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By Klimpel. R. R.

Acknowledgments I am honored to be named as a recipient of the Antoine M. Gaudin Award, and I would like to briefly single out four groups who deserve a special thank-you relative to this award. The first group consists of Professor Kawatra, the many international support references that he gathered and the Gaudin Committee itself. I know such things don't come without considerable effort by many individuals, and I'm grateful for their support. The second group consists of the management of The Dow Chemical Company, the managements of various mining companies and, in recent years, the management of G. S. Technologies, all of whom supported me, financially and otherwise, in my grinding, flotation and other particle-processing work. Operations people are the most overworked and under recognized people in the world, whether they be in the mining or chemical industry. Obviously, without extensive industrial support, especially from operating personnel in the mining industry, almost nothing about the work contained in this lecture would have, or could have, happened. The third group consists of a number of coworkers who have played key roles in specifically implementing the new flotation-product portion of this lecture. Of these, Bob Hansen, Don Leonard, Basil Fee, Simon Isherwood, Fernando Cones and Marty Meulendyke deserve special recognition. They have not only been dedicated, but they have been great to work with. Also in this third group are many members of the international academic community, especially those at Penn State. Professor Leonard Austin must be singled out because he took me as a young mathematically oriented chemist and taught me to be an effective process-oriented engineer. Without this experience, I would not have had the over 30-year career as an engineer, which I have loved. In addition, Len planted the size-reduction-research virus in my blood, which I can't seem to get rid of. I also need to mention Professor Frank Aplan, whose obvious love of mineral processing and froth flotation helped to create my interest in flotation. Thus, for those of you who don't like my approach to flotation, it's all Frank's fault. The fourth, and certainly not the least, group to thank is my wife, Charlene, and our two sons, Richard and Steven. No one could have a more supportive family. In every sense of the word, my wife has been my partner from the beginningwe were married at the end of my freshman year in college. She has sacrificed much for me, and I will never really be able to express the depth of my gratitude for Charlene's encouragement, loyalty and support over the past 36 years. General comments on froth flotation The industrial process of froth flotation: Where is it going? I'm here to tell you loud and clear that froth flotation is alive, well, and growing. The inherent characteristics of high mineral throughput per unit of capital, low operating cost per unit of feed, high mineral specificity and high operating flexibility have guaranteed froth flotation an important past and a growing future. As with any evolving technology, there are always trends at the industrial level that are primarily driven by economics. The current trends include: •the movement towards larger and larger equipment, •more alternative equipment designs (e.g., the column cell),

Jan 1, 1996

By Richard R. Klimpel

INTRODUCTION The process of froth flotation as a means of upgrading the quality of coal by removing water and/or ash and/or pyrite has been receiving increasing attention since the 1960's by the world-wide coal industry. Historically, coal preparation practice has involved the use of primarily gravity based separation techniques, screening, and/or water washing in removing the larger fragments of inert material from coarse raw feed coal. The preparation, handling, and cleaning of raw coal finer than 500 micrometers was deliberately avoided. Most often the fine coal that was produced would be judiciously incorporated into the bulk of the larger sized cleaned coal product or simply discarded. Since the 1960's and accelerating in the 19701s, the desirability of performing some type of coal cleaning or upgrading on the finer raw coal feed materials has been increasing. There are several apparent industrial incentives for implementing fine coal ((500 urn) processing alternatives including the gradually increasing value of the normally lost fine coal product; the additional amounts of fine coal being produced due to changes in mining techniques; the need to appropriately dewater and upgrade fine coal streams in order to meet increasingly stringent environmental regulations on water quality. storage of fines, etc.; the increasing need to lower undesired sulfur and ash contents in some coal end-use applications; and the increasing development of speciality coal use processes that require inherently finer and cleaner coal particles such as in gasification, liquefaction, fuel injection, etc. Looking at the over 70 years of mineral oriented fine particle separation successes at the industrial level, the use of froth flotation in fine coal would also seem to be a natural process to utilize at the industrial level. In the world of strictly mineral processing technology and, more importantly in the world of mineral commodity economics, there are clearly a number of advantages to using froth flotation that are well documented and universally accepted. These include: the relatively low capital and space requirements required when compared to other process options available to upgrade minerals; the extreme flexibility inherent to the froth flotation process both with regard to throughput and differences in raw feed characteristics; the general availability of a wide range of equipment sizes and reliable suppliers of equipment; and the general availability of appropriate and relatively cheap chemicals that are the major driving forces behind mineral flotation. Just how has the implementation of froth flotation fared in coal upgrading over the last 20 years? It is towards answering this question that this paper will be oriented. As such, the discussion will be somewhat philosophical. The comments offered are based on the personal experiences of this author who has worked in the basic side of inventing new flotation chemistry and flotation analysis techniques with the major overriding goal of implementing this new technology successfully at the industrial level in both coal and mineral flotation throughout much of the world. This paper is not intended to be a complete reference to all published scientific work on the subject of industrial froth flotation of coal. As a brief background to what will be expanded upon in this paper, the author would like to offer the following summary comments: 1) the industrial practice of fine coal froth flotation is qualitatively similar but quantitatively different than industrial mineral froth flotation when viewed from either research considerations, operating viewpoints or economic evaluation criteria; 2) it is clearly possible to invent and/or develop new technology (such as new chemistry and/or equipment) that can do statistically better than current plant practice in well-controlled tests in either coal or mineral flotation circuits; 3) there is so much inherent variability to the raw feed coal streams and other operating parameters that even if there is a documented technical incentive to implement some new piece of technology, the people, instrumentation and control methodology required are often simply not available or not economic unless the potential payoff is quite large.

