥ 4.2 m • Panel Raises = Raise-bored at 4 m diameter • Service Winze = Smooth 8.5 m diameter Figure 4 shows a graph of the relative pressure profile for the mature mine case. The relative pressure for this case is the difference between ambient (surface) and locations underground. For the profile a typical circuit has been chosen through the mine, and the trend line plotted against linear distance covered, including ventilation raises. The circuit chosen was: intake portals Æ primary intake drifts Æ service intake drifts Æ far Western panel region Æ service exhaust drifts Æ primary exhaust drifts Æ underground primary exhaust fans Æ exhaust portal. The trend is considered excellent, remaining generally linear and averaging about 333 Pa/1,000 m. The one short steep section represents the "open-split" in the ventilation circuit where the air exhausts Massmin 2004
Scenario/ Location
Total Pres (Pa)*
Quantity Input Pow (kW)* (m3/s)
Op. Cost. (US$/yr)+
Early Mine Main Exhaust Booster Installation Totals For Mine
2,990 2,860 -
1.624 189 1,813
6.474 721 7,195
$ 2,552,000 $ 284,000 $ 2,836,000
Mature Mine Main Exhaust Booster Installation Totals For Mine
3,490 2,990 -
1,699 189 1,888
7,906 753 8,659
$ 3,117,000 $ 297,000 $ 3,414,000
* Systems pressure to be met by fan. Does not incluide fan llosses. * Assumes 75% efficiency for fans and motors + Assumes 0.045 US$/kWh
Figure 4. Relative pressure profile for the 115,000 tpd mature mine case. from the development panel region to the service exhaust drifts. A linear profile is indicative of balanced air velocities and frictional pressure losses, suggesting that drift and raise sizes and numbers are correct for the system. 2.5 Adits Versus Shafts Design Comparison A ventilation study was conducted comparing adits versus shafts for the primary mine infrastructure. An initial shaft layout was developed in which the intake air is brought down two concrete-lined 7.5 m diameter shafts into the orebody area and back out two similar 7.5 m diameter exhaust shafts to surface. Due to the mountainous topography, the shafts would be approximately 1.0 km long. In order to place the shafts far enough away from the cave influence, primary intake and exhaust drifts would still be required, with each drift about 1,500 m long. It was assumed that two sets of four drifts will be developed at 6.8 ¥ 6.8 m connecting the shafts to the production area. The maximum velocities for the 7.5 m diameter shafts would be 18 m/s, which is 25% higher than the economic optimum velocities used as the current basis of mine design (see Table 1). However, it is noted that the economic velocity given in the table is for a rough shaft (2.1 m bored then slashed to size), not a fully concrete lined shaft, which will be considerably more costly to construct per meter, and have a higher optimum velocity.
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The shaft study showed approximately US$69 million in capital allocated to sink the intake and exhaust shafts and develop the shorter drifts. The adit option required approximately US$46 million in capital. The mature mine 115,000 tpd scenario was used for a fan power cost comparison. The results are provided in Table 6. These results indicate that there is a significant power savings (approximately US$580,000/yr) associated with drifting compared to shafts.
Table 6. Power cost comparison – drifting versus shafts. Scenario/ Location
Total Pres (Pa)*
Quantity Input Pow (kW)* (m3/s)
Op. Cost. (US$/yr)+
Shafts Main Exhaust* Booster Installation Totals For Mine
4,110 3,240 -
1,699 189 1,888
9,311 816 10,127
$ 3,670,000 $ 322,000 $ 3,992,000
Drifts and Portals Main Exhaust Booster Installation Totals For Mine
3,490 2,990 -
1,699 189 1,888
7,906 753 8,659
$ 3,117,000 $ 297,000 $ 3,414,000
* Atk. friction factor = 0.0037
kg/m3
(o.0048
kg/m3
stnd) for shafts
In addition to the economic benefits, the drifting option was considered to have certain other advantages, including: • Shafts require specialized equipment and crews to sink. Drift development will be spread over the mine ramp-up years with multiple headings. • Shafts require specialized maintenance equipment and trained crews. Repairs typically require a long time frame. • Historically PTFI has encountered water in long vertical shafts. In an exhaust shaft this counter-current water flow acts as a permanent drag and increases resistance. In such a long shaft it may also result in "blanketing" and cyclic loading (even potential fatigue) of the fans. • The fan installations will be more accessible for construction, maintenance, inspections and power delivery if placed near the proposed portals. Longer-term, access to the shaft collars would be difficult being close to the final cave line and remote from other mine facilities. • In the case of an emergency, personnel can quickly egress to a fresh air intake and if required walk completely out of the mine using that primary escapeway. • The drifts provide large tunnel diameter access to the mine for occasional oversize pieces of equipment that could potentially disrupt the tunnel rail system. 2.6 160,000 tpd Comparison Ventilation analyses were also conducted assuming an expanded production rate of 160,000 tpd. The simulation incorporated an updated equipment list, and the number of panel segments in production increased from 45 to 62. This change resulted in an increase in the main fan airflow to 2,124 m3/s, with a total mine airflow of 2,313 m3/s (when including the GRS BC oreflow ventilation system). Certain infrastructure sizes were increased as follows: • Primary Vent. Drifts = 6.5 ¥ 9.5 m • 2790 m Vent, Service Drifts = 5.5 ¥ 6.5 m Table 7 provides a comparison of the main fan operating requirements for both the 115,000 tpd and 160,000 tpd cases. Although the total airflow increases, there is an actual reduction of about 0.002 m3/s/tpd. The fan pressures are reduced slightly due to the increased drift sizes. 728
Table 7. Fan comparison – 115,000 tpd versus 160,000 tpd (mature mine). Scenario/ Location
Total Pres (Pa)*
Quantity Input Pow (kW)* (m3/s)
Op. Cost. (US$/yr)+
115 ktpd Case Main Exhaust* Booster Installation Totals For Mine
3,490 2,990 -
1,699 189 1,888
7,906 753 8,659
$ 3,117,000 $ 297,000 $ 3,414,000
160 ktpd Case Main Exhaust Booster Installation Totals For Mine
3,240 2,620 -
2,124 189 2,313
9,176 660 9,836
$ 3,617,000 $ 260,000 $ 3,877,000
