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By M. J. Meulendyke, J. D. Purdue

Within the mineral processing industry, a range of grinding conditions exists which include semiautogenous grinding (SAG),rod milling, and conventional ball milling. Each of these mill environments presents a unique environment for grinding media, requiring the application of specific physical and chemical properties for optimum grinding media performance. The environments are characterized by varying degrees and combinations of abrasive, corrosive, and impact wear. An extensive test program has been conducted to determine the extent wear rates vary between these different applications. Test results are related to production results, ball size, and mill operating conditions.

Jan 1, 1990

Little knowledge is available that considers the effects of continuous miner position on the possible respirable dust exposure of the remote mining-machine operator. Such information could be very useful for increasing the awareness of the remote operator as to those parts of an extended cut that generate higher levels of respirable dust. This data could be used to develop guidelines for positioning the remote operator to reduce his potential exposure. This work details the effects of continuous miner position on dust levels measured on the machine and at the remote continuous miner operator location.

Jan 1, 2000

By I. Iwasaki

Rest potential and linear potential sweep techniques were used to study the cathodic-polarization processes on chalcopyrite, simulating chalcopyrite short-circuited with grinding media in a wet grinding mill. Microflotation tests were carried out to investigate the effect of cathodic polarization on the collectorless flotation of chalcopyrite. It was observed that collectorless flotation of a cathodically polarized chalcopyrite sample could he obtained when oxygen was used for aeration, but not if nitrogen was used. A mechanism for the role that oxygen plays in the collectorless flotation of cathodically polarized chalcopyrite is discussed.

Jan 1, 1999

By Henry P. Cole, Andrzej M. Wala

Engineers and psychologists at the University of Kentucky have developed simulation exercises that teach and test proficiency of mining technical personnel for coping with planned and accidental connections between separate air courses in underground mines. The exercises are based upon investigations of actual mine disasters. The exercises are judged as valuable and interesting training materials by experienced mining personnel. Although these specialists generally perform well on the exercises, potentially fatal errors are not uncommon. The use of such realistic exercises may better prepare miners for preventing and coping with underground mine ventilation emergencies.

Jan 1, 1987

[Costs may be classified according to their source and behavior: Scale of aggregation - Total costs - Unit costs Time period for which the costs are charged - Product costs - Period costs Controllability and relevance for decision-making and control - Controllable costs - Noncontrollable costs - Sunkcosts - Opportunity costs Their basic source - Production costs (also termed manufacturing costs) - Nonproduction costs How the costs change with changes in activity - Fixed costs - Variable costs - Semivariable costs Traceability to individual units of production - Direct costs - Indirect costs Use in planning (budgeting) and subsequent control - Standard costs - Normal costs ]

Jan 1, 1999

By Y. Wang, E. Forssberg

Wet magnetic separation methods for the recovery of hematite and chromite fines and ultrafines were investigated. These methods included wet high-intensity and high-gradient magnetic separation, carrier or "piggy-hack" magnetic separation, magnetic field-induced aggregation and magnetic seeding. This investigation indicated that wet magnetic separation is more efficient for fines and ultrafines. This investigation also indicated that other methods can be used to enhance the particle aggregation and/or the magnetic response. The magnetic behavior of these minerals with respect to the magnetizing field, temperature and particle size is also presented.

Jan 1, 1995

By B. J. Arnold, F. F. Aplan, R. R. Klimpel, E. C. Jr. Dowling

Residence-time distribution techniques are useful tools in the analysis, characterization and modeling of the performance of processing equipment and circuits. This paper presents the results of both conventional and more-sophisticated residence-time distribution techniques for industrial copper-flotation-circuit data. While conventional techniques provide a description that is sufficient for macroengineering design requirements, the more sophisticated tracer-technique methods give more realistic descriptions of process flows and system effectiveness. It is demonstrated that these approaches can provide highly useful information about different equipment, as well as its configuration and performance.

Jan 1, 2000

By Sean Gallagher, Thomas G. Bobick, Richard L. Unger

This paper describes research by the Bureau of Mines on alternative materials-handling strategies for reducing the costs and incidence of back injuries in low-seam coal mines. Strategies recommended for redesigning materials-handling tasks include elimination of tasks, mechanization of tasks, and matching lifting tasks to the physical capabilities of underground miners. The report also discusses other methods that can be used by management to reduce the costs associated with back injuries. More detail can be found in Bureau of Mines Information Circular 9235 (Gallagher et al. 1990).