Jan 1, 1988

By Petr Salak, Adrian St. John, Angelos Gakis

"INTRODUCTIONCrossrail will provide London and the South East of England with a world class, high-capacity railway to ease congestion on London’s public transport system and enable better access between business, residential and entertainment districts. At the heart of this £14.8bn scheme lies Farringdon Station, which will form a major interchange between Crossrail, Thameslink and London Underground. The £400m Farringdon Crossrail station was awarded to a joint venture of BAM Ferrovial Kier (BFK) in 2011, and comprises two ticket halls, two platform tunnels and a multitude of connecting tunnels all constructed using open face sprayed concrete linings (SCL). BFK appointed Dr. Sauer & Partners (DSP) as their specialist SCL designer. As the main contractor for both C300/C410, the Western Running Tunnels and Station Caverns, and also for C435 Farringdon Station, BFK were in a unique position to develop strategies to reduce cost, time and risk for the Client. One such innovation involved the continuation of the TBM drives for the running tunnels right through the length of the new Farringdon platform caverns. This created large diameter precast concrete pilot tunnels for the subsequent platform enlargement using SCL. The benefits of such an approach are described in detail by St.John et al 2015. BFK also introduced another alternative strategy to reduce cost, time and risk, by proposing to abandon the TBMs in the ground (“Turn & Bury”) permanently rather than recovering them through the ticket hall shafts. This de-risked the programme, enabling critical ticket hall construction to proceed unhindered. However, one consequence of this revised strategy was an area of increased underground congestion, with two SCL tunnels (RTE2 and ES2) and a bored tunnel (complete with abandoned TBM) in close proximity to a ticket hall excavation (see Figure 1) and a sensitive, listed railway structure. Faced with this complex geometry, and combined with challenging geotechnical conditions (Gakis et al 2015), the team chose to install enhanced monitoring to study the behavior of tunnel RTE2 and provide additional assurance that the tunnels were performing as predicted. Applying advanced 3D numerical modelling and taking into account the residual tensile strength of the SCL helped avoid unnecessary reinforcement installation in primary lining and/or early installation of a secondary lining in RTE2 prior to ES2 construction."

Jan 1, 2016

By Archie M. Richardson, Carl E. Brechtel, Tom R. Kelly, Frank Feero

Recent work to upgrade the ventilation system for the el Mochito lead/zinc mine in central Honduras is discussed. Network modeling and underground measurements were used to evaluate cost-effective alternatives for achieving satisfactory ventilation in a complex and expanding underground operation. Both interim and long-term solutions were implemented to make mining possible under difficult conditions. INTRODUCTION Mining operations at the el Mochito Mine, located in the central highlands of Honduras, Central America, have been virtually continuous since the opening of the mine in 1948. Initially a high-grade silver mine, the mine has expanded along a westward trend of pipe-like orebodies to a distance of roughly one mile from two centrally located shafts. Production of the relatively large San Juan orebody at the western most extension of the mine (Paddock, 1981) led to the introduction of diesel-powered equipment; however, the ventilation infrastructure was insufficient to meet the needs of mechanized mining. A succession of owners/operators in the 1980s allowed the existing ventilation infrastructure to decay to the point that ventilation in the production areas of the mine was very poor. Environmental conditions in some working areas were not conducive to efficient ore extraction because of high dry bulb temperature, high humidity, and diesel emissions. Upon acquiring the mine in 1990, Breakwater Resources of Tucson, Arizona began an aggressive program to refurbish the mine infrastructure to complete extraction of the San Juan orebody and to allow the extension of the mine another 2500 ft to the west for extraction of the Nacional orebody. The program included increasing the capacity of the main ventilation system. This article presents a case history of the process of upgrading the ventilation system in a mine where extensive old workings cause large air leakage. This process has been one of selecting solutions to difficult technical problems that are compatible with the existing mine infrastructure and economic constraints. The initial ventilation system is described in the background section, along with ventilation projections for the mine expansion. Field characterization of the ventilation system for design verification and fan specification is then discussed. The paper describes a series of interim changes to the system to improve ventilation pending completion of new ventilation boreholes. In addition, the temperature/ heat problems in the mine are described. BACKGROUND Initial Condition of Ventilation System The ventilation system is illustrated in Figure 1, which shows the extent of mining with the main ventilation paths superimposed. Early mining around the two shafts opened up vertical connections (stopes and raises) over the entire 2420 ft (737.6 m) of vertical extent, and mining progressed to the west primarily using compressed air and electric-powered equipment. Since ventilation was not a complex problem in the original mining system, the stopes and interconnecting raises were not sealed. The San Juan orebody was much larger than the silver ore zones mined previously, being primarily zinc and other base metals. Its geometry, size, and grade allowed the use of vertical crater retreat (VCR) stoping with diesel mucking and haulage equipment. Its depth, along with the existence of warm groundwater, resulted in a mine climate problem on the lower levels. To establish a complete ventilation circuit, two vertical boreholes (Bonanza Nos. 1 and 2) had been drilled by previous owners from the surface in the vicinity of the San Juan orebody. The system design called for air to be drawn down the intake shafts, across the lower mine levels to the San Juan workings, then up through the San Juan ramps and ore passes to these two exhaust boreholes (see Figure 1). In practice, however, only the Bonanza No. 1 borehole was drawing air through the desired path. Leakage across the old upper levels from the intake shafts, the Caliche tunnel, and from intervening abandoned stopes and raises supplied most of the air flowing to the base of the Bonanza No. 2 borehole. In effect, there were two ventilation circuits in semi-parallel through the mine, of which only one was delivering appreciable air to the San Juan workings.

