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By C. P. Wa1ters

This paper is written primarily from the standpoint of the miner and consumer of Missouri fireclays in refractory products, principally firebrick used in the linings of blast furnaces, hot metal transfer cars, the recuperators of glass melting tanks, and other allied uses involving contact with liquid metals and reactive slags. For these uses the manufacturing plant requires forecast of certain physical characteristics, as well as chemical characteristics, of the raw clays as a determination of practicability of mining, treatment such as pre -firing or calcining, and final usage area. These methods and results may leave something to be desired from an academic view, but are used my operation as an on-running guide to seeking and mining raw material. The high quality fireclays of Missouri are generally differentiated by the user as the softer, somewhat plastic, "bond" clays and the so-called flint clays suitable for a fired or unfired aggregate portion of a brick composition. The semi-plastic "bond" clays are mined in a producing area of northeast Missouri including Montgomery, Audrain, Callaway, Monroe, and Lincoln counties. In this geographic area the clays are the members of the cheltenham formation and occur in rather large lenses in which the usable members may approach 15 meters in thickness and an area of 6o to 8o acres. Overburden is dependent on the surface topography and, in our operations, varies from 8 meters to 25 meters. It can be up to 36 meters in the prairie uplands, but this is rarely strip mined. There was former mining by shaft and drift methods in which the overlying coal was also worked. These clays are primarily used as a binding material for other clay aggregates and is characterized by its lower refractoriness-cone 3o (165o c) to cone 321/2 (172o c), in terms of pyrometric cone equivalent, and an alumina - Al2o3 - content of plus or minus 31%. Physically, these clays are readily slaked by water to a fine grain size and exhibit a degree of

Jan 1, 1977

By R. E. Muller, L. G. Martin, J. M. Stonehouse

Erratic distribution of values in free gold ore bodies presents severe problems in sorting material in large scale open pit mining. However, an intense program of modern blasting techniques and less than modern gold panning results in good sorting at Houston International's Manhattan mine. Blasting technique is important. If the sample location cannot be relocated because of blasted material movement, information acquired while drilling is useless. So short benches are critical to minimize dilution. Though not very accurate, panning allows for rapid analysis. The more accurate fire assay does not provide the needed speed. Panning can be fairly accurate and errors made are mostly on the low side. Samples of mixed and broken stockpile material are used as checks on pit procedure and to develop stockpile content in¬formation. Added tonnages discovered offset the higher labor costs of this careful sorting technique. The mine is located in central Nevada's Toquima Range in the small mining town of Manhattan, 322 km north of Las Vegas. The mine has four epithermal stockwork ore bodies that are mined in two pits. The mine problem was that conventional ore control was not separating ore from waste. Assays of 6-m blast-holes were used on 5-m centers. The blastholes were relocated by stadia survey after blasting. Holes assaying above the cutoff grade were flagged and mined as ore. This system is practical and useful in high concentration ore bodies or in evenly disseminated mineralization. But it was not practical at Manhattan because of large gold particles and small ore pocket sizes. Ore pockets were often missed when drilling a grid area for blasting. Also, large gold nuggets made assay results unreliable. The probability of a nugget being contained in an assay charge is small. So most results were low, compared to actual grades mined based on stockpile samples. When a nugget found its way into a fire assay charge, the result would be far higher than the actual grade of material represented. In addition, disturbance by blasting created problems in relocating ore. Because of these problems, the grade control program was intensified in September 1979. It successfully separated ore and waste in blocked out ore areas and located more ore in "waste" areas. Known reserves mined to date have increased up to 50% on some benches. The methods used are modern blasting techniques, detailed assaying, and traditional panning of samples collected from active mining faces. These methods allow mining flexibility, to match the variable mineral occurrence in ore bodies. The united effort of the geology and engineering departments has been important to the program's success.

