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By Patrick McElhinney, Jacob Carr, Bruce Whisner, Cory DeGennaro, Jingcheng Li

Since 2016, electromagnetic interference (EMI) of personal dust monitors (PDMs) with magnetic proximity detection systems (PDSs) has been observed in underground coal mines. The EMI causes the magnetic field measurements of a PDS to change, which, in turn, alters the calculated location of the miner relative to the machine. Any altered location calculation can potentially cause the PDS to fail to warn a worker who is at an unsafe distance from the machine, arousing a serious concern on safety hazard caused by EMI in underground mines. The search for EMI mitigation strategies led to the development and use of large shielding pouches and boxes to hold the entire PDM to reduce its magnetic emission. Research on these pouches and boxes found that although they were able to reduce the emitted radiation from the PDM, they also disturbed the magnetic field of the PDS, affecting its performance. Researchers from the National Institute for Occupational Safety and Health (NIOSH) have focused on shielding internal PDM components rather than shielding the entire PDM. The PDM air pump motor is one of the PDM components that has been identified as a major source of electromagnetic radiation and has been selected for further study and tests. The measurements show that a small copper or aluminum foil enclosure can effectively reduce the magnetic emission of the motor by between 50 and 85% at 73 kHz. This study compares the test results of the air pump motor with various cost-effective shielding materials. The data provided in this paper can serve as a reference for shielding enclosure design of the PDM air pump motor to reduce its electromagnetic emission as one form of EMI mitigation strategy.

By R. C. Emmett

Gravity sedimentation is one of the most widely used unit operations in mineral processing and hydrometallurgy. Properly selected and operated sedimentation equipment can provide probably the most economical means of concentrating solids or clarifying solutions. Because of the variety of different approaches possible, including the use of modern flocculants, it is easy to overlook the optimum approach in selecting equipment. The different types of sedimentation systems that can be used in the minerals field and the sizing methods that are currently employed will be discussed here. Sedimentation applications usually fall into two areas, clarification or thickening. While the two operations are basically similar, the difference in emphasis (a well-clarified overflow as opposed to a highly concentrated underflow) results in significant differences in testing, equipment design, and operation. Clarifiers may be merely simple settling vessels with no special features, other than ample volume to provide the detention time needed for suspended solids to settle out of the feed slurry. Or, they can be designed to employ elaborate solids-contact techniques utilizing coagulants or flocculants to achieve more rapid clarification through the growth of particle agglomerates. Thickeners also are available in a wide variety of designs, with the main differences concentrated in the design of the raking mechanism. Thickeners operating in series (continuous countercurrent decantation or CCD) are used to recover valuable solution or purify the insoluble solids. Tray thickeners are used for recovering fine particulate solids such as alumina seed, but in other applications have been supplanted largely by unit thickeners or inclined plate units. Thickeners and clarifiers sometimes are used for storage of either solids or overflow liquor, or employed as precipitating reactors, such as in desilication applications. The use of flocculants and coagulants is much more common in modern processing plants, primarily due to the reduction in capital costs rendered possible through the use of these reagents. The terms flocculant and coagulant are often used interchangeably; generally, it is understood that flocculants refer to organic, high molecular weight synthetic or natural polymers which are useful in flocculating most suspended solids, but generally are ineffective on colloidal suspensions. Coagulants refer to chemicals such as those commonly used in water treatment, alum, ferric salts, lime, etc., which are usually far more effective on colloidal solids but cannot achieve the degree of agglomeration that is possible with the polymers. In some cases, both coagulants and flocculants are used in order to achieve optimum results.