Jan 1, 1993

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By G. Whitney, J. Steiner

Introduction The Summitville Mine located in southwestern Colorado is an example of the potential impact that natural weathering and mining activities can have on environmental quality. Acid rock drainage and metals leaching from active and historic mine districts are a major environmental problem facing the mining industry today. Sulfide ore exposed naturally or during mining operations has the potential to produce leachates that contain high concentrations of dissolved metals. These leachates can affect the quality of plant and animal health in surface water used for agriculture, recreation and human consumption. Pintail Systems has developed biological processes for detoxification of cyanide in heap-leached spent ore and process solutions. During application of spent ore cyanide bio-detox processes we have also observed a substantial reduction in many of the leachable metals in biotreatment solutions. These field observations of soluble metal reduction led us to propose that metal bio-mineralization should be a secondary treatment goal of spent ore detox in Focused Feasibility Studies (FFS) for the Summitville Mine. In the FFS (described in a companion paper) for Summitville, spent ore from the heap leach pile (HLP) was loaded into PVC test columns (six inch X ten foot) and was leached with bacteria/nutrient solutions. Two treatment designs were studied: a standard percolation leach (aerobic) and a saturated leach design where the ore was saturated with the heap leach solution to which bacteria and nutrients were applied in a continuous biotreat leach. The bacteria for all column treatment tests were isolated from the spent ore at depth through the Heap Leach Pile (HLP) in the saturated and unsaturated zones. Bacteria isolated from the heap material were tested for cyanide oxidation capacity and were submitted to a bioaugmentation program to improve reaction kinetics and metals tolerance. The final treatment population consisted of several distinct species including aerobic heterotrophs, facultative anaerobes and sulfate-reducing bacteria. The remineralization of soluble metals was observed through a decrease in copper in column leachate solutions and the formation of observable mineral products on ore surfaces in the columns. Several of the tests were run in clear PVC columns to facilitate observation of mineral formation. The column ore contents were photographed before, during and after the biotreatment process which ranged from 10 to 20 days for each test column. Ores were collected from each column after the treatment was complete and were submitted for SEM and TEM study at two laboratories - the USGS SEM Laboratory in Denver, Colorado and the City College of New York SEM laboratory. Treatment Process Description Numerous species of bacteria, fungi and yeasts are capable of accumulating many times their weight in soluble metals. Both living and dead biomass are effective in removing soluble metals from waste streams containing gold, silver, chromium, cadmium, copper, lead, zinc, cobalt and others. Soluble metals may also be immobilized in soils by natural or engineered biomineralization reactions. Several commercial processes using biological reactions are being applied on an industrial scale for metal remediation. Bacteria found in natural and extreme environments have developed a wide variety of metabolic functions to adapt to these environments. These natural microbial functions contribute to global mineral cycling that continuously forms, transforms and degrades minerals and metals in the environment. Biomineralization is described as a surface process associated with microorganism cell walls where the remineralization occurs. The biogeochemical activities initiated by microorganisms in ores, soils, surface and groundwater environments can dominate the formation and transformation of those mineral environments.

Jan 1, 1995

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

By William A. Rhoades

INTRODUCTION Ever since miners were faced with the task of moving ore, some form of scraper has been in use. At first, men and beasts of burden supplied the power to move the scraper, and later, machines were developed for this purpose. In the early 1900s, a few mining properties used small pneumatic single-drum winches to pull loaded scrapers to a raise or ore pocket, and the empty scraper then was dragged back to the muck pile by a miner, as shown in Fig. 1. Just prior to 1920, an improvement was introduced, using two single-drum hoists. As illustrated by Fig. 2, the second hoist was used to return the empty scraper to the muck pile. However, this arrangement still re¬quired two men, one operating each hoist. The next developmental step was to eliminate the second man by locating the hoists side-by-side and using one man to control both hoists. As illustrated by Fig. 3, this involved the use of a tail rope over a sheave to pull the scraper back. The greatest progress in development and the great¬est increase in the use of slusher haulers occurred between 1920 and 1930. In 1921, the Sullivan Machinery Co. designed and built the first two-drum scraper oper¬ating on the principle illustrated by Fig. 4. It was powered by a 4.5-kW (6-hp) Turbinair(r) motor and would pull a 450-kg (1000-1b) load at 0.61 m/s (120 fpm). In 1922, this unit was shipped to the Verona Mining Co. of Caspian, MI, and it experienced immediate success in the Lake Superior iron ore district. Since the two-drum slusher was much less expensive and more efficient than hand mucking, the Lake Superior mines were saved from financial disaster when iron-ore prices fell 25% between 1923 and 1925. Immediately there¬after, a demand developed for slushers that would oper¬ate with electric power, which was considerably cheaper than compressed-air power. In 1923, the Sullivan Machinery Co. responded with the first electric-powered double-drum hoist. During subsequent years, design improvements in¬cluded separate tail-drum gearing to increase the tail¬drum speed, as welt as a number of safety features such as rope guides. One result of these improvements and their utilization by the Michigan iron mines was an in¬crease of 100% in the tons of ore per miner per day in those mines between 1924 and 1929. Since the two-drum slusher was capable only of straight-line mucking, it was not a practical machine for use in open stopes. In 1929, the Sullivan Machinery Co. introduced the first three-drum slusher. As illustrated by Fig. 5, two tail drums and one hauling drum were provided. A tail sheave could be placed at each side of the stope, and the ore then could be loaded and hauled to a central point from the entire width of the stope. During the 1930s, progressively larger slushers were demanded. By 1940, two- and three-drum units were available with motor power as high as 45 kW (60 hp), and slusher power continued to increase after 1940. Be¬tween 1951 and 1952, Joy Manufacturing Co. designed a 112-kW (150-hp) two-drum Blusher for the Climax Molydenum Co. Although there has been no demand for a slusher more powerful than this, 150- to 225-kW (200- to 300-hp) slushers are quite feasible at the pres¬ent time. During the last 30 years, many slusher improvements have been made to the operating life, operational safety, and ease of operation and maintenance. Increased tail-drum speeds have decreased overall scraping times. Rope guards, totally enclosed drums, and operator shields have reduced the hazards of injury due to wire¬rope breakage. Improvements in lubrication have made the slushers relatively maintenance-free, with long operating lives. The introduction of spring-actuated drag brakes prevented uncontrolled unreeling of dis¬engaged drums, allowing the development of practical remote-control slusher operation. Remote control now is available in a choice of all-air, all-electric, or air¬electric slushers. APPLICATIONS Quite simply, slushers are used to load and transport material (ore), generally over a short distance of from