3 RECOMMENDATIONS FOR FURTHER STUDY The work contained in this publication supports a prefeasibility design of the GRS BC operation. As the study progresses to feasibility level a more detailed analysis will be conducted. Certain ventilation issues have been identified for special consideration. 3.1 Mine Total Airflow The total predicted airflow in the mine is 1,888 m3/s, which corresponds to approximately 0.016 m3/s/tpd. This compares to about 0.028 m3/s/tpd for the existing DOZ Mine ventilation system. The value for the GRS BC Mine is significantly lower than this, however the contaminants in the mine are expected to be less than those presently encountered. The GRS BC will incorporate electric panel loaders, electric oreflow handling, low emission diesel development equipment operating on underground mine fuel, and significantly reduced Undercut drilling (room-and-pillar undercutting methods). When combined it is anticipated that these changes will allow compliance with gas and diesel particulate criteria, however it will be important to continue to revisit and revise airflow requirements as the design progresses. 3.2 Airflow Control One of the main ventilation challenges faced by operators of large mines is control of airflow. This can be particularly difficult in caving operations where there are multiple open parallel splits. On Extraction Levels the problem is compounded in longer panels, which often have more than one ventilation zone. Experience at PTFI suggests that automated regulation of the panel raises on the Extraction Level will be difficult due to damage from equipment and drawbell blasting activities. Hence it is likely that remotely controlled heavy-duty regulators will be designed and installed on the Service Level. 3.3 Dust and Diesel Particulates Control of dust in block cave mines can be extremely challenging. Approximately 230,000 tpd of ore are presently moved through PTFI’s site operations via loaders, trucks, crushers, belts and orepasses. In certain areas, such as panel drifts, belt transfers, orepass head chutes and stockpiles particular attention must be given at the design stage to ensure compliance with dust Threshold Limit Values (TLVs). This may consist of a comprehensive plan incorporating weather station data, belt moisture sensors, foggers and sprays tied to the mine PLC network. The mine design will also have to ensure compliance with MSHA’s diesel particulate TLV, and as such the fuels, equipment, and ventilation will factor this in.
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3.4 Fans The predicted main fan operating requirement is 1,699 m3/s at 3.5 kPa fan total pressure, which corresponds to 425 m3/s per exhaust drift. This assumes a single primary exhaust fan installation. Further work will evaluate the potential advantages of primary forcing fans and a push-pull system with both primary forcing and exhausting fans. The characteristic of 425 m3/s per drift can be achieved with either centrifugal fans (portal structure) or underground axial fans (vane or mixed-flow). The power demand will be high, but it is anticipated that losses can be kept to a minimum through careful design of the fans, layout, and electrical supply. The underground booster fan for the conveyor drifts is sized at 189 m3/s at 3.0 kPa fan total pressure. A similar sized fan would also be required to handle the airflow exhausting from the proposed Kucing Liar Mine oreflow area. 3.5 Air Velocity The airflow quantity in the twin exhaust conveyor drifts is about 378 m3/s, corresponding to an air velocity of approximately 6 m/s. This exceeds the maximum design velocity of 4 m/s for a homotropal conveyor system. During further study the GRS BC oreflow ventilation system will be examined. It may be possible to significantly reduce the airflow exhausting the belts, or even intake the belts with a dedicated exhaust to the primary mine fans. Depending on the active segment layout, the air velocity at the East end of the service ventilation drifts may become high (exceed 10 m/s). However, when considering the entire length of these collection drifts the pressure losses are not excessive. It is planned that the two ventilation drifts in each set of three service drifts will be aerodynamically connected with equalizing cross-cuts. 3.6 Upcast Air In the ventilation design airflow upcasts both the lower ramp and Service Winze. Based on experience at PTFI it is likely that condensation will occur with the upcast air (Calizaya et al. 2002). In the ramp this can result in fogging and poor visibility. In the shaft this can cause fogging, water collection, and in some cases more rapid deterioration of the fixed facilities. For the shaft it would be possible to downcast intake air. Further study will be devoted to the determination of shaft resistances and the dynamic effects of cage movement, and the optimization of airflow in the shaft. For the lower ramp fixed fan-heaters may be required to dry the air and maintain good visibility. Heaters have been used to good effect in existing PTFI underground ramp systems. 3.7 AB Tunnels and Mine Fire Approximately 189 m3/s is predicted to intake the lower mine section from the AB Tunnels. This air then either enters the ramp, Service Winze, and shop facilities (dedicated exhaust). This represents a significant volume of air, with a tunnel air velocity of about 5.7 m/s. During future study the airflow requirements for the tunnels will be more accurately determined, incorporating fire simulation to determine contaminant spread in the event of a tunnel vehicle fire. Fire
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analysis is an essential component of an overall mine design. For the pre-feasibility design fire mitigation and control philosophies were incorporated into the mine plan. More detailed "hot-flow" fire analyses will be conducted as the design progresses. For the preliminary design PTFI has: • Made provision for all shops and hot-work areas to be directly ventilated to an exhaust airway and to be equipped with isolation fire doors. • All conveyor belts will have independent airflow splits directed to the belt portals, and will be equipped with smoke sensors and fire-suppression deluge systems. • The air from the AB Tunnels will be used for the lower ramps, and will exhaust via the conveyor drifts. This air will not enter any of the production areas. • The air from the upper ramps will be used for ventilating the conveyor drifts, and will not enter the block cave working areas. • Air critical velocity values (per Froude number analysis) will be sufficiently high throughout the ramps as to prevent backlayering of smoke against the direction of airflow in the event of equipment fire. This is necessary to ensure predictable smoke spread and provide a noncompromised fresh air base for fire fighting. 4 CONCLUSIONS The GRS BC will be a very large mine that will provide many unique engineering challenges. It is anticipated that like most block cave operations, there will be issues associated with the balancing of the air, especially through the production panels. Particular attention will also need to be given to the integration of dust and diesel particulate control strategies and systems within the ventilation design. However, it can be stated that for a pre-feasibility study, the proposed GRS BC Mine layout integrates good ventilation design. The predicted pressure losses and air velocities generally agree with PTFI optimum design criteria, and peak values are within acceptable economic and operational levels. When considering operating cost, there is a definite benefit in developing the mine with portals, rather than intake and exhaust shafts, and the drifting option has been selected as the base case. 5 ACKNOWLEDGEMENTS The authors would like to thank PTFI Management for allowing publication of this paper. REFERENCES • Calizaya, F., Karmawan, K. & Wallace, K.G. 2002. Utilization of Heater Fans to Control Mine Atmospheric Fogging. In Euler De Souza (ed), Proceedings of the 9th NA/US Mine Ventilation Symposium. The Netherlands: A.A. Balkema Publishers. • Decree of the Minister of Mines and Energy Indonesia, 1995. Number 555.K/26/M.PE/1995, General Mining Occupational Safety and Health. • MSHA, Code of Federal Regulations 30, Part 75.