Jan 12, 1980

By L. W. Saperstein, R. J. Sweigard, D. L. Simpson

Deep tillage is widely used as a means of alleviating excessive soil compaction. There has been some concern that the voids created by deep tillage may collapse before a more natural soil structure develops, resulting in recompaction subsequent to deep tillage A laboratory investigation was conducted to determine the feasibility of pneumatically injecting organic material into the voids produced during deep tillage in order to prevent any significant recompaction from occurring. Various amendments were injected into the soil during laboratory tests using a small-scale ripping and injection system. Measurements were made of selected soil proper Pies to evaluate the feasibility of this method.

Jan 1, 1994

By Paul D. Kovscek, Charles D. Taylor, Edward D. Thimons

When blowing ventilation is used on a mining section, a machine-mounted scrubber is often required to maintain face dust levels below 2 mg/m3. Operation of the scrubber also affects airflow patterns at the face and the dilution of liberated methane. Reductions in scrubber flow can result in increased methane concentrations. During mining, loading of the scrubber filter with dust is likely to reduce airflow. Detecting changes in scrubber flow is difficult, especially with extended cutting, because the machine must be operated from a remote location. Tests were conducted to demonstrate how scrubber airflow could be continuously monitored at a remote location. Three variables, static pressure at the fan inlet, differential pressure across the scrubber filter, and scrubber fan motor current, were continuously monitored while the scrubber filter was incrementally loaded to reduce air flow. The results show that monitoring each variable provided a way for estimating scrubber flow.

Jan 1, 1995

By Steve Kral

Philip M. Hocker is president of the Mineral Policy Center (MPC), a Washington, DC-based, non-profit organization formed to "address the environmental problems that are caused by mineral development in both the hard-rock mining and onshore oil arenas." Hocker has been involved with resource development and environmental policy issues since the 1970s. He has been a plaintiff in several lawsuits regarding environmental management of federal onshore oil and gas leasing. He was also active in the 1987 campaign to pass the Federal Onshore Oil and Gas Leasing Reform Act. Hocker served on the board of directors of the Sierra Club from 1980 through 1986. He is also a member of SME and spoke at the welcoming luncheon and keynote session of the SME Annual Meeting and Exhibit, Feb. 14, 1994, in Albuquerque, NM. In his role as president of MPC, Hocker has become a major player in the ongoing battle concerning mining law reform. And, while the mining law issue is the most visible and controversial issue confronting the mining industry and MPC, Hocker's organization is also active in clean water reform and other public policy issues that concern natural resource development and the environment. Hocker and MPC invoke a wide spectrum of reactions from the mining industry, most of it negative. Even so, the editors of ME decided that SME members should hear from Hocker his views on domestic mining, including mining law reform, clean water legislation and other environmental issues facing the industry.

Jan 1, 1995

By Arthur R. MacPherson

Este capítulo revisa las etapas que deben seguirse desde el momento de considerar la molienda autógena en el desarrollo de un nuevo depósito de mineral o en la expansión de una planta existente, hasta la compra de un molino autógeno. En este caso el término autógeno incluye: semi-autógeno, molinos de pebbles secundarios y molino de bolos (trozos), así como los sistemas de molienda autógena total. También se han incluido las alternativas por vía húmeda y seca. Antes de programar el ensayo debe decidirse cuantas muestras brutas son requeridas para calcular propiamente las posibilidades de molienda autógena del depósito. También ha de tomarse la decisión, de como y donde deben tomarse las muestras, por ejemplo: planta piloto situada en el sitio de localización de la planta o en otro lugar establecido; molienda en húmedo y/o en seco....etc. En el mismo programa del ensayo se obtendrán ciertos datos fundamentales. Los ensayos deberán incluir un trabajo suficiente. para determinar cuál de las moliendas, totalmente autógena o semi-autógena, debe utilizarse; y qué tipo produce la máxima eficacia en la potencia consumida. También, los otros datos pertinentes deben ser recolectados, como por ejemplo: potencias, tonelajes, alimentación, carga, tamaño del producto, carga circulante, etc. Si es necesaria una molienda secundaria para obtener el tamano final, la decisión de qué molienda debe emplearse (molino de pebbles o molino de bolas) puede requerirse un ensayo de trabajo adicional. Finalmente, la concentración del mineral puede jugar un papel importante en la selección del circuito y debería ser una parte del ensayo final una vez que se ha establecido la molienda. Para determinar qué tipo de molinos pueden satisfacer los requerimientos establecidos por la planta piloto, debe tomarse una escala para ciertos datos como: potencia del molino, consumo de acero, simple o doble etapa de molienda, tipo de molino secundario. Para asegurar un posterior entendimiento del depósito de mineral y sus características de molienda, deberán efectuarse algunos trabajos adicionales a pequena escala utilizando pequeñas muestras y molinos para producir unos datos comparables de molienda. También, algunos ensayos tales como medio de molienda óptimo, e índice de trabajo, deberán ser completados.