Jan 1, 1992

By T. Stuczynski

In the past ten years about 3000 wastewater treatment plants have been built in Poland. The result of this development has been a massive increase in overall sludge production. According to data supplied by the Ministry of Environmental Protection, 400,000 dry tons of sludge are generated every year in Poland (Koblak-Kalinska 1996). Still it is difficult to estimate the amount of sludge, which is used in agriculture or for reclamation purposes. However, it can be assumed that this comes to no more than 10 percent of all sludge produced in Poland on a yearly basis. Thus it became obvious that an urgent solution to this problem was needed. Local authorities in Silesia realized in the early 90?s that a demonstration was necessary to encourage the proper utilization of sludge. This was the basis for establishing the Biosolids Subpart of the Silesia Project, which addressed the use of biosolids in the reclamation of mining and smelting dumps (Stuczynski et al. 1996). Metal waste sites themselves are known to contain more than 87 million tons of waste (Pistelok et al. 1995). Each year this amount increases by approximately 400,000 tons. We realized that a simple solution for stabilizing these sites would be to cover them with vegetation in order to reduce leaching of toxic elements, as well as to keep metallic fugitive dust from entering the environment. The reclamation of smelter waste sites, which was carried out within the framework of the Silesia Project was a joint effort of local government, industry and international research institutions/agencies, including the US Environmental Protection Agency (EPA), the Center for Research and Control of the Environment (OBIKS), Virginia Polytechnic Institute and IUNG. One of the goals was to evaluate the possibility of adapting the rules of sludge management currently used in the United States to conditions commonly found here in industrial regions of Poland. However, the main objective of the Silesia Project was the development of guidelines concerning all aspects of sludge use for the reclamation of degraded lands and waste sites. Biological aspects related to reclamation of metal waste with the use of biosolids will be discussed in this paper.

Jan 1, 2002

By Andrew S. Lain

Before you begin mining aggregates in Oklahoma, you must first look at the market and ask these questions: (1) Is there a need for additional aggregate production? (2) Can we produce the aggregate efficiently enough to compete? (3) Is the deposit close enough to the market? If you can answer "yes" to each of these three questions, then you are ready to start the process for mining aggregates in Oklahoma. The first consideration to be made is whether the aggregates produced will need to meet state specifications. If you do not want to meet state durability standards, there are numerous formations you can mine and produce aggregates. We will deal only with the formations that will meet state durability standards. These formations are located in the southern half of Oklahoma and in the Tulsa area. Some of the formations that contain limestone, dolomite, and sandstone that will meet state durability tests are West Springs Creek, Kindblade, McKenzie Hill, Butterly-Royer, and Rush Springs Formations. Granite and rhyolite also are found in Oklahoma and can be used as construction aggregates. Once an area has been identified as having durable material, and drilling has verified its properties, a permit must be secured from the Oklahoma Department of Mines, Oklahoma Department of Environmental Quality, and any other applicable agency before construction of a plant occurs. There are many different brands of crushers available, but all work on one of two crushing principles-either compression breakage or impact breakage. Each type has its application in different industrial-mineral types and the desired end product. A typical small crushing plant would start with a jaw crusher for primary reduction, and then employ one standard cone and one shorthead cone for secondary reduction. One grizzly would be in the circuit with one finished-product screen and three to four finished-product stackers. Natural sands and gravels are found in rivers and old river channels. These too need to be tested to see if spec product can be produced. All of our sand operations have very little gravel; thus, we do not process the gravel. The sand is processed through classifying tanks for gradation.

Jan 1, 1999

By Albert W. Evans

Introduction Engineering Health and Safety is concerned with the understanding of the complex interplay between the risk of injury with technological ,environmental, individual, organisational and legal factors. With this understanding, the design and decision making processes can be balanced with the other objectives of industry and employment to arrive at solutions which place appropriate controls on technology and other aspects of human activities to prevent disease and injury to personnel in the coal mining industry and the design of safe production machinery and ancillary equipment. Advances in the detection and diagnosis of causes of diseases and the need for engineering health and safety in the coal mining industry has increased the need for appropriate education and training. Hazards arise in many forms; exposure to toxic chemicals, escape of radio - active materials, explosion of pressure vessels, contact with moving machinery, exposure to dust, environmental dangers such as excessive noise, high temperatures and manual handling etc. As technology advances, these dangers become more complex and less clearly defined and less easily perceived. Consequently, the ways of controlling the dangers also become more complex and wide ranging. The need to educate and train people in coal mining applications of engineering health and safety has been enhanced over the last five years and courses should be established to fulfil the need for competent people to deal with the problems which modern coal mining technology creates together with the legislation that governs engineering health and safety and to carry out research and development work in the coal mining industry. Industrial Involvement The engineering health and safety experts must be prepared to take a more active part in monitoring conditions that exist in the mining industry, and not design from information obtained from the mechanical, electrical and hydraulic expertise and off-the-shelf components from other industries. Surveying and control of the operating environment must be essentially the concern of the engineering health and safety experts and, whilst it is not possible to create a perfect environment for each machine, every effort must be made to minimise failures and hence plan for designs which will maximise