Jan 1, 1986

By Tetsuo Yamazaki

Natural cold seepages are characterized as rapid upward transports of methane from deeper parts of geological structures to the seafloor. Thermogenic and/or biogenic methane generated in deep sediments moves up to seafloor, and major part of methane is consumed in anaerobic oxidation of methane (AOM) and hydrogen sulfide is created in anaerobic sulfate reduction by consortium of microorganisms. Then the hydrogen sulfide is utilized by chemosynthetic communities near the seafloor, such as Acharax, Calyptogena, Bathymodiolus, and the mat-forming bacteria Beggiatoa (Fig. 1). When the supply of methane is large or rapid, remaining methane escapes from seafloor, and it is aerobically oxidized in seawater. A mass balance ecosystem model of chemosynthetic community around natural cold seepages in sediment layer and on seafloor has been created and improved by the authors (Yamazaki et al., 2005; Takeuchi et al., 2007). The model is constructed from 3 main processes as schematically shown in Fig. 2: (1) Methane supply process; (2) Ecosystem process; (3) Water column process. The part of chemosynthetic communities is in the ecosystem process. A numerical model CANDI (Carbon And Nutrient Diagenesis) was developed by Boudreau (1996) to simulate the biogeochemical processes such as organic matter, oxidant, nutrient, bi-product, and pH diagenesis in aquatic surface sediments. Then, the improved version named C. CANDI was developed (Luff and Wallmann, 2003; Luff et al., 2004). The authors created the same type carbon and nutrient diagenesis process model with calcium carbonate formation like C. CANDI (Yamazaki et al., 2005), and have added the sulfur oxidizing and immobilization processes (Takeuchi et al., 2007). The sulfur oxidizing and immobilization processes are connected with the carbon and nutrient diagenesis process model.

Jan 1, 2008

By Roland Hoet

Mine operators, conveyor system designers, consultants, and manufacturers have had opportunities over the past decade to utilize various actively controlled technologies to drive complex underground conveyor systems. These technologies included several electrical drive solutions, mechanical drive solutions, and hydromechanical (fluid coupling) drive solutions. All of these technologies have their inherent advantages and disadvantages, and only after several years of operation can mine operators fairly evaluate the overall benefits (short term performance and cost vs. longer term operational and maintenance costs) of any particular drive. Far too often in today's world, decisions for new equipment are made with inadequate regard to issues concerning serviceability and maintainability. Many of the newer electrical technologies (primarily AC variable speed drives) which have been used in recent years were selected based on the perceived desirable benefit from variable speed operation. This technology, however, which differs sign8cantly between various manufacturers, is in a constant state of evolution which puts into question the manufacturers' ability for long-term product support. In addition, various forms of this technology have also generated some undesirable operational issues, including reduced motor life, harmonics for the power system, and mine communications problems. This technology is also extremely current-limited, which results in problems handling overload conditions, a frequent occurrence often resulting from a decline in the coal recovery rate (conveying more rock than coal). The fluid coupling, however, is based on a technology which has been used successfully in underground conveyor applications for decades. Although the base hydrodynamic technology is well proven and has inherently beneficial characteristics on conveyor applications, Voith has continuously improved on this technology to meet the increasing demands of today's challenging conveyor haulage systems. Recognizing a need in the market for a reliable, actively controlled drive system, Voith Turbo began development in 1996 of a new, actively controlled coupling which was targeted specifically for the higher powered, complex underground conveyors. Voith introduced the new TPKL Fill Controlled coupling product to the market in 1997, and the first two applications for this new technology were on multiple drive mpper booster applications on longwall panel conveyors. Since the successful startup of these two conveyor systems in 1998, Voith has supplied over 200 TPKL couplings to the mining industry worldwide, with over 150 in US coal mining applications alone. This paper will provide an overview of the TPKL's capabilities, performance, and reliability based on the TPKL's impressive installation base and operational history.

