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By M. C. Roberts

The consumption of industrial minerals, like most other commodities produced from the earth, depends on the demand for the goods and services that use these materials. Minerals are not generally used directly by final consumers (table salt is a notable exception) but instead are used to either make products that are then purchased by customers, or the minerals are used to make industrial products that eventually contribute towards producing other goods and services. Therefore the demand for industrial minerals in most uses is derived from the ultimate demand for goods and services by consumers. The problem for producers of industrial minerals is that the demand for their mineral product fluctuates greatly in response to the changes in the production of goods and services over the course of the business cycle. In this respect, the producers of industrial mineral are innocent victims, subject to market fluctuations outside of their control. Because the demand for most minerals depends on the demand for the product and because minerals typically make up a minor portion of the costs of producing the product, changes in the price of the mineral often do not influence demand for it to a significant degree in the short run. This is described as a very low price elasticity of demand, as shown by the steep slope of the hypothetical demand curve in Figure 1. This demand curve shows that changing the price of the mineral does not alter the quantity demanded very much, but shifts in the demand curve will be the important determinant of the quantity of the mine.ra1 that is sold. The shifts are caused by the changing demand for the products using the mineral over the course of the business cycle.

Jan 1, 1994

By Ernst Langer

INTRODUCTION Morro do Niquel is located near the small town of Pratdáolis in the southwest part of the State of Minas Gerais, Brazil, about 300 km in a straight line north of São Paulo and 300 krn west of Belo Horizonte. The deposit is known since 1935. Systematic exploration started, however, only in 1957 and production with one furnace in 1962. Since 1969, after the installation of a second furnace, the mine produces about 2 500 tons of nickel, contained in ferronickel, per year. This production is now consumed totally by the Brazilian steel industry. The "Morro do Niquel" is a serpentinite stock of 900 by 400 m which rises 200 m high above the surrounding metamorphic rocks of Algokian age (see Figure 1). On top of the hill there is a leached barren zone of about 200 by 400 m, generally 20-30 m deep, which consists practically only of a boxwork of quartz colored brown by limonite. Surrounding this zone on the sides of the hill is weathered, decomposed serpentinite which contains about 213 of the economic nickel mineralization. Roughly 113 of the ore reserves lie below the leached overburden representing ore of a different type. The geological map and cross-section, Figure 2, give a clear picture of the occurrence.

Jan 1, 1979

By P. Albert Carpenter, Wallace B. Howe, Michael N. Greeley, E. C. Bingler, Gilbert R. Eakins, Christy Bell, John Callahan, Gary B. Glass, T. J. Crabs

Although gold's market price dropped considerably from 1980's all time high, exploration for the precious metal remained a particularly bright spot last year, especially in the western US. For example, in Nevada-the nation's largest gold producer-further exploratory drilling at Louisiana Land & Explorations Co.'s deposit at Round Mountain and Newmont Mining Corp.'s find near Carlin resulted in sharp boosts in estimated reserves at both sites. Louisiana Land said its deposit is well on the way to being one of the largest gold reserves in the US. Based on further drilling, Newmont said its Gold Quarry property near Carlin could now produce 4.7-6.2 t/a (150,000-200,000 oz/ year). Reserves at Newmont's two properties near Carlin and a third property in southeast California are reportedly at least 218 t (7 million oz), up from the previously announced 128 t (4.1 million oz). Gold exploration activities were brisk in many other states, including a noticeable effort to explore and reopen dormant properties. Although the return on investments didn't appear as favorable last year as they did in 1980, the gold price was apparently sufficient to spur healthy exploration activity. Outside precious metals, market conditions, and subsequent exploration activity, appeared much more bleak. In the molybdenum market, for example, prices continued to slide, and Amax Inc. tightened its exploration budget. (Amax supplies 60% of the domestic market and 45% of the world demand for molybdenum.) The company spent 16% less on exploration during the first six months in 1981, compared to the same period in 1980. Copper and many other base metal producers appeared in similar straits throughout the year. Uranium exploration activities continued to suffer last year also due to depressed market conditions. Among others, Gulf Minerals Resources and Amoco Minerals Co. suspended uranium exploration activities in the West. Despite current unfavorable conditions, some companies continue uranium searches, confident that market conditions will rebound. On a brighter note, interest in exploring for some industrial and strategic minerals appeared on the upswing last year. Minerals used for oil and gas drilling, such as barite and special sands, were especially noteworthy. In geophysical activity, significant developments were achieved for electromagnetic techniques in both frequency-domain and time-domain systems. Geophysical logging activity declined due to decreased uranium exploration, and computer hardware continued to impact the way mining geophysical data are being acquired. Many presentations and papers were given last year to help exploration geochemists develop new techniques to locate concealed ore bodies. Among techniques described were statistical and computer modeling for enhancing anomalies over blind deposits.