Jan 1, 1986

By Paulo Roberto Pereira

This paper is motivated by the belief that increasing drilling and blasting costs will not necessarily bring about better digging or crushing productivity. As a matter of fact it can be counter productive as the ensuing discussion will show. In the two case studies presented herein when powder factor increased, productivity at a rope shovel and at a primary crusher decreased. Experience has shown that equipment (shovel and primary crusher) productivity increases as the powder factor increases and concurrently operating costs decrease. This is illustrated by Figure 1where the graph shows (assuming drilling costs constant) that for increases in powder factor initially operating costs might fall till a certain point and then they start to go up again. By the same token we need to understand what is the behavior of the operation’s productivity. It seems intuitive productivity goes up for increases in powder factor but when an operation reaches the point of minimum operating cost one could say that productivity reached its maximum. For continuing increases in powder factor can productivity still go up or does it go down? This paper presents two examples where the productivities of a surface coal mine and of a quarry are reduced for increases in powder factor. This indicates that these two operations could be operating to the right of their minimum operating costs and maximum productivity and therefore wasting precious drilling and blasting dollars and resources. Moreover, their productivity is following the red dotted line in Figure 1 and the question becomes can they follow the blue dotted line for increased drilling and blasting costs? The industry owes to itself to determine this minimum operating cost point and what can be done in case that the productivity goes down afterwards despite an increase in drilling and blasting costs. In other words, utilize your drilling and blasting resources correctly. In the two examples that follow single explosive loading strategies make powder factor a valid means of comparing the results of productivity as a function of powder factor. Also, external factors such as equipment operator influence are eliminated due to the long time span considered in the analysis

Jan 1, 2005

By Omar A. Muhtadi

8.1 INTRODUCTION Several methods currently exist by which gold may be recovered from a pregnant leach liquor. Two of these methods are discussed in this chapter, the Merrill-Crowe (zinc cementation) process and the carbon adsorption process. These two methods represent the most frequently used techniques for gold extraction and are, -in fact, used almost exclusively in U.S. heap leaching operations. Because of the efficiency and ease of use of these two methods, it is extremely likely they will continue to be quite prevalent within the gold mining industry in years to come. The remainder of this chapter includes information on the history of the two processes, descriptions of the actual processes, criteria to consider in selecting a recovery process, and commercial design and construction considerations for both of the gold recovery methods. As with other chapters in this book, some fairly technical information is presented for the benefit of those readers needing a greater level of detail. However, the overall intent of the chapter is to provide an overview of the processes likely to be used in gold extraction, thus giving the reader a firm basis for more technical research into the topic, should this be needed. 8.1.1 History of Zinc Cementation During the late 18901s, zinc cementation was introduced for the precipitation of gold and silver from cyanide solutions. This occurred at about the same time as the introduction of the cyanidation process, the history of which is recapped in Chapter 1. The initial cementation process involved introduction of a gold-bearing cyanide solution onto a bed of zinc shavings. It proved to be quite inefficient because the reaction rate was very slow. The zinc quickly became "passive", inhibiting further gold deposition. Shortly after this first introduction, zinc precipitation was improved by adding a lead salt (usually lead nitrate) to the zinc. This allowed a zinc-lead couple to form on the surface of the shavings, eliminating passivation of zinc surfaces and thereby allowing continued deposition of gold. Further improvements were shortly forthcoming. The first of these involved the use of zinc dust rather than zinc shavings. This provided a much