Jan 1, 2004

By Kenneth S. Johnson

Oklahoma is the only; state in the nation that is currently producing iodine. The iodine is extracted from iodine-rich natural brines being pumped from deep wells drilled into Pennsylvanian sandstones in the Anadarko basin of northwestern Oklahoma. Woodward Iodine Corp., started in 1977, was the first plant in the United States built for iodine recovery since the 1930s. At present, three companies are operating four iodine-producing facilities in northwestern Oklahoma. The combined production of these facilities, 1,321,000 kg in 1997, is about 26% of the U.S. annual consumption. Iodine is present in concentrations ranging from less than 100 ppm (parts per million) to as much as 1,560 ppm in brines in various Paleozoic strata in the subsurface of northwestern Oklahoma; these concentrations are exceptionally high, and are well above those of any other known brines in the world. Strata with the highest iodine content in the area (Mississippian limestones) have low permeabilities, and thus would yield only small amounts of brine. On the other hand, the lower Morrow sandstones (Lower Pennsylvanian) have a relatively high permeability and are capable of yielding large quantities of brine that average 300-350 ppm iodine. The Morrow brine reservoir was deposited in shallow-water-marine and intertidal-shelf environments and now is buried 6,000-10,000 ft below the land surface. Production is localized to Morrow channel deposits confined to an Early Pennsylvanian paleovalley that is 1-2 mi wide, is about 70 mi long, and is referred to informally as the "Woodward trench." All three companies are producing iodine-rich brines from Morrow sandstones in three different parts of the Woodward paleovalley, and one of the companies also is extracting iodine from oil-field brines coming from several petroleum reservoirs near Dover, about 80 mi east of Woodward. The basic extraction method employed by Woodward Iodine Corp. was used in Louisiana and California operations in the late 1920s. This method is commonly called the air-stripping process. Lower Morrow brine containing dissolved sodium iodide is pumped to the plant through a series of pipelines, where by-product natural gas is separated and processed. The brine is then acidified and subsequently oxidized with chlorine to convert the iodide to iodine. The iodine is air stripped and reabsorbed with sulfur dioxide and water, converting it to hydriodic acid, or process liquor. The process liquor is then processed, using chlorine once again, to crystallize the iodine as a solid. This is then melted and frozen to form a solid product, which is packaged in 110-lb (50-kg) drums and shipped to end users throughout the world.

Jan 1, 1999

By R. Allan Wallis

In the past mining engineers have tended to rely heavily on the advice of fan manufacturers in relation to fan equipment issues. A more adequate approach was advocated in a paper [ Wallis, 1965 ] submitted to the Eighth Commonwealth Mining and Metallurgical Congress held at Surfers Paradise, Australia during 1965. At that time Mount Isa Mine were in the process of installing 3 large 6096mm (20ft) diameter vertical axial flow surface fans of an advanced type [ Mathews et al, 1967 ]. The preliminary feasibility studies prior to preparing the detailed job specification for this project were extremely thorough with the result that it was possible to include detailed information and instructions in the document, ensuring the procurement of equipment that was satisfactory in all respects. Full advantage was taken of expert services provided by external organizations during the large coordinated multi-disciplinary exercise. My continuing activity and development in relation to the technology of axial flow fans is now at a stage where all known design and performance data are accessible through CAD programs [ Wallis and Kipp, 1987 ]. The extensive calculations that were necessary in 1983 can now be carried out in a mere fraction of the time, and with complete confidence. Model testing in order to validate the fan and duct design proposals is, in my view, no longer essential. The first of the computer programs enables a mine ventilation engineer to study a large number of alternative axial fan arrangements for comparison with the centrifugal alternative. By entering the volume flow rate, fan pressure and air density, he can then with ease, and rapidly, investigate the effect of various blading arrangements, downstream evasee area ratios, and hub ratios, on fan diameter and speed, with their related attainable fan efficiencies. This procedure had its origin back in the 60s when I was advising mining companies on new fan equipment. At that time such an investigation was quite lengthy and lacked the precision and completeness that is now available within the present programs. To obtain the major benefit from these outputs, it is most important not to be disconcerted by the three very important nondimensional parameters, namely, the tip flow coefficient (LAMt), the fan pressure coefficient (Kt or Kit), and the swirl coefficient (Es and/or Ep) as these are important when assessing the various design proposals. When seeking an understanding of the role and importance of each of these variables, ample help is provided in the user manual; the design recommendations are summarized in the help screens which are readily accessible within the design programs.