Jan 1, 1982

By Angelus S. C. Owili-Eger

The past decade has witnessed an unprecendented proliferation of groundwater computer models. Some of these numerical models were developed for the purpose of simulating groundwater movement in mining environments and predicting inflow rates into active mines. However, a general acceptance of these models as routine planning tools by the mining industry has been slow. Until recently, the major use of these models in the industry is by consultants. There are several reasons that have contributed to this situation. Some of these factors - include scarcity or non-existence of an appropriate data base information required for executing these models, the difficulty and costs involved in acquiring data, and the fact that the instructions for using these models are often too involved and time consuming. Today, however, the state-of-the-art technology has been advanced to a level where numerical models can be routinely used to simulate and predict inflow rates into active mining environments. Such environments include highly fractured (longwall) and deforming flow media. Features integrated into these comprehensive and yet easy to use models are double-porosity-pressure, multiphase, and non-darcian flow conditions. This paper reviews the role that computer modeling has played in mining-related operations by presenting several case studies and the state-of-the-art tools currently in use. Included will be some of the practical field applications that have been compiled during the past decade.

Jan 1, 1986

By Edward C. Dowling

Cleveland Cliffs and affiliated companies have established a commitment to the development and use of integrated process control. Substantial capital investments in process control, coupled with the dedication of human resources to effectively use the information, have generated meaningful business results. Improvements in productivity, operating costs and product quality have been realized. This paper presents an overall process control and information strategy, specific case studies, lessons learned and the current view of the next steps in the continued journey toward integrated process control and information management.

Jan 1, 2000

By Barry Silverman

As a significant mineral supplier, consumer of primary fuels, and producer of wastes and emissions to air, water, and land, the domestic copper industry through its investment and operating decisions influences the nation's economy and the environment. In turn, the copper industries' decisions are influenced by various factors: government energy and environmental policies, economic conditions, and technological considerations. The authors have developed a framework for assessing in an integrated manner the impact of these factors on the future investment and operating decisions made by copper firms. Specifically, this model is designed to capture the technology forcing nature of regulatory constraints, the differential situation of individual copper deposits and plant locations, and the price of factor inputs and conditions in the nation's economy. The capability to do this has only become real because of the decision of the U.S. Bureau of Mines to create Mineral Availability system1 or MAS. This system contains detailed but confidential information about each mine and mill in each of 35 mineral industries." Without MAS or a similar information system, the utilization of the methodology we offer would be impractical, but with. MAS it offers a unique technique of dealing with environmental-resources policy questions.