Jan 1, 1982

Introduction The objective of the design and operation of a mine is to make as much money as possible within certain reasonable and responsible constraints. The imposition of constraints on the objective of maximizing profits can result in an optimum mine design. This paper deals with the maximization of profits. Optimization will be discussed in a future paper. In the mining industry, it is generally accepted that the maximum economic recovery is defined by identifying the reserves and the sequence of mining that results in maximum net present value (NPV) or discounted cash flow-rate of return (DCF-ROR), as indicated in Fig. 1 (see Lemieux, 2000). Current parametric analyses techniques, such as varying the mining costs; profit reservation; sales price or metal content, to define pit limits and mining sequence technically referred to single, double or triple parameterization, do not assure that the curve in Fig. 1 will be maximized. In defining the sequence, the first material to be mined should be the first grouping of recoverable mineral that has the highest after-tax unit value and that forms a feasible mining unit. The assumed mining of the first pit should be followed by the assumed mining of the second most profitable and practical pit or pit expansion. If this process of grouping and sequencing is followed until the profitability approaches zero, the general sequence required to extract the maximum economically recoverable reserve is identified. The curve in Fig. I is constructed by calculating the NPV for a series of sequences, each assuming operations are terminated on successively lower profit pits. If the operation ceases on a pit with too high of a profit, opportunity is lost, and the NPV is lower than the maximum. The NPV curve peaks and then starts to decline before the zero-profit pit is assumed mined. This occurs because the money invested in advanced stripping would have produced a greater return if it was invested at the dis¬count rate. The maximum economically recoverable reserve is defined by the pit limit corresponding to the maximum NPV or DCF-ROR. Pit and phase definition The real challenge of reserve definition is identifying the phasing and after-tax profit per ton, so that the curve in Fig. 1 can be plotted. The after-tax profit cannot be properly addressed until after the pits are designed, the production is scheduled and the cash flow is estimated. Throughout the years, many investigators have ad¬dressed this problem. A major thrust of the famous paper in which Lerchs and Grossman (1965) presented their algorithm was to identify a sequence that would produce a result similar to that identified in Fig. 1. Whittle's (1988) four-dimensional analysis is designed to address this issue. The author addressed this same issue in a paper presented to the 1968 Canadian Institute of Mining national convention (Lemieux, 1968). Without a relatively simple methodology to con¬struct the curve in Fig. 1, the number of iterations in a trial-and-error method becomes prohibitively burden-some. The pit designer seeks a simple pit-planning parameter that will serve as a proxy for the after-tax profit per ton used in defining the curve in Fig. 1. The designer's goal is to apply this pit-planning parameter using standard pit-wall location techniques to design a series of pits that will provide a guide to the sequencing. Common techniques used to position pit walls are strip-ratio limits, highwall incremental analysis, floating-cone methods or "maximizer" applications1. Costs, revenues, noncash charges, taxes and profits are usually structured on a unit basis for use in the pit-wall location analysis. Application of the proxy parameter using the pit-design tools should, ideally, result in the definition of pits and sequences that will maximize the curve in Fig. 1. 1 The pit-design tool frequently referred to in the literature as an "optimizer" defines a pit boundary that maximizes the value contained within based on the input parameters and the safe pit-wall angle. The optimization of the pit design involves selection of rate of production, cutoff grade, product quality and other considerations. These considerations constrain the maximization of value. Therefore, pit-design tools that maximize value should be called "maximizers."

Jan 1, 2000

By N. J. Peberdy, C. N. Moreton, K. C. Garner

Introduction During the past decade a major objective of the process industry has been to use digital computer technology to improve plant operating efficiencies. This objective implied some form of optimization, a concept that has various interpretations depending on the view of the prospective user. For the purpose of this paper, optimization of a process plant is defined as the establishment and setting of plant operating conditions that maximize some mathematical yield function, i.e. maximum profit, minimum residue, etc. Analysis of these objectives and the available design and implementation techniques led to the conclusion that digital computer and optimization techniques are not the stumbling blocks, but rather the development and derivation of the mathematical models of the unit operations and process plants to be optimized. Such models should not only describe the optimized (steady-state) objective, but also how one steers to this state (control algorithm). Due to the multidisciplinary nature of the skills associated with the design and operation of process plants, the development of suitable models by a single discipline, such as the process control engineer, was found to be not only difficult but often impossible, due to budget and human resource limitations. To over-come these limitations, a computer aided design (CAD) tool has been developed. It aims to provide a productivity tool to the various disciplines, at the same time coordinating the technical input from each. The system described is but the starting point in an evolutionary development of a tool that, with use, is becoming more efficient and cost effective to use. Development has become an application engineering activity rather than the preserve of the computer specialist. Project phasing The development of a mathematical description of a process plant requires coordination of information from conceptual design to operation management. The activities required to build and operate a process plant are divided into four basic chronological activities or phases. These activities are often undertaken by different organizations and disciplines. As a result, continuity is often lost with the resultant loss of critical design data. The major activities are considered to be: conceptual and flowsheeting; detailing around the P & ID; building and commissioning; and plant operation. The CAD system described provides a design tool to be used for each of these activities, as well as providing continuity between the activities and the disciplines involved. The heart of the system is the dynamic simulation of the flowsheet. Each of the activities will be discussed, leading to two simple examples that demonstrate the use of the simulator. Figure 1 shows a schematic format of the various activities and the path followed by the dynamic flowsheet simulator in the life of a project. Flowsheet development The prime requirements in the design and develop¬ment of a process flowsheet are • selection of the correct unit operations to achieve the most economic (capital and operating) beneficiation of the specified reserve ; • the sizing of the unit operations to achieve the desired results, as a function of the projected feed rates etc., to handle the time related (dynamics) of the process; and • the production of a set of engineering documents showing the drawn and labeled flowsheet with an equipment list and process specification for each of the unit operations. The question may well be asked at this stage why dynamic flowsheet simulation should be considered when steady state modeling has been found to be adequate to date. With the increases encountered in the cost of capital, one often cannot afford the luxury of designing around the compounding worst case technique. Further, a more accurate design of the control surges can be achieved. No information is lost in that the steady state solution is in fact a subset of the dynamic model. In generalized state space modeling, the differential equations describing the process dynamics are illustrated in the following matrix notation: XDOT=A.X+B.U(1) Y =C.X+D.U(2) where XDOT describes the set of first order derivatives of the system state Vector, and X- is the system state Vector; A - is the system matrix operator which in the general nonlinear case is both a function of X and time ; U- is the process input vector; B - is the input mapping matrix; Y - is the set of observations; C - is the output mapping matrix which maps X - onto Y; and D- maps the input onto the observations. Thus, by time integration of the system dynamic equations, described in (1), the dynamic trajectory away from any set of initial conditions can be deter¬mined. Further, by finding the conditions at which XDOT = 0, the steady state solution can be determined.