Jan 5, 1982

By Gerard P. Hand

The use of graphite in industrial applications is one of the oldest in the world. Yet, it is one of the fastest-changing technologies today. Graphite was first used centuries ago in developing writing instruments and as a raw material in crucibles used in melting metal. Today, graphite is used in some of our most demanding industries. These include aerospace, electronics and automotive applications. Graphite usage has historically been split between the natural and synthetic forms. Figure 1 shows the breakdown of natural graphites and synthetic graphites used. The usage of natural graphites and synthetic graphites is split fairly evenly, although the applications are quite different. Figure 2 shows the breakdown of applications using natural graphite. The largest application is the refractory industry. This includes refractory brick and plastics, crucibles, shrouds and other refractory parts. This main application uses roughly 50% of all natural graphite. The refractory industry is an important application for natural graphite. However, the refractory markets have been struggling for several years. As demonstrated in Fig. 3 (Semler, 1996), the production of refractories in tons per year has declined by more than 32% during the last 25 years. Refractory prices have actually risen on a dollar per ton basis during that same period as the product mix has shifted to the value-added side. Raw material and production cost increases, inflation and other related costs have kept profitability at low levels for the industry. Furnace linings that once lasted 1,000-2,000 heats now last 8,000-12,000 heats. This means that tonnage requirements have gone down and quality requirements have gone up for graphite suppliers. Therefore, new opportunities and applications must be developed if the graphite industry is to grow. Synthetics are the other major forms of graphite. There are two forms of synthetic material currently used. One is the traditional type. It is made from the turnings of graphite electrode production. The second was developed in the mid-1970s in response to a shortage of conventional synthetic graphite. This occurred by a breakthrough in thermal technology. This new technology purifies, desulfurizes and graphitizes petroleum coke on a continuous basis at temperatures of more than 2,500° C (4,532° F). This product is called desulfurized petroleum coke (Desulco). It now accounts for about half of the synthetic type material used in the United States. Figure 4 shows the breakdown in the use of synthetic types of material by application. The largest requirement is for use as a carbon addition in iron and steel. This is followed by alkaline batteries and friction materials. Current technological opportunities New opportunities for graphite products are now being developed. These new areas are made possible due to the advances in graphite thermal technology. The ability to purify and modify graphite and carbon products is the key to future growth in the graphite industry. New, higher-purity, more consistent graphite products are leading to enhanced performance in current applications. They are also opening the door to new applications.

Jan 1, 1997

By J. M. Keane

Critical to successful dump and heap leaching for copper is an adequate testing program for evaluation of the applicable parameters. This paper deals with testing methods to establish best conditions for dump and heap leaching of copper so that commercial operations can proceed on a rational basis. Discussion is also offered of ore occurrences suitable for dump and heap leaching.

Jan 1, 1985

By K. J. Sollman

With an ongoing concern for safety and health in the workplace, NYCO® requested three medical evaluations be undertaken to study possible health hazards of wollastonite. Two of the evaluations were conducted by the National Institute of Occupational Safety and Health (NIOSH). One was in 1976 (Ref. 1), and the second in 1982 (Ref. 2). The aim of the two studies was to answer the following two questions: 1. Do wollastonite workers exhibit more respiratory abnormalities than workers not exposed to wollastonite? 2. Is there a relationship between prevalence of respiratory abnormalities and accumulated expo-sure time to wollastonite?