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Fluidized-bed biofilms for process and environmental applications in mining and metallurgy Jaakko A. Puhakka, Anna H. Kaksonen and Päivi H.-M. Kinnunen, Institute of Environmental Engineereing and Biotechnology Tampere University of Technology, Finland
Abstract Biofilm based fluidized-bed processes were developed for ferric sulphate production and regeneration in biohydrometallurgy and for treatment of acidic metal-containing wastewaters. High-rate ferric sulphate generation with iron oxidation rate of up to 26.4 g Fe2+ L-1 h-1 and a hydraulic retention time of 0.2 h were obtained and long-term maintained in Leptospirillum ferriphilum-dominated biofilm systems. The treatment of acidic wastewater (pH 2.5 to 5) containing sulphate, and high-concentrations of zinc and iron was studied in a sulphate-reducing fluidized-bed reactor. The FBR feed was supplemented with lactate or ethanol to support biological sulphate reduction. During long-term operation, this system achieved the following metal precipitation rates: 360 mgL-1d-1 for Zn and 86 mgL-1d-1 for Fe with over 99.8% Zn and Fe removal. The alkalinity produced from ethanol and lactate oxidation increased the wastewater pH from 2.5 to 7.5-8.5. In summary, the fluidized-bed biofilm process can be used for high-rate ferric solvent production in biohydrometallurgical applications and the sulphate-reducing and metal precipitating biofilm process has potential for wastewater treatment applications producing a good quality effluent with metal concentrations below 0.1 mgL-1.
1 INTRODUCTION The use of biotechnology in mining and metallurgy is continuously increasing and includes both production process and environmental engineering applications. New biohydrometallurgical methods as compared to traditional pyrometallurgical techniques have many advantages, including no emission to air, simplicity, low cost, and applicability to low-value ores and mineral resources that have not been feasible by conventional mining (e.g., Bosecker 1997; Hsu and Harrison 1995). The economical value of the global bioleaching processes in recent years has been over 10x109 US dollars (for reviews, see Brandl 2001; Johnson 2001). The leaching of sulphidic ore is performed by noncontact, contact, or cooperative leaching mechanisms, in which iron and sulphur compounds are oxidized to ferric iron and sulphuric acid, respectively (for a review, see Rawlings 2002). The exploitation of sulphide minerals results in the oxidation of iron and sulphur, and thus in the production of acidic metal-containing wastewaters (Foucher et al. 2001). Recently, the interest in the application of sulphate reduction as the dominant step of wastewater treatment has been growing (for a review, see Hulshoff Pol et al. 2001). The process is based on biological hydrogen sulphide production (equation 1) by sulphate-reducing bacteria (SRB), metal sulphide precipitation (equation 2) and neutralization of the water with the alkalinity produced by the microbial metabolism (equation 3) (Dvorak et al. 1992; Christensen et al. 1996): SO42- + 2CH2O → H2S + 2HCO3where CH2O = an electron donor
(1)
HCO3- + H+ → CO2 (g) + H2O
(2)
H2S + M2+ → MS(s) + H+ where M2+ = metal, such as Zn2+
(3)
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In this work, fluidized-bed biofilm processes were developed for the following: 1) ferric sulphate generation, and 2) for treating acidic metal- and sulphate-containing wastewater. The process performance of these FBRs was evaluated under various operational conditions with continuous flow experiments. 2 MATERIALS AND METHODS The ferrous sulphate oxidation experiments were performed in laboratory-scale FBRs (Kinnunen et al. 2003,
Figure 1. Schematic diagram of a fluidized-bed reactor used for iron oxidation
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The FBR systems used for sulphate reduction and metal precipitation were as shown in Figure 2 and the experimental designs and feed solutions as earlier described (Kaksonen et al. 2003a; 2003b; 2004). The analysis methods used were as summarized in Table 1. 3 FERRIC SULPHATE GENERATION Iron oxidation was studied in FBRs at different temperatures and with microorganisms enriched from various mine environments. At 37ºC with a Leptospirillum ferriphilum dominated biofilm (Figure 3) and with pure oxygen aeration and HRT of 0.2 h the iron oxidation rate was 26.4 Fe2+ L-1 h-1 (Table 2). With moderately thermophilic acidophilic enrichment cultures and dominated by Sulphobacillus spp. enriched from an open-pit mine in Collie, Western Australia, the iron oxidation rates always remain significantly lower than at 37oC.