Jan 1, 1979

By C. J. Dungey

INTRODUCTION The County of Cornwall is situated in the South Western extremity of England, see Fig. 1, and has for hundreds of years been an historical centre of mining activity. In the last century many thousands of Cornish miners emigrated to Canada, U.S.A., Australia and numerous other countries throughout the world to form the nucleii of mining industries in those lands. At present there are four active underground mines within the county, all mining tin, together with a relatively large china clay industry. All the operations are directly or indirectly associated with the three main granite intrusions, namely the Land's End Granite, the Carnmenellis Granite and the Bodmin Moor Granite. The report which follows relates to two mines, owned by the South Crafty Group of Companies, namely the South Crofty Mine itself and its smaller subsidiary, Wheal Pendarves. Both these mines are adjacent to the Carn Brea Granite ridge, see Fig. 2, which is an off-shoot of the main Carnmenellis Granite mass. The Carn Brea Granite has provided one of the most intensive areas of copper and tin mineralisation anywhere in the world. Many of the mines worked during the last two centuries and the early part of this century were situated close to the metamorphic aureole surrounding the granite intrusion. Both South Crofty and Wheal Pendarves have their shafts collared in the overlying slate, locally known as killas, but the major part of their workings are in the granite beneath. The ore bodies, or lodes, of both mines are generally steeply dipping and approximately one to four metres in width. They have been, and continue to be, worked by either back or under-hand stoping with the major level interval being approximately 37 metres. A feature of both stoping systems is that they require no back filling and in many cases the completed working areas have remained open for many years due to the competent wall rock conditions. The stoping operation necessitates having box-holes or draw points at regular intervals along the strike length at 4 to 7 metres centres; thus a stope some 50 metres long may typically have ten box-holes or draw points through which air can percolate into the empty stopes above. The granite country rock contains traces of uranium minerals, namely pitchblende or uraninite, and it is the decay of this uranium mineral which leads to the formation of the mobile radon gas which, together with its short lived daughters, may create a hazard. The presence of uranium minerals has long been known, but the quantities being so minute, it was not appreciated until some ten to fifteen years ago that these might present a hazard. The Mines and Quarries Inspectorate approached Management approximately ten years ago with a view to taking samples in areas of the mine that they felt might contain the highest concentrations of radon. Samples were thus taken in several areas of both mines, normally well away from working areas and where there was no through ventilation. The results of these initial tests were published by the National Radiological Protection Board in a paper entitled "Radon and Its Daughters in Various British Mines". The results of the initial surveys made it apparent that improved ventilation was the most important factor in reducing radon daughter concentrations within the mine atmosphere, with increased flow rates having three beneficial functions, viz. (1) To dilute radon daughter concentrations. (2) To prevent radon and its daughters reaching a state of equilibrium. It is not uncommon for radon and its daughters to have a disequilibrium factor of 1,000 in some uranium mines.

Jan 1, 1981

By Peter K. Achireko, Samuel Frimpong

Design and optimization of open pit limits are of paramount importance because they provide information for evaluating the economic potential of a mineral deposit and for developing short and long range mine plans. Many algorithms and their modifications have been used to design and optimize open pit limits. These algorithms have provided mine planning engineers with pertinent information in designing, optimizing and extracting ore reserves by open pit technology. However, they do not address the random field properties associated with the ore grades and reserves and commodity prices, and thus, fail to yield the truly optimized pit limits in any time horizon. In this paper, a new algorithm, MCSIMFNN, which overcomes these limitations is proposed and used to optimize open pit limits. The random field properties of the ore grade and reserves have been modelled using the modified conditional simulation based on the best linear unbiased estimation and local average subdivision techniques. Artificial neural networks are used to classify the blocks into classes based on their conditioned values. The error back propagation algorithm, in the neural networks, is used to optimize the pit limits by minimizing the desired and actual outputs error in a multi- layer perceptron under the pit wall slope constraints. The optimized pit value obtained by this algorithm has been compared to that from the Lerchs-Grossman's algorithm in a case study. The results are the same, but in random multivariable states, the MCSI MFNN algorithm is the most suitable for pit optimization. The method is also fast since it mimics the operation of the brain in the solution of the pit optimization problem.

Jan 1, 1996

By Henry R. Tiedemann, Eugene H. Skinner, George E. Wickham

The need for ground support research in underground excavation has been well identified in recent years by many studies in the subject area (1, 2, 3, 4, 5). A typical conclusion from these reports (1, 2) is that the drill-and-blast (D & B) mining or the mechanical tunnel boring machine (TBM) methods are expected to be continued in use for the needs of tunneling in the immediate future. The Organization for Economic Co-operation and Development (OECD) sponsored in June, 1970 advisory conferences on Hardrock and Soft Ground Tunneling (3,4). These two reports were based on the combined response received from approximately six hundred inquiries from 17 nations in the summer of 1969 and which concerned all aspects of present-day inadequacies of, and the needed improvements in the technology of tunneling. Some specific problem areas in ground support and tunnel linings were: 1) the ability to successfully predict rock loads that must be carried by the support system; and 2) inadequate methods for determining the load-carrying capacity of the various support systems. Therefore, the problem area is clearly defined by these two objectives: 1) To find means of predicting rock loads; and 2) To develop methodology for determining safe, efficient, and economical ground support design practices. Noting these research needs from a great number of sources, the United States Bureau of Mines, Spokane Mining Research Center (SMRC), acting as agent for the Advanced Research Projects Agency (ARPA), undertook contract research effort in these problem areas. Work by Jacobs Associates has studied a methodology for specification of temporary artificial support based on the dependent sequence of mining operations and prediction of various rock conditions in the pre-excavation stage. The major thesis is that rock behavior is the