Jan 1, 1988

By F. Rosenblum, P. Cancilla

The manual moisture determination methods employed today by mineral processing plants falls short in providing timely information required for automatic control. The costs associated with transporting and handling concentrates still represent a major portion of the overall treatment price. When considering the cash flow of a mining operation, that is governed by both the smelter contract, with moisture penalties and the quantity and quality of the concentrates shipped, an efficient method of on-line moisture content would be a welcome tool. A novel on-line determination system for ore concentrate moisture content would replace the normally tedious classic manual procedure. Since the introduction of microelectronic¬based control systems, operators have strived to reduce the treatment costs to the minimum therefore a representative and timely determination of on-line moisture content becomes vital for control set points and timely feedback. Reliable sensors have long been on the "wish list" of mineral processors since the problem has always been that you can only control what you can measure. Today, the task of moisture determination is still done in the classical technique of loss in weight utilizing uncontrolled procedures. These same methods were introduced in the earliest base metal concentrators. Generally, it is acceptable to have ore concentrate moisture content vary within a range of 7 to 9%. Many times, delays in manually achieving reliable feedback on the moisture content results in the moisture varying from 5 to 12% before corrective actions can be made. This paper first reviews the traditional and widely available methods for determining moisture content in granular materials by applying physical principals and properties to measure moisture content. All methods are in some form affected when employed on mineral ore concentrates. This paper describes a unique and promising on-line moisture sensor in two mineral processing applications, which not only automates the tedious tasks but also results in reliable moisture feedback that can be used in the optimization of the de-watering process equipment such as pressure or vacuum filters and fuel-fired driers. Finally, two measurement applications will be presented to indicate the usefulness and to summarize the measurement requirements for the proposed method of employing drag force and mechanical properties of the material itself to determine the moisture content.

Jan 1, 2002

By Richard L. Breithaupt

INTRODUCTION A shuttle car is a self-propelled rubber-tired haul¬age vehicle designed specifically for underground min¬ing, primarily in coal mines (see Figs. 1 and 2). The drive mechanism is located on the sides of the vehicle, and the load is carried in the center. A chain-and-flight conveyor distributes the load when the car is being loaded by a loading machine or continuous miner and also discharges the load onto a conveyor belt or into a mine car. As the name implies, the shuttle car shuttles back and forth between the working face and the un¬loading point and is not required to turn around. Shuttle cars have served as primary haulage in this country since 1940 and still perform this function in most coal mines today. TYPES OF SHUTTLE CARS Shuttle cars may be battery-powered, diesel-powered, or cable-reel electric powered. Battery-powered cars were used between 1940 and 1950, but are almost out of existence today because of their limited power. Diesel cars have been used in coal mines, but are used mostly in noncoal mines where rules and regulations are not as stringent. Also, objections by organized labor have limited the use of diesel-powered shuttle cars in coal mines. Today, the largest number of shuttle cars are cable-reel cars. Electric power is fed through a cable stored on a cable reel that may have a capacity from 152 to 244 m (500 to 800 ft), depending upon the size of the cable and the cable reel (see Table 1). This, of course, limits the travel of a shuttle car. However, time studies have shown that for the most efficient haulage cycle, the length of haulage should not exceed 91 m (300 ft). As haulage distance increases, the waiting time for the loader or miner becomes longer, and haul¬age efficiency suffers. Cable-reel cars are manufactured in standard hand and opposite standard hand models. A standard car has the operator's control on the left front side and the opposite standard car has the opera¬tor's controls on the right front. Generally a loader or miner operates with two shuttle cars, one of each type. This allows the trailing cable to lie on either side of the entry and prevents the cables from fouling each other. Shuttle cars are designed to operate with either dc or ac power. Cars using dc power are designed for 250-v supply and ac cars are designed for either 440, 550, or 950 v, three-phase voltage. Most coal mines that are relatively level can use dc cars where dynamic brak¬ing is not required. For mines having steep grades (in excess of 15%) ac cars are used because ac traction motors have inherent dynamic braking and are safer when tramming downgrade. A recent innovation in shuttle car design in the last five years has been the introduction of an ac/dc car. This design utilizes ac input power through the cable, and the power is rectified on the car to 250-v do for traction motors. The pump motor and conveyor motor are powered by ac voltage. The main advantage of this design is efficiency-a much smaller cable may be used and the smooth operating characteristics of the dc traction motor is obtained. This prevents shock loading on the traction drive train, which is an undesirable feature of ac traction motors. SHUTTLE CAR FEATURES Elevating Discharge Today, all shuttle cars are equipped with an elevating front conveyor that allows the shuttle car to discharge into a mine car or belt feeder without the use of ramps. Hydraulic Power System Today, all shuttle cars are equipped with an hy¬draulic power system. An hydraulic pump is driven by an electric motor which provides hydraulic power for steering, elevating the front conveyor, and powering the

Jan 1, 1982

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader's orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature's effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By A. L. Frank, Benton