Jan 1, 1989

By John Rushton

Surface mining, from ore extraction to processing, requires many types of equipment to obtain the final product. Mine equipment often includes shovels, drills and haul trucks. Ore processing equipment, including plants, mills and concentrators, are also part of surface mining operations. Surface mining operations emphasize production. Production is where the bills are paid and fortunes are made. However, there must be support to keep the equipment operational. To most production people, equipment maintenance is a necessary evil. Maintenance is not glamorous work. It is often neglected when praise is being passed around. It is easy to recognize higher production, greater capacities, faster cycle times or more loads hauled. But who cares if an engine is replaced in 12 hours instead of 48? The answer is, everyone should - from the hourly miner to the mine manager. Surface mining equipment needs constant maintenance work. It must be lubricated, components replaced, electrical problems repaired and other problems addressed. Repair work for any surface equipment takes time away from production. It costs money. Maintenance costs cannot be avoided. However, they can be minimized. Minewide communication should include operations cooperating with maintenance. Traditionally, operations resist surrendering equipment to maintenance if it can continue to produce. The rationale is, every load that can be hauled today is money in the bank. If the unit is taken down, the money is lost. This thinking is prevalent in many surface mining areas, including the mill. "Not on my shift," the operator thinks. "Let the next crew worry about it. The problem with this thinking is in probability and statistics. The longer equipment goes without maintenance work, the higher the probability of equipment failure. When the failure does occur, it is certain to cause the equipment to remain down longer than if the maintenance department had been allowed to take it down. Many mining operations have closed early in their lives because of poor maintenance practices. This has become the industry standard because substandard levels of performance are acceptable. Management believes it gets five to six hours of work from maintenance people during an eight-hour shift. Actually, the industry average is less than two hours of work per shift. Real productivity levels of less than one hour are not unusual. There are few exceptions. However, a daily work schedule (DWS) will increase productivity at many properties by 50% to 100%. A 200% improvement is not unusual. The DWS will work. Maintenance people are among the most intelligent and motivated people on the property. They deserve the chance to be productive. [ ] In preparing a DWS, management needs to understand two productivity principles - the productivity spread and the real maintenance workday.

Jan 1, 1997

By Scott G. Britton

The first section of this text delves into construction engineering techniques for successfully completing necessary underground production support projects. I will discuss alternative design concepts, installation practices, materials and material costs, completion time, and necessary labor for individual projects. The emphasis of this section is on describing each project thoroughly and providing a practical guide for the supervisor or engineer. There are four main construction categories for typical underground mining projects. These are: 1) Transportation of men and supply materials. This refers to track and wire installation and roadbed grading and ballasting. 2) Ventilation, composed of overcasts, man doors, regulators, and stoppings. 3) Drainage, including sumps, pumping stations, drainage networks, pumps, and face drainage. 4) Haulage of coal and rock from the face to the outside. The focus here is in the installation of belt lines, belt drives, transfer points, and vibrating bins/feeders. The first section shall fully discuss these four categories in regard to the construction engineer or supervisor designing, building, or improving a system. In addition, special projects such as surge bins, rotary dumps, and underground offices not classified under these categories will be discussed. For manageable purposes, I have limited the installation discussions to one or two approaches I feel have merit. The variation of installations in coal mining is endless, each with its own advantages and disadvantages. Therefore, the reader should not dismiss or restrict his thinking to only the approaches in the text. It serves as a guide only.