Jan 1, 1997

By S. G. Malghan

This paper reviews uses of sodium sulfide in the flotation of oxide minerals of copper, lead, and zinc. The activation and depression effects of sodium sulfide are of particular importance because of their applications with oxidized lead and copper ores. In spite of its industrial importance, the nature of the heterogeneous reactions between sulfide ions in solution and the surface of oxidized minerals has not been fully understood. Hence, proposed mechanisms of the sulfide action in flotation, methods of measuring and monitoring sodium sulfide in flotation pulps, and variables influencing the sulfide addition are analyzed. Recent studies on sodium sulfide addition to flotation pulps, monitoring its concentration, and the optimum dosage control using platinum and/or sulfide specific ion electrodes are discussed.

Jan 1, 1987

By B. J. Arnold, F. F. Aplan

A model incorporating rate and residence time to describe the froth flotation process was used to provide data for the evaluation of various kinetic flotation models and to provide scale-up information from laboratory to industrial scale coal froth flotation. Rate constants were determined in laboratory and plant froth flotation tests and the residence time distribution was measured for a bank of four 21 ft 3 (0.6 m3 ) WEMCO cells treating Upper Freeport seam coal and a bank of four 300 ft 3 (8.5 m3 ) cells treating Lower Kittanning seam coal. It was found that a lag time for continuous flotation should be incorporated into scale up models and that a kinetic froth factor ratio of between about 2.0 and 2.8 should also be used to scale up data from laboratory to plant scale.

Jan 1, 1989

By Thomas Fenner

Surface and underground mining operations are often hindered by unknown tunnels and slopes and the precise locations of adjacent workings. Mine safety and planning are adversely affected especially in previously undermined areas where historical records are incomplete or nonexistent. The results of several ground penetrating radar (GPR) studies for the detection of known mine tunnels in underground coal, salt and gold mines are presented. A case study to locate unknown, abandoned mine workings at a surface gold mine successfully located all workings to a depth of 11 m (36 ft).

Jan 1, 1996

By Earl Cook

Introduction Men like physical power because it gives control over nature and over other men. Physical power comes in two common forms: explosive and non-explosive. In both forms, the most powerful devices yet invented make use of nuclear energy: on the one hand, the hydrogen or fusion bomb, and, on the other hand, a nuclear power plant. For some nations, nuclear power offers a measure of independence not otherwise available in a time of diminishing and unevenly allocated fossil fuels. Consequently, it has great attraction. Unfortunately, the application of nuclear energy (especially fission energy) to either explosive or electrical power entails a singular risk, that of injury from ionizing radiation, both natural and induced; this risk is found in all stages of the nuclear-fuel cycle, from mining of fissionable materials to the disposal of wastes produced in the "burning" of nuclear fuels. The health costs of nuclear power were not fully assessed in the early stages of the development of nuclear technology and some of them were not properly guarded against, either in the USA, which led the world into the nuclear age, or in Canada, which followed an independent path. What I shall do in this paper is to sketch the history of nuclear regulation in our two countries, to point out what I consider similarities and differences, and to suggest that regulatory history, as perceived by a concerned public, has a great deal to do with what is possible today and probable tomorrow, in other words, with public acceptance or rejection of the nuclear option. The U.S. situation As we sow, so shall we reap. In the USA, further expansion of the nuclearpower industry has been almost stymied by government actions that reflect mounting public concern. There is no national storage facility for spent reactor fuel and it has been decided not to reprocess it; consequently, existing power plants are about to choke on their own wastes. Since the Three Mile Island accident, no new nuclear plants have received operating licenses, although several are ready to start up. The uncertainty of future government action has caused the abandonment of plans for a number of new nuclear plants. In addition, many research projects related to medicine and agriculture, which produce low-level radioactive wastes, are about to be shut down for lack of licensed disposal facilities. Some would maintain that this situation has come about because of the low benefit-to-cost