INTRODUCTION The selective sensitivity of plastic nuclear track detectors (PNTD's) to low energy 4He particles in an environment which also contains gamma and beta radiations has made these detectors prime candidates for the dosimetric measurement of the concentrations of radon and its daughter products in mine air. Their passive, integrating mode of measurement, small size, low weight and inexpensiveness are attractive characteristics for large scale personnel dosimetry. The detectors can be used in either active or passive dosimeters. In active devices, the PNTD is placed in close proximity to a sampling filter. The filter collects, from a calibrated air flow, all the daughter nuclei which are in suspension. As the daughter nuclei pass through the decay chain, through 214Po, a fraction of the 4He particles emitted are registered as latent tracks in the plastic. In passive devices, the PNTD is placed in direct contact with the ambient air containing the radio-nuclei concentrations to be determined. The active dosimeters have the advantage of excellent sensitivity, and the measured track densities yield very close approximations to accumulated Working Level (WL)** exposures. They have the disadvantage that the simplicity of the passive PNTD is lost, since a battery-driven constant flow-rate air pump is a necessity. The passive dosimeters are simple in construction and use, but they have the disadvantage that WL exposures are not directly measured and certain assumptions concerning radon and daughter equilibrium conditions must be made in determining WL exposures from the measured track densities. Also the sensitivities, in track densities per Working Level Hour (WLH) exposure, are much less than for active dosimeters. Dosimeters of both types have been investigated. Passive devices have been tested extensively both in the laboratory and in mines. Active devices have been laboratory tested. In our earlier measurements, cellulose nitrate plastic detectors were used exclusively, since this material had the highest sensitivity of the PNTD's then in use. When the properties of CR-39 plastic were discovered (Cartwright, 1978; Cassou and Benton, 1978), this material was used, where possible, to take advantage of its superior sensitivity. The results of our earlier work have previously been published (Frank and Benton, 1977). * Research sponsored by the Bureau of Mines, U.S. Department of Interior, under Contract No. JO-188003. ** A Working Level is defined as any combination of the short-lived radon daughters containing 1.3 x 105 MeV of potential 4He-particle energy per liter of ambient air. PASSIVE DOSIMETRY In the earliest testing of PNTD's for passive WL dosimetry, a single strip of cellulose nitrate plastic was used to determine the total cumulative 4He-particle activity in the air of simulated uranium mine and uranium mine environments (Rock, 1968; Rock [et al], 1969; White, 1969). The measurements yielded a close relationship between track densities and WL exposures in a controlled, static environment, but poor accuracy in the active mine tests. It was assumed that equilibrium differences contributed largely to the variations found. The response of PNTD's to airborne 4He-particle emitters, derived by Lovett (1969), demonstrated that detectors such as cellulose nitrate, which has a sensitivity cutoff at 4He-particle energies below the emission energies of radon and its daughters, are equally sensitive to the activity concentrations of radon, 218po(RaA) and 214Po(RaC'). Since radon does not contribute to WL as it is defined, and since, under normal ventilated uranium mine conditions, the radon activity constitutes about 40% to 70% of the total 4He-particle activity, the WL exposures calculated from measured track densities are very sensitive to the particular radon daughter equilibrium conditions. Also, the detector does not weight the individual daughter activities in proportion to their importance to WL. The equation for WL is WL = 0.00103C2 + 0.00507C3 + 0.00373C4 (1) where C2, C3 and C4 are the activity concentrations of RaA, RaB and RaC-C', respectively, in pCi/[L]. The detector leaves C3 unmeasured and weights C2 and C4, equally. [This problem has been approached by Domanski et al (1975, 1979), for some non-uranium mines, by measuring equilibrium conditions throughout the mines to determine a mean value for A, the inverse of the Working Level Ratio (WLR = 100 WL where C1 is the activity of radon-222 in pCi/[L].) The distribution of values allowed Domanski to determine probable errors in calculating WL exposures from track densities. However, measurements of equilibrium conditions in U.S. uranium mines (Breslin et al, 1969; Holub and Droullard, 1978) indicate that this method would not be accurate enough for uranium miner personal dosimetry.] A two-element dosimeter was designed at our laboratory to compensate for equilibrium variations. The first element is a radon detector; the second is a detector for the total 4He-particle activity in the ambient air, just as in the single element dosimeters cited above. The addition of the radon detector allows the equilibrium conditions for the individual nuclei to be determined, given certain assumptions concerning the interrelationships between the nuclei concentrations in mine air. The two detectors and the problems involved in calculating WL exposures from their measurements are discussed below.