Jan 1, 1987

By Benno Ring

"When designing segmental lining at a cross passage, modern solutions try to bear the loads transferred around the opening within the rings by using shear elements in the ring joints. Commonly used shear systems providing coupling are cam & pocket, shear keys, bicones and steel dowels. These solutions have advantages and disadvantages, but most importantly they have limits in their capacity to stabilize the lining at cross passages. With the help of 3D-FE-analyses for an opening of one segmental ring these different solutions are compared taking various ground conditions into account. As a result the capacity of the individual system is determined. Notice is made regarding the applicability of the system as well as the consequences for the design and the verification. INTRODUCTION The design of the lining at a cross passage is among the most difficult tasks when designing TBM-tunnels. At one time the bedding of the lining disappears due to the excavation behind the lining, at another the ring-structure is interrupted when parts of the lining are removed for the opening. The trick is to transfer the loads around the opening into other segments and to areas of the lining which are still supported by the ground. In the rings immediately at the cross passage this creates a distribution of forces within the segments that strongly deviates from that of the regular lining. The design needs to consider the boundary conditions of the respectively chosen solution for stiffening of the opening and its peculiarities in the modeling as well as in the verifications. In the past segments of steel or an internal strutting of steel (DAUB 2013) placed around the later openings were commonly used, see Figure 1. The application of a strutting from the outside to the rings is restricted to such circumstances when for example a connection from a shaft to the tunnel can be made in suitable ground conditions (Ring and Glomp 2012). Though steel segments provide a robust solution, they are expensive due to their laborious manufacture, the high proportion of construction steel required and the time consuming installation in the tunnel (Häfliger 2013). A strutting of steel requires extensive design and verification. The resulting structure usually weighs several tons and must be erected as well as demounted within the confined space of the tunnel, requiring suitable transport and assembly equipment. The cross section remains restricted, thus causing a bottleneck for TBM operations and hindering the works and installations for the cross passage works."

Jan 1, 2016

By Robert F. Kaufmann

INTRODUCTION In response to the Uranium Mill Tailings Radiation Control Act (Public Law 95-604, Section 114c), a report has been prepared by the U.S. Environmental Protection Agency (1980) which describes the potential health and environmental effects associated with uranium mine wastes. The author was responsible for the geology, hydrology and aqueous pathways analysis portions of the report. All potential sources of contaminants, solids, liquids, and airborne emissions, which might cause adverse impacts upon the environment or health of nearby populations were considered. Wastes at both active and inactive uranium mines (Hans and O'Connell, 1981) are described and assessed with respect to dose and health effects (Blanchard et al. 1981). The emphasis of this paper is on the liquid emissions and the aqueous pathway. Model uranium mines were developed for surface and underground mines. The model underground mine was placed in New Mexico while the model surface mine was in Wyoming. Development of a generic assessment was complicated due to the inherent variability of uranium mines. Also, most information available relates to uranium mills rather than to mines, or reported measurements were often influenced (biased) by wastes from a nearby mill and did not specifically reflect mine discharges. Consequently, the assessment presented is often based on conservatve (maximizing) assumptions which are identified as such, and which were based on knowledge from field investigations, available literature, and consultations inside and outside the U.S. Environmental Protection Agency. Maximum use was made of available data, supplemented with limited field studies in Colorado, Texas, New Mexico, and Wyoming. PREVIOUS INVESTIGATIONS Stream survey work in Colorado by the Water Pollution Control Commission and the U.S. Geological Survey (Moran, 1974; Wentz, 1974) mentioned a portion of the Uravan Mineral Belt and uranium mines therein, but the work did not emphasize uranium mines. Significant amounts of acidity and total trace metal concentrations were found in streams from 18 different mining areas. Dilution and chemical precipitation below mine drainages decreased concentration and increased the pH. Given enough time and distance, the streams recover naturally, but the accumulations of trace metals in the sediments increase. Field observations in 1971-72 of streams in most of Colorado indicated that approximately 724 km of streams in 25 different mining areas were adversely affected by mine drainage (Wentz, 1974). Discussions of the impacts of uranium mining on water quality or quantity are incidental in numerous impact statements and environmental reports prepared by industry and (or) the U.S. Nuclear Regulatory Commission as an integral part of licensing or relicensing uranium mills. Coverage on mining is usually minor as the principal focus is on milling impacts. The same is true for the recently prepared generic EIS on regulation of uranium milling (U.S. Nuclear Regulatory Commission, 1979). Radiochemical assessment of surface and groundwater in uranium mining districts of New Mexico is done by self-monitoring programs associated with NPDES permits. Also, radiochemical assessment studies have been funded by the New Mexico Environmental Improvement Division and U.S. Environmental Protection Agency. Unfortunately, there has been no concerted effort to prepare broad assessments of the cumulative impacts of mining and milling in areas of interest in Texas, New Mexico, and Wyoming. Critical review and synthesis of these types of data can produce rather useful information. For example, the publication "Water Quality Impacts of Uranium Mining and Milling Activities in the Grants Mineral Belt, New Mexico" (U.S. Environmental Protection Agency, 1975) addresses the groundwater and surface water changes as the result of extensive uranium mining and milling production in a relatively confined area. A more recent report by the Wyoming Department of Environmental Quality summarizes 16 years of aqueous radium and uranium data. The study reports significant amounts of radium-226 and uranium were present in surface water in the Shirley Basin as a result of inadequate mine water treatment (Harp, 1978). In Texas, surface water and groundwater monitoring conducted by industry, as well as by State and Federal agencies, reveals little or no change of chemical quality attributable to uranium mining and milling (Gesell, 1977; Kallus, 1976). This conclusion is based on 586 samples collected from 198 stations over a period of 39 years, but primarily from 1961 to 1975. The State monitoring program by several agencies is continuing, but data reporting and interpretation is lagging. One survey (Itin, 1975) showed that none of the mine water from the 10 lakes that were sampled was suitable for human use. The lakes were also unsuitable for irrigation due to mineralization of the water by sulfate, chloride, and TDS. One of the 10 was suitable for stock watering.