Figure 2. Schematic diagram of the configuration of the sulphidogenic fluidized-bed reactor Kinnunen and Puhakka 2004) (Figure 1). The aeration system was connected to the recycle flow line, and the aerated liquor was pumped to the bottom of the FBRs. Activated carbon was used as the biomass carrier. The FBRs were operated at 37, 50, 60 and 70oC. The composition of the feed and sources of microorganisms were as previously described (Kinnunen et al. 2003, Kinnunen and Puhakka 2004). Figure 3. Scanning electron micrograph of the iron-oxidizing culture (Kinnunen and Puhakka 2004).
Table 1. Summary of the analyses employed in this study. Analysis
Instrument(s)
Ref.
Sulphate
HPLC, IC
a,b,c
Dissolved organic carbon (DOC)
TOC analyzer
a,b
Acetate, Ethanol
GC-FID
a,c
Soluble metals
AAS
d,e
Dissolved sulphide
Spectrophotometer
f,g
T (ºC) Aeration
Ferrous iron
Spectrophotometer
h
pH
pH electrode
a,b,c
Oxygen
Oxygen meter
Total alkalinity
Potentiometric titration
h
Total suspended solids (TSS)
Oven, balance
b
Volatile suspended solids (VSS)
Furnace, balance
a,b,c
Total solids (TS)
Oven, balance
a,b,c
Volatile solids (VS)
Furnace, balance
i
Minerology of the metal precipitates
X-ray diffractometer
b
Elemental composition of the metal precipitates
SEM with energy dispersive x-ray spectrometer
b
37
Air
HRT (h) Aeration-% 0.6
Iron oxid. (g Fe2+/l·h)
91
8.1
37
95.5% O2
0.2
82
26.3
50
Air
9.8
71
0.7
60
Air
4.0
85
0.5
70
air
3.1
11
0.1
The iron oxidation rates obtained in the 37oC FBRs compared favourably with iron oxidation rates obtained in other studies published earlier.
4 WASTEWATER TREATMENT
References: aKaksonen et al 2003a, bKaksonen et al. 2003b, cKaksonen et al. 2004, dSFS 3044, eSFS 3047, fTrüper and Schlegel 1964, gCord-Ruwish 1985, hSFS-EN ISO 9963-1, iSFS 3008 hAnon. 1992.
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Table 2. Biological iron oxidation rates in FBRs operated at different temperatures and with different enrichment cultures (Kinnunen et al. 2003; Kinnunen and Puhakka 2004).
The FBR process performance was evaluated under various operational conditions with continuous flow experiments. In terms of sulphate reduction, sulphide production and effluent alkalinity, the start-up of the FBR with 10 % fluidization rate was superior to the FBRs with 20-30 % fluidization rates. However, the FBRs with 20-30 % recycling rates performed better at the highest loading rates. The substrate feed of the FBR could be changed from lactate to ethanol without significant change in the FBR performance. The robustness of the sulphidogenic FBR was studied by increasing stepwise the zinc, sulphate and substrate feed
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concentrations and decreasing the feed pH, as well as decreasing hydraulic retention time (HRT) until process failures occurred. Lactate- and ethanol-utilizing FBRs precipitated 350-360 mg Zn l-1 d-1 and 86 mg Fe l-1 d-1 from acidic wastewater containing 230 mg Zn l-1 and 57 mg Fe l-1 at a HRT of 16 h (Table 3). Percentual metal precipitation was over 99.8 % resulting in effluent soluble Zn and Fe concentrations below 0.1 mg l-1. The alkalinity produced in the sulphidogenic lactate and ethanol oxidation increased the wastewater pH from 2.5 to 7.5-8.5. Maximum metal precipitation rates were obtained with ethanol-fed FBR at a HRT of 6.5 h from acidic wastewater containing 176 mg Zn l-1 and 87 mg Fe l-1(Table 3). The rates were over 600 mg Zn l-1 d-1 and 300 mg Fe l-1 d-1, percentual metal precipitation over 99.9% and effluent metal concentrations below 0.08 mg l-1. The alkalinity produced in ethanol oxidation increased the initial pH of 3, resulting in effluent pH of 7.5-8.5. At the HRT of 6.5 h, ethanol oxidation was incomplete and acetate accumulated in the FBR. The operation of the FBR at HRTs below 6.5 h was limited by partial acetate oxidation, since alkalinity production by acetate oxidation is necessary for wastewater neutralization. Zinc and iron precipitated predominantly as ZnS, FeS2 and FeS. The metal precipitates did not cause clogging of the reactors during the continuous operation of over 1200 days. Table 3. Summary of the process performance of the sulphidogenic FBR. Electron donor
Lactate
Ethanol
Ethanol
16
16
6.5
Eb
2.5±0.1 7.8±0.2
2.5±0.1 7.7±0.2
3.0±0.1 7.9±0.1
Feed (mg l-1)
Zn Fe
233±4 58±3
240±10 57±1
176±6 87±5
Effluent (mg l-1)
Zn Fe
0.07±004 0.05±0.04
0.09±0.06 0.04±0.04
0.02±0.02 0.04±0.03
Precipitation (%)
Zn Fe
99.97±0.02 99.92±0.07
99.96±0.02 99.93±0.07
99.99±0.01 99.96±0.04
Precipitation (g m-3d-1)
Zn Fe
350±6 86±5
360±15 86±1
635±50 315±22
Kaksonen et al. 2003b
Kaksonen et al. 2003a
Kaksonen et al. 2004
HRT (h) pH change
Fa
Reference
aF
= Feed, bE = Effluent
5 CONCLUSIONS This study demonstrates a high-rate fluidized-bed bioprocess for ferric sulphate production by an enrichment culture dominated by L. ferriphilum. Further, this study also demonstrated the feasibility of a sulphidogenic FBR for concomitant removal of acidity, metals and sulphate from wastewaters. Both FBR processes offer viable alternatives for existing processes. ACKNOWLEDGEMENTS National Technology Agency of Finland, Outokumpu Oyj, Finnish Graduate School of Environmental Science and Technology and the Academy of Finland are gratefully acknowledged for their financial support.