Jan 1, 1983

By Henry N. McCarl

In the period 1982-1992, world trade has increased in real terms about 65% (Anon., 1992b). The nominal value of this trade in the early 1990s is approximately $5 trillion. In 1970, mineral commodities constituted approximately 20% of the value of world trade. By 1980, despite rising petroleum prices, mineral commodities made up about 18% of the total value of world trade, with crude oil amounting to 16% and other minerals the remaining 2%. By 1987, the major worldwide economic slowdown and falling oil prices meant that minerals constituted only about 10% of the total value of world trade, or about $500 billion in 1987 (Kimbell and Zajac, 1988). The importance of minerals in world trade since 1987 has remained at about these levels. Approximately 40% of world mineral commodity production is traded internationally. Some economists predict that minerals and other commodities are likely to continue to decline in relative importance in terms of the value of world economic output, as well as world trade, into the foreseeable future (Thurow, 1992). This projected decline in commodity demand may also be related to much slower global economic growth over the next decade (Alperovitz, 1992). Industrial minerals and rocks constituted roughly 2.5% of total mineral commodity exports in the late 1980s, but their relative importance as a portion of mineral commodity exports nearly doubled in the decade 1977-1987 (Kimbell and Zajac, 1988). Industrial minerals constitute the most rapidly growing share of total mineral commodity exports. In addition, price declines among industrial minerals were uncommon in the decade 1977-1987 (Kimbell and Zajac, 1988). While prices of sulfur and most of the fertilizer minerals have seen some erosion and slow growth in markets in the period 1987-1992, other industrial minerals have increased in both value and share of the export market (USBM, 1992). The place of industrial minerals in world trade becomes even more significant when one looks at shipments of chemical fertilizers and basic chemicals that are counted as manufactured products rather than industrial mineral commodities. The United States is the major world producer of industrial mineral commodities and products made from these nonmetallic minerals and rocks. The former Soviet Union, now The Commonwealth of Independent States (CIS), and China have very significant deposits of industrial minerals. Japan exports significant quantities of cement. The US, CIS, and Morocco export large amounts of phosphate rock. Canada and Germany export substantial amounts of potash. Sulfur from Poland, industrial diamonds from South Africa, Zaire, and Brazil, and fluorspar from Mexico are all important exports.

Jan 1, 1994

By J. H. Rowland

Four different types of fire suppression systems were evaluated in large-scale fire experiments to study the effect of air velocity on the effectiveness to extinguish conveyor belt fires. The four different types of suppression systems were water sprinkler, deluge type water spray, and two different dry chemical fire suppression systems. The large-scale fire tests were conducted with 72-in-wide fire resistant rubber belt that met the Mine Safety and Health Administration flame resistant requirement specified in Title 30, Code of Federal Regulations, Part 18, section 18.65 (also known as the 2G Flame Test). Each fire suppression system was tested at air velocities of 500-550 feet per minute (fpm) and 1350-1500 fpm. Two tests were conducted at both air velocities for each fire suppression system. Both of the water-based fire suppression systems were able to suppress the fire to the point that a miner could extinguish it with a fire extinguisher. However, the amount of water needed to suppress the fire to the point where a miner could walk up to extinguish any smoldering belting was greater than the current MSHA regulations require. MSHA regulations only require 10 minutes of water supply to the suppression system. In the test set up neither of these systems would have suppressed the fire had the water been turned off to the system after 10 minutes. The dry chemical suppression systems produced mixed results. Dry chemical suppression system A did not suppress the conveyor belt fire at either air velocity. Dry chemical system B suppressed the fire at 500 fpm air velocity, but produced mixed results at the higher air velocity. Details are presented on the large-scale fire test set-up, arrangement of the fire suppression systems, and conclusions regarding the effect of air velocity on fire suppression system design and performance. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of NIOSH.

Jan 1, 2009

By John A. Cordes, John J. Schanz

INTRODUCTION The economic structures of industrialized countries require a continuous supply of reasonably priced and appropriately processed nonfuel minerals. As a result, national policy makers are sensitive to real or perceived threats to the effective operation of these commodity markets. Their most persistent concerns can be subsumed into the related political economy issues of supply availability and marketplace stability. These concerns are not new. For over a century the "global struggle" for mineral resources has influenced national policies and international relations (Eckes, 1979). During the past two decades the complexity of these issues, if not their place on national political agendas, has increased. Dramatic changes in the structure of mineral industries and in the nature of international economic relations have redefined policy needs and response options. This paper focuses on how selected OECD countries have responded to nonfuel mineral industry issues over the past twenty-five years. Particular emphasis is placed on the impact these policy decisions have had on the size and location of productive capacity and its utilization rates. The results, perhaps, are not surprising for this group of market- oriented economics. While individual cases of significant impact are easy to identify at various times in most countries, the overall importance of policy decisions has not been great. This is especially true for the past decade when excess capacity reached unusual highs of magnitude and duration. Macroeconomic forces and market structure changes played much more important roles. PERSPECTIVES ON THE ISSUES OECD nations dominate the international nonfuel minerals marketplace as producers, consumers and traders. In 1986, these twenty-four countries accounted for 75 percent of the Western world's export trade value in minerals, ores and metals. They imported 88 percent of the total value. During 1966, the respective percentages were 70 and 94. Japan was the world's largest importer in 1986 followed by the US, FRG, UK and France. The same five countries led the 1966 list except that, then, Japan was ranked fourth. Table 1 ranks the fifteen largest Western exporting nations in the two specified years. It also indicates the growth in current export dollars earned in 1986 compared to 1966.