Jan 1, 1988

By W. A. Haley

The use of diesel engines in underground hard-rock mines dates back to the late 1940s. For the first several years, they were used only occasionally, being limited to a few metal mines that experimented with crawler¬mounted front-end loaders, tractor-trailer hauling units, a few tractors for drill-compressor mounts, and utility cleanup machines. By the mid-1950s, track loaders had become commonplace in limestone mines and uranium mines on the Colorado Plateau in the United States, as well as in Canada. Use of crawler-mounted tractors as drill and compressor mounts also increased. By the end of the 1950s, rubber-tired loaders and some haulers began to replace the track-type machines and rail-mounted cars that had been in use. About 1960, the rubber-tired machines brought about a new era of underground mining mobility and flexibility, centered on a method commonly known as "trackless mining." Ultimately, many of the underground rail-type systems for loading and hauling were replaced by the trackless mining technique. ECONOMIC CONSIDERATIONS The size and nature of mineral deposits, plus ground control techniques, historically had dictated small open¬ings to the surface from many underground mines. The small mine openings led to the development of special rubber-tired loaders and haulers designed specifically for access through the small openings. However, some mines, particularly those in massive mineral deposits, are able to excavate and maintain very large openings, and some use modified room-and-pillar systems. With the large mine openings, the use of larger, more produc¬tive equipment such as that commonly found in surface mining becomes economical. In fact, productivity gen¬erally increases at a more pronounced rate than machine size increases because many of the larger machines were designed for heavy-duty shot-rock applications in surface mines and construction sites where the handling of blasted rock is common. Table 1 can be used as a very Table 1. General Productivity Comparison for Conventional Machines In Underground Use (Shot-Rock Conditions) 2.3 m3 (3 cu yd) 4.6 m3 (6 cu yd) Loader Loader Expected Surface 230 t/h 540 t/h Production (250 stph) (600 stph) Expected Underground 90 t/h 270 t/h Production (100 stph) (300 stph) Expected Total Efficiency: Surface 40%-60% 50%75% Underground 25%.-40% 30%50% Expected Useful Machine 8000 hr 12,000 hr Life Before Replacement general comparison of the production and efficiency between small and large machines. Combining greater productivity often inherent in larger machines, with reduced downtime resulting from using fully developed machines with fast parts and service backup, some mine operators have been able to reduce material handling costs appreciably while reduc¬ing manpower requirements for operators and main¬tenance men. Large mine openings increase the amount of rock that must be handled in the development work, and they sometimes increase the dilution in stopes or rooms, de¬pending upon the dimensions of the ore zone. Providing adequate space for the unrestricted operation of large surface mining machines could, therefore, lead to more waste segregation and handling costs. It could also cause greater ore dilution that would result in a lower grade of ore being delivered to the processing plant. The tradeoffs between opposing cost factors must be reconciled and balanced to achieve the best overall cost of the crude ore, concentrates, or product. EQUIPMENT MODIFICATIONS Loaders and haulers designed for surface mining are seldom used underground in their standard con¬figurations without some modifications. If done, the modifications generally are made by the equipment dealer and/or the user, and the modifications usually include one or more of the following items: 1) The exhaust stack is lowered, and its direction is changed. Usually, it is repositioned horizontally to the rear, or it is fed into the engine fan to diffuse the exhaust gases. 2) The operator's position is lowered by either lowering the seat or changing the seat to a side mount. 3) The operator controls are adjusted to fit the new operator position. 4) Other components, such as the radiator and loader tower, are lowered. 5) Special bumper guards are mounted at the base of the radiator area. 6) An exhaust conditioner is mounted and con¬nected, using either a catalytic or a water-type condi¬tioner, or both. This usually is controlled by the safety and health regulatory authority having jurisdiction. 7) The positions of other components are rearranged