Jan 1, 1981

By Richard E. Ray

INTRODUCTION Main mine fan installations at underground coal mines must be designed to comply with the requirements outlined in 30 CFR (Code of Federal Regulations) Part 75. However, very little emphasis is typically placed on the aerodynamic optimization of shaft collar designs, ductwork configurations, and the selection of a fan isolation, or "closing door," system. Field testing has shown that poorly designed systems may produce large pressure losses between the shaft and fan inlet, as well as air turbulence problems which adversely impact fan efficiency and sometimes fan structural integrity. This paper analyzes each component of a surface exhaust fan system for an underground coal mine: • The shaft collar arrangement, • The 45 or 90-degree turning duct that isolates the fan motor from the return airstream, • Alternative ductwork arrangements for parallel fan installations, • The self-closing doors or dampers that are required for multiple fans, and • Horizontal and vertical evas6 arrangements on the discharge side of the fan. While discussion is limited to axial fan installations at shaft mines, most of the recommendations would also apply to centrifugal fan installations and to main mine fans for drift mines. FEDERAL REGULATIONS The mandatory safety standards for underground coal mines that apply to the installation of main mine fans are listed in 30 CFR Part 75.310. Stipulations which impact the design of the fan sys- tem arrangement are contained in paragraphs (a)(l), (a)(5), (a)(6),(d), and (1) of 75.31 0: • Each main mine fan shall be installed on the surface in an incombustible housing; • Each main mine fan shall be protected by weak walls or explosion doors in direct line with possible explosive forces; • Explosion doors or weak walls shall have a cross-sectional area at least equal to the area of the return air shaft or drift opening through which the pressure from an underground explosion would be relieved; • Each main mine fan shall be offset by at least 4.57 m (15 feet) from the nearest side of the mine or shaft opening; • Automatic closing doors shall be provided to prevent possible air reversals through the fans in mines ventilated by multiple main mine fans. The installation of main mine fans underground is prohibited by 30 CFR Part 75.310, (a)(l). Providing explosion doors or weak walls in line with a possible underground explosion and offsetting the fan 4.57 m (1 5 feet) from the edge of the shaft preclude the installation of the fan directly over the shaft. In addition, 30 CFR Part 75.507-1 specifies that all electrical equipment used in return air courses be permissible, which requires locating the fan motor outside of the airstream exhausted through the fan. To comply with the Part 75.310 regulations, explosion doors are typically located directly over the shaft as part of the shaft cover or 'bonnet." A sufficiently long section of ductwork placed between the shaft enclosure and fan, equipped with a 45 or 90- degree turn, enables the fan to be offset at least 4.57 m (1 5 feet) from the shaft and the motor to be isolated from the airstream. The need for the turn in the ductwork can be eliminated by driving the fan from the discharge side, instead of from the inlet side, and providing for a 90-degree bend in the evas6. For multiple main mine fan systems, self-closing doors can be located on the inlet or discharge side of the fan. FAN SYSTEM COMPONENTS A poorly designed exhaust fan arrangement will produce turbulent conditions at the fan inlet and significant pressure losses between the shaft and fan. Field tests have shown that pressure losses through inlet ductwork at some installations may constitute as much as 30 percent of the total mine head. Unfavorable flow conditions may cause a change in the operating point on the fan performance curve, resulting in reduced fan efficiency and pres- sure generation capability. In some instances, poor airflow distribution through the fan may create mechanical stresses that lead to fan blade or fan housing fatigue failures. The objective in designing an air duct arrangement for a fan system is to provide a low cost duct that will move the air from the shaft to the fan in a uniform air distribution and with minimum energy requirements. The optimum solution is a compromise between technical and economic factors. Aerodynamic design considerations for the shaft collar and cover, fan inlet ductwork for both single and parallel fans, self-closing doors, and evas6 are discussed in the sections below. Shaft Collar/Cover To prevent excessive turbulence and non-uniform air distribution through the fan, a smooth transition for the change in direction of airflow at the top of the shaft must be provided. However, the concrete lining of the shaft is abruptly terminated at the surface at many coal mine installations (see Figure I), creating a sharp 90-degree bend. The airflow exiting the shaft becomes heavily concentrated in the upper portion of the horizontal fan in-

Jan 1, 1997

By R. R. Klimpel

Introduction The grinding of coal-water slurries has received increasing industrial attention during the last decade. In particular, there is long-term interest in the use of pulverized coal-water slurries to replace oil in combustion equipment and in the development of coal gasification/liquefaction processes that require coal-water slurries as feed. More specifically, the use of coal-water slurries in gasification requires grinding a high-density slurry containing the smallest amount of water consistent with slurry pumping and spraying. As part of a fundamental engineering research support program aimed at the industrial implementation of dense coal-water slurry grinding, this author has published several papers on how specific rates of breakage vary as a function of slurry rheology (Klimpel, 1982,1982/83). These papers demonstrated that there is a consistent pattern of change in specific rates of breakage of coal in dense slurries with controlled variation in slurry rheology. By matching rheological data with laboratory grinding results, it was possible to identify directly slurry conditions that correspond to: 1) slowing down of breakage rates, 2) the occasional acceleration of breakage of some sizes, and 3) conditions where chemical additives will increase rates of breakage. In brief, these conditions were analyzed using two different criteria: a) the net production rate of material less than some specified size (e.g. kg/min of minus 325 mesh) in a standard batch laboratory mill test as a function of controlled changes in grinding conditions, and b) the use of the one-size-fraction feed method, which consists of following the disappearance of this largest size over grinding time in a batch laboratory mill to arrive at well-known specific rates of breakage (Austin et al., 1984). Detailed references to the methodology used as well as the conclusions are available (Klimpel, 1982, 1982/83) and will not be repeated here. The purpose of this paper is to further demonstrate several additional characteristics of dense coal-water slurry grinding that were shown in a simplified sense in the earlier publications of the author but which have clearly demonstrated themselves as being very important in the industrial simulation and scale-up of such coal-water grinding systems. In particular, this includes the clear and unambiguous demonstration of how the simultaneous acceleration of breakage of some size fractions and slowing down of the breakage of other size fractions is occurring as a function of changes in coal-water slurry density. In the earlier publications (e.g. Klimpel, 1982), it was shown by specially designed experiments that the addition of fine material and/or the use of a chemical thickening agent accelerated the specific rates of breakage of coals of coarser size fractions using the one-size-fraction method. There were also numerous examples given of non first-order breakage (the slowing down of coal breakage rates) using also the one-size-fraction method due to the presence of excessive amounts of fines which corresponded to the development of a rheological yield value. The problem with the simulation and scale-up of any laboratory and/or pilot-scale mill data to an industrial scale using the mechanistic modelling approach involving specific rates of breakage and breakage distribution parameters (e.g. Austin et al., 1984) is the number of assumptions involved in translating the smaller mill breakage parameters to the predicted larger mill breakage parameters. It is apparent, at least to this author, that to accurately simulate and predict larger scale equipment performance from smaller scale data (given that the larger scale data performance is known and hence predictions can be thoroughly checked) requires a better knowledge of breakage parameters than is currently available. More specifically, it was felt that one of the chief problems was the inability of the one-size-fraction method of determining breakage parameters to sufficiently represent the actual magnitude and sometimes even the directions of. the complicated interactions involved with slurry density changes in coal-water slurry grinding. Thus, a special set of experiments was conducted in a somewhat larger batch ball mill (0.457 m diam x 0.610 m length) than the 0.203-m-diam mill used in the original rheology characterization paper (Klimpel, 1982) so as to minimize any unusual effects due to wall-ball interactions (2.54-cm-diam balls used in both mills). More importantly, the measurements of specific rates of breakage were done using a proprietary tracer method on a portion of a given size fraction, which was then remixed into a natural feed size distribution before grinding. The experimental procedure and analysis of subsequent data was done in exactly the same manner as the radioactive tracer technique on coals as originally developed by Gardner (1962). The advantage of such an approach is that it makes no assumptions such as the independence of the specific rate of breakage of any size on the absolute sizes and amounts of other sizes present (both larger and smaller) in the mix of natural feed material. It will be shown that the measured rates of breakage using the direct tracer technique and the one-size fraction method on the same coal are indeed different. In fact, an accurate assessment of what is happening to the rates of breakage as a function of changing slurry density can only he made by measuring particle breakage under grinding conditions approximating the size distributions actually being produced in practice. Experiment procedures and results The pilot mill used was 44 cm diam x 60 cm long with a volume of 91,250 cm3 and was fitted with six 0.5-in. lifter bars.