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REFERENCES • Anonymous. 1992. Standard methods for the examination of water and wastewater. 18th ed. In: Greenberg AE, Clesceri LS, Eaton AD, editors. American Public Health Association. • Bosecker K. 1997. Bioleaching: metal solubilization by microbes. FEMS Microbiol. Rev. 20: 591-604. • Brandl H. 2001. Microbial leaching of metals. In: Biotechnology, Vol. 10, Rehm H-J (ed.), Wiley-VCH, Weinheim, pp. 191-224. • Christensen B, Laake M and Lien T. 1996 Treatment of acid mine water by sulfate-reducing bacteria: Results from a bench scale experiment. Water Res 30: 1617-1624. • Dvorak DH, Hedin RS, Edenborn HM and McIntire PE. 1992. Treatment of metal-contaminated water using bacterial sulfate-reduction: Results from pilot-scale reactors. Biotech Bioeng 40: 609-616. • Foucher S, Battaglia-Brunet F, Ignatiadis I and Morin D. 2001. Treatment by sulfate-reducing bacteria of Chessy acid-mine drainage and metals recovery. Chem Eng Sci 56: 1639-1645. • Hsu C-H and Harrison RG. 1995. Bacterial leaching of zinc and copper from mining wastes. Hydrometallurgy 37: 169-179. • Hulshoff Pol LW, Lens PNL, Weijma J and Stams AJM. 2001. New developments in reactor and process technology for sulfate reduction. Water Sci Technol 44(8): 67-76. • Johnson DB. 2001. Importantce of microbial ecology in the development of new mineral technologies. Hydrometallurgy 59: 147-157. • Kaksonen, AH, Franzmann, PD and Puhakka, JA. 2004. Effects of hydraulic retention time and sulfide toxicity on ethanol and acetate utilization in sulfate-reducing fluidized-bed reactor. Biotechnol. Bioeng. 86: 332-343. • Kaksonen, AH, Franzmann, PD and Puhakka, JA. 2003a. Performance and ethanol utilization kinetics of a sulfatereducing fluidized-bed reactor treating acidic metalcontaining wastewater. Biodegradation 14: 207-217. • Kaksonen, AH, Riekkola-Vanhanen, M-L and Puhakka, JA. 2003b. Optimization of metal sulphide precipitation in fluidizedbed treatment of acidic wastewater. Water Res. 37: 255-266. • Kinnunen PH-M, Robertson WJ, Plumb JJ, Gibson JAE, Nichols PD, Franzmann PD and Puhakka JA. 2003. The isolation and use of iron oxidizing moderately thermophilic acidophiles from the Collie coal mine for the generation of ferric iron leaching solution. Appl. Microbiol. Biotechnol. 60: 748-753. • Kinnunen PH-M and Puhakka JA. 2004. High-rate ferric sulfate generation by a Leptospirillum ferriphilum-dominated biofilm and the role of jarosite in biomass retainment in a fluidized-bed reactor. Biotechnol. Bioeng. 85:697-705. • Rawlings DE 2002. Heavy metal mining using microbes. Annu. Rev. Microbiol. 56: 65-91. • SFS. 1980a. SFS 3044: Metal content of water, sludge and sediment determined by atomic adsorption spectroscopy, atomisation in flame. General principles and guidelines. Finnish Standards Association, SFS. 8 p. • SFS. 1980b. SFS 3047: Metal content of water, sludge and sediment determined by atomic adsorption spectroscopy, atomisation in flame. Special guidelines for lead, iron, cadmium, cobalt, copper, nickel and zinc. Finnish Standards Association, SFS. 6 p • SFS. 1990. SFS 3008: Determination of total residue and total fixed residue in water, sludge and sediment. Finnish Standards Association, SFS. 3 p. • SFS. 1996. SFS-EN ISO 9963-1: Water quality. Determination of alkalinity. Part 1: Determination of total and composite alkalinity. Finnish Standards Association, SFS. 16 p.
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An audit methodology used to assess the current and future capabilities of the backfill system at Konkola No 3# Nic James, Mine Manager, Konkola Division Alan Naismith, Group Geotechnical Engineer, KCM, Zambia
Abstract Mining with backfill was introduced to Konkola as part of the Konkola Deep Mine Project (KDMP). A temporary pilot plant was commissioned in November 2002 to produce cyclone classified tailings fill to be used with the Overcut and Bench Mining method (OCB) being practiced. At that time mining between 1850 level and 1660 level was suffering from deteriorating ground conditions and planned recoveries and production rates were not being achieved. In addition to providing fill for concurrent operations, the plant was required to produce fill for previously unfilled areas. Backfill has been incorporated as a concurrent part of the mining cycle in areas below 1850 level. An audit was carried implemented during July and August of 2003 to critically assess the current status of backfilling practice at Konkola and to determine the ability of the system to supply future operations. An unusual aspect of this audit was that it was mine staff, under the guidance of an external consultant, that carried it out, and not an independent external team. A valuable outcome of this approach was to internalize the understanding of the backfilling process and create a more team based approach to achieving filling objectives. Three essential areas were identified. These were; - System Objectives and Design; Fill Preparation, Transportation and Placement and System Management. Thirteen key elements within these areas were identified that defined the backfilling process. Within each element, the status quo was critically assessed and compared to a perceived "best practice" standard. Potential risks associated with failure of the system to meet the "best practice" guidelines were highlighted and action plans developed to address shortcomings. Completion of the audit necessitated collation of information from a number of sources together with additional information obtained from testing. The mine is now in a good position to finalise a Code of Practice for Backfilling. This paper describes the audit format and its application to the Konkola backfilling system. It then summarises the most important findings and describes action being taken to meet the "best practice" standard.