Jan 1, 1982

By Egon Pohl, Friedrich Steinhäusler, Werner Hofmann

INTRODUCTION Anticipated increase of nuclear fuel production in the future coincides with growing concern about the occupational health risk of miners from inhaled radon decay products. As a consequence it has been suggested to even lower presently permitted exposure levels (NIOSH, 1980) and implement more stringent control on measurement programmes. Compliance with these regulations requires large investments in new or modified ventilation systems as well as in increased expenditure for staff and instrumentation for health physics operational monitoring. Since both costs are directly related to the overall ore production costs, this can have far reaching implications with regards to the economic feasibility of certain mining operations and consequently reduce estimates of low-cost minable ore reserves. In addition there is increasing evidence that the radon problem is not limited to the nuclear fuel cycle only, but can also represent a significant hazard to non-uranium miners. A key component in the cost-effective implementation of legislative control measures is the monitoring system employed. The choice of system is decisive for the total costs of installation and maintenance, manpower requirements and accuracy of nuclide determination. In the following operational specifications are defined for monitors in mining and milling environments. Different types of available instrumentation are discussed with regard to their suitability for practical radiation protection in underground mines, open cut mines and mills. ATMOSPHERIC CHARACTERISTICS Underground mines Radon in the air of underground mining operations is exhaled from surrounding rock surfaces, crushed material and, to a lesser extent, from water seepage. Whilst in uranium mines radon releasing ore bodies are generally localized in distinct areas, radon sources in non-uranium mines can be very dispersed throughout the system of mine tunnels. The ventilation scheme used influences the absolute atmospheric level of radon as well as the equilibrium conditions between radon and its decay products (factor F). In uranium mines mechanical forced-air ventilation is normally the only way to achieve and maintain legally required nuclide levels. This causes the F-factor to be rather low, e.g. in French CFA-mines F is of the order of 0.2 (Francois, Pradel, Zettwoog, 1975). At the same time the fraction of unattached radon decay products (fp) can increase due to the high air velocities employed. However, it is possible to find non-uranium mines with either natural draught ventilation only or assisted on demand by forced air ventilation during special operations or climatic conditions. Thereby the F-factor is more dependent on seasonal changes of temperature differences between outdoors and mine atmosphere and work routines. As a result F can cover a wide range from 0.02 to 0.95 (Steinhäusler, 1976). The use of filters or electro-precipitators in mine ventilation systems can modify the atmospheric characteristics twofold as it generally decreases the content of radon decay products, but at the same time increases the content of the unattached fraction fp . Average concentration levels of radon decay products are mostly lower in mechanically ventilated non-uranium mines than in equally ventilated uranium mines and are below 0.3 Working Level (UNSCEAR, 1977). However, some working places in non-uranium mines, specially with only natural draught ventilation can occassionally approach maximum permissible levels as defined for uranium mines (Strong, Laidlaw, O&apos;Riordan, 1975; Snihs, 1976; Sciocchetti, Scacco, Clemente, 1981). Open pit- and surface mines Radionuclide levels of radon decay products in the atmosphere of these mines are mostly too low to represent a significant inhalation hazard for miners, ranging typically from 0.03 to 0.1 Working Level (Steinhäusler, 1976). However, personnel using airpurifying respirators or working in cabins ventilated with filtered air can be exposed to a radon atmosphere with low value for the F-factor (F [<] 0.1) and high fp-value up to 80 % (Leach and Lokan, 1979). Mills Atmospheric radon concentration in crushing, grinding, drying and packing sections depends on the radium 226 content of the ore, ore storage methods and ventilation system employed. Providing adequate ventilation ([>] 2 air changes per hour) and control of dust production radon and its decay products represent only a minor problem (Saconney, 1979). MONITORING OF OCCUPATIONAL EXPOSURE Objectives Operational monitoring of the working places provides information on: - confirmation of appropriate control of the routine mining methods employed - indication of abnormal conditions. Although this type of monitoring enables the location

Jan 1, 1981

By John E. Auten

The computer program "Spike" is an "expert-rule based system" designed to fulfill two requirements. 1) acquaint unfamiliar personnel with big hole drilling techniques and 2) to stimulate the "old hands" in solving drilling problems. "Spike" will not make anyone a drilling expert The best way to understand a computer program is by example, but first some explanation of an? "expert-rule based system", As the name implies, the program tries to imitate a human expert. Unlike a normal computer program which uses a list of instructions to follow in a set order, a rule-base program has a collection of independent rules without an explicit list of things to do, one after the other. This type of program was mandatory to accommodate the ever changing situations and corrective actions that more accurately duplicate actual drilling operations and the advise of an experienced drilling expert.

Jan 1, 1988

By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By L. P. Meade

The particular impact of imported industrial minerals varies within the Industry and is dependent upon many diverse factors. In a similar way, the impact on individuals has been varied and the degree of impact will depend partially on the individual&apos;s ability to handle change. Just as many corporations are adjusting their marketing strategy and production methods to compensate for imported mineral products; individual managers, mining engineers and geologists are making adjustments to their own personal job marketing techniques and on-the-job operating procedures. Modifying the degree of impact for both industry and individuals depends upon the ability to adapt.

Jan 1, 1988

By D. A. Dahlstrom

Crushing and grinding consumes about 1.3% of the nations produced electrical energy.( Furthermore, only one in the first ten highest consumers comes from an industry other than mineral processing. This is the grinding of wheat to make flour. The other nine are all in the metallurgical industry and are easily the largest consumers.