Jan 1, 1980

By S. K. Kawatra

For efficient control of particle size in a wet grinding circuit, three alternative methods for monitoring pro- duct particle size are available. These methods are based on: 1. Ultrasonic wave absorption. 2. Low-angle forward scattering of a laser beam. 3. Inferential sizing techniques. The relative advantages and disadvantages of each technique are discussed. A particle size monitor can be used on a single grinding line to control the product particle size. To minimize the number of particle size monitors at an operation consisting of many similar grinding lines, a portable multi-stream on-line analyzer can be used. The paper outlines the design and the operational characteristics of the portable multi-stream online analyzer.

Jan 1, 1979

By Xi Huang

The coal workers' pneumoconiosis (CWP) has been considered as the most serious health problem in the coal mining industry. The development of CWP in a worker lung depends on the average concentration and toxic components of the coal dust and the exposure length. Toxicity studies on coal dust shows the bio-available iron (BAI) in the coal dust plays an adverse role in the CWP development. It is also found that presence of calcite in the coal seams can significantly reduce the CWP cases as evidenced by the significant difference between the coal miners in the western and Appalachian coal regions. Based on this finding, it is proposed to artificially introduce calcite to BAI -containing coals in the effort to mitigate the CWP disease in coal industry. This paper introduces the scientific bases, the main geological factor that makes the differences, the proposed methods to introduce calcite to coal dust or to coal seams. Feasibility for the proposed methods is also discussed.