LOCATION OF KONKOLA COPPER MINES The operation at Konkola is one of three copper production operations owned by KCM and lies in the rich Copperbelt province located in the North Eastern portion of Zambia. The three divisions of KCM all lie between 12º and 13º South and 27º 30’ and 28º 30’ West. Two operating Konkola shafts are located next to the town of Chililabombwe, just 12 km from the DRC (Democratic Republic of the Congo) border, about 450 kilometers northwest of Lusaka. Konkola is the most northerly of the Zambian Copperbelt mines. There are three main ore-bodies at Konkola Division; these are the Kirila Bombwe South (No. 1 shaft), the Kirila Bombwe North (No. 3 shaft) and the Konkola (No. 2 shaft). These lie on a north – west, south – east trending Kirila Bombwe anticlinal structure that plunges towards north – east. The South ore-body forms the southwestern flank of the anticline, while the north ore-body lies in the nose of the fold. The ore-bodies range greatly in thickness from as little as 3m to as great as 15m. The dip also varies greatly from a low of 10º in the nose area to a high exceeding 75º on the fold limbs. In-situ copper grades of about 3 to 4% total copper are typical. There is also cobalt mineralisation averaging between 0.1 – 0.2 % total cobalt. No other significant mineralisation exists. Production is from Nos.1 and 3 shafts. No.2 shaft, accessing the Konkola ore-body, has been closed since 1958. Sub – level open stoping is practiced extensively at Massmin 2004
No 1 shaft with ore flow being assisted by scrapers in areas of decreasing dip. In - ore –body mining has been practiced in the wide, flat dipping areas of No 3 shaft since 2000 initially using post pillar cut and fill and, more recently, over – cut and benching methods The Konkola mines are reputed to be amongst the wettest in the world. Significant aquifers occur in both hanging-wall and foot-wall sequences and are amply recharged by an annual rainfall generally exceeding 1500mm and a network of surface streams and rivers. Dewatering is achieved through a comprehensive system of drainage holes, crosscuts and drives. The mine is currently discharging approximately 280 000m_ of water per day. THE OVER-CUT AND BENCH (OCB) MINING METHOD PRACTICED AT KONKOLA NO 3 SHAFT Access to the ore body is achieved from foot – wall infrastructure via twin inclines on an apparent dip of approximately 7º developed at 150m intervals along strike. Both inclines lie within the ore – body. One incline is developed against the assay hanging wall (AHW) to give access to the over-cut. The other lies on the assay footwall (AFW) and provides access for ore extraction during the benching phase. These inclines lie 10m apart and are protected by 10m wide pillars on either side. The initial OCB mining was confined to an area between the 1660 foot level and 1850 foot level. Over-cuts, 7m wide, are mined along strike against the AHW. Strike pillars, 9m wide, separate adjacent over-cuts. Cross cuts, 7m wide, are developed through the strike
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pillars. Once an over-cut is complete it is benched down to the footwall contact to leave 9m x 9 m square pillars over the full ore – body width. Hydraulically placed backfill was not incorporated into the initial OCB mining plan as it was thought that pillars would remain stable over the working life of the block. It was believed, however, that backfill would be required at deeper levels and a pilot plant was constructed. This was commissioned in November 2002 when deteriorating ground conditions in the upper block necessitated introduction of fill immediately the benching phase had been completed. Backfill was also introduced to benched areas in which deteriorating conditions posed the threat of pillar collapse. AUDIT OBJECTIVE During July 2003, KCM invited Steffen, Robertson and Kirsten Inc (SRK) to conduct a technical audit of the Konkola operations as part of their "Buyantanshi" (way forward") initiative. This initiative has the objective of increasing KCM production to 250 000 tonnes of finished copper per annum at a cost of $0.65/lb. Key elements of the mining value chain were to be considered. These were Development, Stoping, Logistics, Hydrology, Geotechnics, Dilution, Backfilling, Costing and Human Resources. The objective of the audit was to establish the status of each element as it is currently practiced at Konkola and to determine its ability to service current and future operations. This would have the joint advantages of providing management with confidence in the ability of Konkola to deliver the copper called and to identify possible threats to the achievement of targets. Mindset changes introduced during the audit and followed up with subsequent actions would also instill confidence in Konkola personnel that they were capable of meeting targets. AUDIT STRUCTURE AND MODUS OPERANDI The audit was structured in three main phases as illustrated in Figure 1. The first part involved defining all elements in the process to be audited and identifying the status quo, best practice and risk associated with each element of the process. Key performance indicators were also identified to indicate the level of performance required in each element. Status quo is defined as reflecting the current state of understanding or practice. Best practice reflects the requirements needed to meet Konkola objectives in the short term (12 to 18 months) and, where appropriate, requirements dictated by industry experience. Risk reflects an assessment of the nature of potential loss that could be faced by KCM should performance in any process element not meet the required standard. The findings from this phase are indicated for the Backfilling process in Table 1. The second phase involved creating and implementing action plans to address any shortcomings between perceived best, and current, practices. A template used to prioritise action is shown in Figure 2 and uses the dimensions of "Importance" and "Urgency" as a prioritization guideline. An example of this approach as applied to Element 9: Design and Construction of Bulkheads is shown in Table 2. Measurement of progress towards completing required actions is achieved using the Action Plan Monitor. Table 3 illustrates an application of this monitor to Element 4: Fill Volume Requirements. Key factors in ensuring that action is taken include clearly defining the expected output, nominating a responsible person who’s role it is to report back by a specified time and identifying resources required. Although convened and managed by an external consultant, the audit was implemented by KCM staff. A team 734