Jan 1, 1988

By C. Williard

Computer based mine monitoring and control systems represent a relatively new technology to be applied in the U.S. mining industry. While computer assisted engineering, design, and planning have been accepted as tools for specific technical and engineering jobs within the industry, integrating mine-wide monitoring and control systems into mining operations has been slow. This paper reviews the state of the art of mine monitoring and control systems and provides insights into future trends and developments.

Jan 1, 1988

By R. O. Burt

The Tanco pegmatite is located in the Bird River greenstone belt of south-eastern Manitoba approximately 160 km northeast of Winnipeg (Figure 1). The orebody was discovered in 1929 and subsequently attempts were made to exploit it for tin by Jack Nutt Mines Ltd. and later in the middle 1950&apos;s by Montgary Oil Ltd., who extensively developed but did not produce from its lithium deposits. In 1966 systematic investigation of the pegmatite as a potential source of tantalum led to the founding of Tantalum Mining Corporation of Canada Ltd. (TANCO), and production of tantalum concentrates from the Bernie Lake mine commenced in 1969. Tanco&apos;s currently jointly owned by Hudson Bay Mining and Smelting Co., Cabot Corporation, and the Province of Manitoba.

Jan 1, 1988

By C. E. Jordan

The Bureau of Mines conducted research to determine the effect of fine bubbles on the flotation of fine-sized particles (minus 20-um size). Turbulent flow collision models for fine particles and fine bubbles were reviewed. Velocity distributions within the flotation cell along with measured bubble and particle size distributions were considered in the models to predict the flotation response. The models were corroborated with laboratory fine-bubble flotation tests to verify the expected increase in flotation efficiency.

Jan 1, 1988

By J. Bartrum

Semi-autogenous, (SAG), milling operations can realize substantial Increases in mill throughput by removing and crushing critical size, slow grinding, material from the SAG mill. Full-scale plant testwork showed that modifying the Kidston SAG/ball circuit to a semi-autogenous-ball milling-crushing (SABC) circuit would increase plant throughput by 25 to 30%. The methods used to determine the value of the SABC circuit and the design of the circuit are discussed.

Jan 1, 1988

By R. J. Smith

Hornestake&apos;s Mclaughlin Mine is the first commercial application of pressure oxidation pretreatment of gold prior to cyanidation. Inherent to the autoclave process are several liquid/solid separations. Flocculant consumption represents a significant portion of the consumable reagent costs at McLaughlan. This paper focuses on the efforts of the Operating, Metallurgy and Engineering groups to improve thickener operation and minimize flocculant consumption. An operating philosophy incorporating feed solids control, increasing flocculant dilution water and minimizing flocculant solution shear has resulted in a reduction in 1986 flocculant costs of $0.054/t ($0.06 st) versus 1985 costs. Additional changes are projected to result in a further cost reduction of $0.20/t ($0.22/.st) in 1987.

Jan 1, 1988

By W. K. Fletcher

As part of an ongoing programme on behaviour of heavy minerals In streams, the dispersion of Au as free gold (Day and Fletcher, 1987), Sn in cassiterite (Fletcher et al, 1987) and W in scheelite (Saxby and Fletcher, 1987) have been investigated downstream of mineralization. The studies of Au and W were undertaken, in British Columbia and the Yukon Territory, respectively, in streams that are strongly influenced by spring snowmelt floods. Sn was investigated under tropical rainforest conditions in Peninnsular Malaysia.