Jan 1, 1989

By W. Haupt, M. Thormé, H. J. Ertle, K. Reim

Though Germany has a very old mining history, her experience in uranium mining only started in the late sixties. Today in the Federal Republic of Germany (FRG) there are only 5 small mines in the exploration stage, namely Menzenschwand and Müllenbach in the Black Forest and Mähring, Popenreuth and GroBschloppen in northern Bavaria. In all these exploration mines much development is required to define shape and size of ore bodies. Mining engineers, government authorities and environmentalists alike are very much concerned about environmental safeguards and protection of mine workers against radon and its daughter products. Miners are affected by internal and external radioactive exposure. In general the external [f3]- and [µ] -exposures can be considered as small in comparison to the internal exposure and are therefore not discussed in this paper. The internal exposure is caused by the inhalation of radon and its daughter products and is relevant to the lung as the critical organ. The nobelgas radon is inert and does not from compounds with other elements easily. Contrary to the physical half-life of 3,8 days the biological half-life of radon is only 10 to 20 minutes. During the rentention time in the body only approximately 1 % of the long-lived decay products are retained in the system (SCHOGER, 1976). In this connection is it important that the radon resorption as well as the radon secretion varies individually and depends to a large extent on the age of the person (SCHOGER, 1960). RENOUX et al. (1978) have determined that the greatest part of the a(-activity in the air is attached to aerosoles with a diameter of 0,4/14. The radiation exposure of the lung by radon is negligible if compared with the exposure by short-lived decay products. For this reason one has introduced the concept of the potential [o](-energy (P) of the short-lived radon decay products, which means the total [ct]-energy, which is produced when an atom decays in the decay-series down to the long-lived lead-210. [ ] The unity of the potential [o]-energy concentration in the SI-system is 1 j/m3 air. In radiation protection however a specific unit is used namely the working level (WL): [ ] The WL-unit, though internationally recognised has not been introduced officially in Germany. Because of the large amount of tunneling still to be done in the FRG, the various mines departments responsible for safety and health in mining ask searching questions such as: - What are the laws governing radon emanation in tunneling? - Can workers be adeguately protected against radiation hazards? - How should one drive through high grade uranium ore? - What happens if the ventilation system breaks down?