Jan 1, 1981

By William F. Eichfeld, Robert J. Wenzel

INTRODUCTION The phrase "subsidence control planning" can refer to the use of specific mining methods and-specialized techniques to control subsidence events or it can refer to planning for subsidence by predicting its magnitude and timing and further planning mitigation strategies. In the first case the objective is actually subsidence prevention while in the second it is a design procedure that seeks to maximize the use of both surface and subsurface resources. Neither technology is well developed at this point. The purpose of this presentation is to briefly outline the Department of Energy's Subsidence Program. This program is, in general, bases on the premise that subsidence control technology, by any definition, can be accomplished only through a thorough and rigorous understanding of subsidence phenomena. This understanding must come from many well planned and executed studies of stress redistribution and strata deformation as they occur during mining activities. There is presently considerable incentive to improve our subsidence engineering capabilities due to the desire for increased coal utilization with acceptable environmental impact. Considerable research in this area is being sponsored by federal and state governmental agencies, universities, and the coal industry. Federal sponsorship has been borne by the Department of Energy, the Bureau of Mines, the Office of Surface Mining and the U.S. Geological Survey for the most part. These agencies have carried out basic research in their laboratories and have contracted with other organizations for additional work. At this time most of the available resources are directed toward contract research, especially in DOE. It is the DOE research plan which will be addressed here.

Jan 1, 1982

By Victor F. Hollister

This section includes examples of discoveries of Kuroko- and Anyoz-type massive sulfides, and shale-hosted, zinc-lead deposits, and describes exploration models for each group of case histories. Exploration or empirical models can differ from genetic models, because ore guides and controls are emphasized in exploration models. Genetic models may form the scientific basis for exploration or prospecting models, although factors that are important to the localization of ore are stressed in exploration models. The paper by Sato (abstracted herein) on Kuroko deposits can be used as a general exploration guide for that type of deposit. It is clear, however, that no unambiguous stratigraphic control for Kuroko-type ore is given in the Sato paper. The paper by MacIntyre (an excerpt herein) describing Gataga, Midway, and Windy Craggy includes most of the elements of the shale-hosted and Anyox models. Although more recent work by the Geological Survey of Canada has modified the Windy Craggy stratigraphy, the Anyox model as applied to Windy Craggy is well described in the MacIntyre (1983) paper. Discoveries of ophiolite-, Ducktown-, volcanic-associated Ni- Cu-, and Besshi-type deposits are not included in this section. Conformable massive sulfide deposits are generally assumed to be syngenetic. Sulfides in the crosscutting feeder zones, however, are classed as epigenetic. Therefore, volcanogenic massive sulfides can contain both a syngenetic and an epigenetic component. Vital to the development of the epigenetic component is a heat source that propels the convection cycle that forms both components. In all massive sulfide discoveries described in this section, it is assumed that a volcanic magma fueled the hydrothermal fluids.

Jan 1, 1991

By Garland R. Dever

Extensive deposits of high-calcium limestone occur in the Upper Mississippian Warsaw and Ste. Genevieve Limestones in the region of the lower Cumberland, Tennessee, and Ohio Rivers in western Kentucky. Analyses of foot-by-foot samples from 20 quarries and cores in 8 counties show variations between the chemical composition of high-calcium deposits in the 2 formations. Knowledge of probable variations in composition should aid in prospecting for high-calcium stone for a specific use and in evaluating potential deposits. High-calcium strata in the Warsaw consist of bryozoan- and crinoid-rich bioclastic limestones, and in the Ste. Genevieve, predominantly oolitic limestones. Values used for comparison were obtained by averaging the foot-by-foot analyses of each high-calcium deposit, 10 in the Warsaw and 15 in the Ste. Genevieve. Average MgCO3 and Fe2O3 contents are generally lower in the Warsaw than in the Ste. Genevieve. Average SO2, S, and P contents are- generally higher in the Warsaw. No characteristic differences are apparent between the average CaCO3 and A1203 contents of the two formations. The higher S and P contents of the Warsaw are believed attributable to the predominance of biogenic material in these limestones. Within the Warsaw Limestone, higher average S and A1203 contents and lower average Fe2O3 contents were found in deposits in the northern half of the region. In the Ste. Genevieve, vertical variations in composition occur within individual high-calcium deposits containing oolitic and nonoolitic limestones. Lower average CaCO3 contents and higher average contents for the other constituents were found in the nonoolitic intervals.

Jan 1, 1972