comprising metallurgical, engineering, planning, production and geotechnics personnel met on a regular basis to achieve consensus which was reported to the consultant and KCM Executive management. AUDIT FINDINGS AND BENEFITS The audit was able to settle a number of misperceptions held about backfill and the backfill system and was able to identify areas where additional focus was required. The more important findings included: 1. To create fill with a percolation rate of 100mm/hour, the preparation system is such that only 40% of tailings material is suitable. Despite this constraint, adequate hydraulic fill is available to meet current and future demands providing tailings capture remains above 33%. Utilisation of waste fill will reduce the hydraulic fill demand. 2. Even at a placed specific gravity of 1.45, the backfill provides sufficient lateral reinforcement to stabilize pillars and to enhance regional stability. 3. Although there is a risk associated with single stream systems incorporated into the pilot plant design, it is adequate to meet existing and future demands. Ongoing maintenance will be necessary to retain plant reliability. 4. Demands on the system should be simulated to assess the probability of its being unable to supply when required and to confirm (or otherwise) design parameters. 5. Underground reticulation should be located where practicable in secure and stable excavations. HDPE piping should be used in working areas to improve flexibility. 6. Bulkhead design and construction is critical to safety, meeting production schedules and minimizing fines contamination. Although no significant bulkhead failures have occurred to date, a more rigorous design and construction process will be needed in modified layouts where hydraulic heads in excess of 25m are possible. 7. The decentralized management system has proved adequate to date but may need to be reviewed once filling is fully incorporated into the mining cycle. 8. Participation in the audit resulted in a collation of, and sharing of, knowledge of system operation. It further engendered a belief in mine personnel that the system is capable of meeting both current and future requirements. This belief was reinforced when examination of January to June placement figures indicated a general improving trend (Table 4). POSTSCRIPT Due to an extreme deterioration in ground conditions towards the middle of 2003, mining of OCB stopes in the Northern section was suspended. Limited production came from OCB stoping in other areas while Sub-Level Open Stopes continued to produce. The need for concurrent backfilling reduced and the backfill system continued at a very low production rate. The phase of mining that is now taking place in the underlying block has incorporated several modifications but remains essentially Over Cut and Bench. Backfill has been included as an essential and concurrent part of the mining cycle. It is expected that the first fill will be placed in June 2004. ACKNOWLEDGEMENTS The authors acknowledge the assistance provided by all members of the Buyantanshi team in both participating in the audit and implementing its recommendations. Further,
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the authors acknowledge and thank KCM management for their assistance in facilitating audit recommendations and their permission to publish this paper. REFERENCES • Carvill P.G. "A review of the use of backfill in the OCB stoping method for the 1660 – 1850 mining block" Anglo American Report to Konkola Copper Mines, September 2000. • Covey S.R. "The Seven Habits of Highly Effective People" Simon & Schuster, 1992.
• Leach A.R. "Assessment of mining in the nose and west limb area at Konkola 3 shaft" ITASCA Report, November 2003. • Harr M.E. "Reliability Based Design in Civil Engineering" McGraw Hill, 1987. • Techpro M & M "Konkola Deep Mining Project" Report to ZCCM, November 1994. • Thomas E.G et al "Fill Technology in Underground Metalliferous Mines" International Academic Services Ltd, Kingston, Ontario, Canada, 1979.
Figure 1: Buyantanshi Audit: Implementation Process.
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BUYANTANSHI AUDIT: ACTION PLAN DEVELOPMENT TEMPLATE Audit Element: Date of this plan: -
Date of next update: Urgent
Important
Not Urgent
Item
Action
Item
NOW!
0 – 3 Months! Middle term Some planning
Short term and Reactive! Not Important
Item
Action
Action
Item
+ 3 Months!
Action DON’T WASTE VALUABLE RESOURCES!
Longer term Plan to utilize resources effectively!
Urgent Important Not Urgent Not Important
= requires immediate attention because if it is not done now there will be an impact on safety, cost or production in this measurement period (day, week or month) influences safety, cost or performance in the longer term (month, quarter) = the influence of this action will only be felt outside this measurement period (day, week or month) = the influence of this action will only be felt in the very long term (quarter or longer)
Figure 2: Buyantanshi Audit: Action Plan Development Template. Table 1 Buyantanshi Audit: Backfilling Elements and initial findings ELEMENT
STATUS QUO
BEST PRACTICE
RISK
ACTION
KEY PERFORMANCE INDICATOR
1. Fill Benefits (to meet mining objectives
Stated, but not quantified, that fill will improve recovery of ore resource
Economic analysis that compares best recoveries and profitability with and without fill.
1.Loss of reserves 2. Additional mining costs
Preparation of mining plans for best alternatives and economic analysis
NPV or IRR figures for alternative approaches.
2. Fill Performance requirements (to achieve benefits)
Must drain sufficiently to support the operation of trackless equipment within 72 hours
Support given to pillar sides is quantified
1.Unsafe working conditions 2. Loss of reserves 3. Low productivity rates
1. Revisit equipment requirements. 2. In-situ testing. 3. numerical modeling to quantify support performance
Performance specification for the fill.
3. Fill material properties (to meet performance requirements)
Percolation rate specified at 100mm/hr
Mechanical properties of fill are determined. Percolation rates optimised
1. Reduction in mass recovery. 2. Inefficient system design. 3. Pipeline wear
1. In-situ and laboratory testing of fills (including cemented fills)
Test reports on fill performance.
4. Fill Volume Requirements
Based on measured in-situ specific gravity = 1.45, 1.2m_ of wet fill is required to fill 1m_ in –situ .i.e. 1.1 tonnes of solids/m_
Fill demand-supply balance is known. Alternative sources of fill are identified and quantified
1. Ground deteriorates if stopes are not filled timeously. 2. Production is compromised
Short and long term fill demand and supply schedules created
Ongoing specific gravity checks. Up dated filling plans and schedules
5. Fill Preparation
Pilot plant rated at 2000 tonne /day of solids. This is equivalent to 90m_ of fill/hour
Plant design meets production requirements
1. Delays in feed from concentrator
Fill preparation in accordance with mining demand. Mechanical agitation to be introduced
Delivery of fill on demand
736
Santiago Chile, 22-25 August 2004
Massmin 2004
Table 1 Buyantanshi Audit: Backfilling Elements and initial findings ELEMENT
STATUS QUO
BEST PRACTICE
RISK
ACTION
KEY PERFORMANCE INDICATOR
6. Fill Transportation (surface)
Transfer from plant to shaft is 3.54km using single 125mm pipeline. Delays due to preand postflushing
Minimise transportation distance or provide back up system
1.Pipeline pressures estimated at 4,5MPa. 2. Single stream system
Quantify reliability of transportation system. Establish procedures to deal with surface spillage.