Jan 1, 1988

By S. Horton, F. L. Stanley, P. J. Bethell

Introduction The McClure River Processing Plant is located at McClure in Dickenson County, VA. The plant was originally constructed in 1979 as a 600 st per hour facility. The plant is fed by coals from numerous mines in the area as well as a major longwall mine onsite. Coal is processed predominantly for the metallurgical coal market. The original plant was comprised of three separate processing circuits incorporating a heavy medium vessel for the coarse material (+10 mm). Intermediate sized material (10 x 0.6 mm) was beneficiated in heavy medium cyclones, with the remaining fines (0.6 mm x 0) passing to froth flotation for upgrading. The products from all three circuits were combined prior to thermal drying and then stored in silos or ground storage prior to shipment to the metallurgical coal markets. A middlings heavy medium cyclone circuit was also present to rewash the vessel and primary heavy medium cyclone rejects. The original circuit is as shown as Fig. 1 (Editor&apos;s note: All figures are reproduced at the end of this paper.) In reasonably close proximity to the McClure Plant is the much older, less efficient Moss 1 Plant, (Chance cone, table, water only cyclone, froth flotation, Fig. 2), which prior to the McClure Plant expansion was processing considerable tonnages of coal. At that time, the McClure River Plant was processing coal at the maximum possible feed rate, and the Moss 1 Plant was finding increasing difficulty in maintaining the stringent quality specifications on its major metallurgical coal accounts. Given the fact that many of the feeder coals to Moss 1 could be moved with minimal difficulty and cost to McClure, an exercise was conducted to evaluate the potential of expanding the McClure Plant. This would result in transferring coals from Moss 1, thereby increasing yield, reducing operating costs, and enabling customer specifications to be met more easily. After evaluating many possible ways of expanding the McClure Plant, we eventually settled on an approach based on the idea of Gallagher at German Creek in Australia (Gallagher, 1986). Our objective was to increase plant capacity by removing the coarser size fractions (0.60 x 0.15 mm) from froth flotation and at the same time eliminating the finer fractions (1.25 x 0.6 mm) from the heavy medium cyclone circuit, thereby allowing these circuits and therefore the plant as a whole to increase its capacity. These size fractions (0.60 x 0.15 mm from froth flotation and 1.25 x 0.6 mm from the heavy medium cyclone circuit) would be treated in a new spiral circuit. After desired evaluation of the various circuit capacities, including desliming screen, clean coal filter and thermal dryer capacities, etc., it was concluded that we should be able to maintain high levels of processing efficiency in the existing circuitry while at the same time expanding plant feed rate from a nominal 600 st per hour to 1000 st per hour, thereby achieving the objectives The circuit that was finally settled on is shown in Fig. 2. Several unique features were incorporated in this circuit in an attempt to maximize processing efficiency and reduce maintenance costs. First, the idea of a single-stage spiral circuit was abandoned in favor of a two-stage circuit incorporating rewashing the primary spiral middlings in a secondary spiral (Bethell, 88). The spirals selected were Mineral Deposits Model LD4 units. It was felt, after much previous testing, that spiral middlings rather than being a true middling gravity fraction were in fact a combination of true middlings with a preponderance of misplaced light coal and high-density rock. Consequently the insertion of this material directly into either clean coal or refuse streams (per normal practice) would only lead to loss of efficiency. The operating philosophy behind the design of the circuit was to produce the purest coal and rejects possible by adjusting the spiral splitters. Misplacement of refuse into the clean coal and vice versa would be minimized by taking a fairly large middlings cut. By reprocessing this material (10%-15% of primary spiral feed) in a secondary bank of spirals the misplaced material in the primary middlings would be reprocessed and its probability of reporting to the correct stream greatly enhanced. To minimize fines loading on the spirals, fine aperture sieve bends were to be included. A further novel feature of the circuitry is the desliming/ dewatering system. A philosophy of accepting inefficiency but forcing it to work in our favor was developed. The dewatering screens (Linatex Combis) were chosen not for their ultimate screening efficiency but for their ability to remove slimes (tolerating a fair quantity of oversize in the effluent) and their enhanced maintenance characteristics. Any oversize passing into the effluent, together with the fine coal centrifuge effluent, would pass to classifying cyclones, the objective of which was to reject to overflow (and hence froth flotation) as many fines as possible (-150 µm) without running the risk of losing coarse (1.25 x 0.15 mm) material to oversize. Hence, by a process of producing a "slime free" screen product coupled with a "coarse free" cyclone overflow, sizing efficiency could be maximized. Obviously, recirculating load had to be factored into the Combi Screen design capacity. The spiral circuit as shown in Fig. 2 was installed at McClure River in January 1988. The spiral circuit efficiency was very high immediately, and, after the normal debugging, plant capacity has been increased to the point at which we are currently processing of the order of 1100 st per hour. Test program The performance of the spiral circuit has been evaluated by two major samplings of the entire new circuit, coupled with routine bi-monthly tests on all other major components of the plant circuitry.

Jan 1, 1992

By A. W. Khair

This paper presents an assessment of modeling techniques in predicting subsidence over longwall areas. Extensive instrumentation, both at the surface and subsurface of three longwall mines in northern West Virginia have been made. Ground movements were monitored during" a three year-period until subsidence ceased in these areas. A comparative analysis of empirical subsidence prediction techniques including, graphical {NCB}, profile function, influence functions (USBM, BAL&apos;S, KNOTHE&apos;S, and zone area) will be presented. A brief geologic and topographic description of each site along with the instrumentation and monitoring program will also be presented.

Jan 1, 1989