Jan 1, 1992

By C. J. Monahan

INTRODUCTION With the introduction of large diameter hole drilling to underground mining, a need for the development and utilization of new blasting technology has arisen. This new technology has resulted in more efficient blasting and in easier and cheaper mining. The application of spherical charges to the crater blasting method is an im¬portant part of this technology. Spherical charges, or their geometric equivalent, make for highly efficient use of explosives in the cratering application. Their geo¬metrical configuration, normally a length :diameter ra¬tio of <6:1, limits their charge weight size. For exam¬ple in a 165-mm (61/2-in.) hole a spherical charge weighs approximately 34 kg (75 lb). These dimensional and weight restrictions require careful engineering design and control in the production application. Strathcona mine, a large tonnage mechanized cut¬and-fill and blasthole stoping operation located on the north rim of the Sudbury Basin, has successfully em¬ployed the crater blasting method in production blasting for the past two and a half years (as of 1979). As a result, this method, using spherical charges, has made possible the development and implementation of a verti¬cal retreat blasthole mining method which is being used for rib pillar recovery. The crater method used at Strathcona essentially in¬volves blasting off horizontal slices of ore into an under¬cut while retreating vertically to the pillar overcut elevation. Loading and priming of the holes is achieved from the overcut and the blasted muck is recovered by load¬haul-dump (LHD) units through a drawpoint system. Valuable experience on the behavior and performance of explosives in large diameter holes has been gained in both the operating and engineering fields from the appli¬cation of this blasting method. It has been used in drop raising and is currently being evaluated for use in pri¬mary stoping operations. Also, serious consideration is being given to its use for the recovery of post pillars in mined-out and backfilled cut and fill stopes. The method has significant advantages over conven¬tional methods of pillar recovery from economic, ground control, and safety points of view. These include cost benefits due to a minimum stoping development, less ground support, reduced labor requirements, and faster mining rate. Research on the blasting mechanism indi¬cated the method should result in less damage to fill walls than would be normally expected with conven¬tional longhole or large diameter hole cylindrical blast¬ing (benching). This is primarily due to the limited charge weight per delay because of the size dimensions of a spherical as compared with a cylindrical charge and also because the large diameter holes experience less de¬viation than standard-sized longholes. Experience has shown that wall damage has been minimized with the use of this method. The safety aspects of the working environment are also improved as there is no require¬ment for manpower to go under exposed backs after blasting, which is a significant improvement over the cut-and-fill method. CRATERING THEORY General The term cratering, in blasting, is applied to the for¬mation of a surface cavity in a material as the result of detonating an explosive charge in that material. This blasting concept was initially used as a tool in the eval¬uation of explosives performance; however, it has been utilized more recently on surface and to a lesser extent in underground blasting operations. Explosive charges used in crater method are nor¬mally spherical or the geometric equivalent, as research into the application of this breakage mechanism to rock indicates that spherical charges or their equivalent pro¬duce optimum results. In blasting practice, spherical charges have been defined as having a length to diameter (L: D) ratio of 4:1 or less, and up to, but not exceeding a L: D = 6:1. Thus, for holes 165 mm (61/2 in.) diam, a charge 165 mm (61/2 in.) diam, and 990 mm (39 in.) in length would constitute a spherical charge. Crater Testing Crater testing for underground application is nor¬mally accomplished by drilling smaller than production sized horizontal holes into drift walls, placing relatively small charges of explosives in the holes at various depths of burial, firing the charges, and measuring the results. The results are scaled to develop charge weight size and depth of burial parameters for production hole sizes. In theory, this procedure is valid and is appli¬cable for surface blasting; however, in underground applications where the craters are inverted, there is some question as to whether extrapolation of the test results is appropriate. As well as the fact that craters are in¬verted with resultant gravity stresses, testing is often done in a waste rock drift where the properties differ from those in ore rock. There is also some question as to whether a 51 or 76-mm (2 or 3-in.) diam charge behaves similarly to a 165-mm (61/2-in.) charge in a confined borehole. Crater testing at Strathcona involved drilling 76-mm (3-in.) diam holes at various depths ranging from 1 to 3 m (3 to 10 ft). The test holes were drilled in the wall of a predominately waste drift as well as in a stope wall. Charges were placed at various depths of burial, were confined, fired, and the results measured. Although the results were somewhat erratic, possibly because of the aforementioned reasons, they did provide guidelines for basic design for production application. MINING METHOD APPLICATION FOR PILLAR RECOVERY Following initial crater testing, assessment, and tech¬nical evaluation, it was decided to use this blasting tech¬nique in a production application, and to incorporate