Jan 1, 2013

By K. J. Henderson

INTRODUCTION Shrinkage mining at Inco Ltd.'s Crean Hill mine can best be defined as a horizontal breasting method utilizing the broken ore for the mining floor and to provide wall support, with draw from a series of boxholes or draw¬points at the base of the stoping block. In choosing to shrink a given block of ore, the initial investigation centers on the ore configuration and grade, the expected ground conditions in the stoping area, the expected profit compared to alternative mining methods, the future production requirements, development lead time, the availability and type of equipment in the area, and the availability of experienced miners. These factors are weighed against other mining methods and shrinkage mining is chosen when the advantages outweigh the dis¬advantages. These will be discussed in more detail later. At Inco Ltd.'s Crean Hill mine, shrinkage mining makes up 17% of total mine production and is decreasing due to the advent of advanced blasthole techniques, notably in-the-hole drilling. Shrinkage is an adaptable mining method that is currently employed in narrow irregular ore zones, as undercuts for in-the-hole blast¬hole stopes, in ore bodies that do not extend from level to level, and in areas where quick production is required with a minimum of capital outlay or minimum develop¬ment. Two major types are currently in use which are classified on the basis of type of draw system employed. These are the utilization of load-haul-dump (LHD) equipment drawing from drawpoints, and the slusher and train method drawing from boxholes into a slusher trench and loading into a train. A third method being phased out is drawing from chutes directly into a train. In all three methods the actual stoping operation remains the same. In the initial planning of a mining block, the ore body is first located using an exploration diamond drill¬ing program. This is essentially done at a 130-m (400¬ft) spacing and it gives an indication of the amount of ore, grade of ore, the configuration of the ore body, as well as other geological features such as ground condi¬tions which will affect future mining. This drilling is done from an exploration drift 165 m (500 ft) long in the hanging wall and covers a vertical extent of 330 to 670 m (1000 to 2000 ft). However, these data are not accurate enough for detailed planning. When the de¬cision has been made to mine, a preliminary develop¬ment drill drift is driven in such a position that it can serve the dual purposes of allowing the ore body to be drilled off in detail, or of later serving as a haulage or access drift. Diamond drilling from this drift is usually on a 15-m (50-ft) grid spacing and gives adequate information for detailed planning. Utilizing the drill data, a decision is made on the mining type. For a shrinkage operation, the stoping block is ideally 49 m (150 ft) in length and a maximum of 13 m (40 ft) in width. Lengths of 65 m (200 ft) are considered a practical maximum because of the length of time the back of the stope remains open during the min¬ing of a cut. A minimum length of 13 m (40 ft) is used because the amount of time spent in opening a new cut increases in comparison with the total stoping time. The maximum width of 13 m (40 ft) is used because ex¬perience with the ground indicates that generally stope backs tend to be unstable over greater widths. However, widths of 33 m (100 ft) have been mined. A minimum width of 5 m (15 ft) is used to allow an effective draw, but in unusual cases, a width of 3 m (10 ft) has been mined successfully. The dip of the ore should not be less than 1.04 rad (60°). Angles flatter than this will not pull properly resulting in inefficient mining because of ore hang-ups on the footwall side of the stope and the need for large stagings on the hanging wall. In considering shrinkage, mining ground conditions play an important role. Strong shearing, cross faulting, or dikes are avoided. These usually cause dangerous mining conditions and an alternative safer mining method which will give adequate support to these adverse ground conditions is employed. Production requirements are taken into considera¬tion when planning a shrinkage mining program. In it¬self, shrinkage mining is capable of producing ore in a relatively short time with little capital development. If quick production is required, this method can produce ore rapidly as all that is required is that the draw system be completed and the miners with jacklegs and stopers produce a limited amount of ore. The obvious disadvan¬tage is that not all the ore broken can be pulled from the stope immediately. However, a mining horizon with sev¬eral shrinkage stopes in different stages of production can adequately maintain scheduled tonnage require¬ments, the cycle being stopes under development, stopes actively mining, and stopes completed and pulling. When the decision has been made to use shrinkage mining in a given block, the haulage and draw system is driven. The proximity of the orepass system dictates whether the system is to be a LHD operation or a slusher trench operation. If the orepass system is in close prox¬imity to the stoping block, LHD equipment is employed because of its ability to operate efficiently over short haulage distances. A long tram to the orepass usually dictates a slusher trench operation with haulage by train. Internal orepass systems are used effectively with LHD equipment with a load-out and retramming taking place on a lower horizon. In addition, stoping blocks in older areas of the mine are developed track because of the capital cost involved in converting an existing level to trackless mining. DRIFT DEVELOPMENT USING LHD EQUIPMENT The haulage drift for a draw system using LHD equipment is driven parallel to the ore body in the foot¬wall at a sufficient distance to allow an 11.5-m (35-ft) drawpoint to be driven into the ore (Fig. 1). The actual position of the drift depends entirely upon the ore con¬figuration and the drift size depends on the equipment to be used. Drift sizes of 4.6 x 4.1 m (14 x 121/2 ft) for an St-4 and 3.8 x 3.6 m (111/2 x 11 ft) for an St-2 meet company and government clearance standards and allow sufficient space to accommodate a ventilation pipe. The

Jan 1, 1982

By T. Shungu, N. Vermaut

The various copper-cobalt ores mined by Gecamines, together with typical plant applications and results, are briefly described. Intensive laboratory investigations have been undertaken in recent years to improve the metallurgical performances of the various concentrators. As a result of these tests, it was shown that the flotation of the siliceous oxide ores could be improved by increasing the efficiency of the palm oil collector (energetic conditioning, chemical emulsifier, petroleum sulfonate addition) and by varying the froth depth in the cells. A new flowsheet incorporating these modifications and a regrinding stage has been approved for the rehabilitation of the Kolwezi concentrator. For what affects dolomitic oxide ores, hydroxamates were tested as secondary collectors and, based on successful laboratory results, plant trials were run but failed due to excessive frothing. In other plant tests, xanthates were replaced by a xanthate- fatty acid mixture; plant trials over more than one year proved successful for copper, but not for cobalt. Mercaptans were tested as secondary collectors for the sulfides flotation from sulfide or mixed oxide-sulfide ores; laboratory results proved successful but, to date, no plant test has been run.

Jan 1, 1987