Delivery of fill on demand
7. Fill Storage (surface only)
720m_ live storage at plant + 360m3 at shaft.
System incorporates optimum storage volume
1. Agitation in storage tanks causes degredation. 2. Fluctuation of air pressure during agitation
Tanks rubber lined to reduce wear. Install mass flowmeter at shaft discharge
Delivery of fill on demand
8. Fill Reticulation (underground)
Full flow condition in 150mm vertical pipeline. Flow rate estimated at 110m3 to 116m3/hour. Steel pipelines to working faces
Place pipelines in secure areas. Replace steel with HDPE piping. Simulate the complete reticulation system to determine flow pressures and velocities.
Delivery of fill on demand
Reticulation system 1. Pipeline blockages. is appropriate for the 2. Damage due filling environment to falls of ground. 3. Lack of flexibility with steel pipe
9. Bulkhead Design and Construction
Initially waste Minimise excavation rock with Geofabric. size at bulkhead Revised to gumplanks locations. Designs with wiremesh. meet specific Cable based design engineering criteria. under trial Construction is approved before filling commences
1. Bulkhead failure and fill inrush in to working areas. 2. Leakage of solids into the drainage system. 3. Delays to production cycle
Suite of designs to be completed. Examination and approval procedures to be developed. Monitor hydrostatic pressure acting on bulkheads during and after filling.
Construction time per installation. Cost per installation
10. Fill Placement
Fill volumes estimated, System utilization is not measured. Fill such that required rise per pour volumes are placed restricted to 3m. to schedule. Contingency fill areas are available
1. Production delays. 2. Deterioration in ground conditions.
Develop short and medium term filling schedules to maximize system utilization. Estimate volumes to be filled using CMS technology.
Volume filled daily is according to schedule. Utilisation %
11. Drainage and Fines Loss
Decant and drainage Water and solids systems facilitate balance system in-stope dewatering employed. Drainage (0.45m3 of water controlled to liberated per 1m3 minimize impact of fill placed (excluding on other mining flushing water). No activities formal monitoring of water or fines loss. Reports that drains are becoming blocked with fines.
1. Solids block drainage systems. 2. Fines cause wear in pumps. 3. Fill does not drain and bulkhead stability decreases.
Establish drainage and fines loss profiles. Establish impact of additional water on other mining activities
Drainage rates during and after filling. Fines loss/m3 of fill placed
12. System Management
Responsibility is Written procedures shared between for all aspects of concentrator, system operation. plant and underground Personnel involved personnel. There in system operation is no single responsible are deemed person. Communication competent. occurs daily between plant and underground personnel. One communication meeting is held weekly during which short term demand and supply schedules are agreed.
Low system efficiency.
Establish radio links as back up to land lines. Review cost reporting procedures. Establish training and competency assessment programmes.
Sufficient competent personnel are available. Costs remain within budget.
Massmin 2004
Santiago Chile, 22-25 August 2004
737
Table 1 Buyantanshi Audit: Backfilling Elements and initial findings ELEMENT 13. Quality Assurance
STATUS QUO
BEST PRACTICE
Specific gravity and Control systems are percolation rate is in place to measure checked regularly at key parameters plant and shaft Mass such as Void ratio; flowmeter records Moisture content; receipts at shaft PSD; Specific gravity; (but not delivery to Pump output; underground). No Pipeline wear; survey reconciliation Volumes pumped; of demand and supply. Rate of rise; Few in-situ tests. Drainage rate; Fines loss.
RISK
ACTION
KEY PERFORMANCE INDICATOR
Sub standard system performance.
Establish quality assurance programme.
Parameters measured according to a schedule and comply with specifications.
Table 2: Buyantanshi Audit: Action Plan Development Example applied to Bulkhead design. Audit Element: Backfill: -
(9) Bulkhead Design and Construction
Date of this plan: -10-09-2003
Date of next update: Urgent
Important
Item
Not Urgent Action
Monitor pressure acting on selected (critical) bulkheads
Item
Action
Establish cost of bulkhead construction. Examination and sign off procedure to be developed for completed bulkheads Risk management plan developed to deal with bulkhead failure and fill inrush.
Not Important
Item
Action
Item
Action
Suite of designs to be completed and signed off by "specialist" Examination and sign off procedure to be developed for mined out stopes to pass them for filling.
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Table 3: BUYANTANSHI AUDIT: ACTION PLAN MONITOR Example applied to Fill Volume Requirement Estimation. (4) Fill volume requirements Action
Confirm volume requirements and schedule for future mining areas above 1850
Expected Output
Responsible Person
Milestones
Resources Required
Short and medium term backfill schedule with contingency plans
Senior Geotech Engineer reviewed by Group Geotech Engineer
Early October
Planning to provide proposed mining schedule
Accurate plans
Group Geotech Engineer to liaise with Senior Surveyor
End of October
Survey
Status (%complete)
Define alternative fill areas for contingency disposal (backlog areas) N.B. Add 5% to demand schedule to allow for fill loss into unplanned areas, through bulkheads and with flushing
Filled areas to be depicted on plans
Table 4: Fill volumes placed. January to June 2003
Tailings Fill (m3) Development Waste (m3) Total fill placed (m3)
Jan
Feb
Mar
Apr
May
June
Total (m3)
18 676
20 505
28 835
24 853
21 975
34 179
149 023
6 112
5 968
8 833
3 023
1 303
1 452
26 691
24 788
26 473
37 668
27 876
23 278
35 631
175 714
149 000m3 of tailings fill placed equates to 124 000m3 of void filled from which 333 000 tonnes was mined. 26 000m3 of development waste fill equates to at least 22 000m3 of void filled from which 60 000 tonnes of ore was mined
Massmin 2004
Santiago Chile, 22-25 August 2004
739
740
Santiago Chile, 22-25 August 2004
Massmin 2004