Jan 1, 1982

By Stephen B. Castor

The rare earth elements (REE) which include the 15 lanthanide elements (Z = 57 through 71) and yttrium (Z = 39) are so called because the elements were originally isolated in the late 18th and early 19th centuries as oxides from rare minerals. Most REE are not as uncommon in nature as the name implies. Cerium, the most abundant REE (Table 1), comprises more of the earth&apos;s crust than copper or lead. Many REE are more common than tin and molyb¬denum, and all but promethium are more common than silver or mercury (Taylor, 1964). Promethium (Z = 61) is best known as an artificial element, but has been reported in very minute quantities in natural materials. Lanthanide elements with low atomic numbers are generally more abundant in the earth&apos;s crust than those with high atomic numbers. In addition, lanthanide elements with even atomic numbers are two to seven times more abundant than adjacent lan¬thanides (Table 1) with odd atomic numbers. The lanthanide elements traditionally have been divided into two groups: the light rare earths (LREE), lanthanum through eu¬ropium (Z = 57 through 63); and the heavy rare earths (HREE), gadolinium through lutetium (Z = 64 through 71). Although yttrium is the lightest REE, it is usually grouped with the HREE to which it is chemically and physically similar. The REE are lithophile elements (elements enriched in the earth&apos;s crust) that invariably occur together naturally because all are trivalent (except for Ce+4 and Eu+2 in some environments) and have similar ionic radii. Increase in atomic number in the lanthanide group is accompanied by addition of electrons to an inner level rather than the outer shell. Consequently, there is no change in valence with change in atomic number, and the lanthanide elements all fall into the same cell of the periodic table. The chemical and physical differences that do exist within the REE group are caused by small differences in ionic radius, and generally result in segre¬gation of REE into deposits enriched in either light lanthanides or heavy lanthanides plus yttrium. The relative abundance of individual lanthanide elements has been found useful in the modelling of rock-forming processes. Comparisons are generally made using a logarithmic plot of lanthanide abundances normalized to abundances in chondritic (stony) meteorites. The use of this method eliminates the abundance vari¬ation between lanthanides of odd and even atomic number, and allows determination of the extent of fractionation between the lanthanides because such fractionation is not considered to have taken place during chondrite formation. The method is also useful because chondrites are thought to be compositionally similar to the original earth&apos;s mantle. Europium anomalies (positive or negative departures of europium from chondrite-normalized plots) have been found to be particularly effective for petrogenetic modelling. REE were originally produced in minor amounts from small deposits in granite pegmatite, the geologic environment in which they were discovered. During the second half of the 19th century and the first half of the 20th century, REE came mainly from placer deposits. With the exception of the most abundant lanthanide el¬ements (cerium, lanthanum, and neodymium), individual REE were not commercially available until the 1940s. Since 1965, most of the world&apos;s REE have come from two hard rock deposits: Mountain Pass, United States, and Bayan Obo, China. GEOGRAPHIC DISTRIBUTION OF REE DEPOSITS More than 70% of the world&apos;s REE raw materials come from three countries: China, the United States, and Australia. China emerged as a major producer of REE raw materials during the 1980s, while Australian and United States market share decreased dramatically (Fig. 1). Table 2 gives recent annual production figures along with estimated reserves by country, and Fig. 2 shows loca¬tions of significant REE mining. MINERALS THAT CONTAIN REE Although REE comprise significant amounts of many minerals, almost all production has come from less than ten minerals. Table 3 lists minerals that have yielded REE commercially or have po¬tential for production in the future. Extraction from a potentially economic REE resource is strongly dependant on its REE miner¬alogy. Minerals that are easily broken down, such as bastnasite, are more desirable than those that are difficult to dissociate, such as allanite. In general, producing deposits contain REE-bearing min¬erals that are relatively easy to concentrate because of coarse grain size or other attributes. For more thorough discussions of REE¬bearing minerals see Mariano (1989a) and Cesbron (1989).

Jan 1, 1994

By P. W. Hynes

Tg Soda Ash has operated a shortwall mining system in the Green River, Wyoming trona district since August. 1982. The system has proven to be highly productive. In an effort to reduce the relatively expensive development mining and Improve mine planning flexibility, a method to mine the return air tailgate during the shortwall retreat was tested beginning In April, 1987. As of June. 1988 four shortwall panels had been completed utilizing the advancing tailgate, producing over 815,000 tons.

Jan 1, 1989

By Y. A. Attia

Auriferous sulfide ores often contain very finely disseminated gold and other precious metal particles inside the sulfide crystals. The encapsulation of precious metal particles gives rise to the refractory nature of such ores to current extraction methods. As a result, these fractions are frequently lost in the tailings. The purpose of this work was to investigate the feasibility of pretreating these refractory tailings with bioleaching before employing standard extraction techniques such as cyanidation. Leadville gold ore tailings containing auriferous pyrite were first concentrated by froth flotation. The pyrite concentrate was treated for different periods of time with 12 weeks adapted thiobacillus ferrooxidans before cyanide leaching. The bioleaching of the concentrate before cyanide leaching increased the gold extraction from 32% to 95% and that of silver from 48% to more than 98%.

Jan 1, 1987

By W. Hu

Traditionally, coarse wolframite ore processing has been done almost exclusively by gravity concentration methods. However, there has been a depletion of rich and coarse ore reserves and a proportional increase in the mining of low-grade, fine-grained complex ore feeds. This change has resulted in the need for better slime-processing technology. Research and development activities in this area have gained momentum, and recent innovations appear promising. This paper covers the recent innovations that have been made in dealing with tungsten ore slime processing. The new techniques include the application of a centrifugal cone separator, high-intensity wet magnetic separator, the use of phosphonic and arsonic acids as collectors in ultrafine wolframite flotation, and combined processes.

Jan 1, 1989

By A. T. Yu

Bulk sampling is define&apos; as the ?process of extracting a small fraction of material from a large bulk ...sufficiently representative ...for the intended purpose.? (ref. 1). Today&apos;s pressing need for simple, effective and reliable sampling systems is intensified by the growing demand for quality control as a means to cope with inflation, competition and rapidly increasing consumption of raw materials. Much has been written on bulk sampling (ref. 2 thru 9). It involves a wide range of subjects cutting across the domains of physics, chemistry, geology, mining, materials handling and processing, as well as probability and statistics. Since the problem itself is complex, any in-depth treatment cannot avoid a somewhat lengthy coverage. Of all of the related topics. Arbitrary, haphazard studies, on the other hand, can leave out important details and obscure or confuse the issue.

Jan 1, 1970