Jan 1, 1981

By M. Tirmarche, J. Brenot, J. W. Piechowski

INTRODUCTION Statistical analysis of results obtained by workers survey is very often limited to the determination of the collective dose and mean annual values. This information is not sufficent by itself for radiation protection purposes since the cumulative exposure of a worker during his working life is not taken into account. A more complete analysis would be obtained by considering both: [ ] These two notions are different. The former depends on the evolution of the protection with time and is related to annual limits (ICRP, 1977 and ICRP,1977 a). The second depends mainly on the works performed during the period of survey, the interference of protection evolution is weakened. So it is useful to foresee how the individual average annual doses would be when considering survey periods of 5, 10 years, or more. STUDIED POPULATION In order to analyse the two previously mentionned notions it is necessary to follow up a well defined population of workers which is continuously exposed to the radiological hazards on a long enough period. The selected population is that of the workers of the French uranium mines always present during the 1971-1979 period (9 years). The number of individuals is 937. The radiological hazards are: 1. exposure to radon and short-lived daughters 2. exposure to long-lived ore dust 3. external irradiation. The workers of the uranium mines are exposed to these radiological risks in different ways according to the work performed. It is therefore necessary to subdivide the population into more homogeneous groups with respect to the radiological risks. Three groups are considered: - Group 1 : 227 individuals whose radon and dust exposures are always nil - Group 2 : 449 individuals whose radon and (or) dust exposures are different from 0 every year - Group 3 : 261 individuals represent a mixed category. Some features of these groups are given in table 1. [ ] The 937 followed individuals represent approximatively the half of the whole population of the French uranium mines. Taking into account the procedure adopted to select the studied population, it appears clearly that this population is not representative of the whole miner population. DOSIMETRIC RECORDS Three risk indicators are considered .Individual annual exposures to radon and its shortlived daughters. They are determined by ambient monitoring based on radon concentration measurements and recording of the times elapsed in the various workplaces. For each worker the annual exposure is represented by the mean annual air concentration and is expressed as a fraction of the maximum annual concentration (MAC). Given the administrative arrangements and the effective state of equilibrium between radon and its daughters, the maximum annual concentration is practically equivalent to 1 Working Level = 2.08 x 10- 5 jm-3. (UNSCEAR, 1977). . Individual annual exposures to long-lived ore dust which are also expressed in fraction of the maximum annual concentration. . Individual annual doses of external irradiation (X and [y] rays) expressed in mSv (= 10-3 Sievert). Irradiation monitoring is based on use of individual film badges.

Jan 1, 1981

By J. M. Scharer, Bruce E. Halbert, Pettit. Carol M., Lysane Bolduc

The ratio of neutralization potential (NP) to acid production (AP) ratio is often employed for predicting acidic discharges from waste rock dumps. The available NP was evaluated in laboratory studies of limestone dissolution in simulated acidic mine water. The available neutralization potential strongly depended on the particle size. Using limestone particles of 114" (6.4 mm) or greater, less than 20% of the total neutralization potential was exhausted at the onset of acid conditions. Apparently, the dissolution of limestone became mass transfer limited. Under mass transfer limitations, the rate of pyrite oxidation may exceed the rate of neutralization by buffering minerals. The implications concerning the controls on limestone dissolution were examined using a geochemical predictive model for sulphide mineral oxidation in waste rock. The results have shown that in case of heterogeneous waste rock piles the NP/AP ratio is a reliable indicator for short-tam predictions only. Kinetic data on the depletion rate of the neutralizing minerals and geochemical modelling suggest that waste rock environments with NP/AP ratios as high as 5.0 may turn acidic in the long term.

Jan 1, 2000

By K. K. Cheng, Z. M. Chen, A. L. Mular

INTRODUCTION Since the nineteen fifties, process engineers have been seriously concerned with process models for various purposes (1,2,3). Models serve as subroutines to mathematical flowsheets which simulate circuits of interest. Simulators can serve to increase process know- ledge, assess circuit performance, investigate flowsheet modifications, evaluate new flowsheets, perform off-line optimization, and compare alternative control strategies. This chapter describes a methodology to develop a simulator of a bulk Cu-Mo circuit comprised of roughers, scavengers and 3 stages of cleaning. The simulator was employed to evaluate the influence of water addition at circuit points and to estimate the optimal number of cells for each stage in the circuit. DESCRIPTION OF CIRCUIT Figure 1 is a flowsheet of the Brenda Mines Ltd. (located near Peachland, B. C.) bulk Cu-Mo flotation circuit as simulated prior to an expansion/renovation which involved the installation of large, new flotation cells. Referring to the flowsheet, feed to rougher cells includes fresh feed, scavenger concentrate and 1st cleaner tailing. Rougher con, 1st cleaner (sect ion 2) con and 2nd cleaner tailing are fed to 1st cleaners (section 1). 1st cleaner (section 1) concentrate and 3rd cleaner tailing serve as 2nd cleaner feed, while 2nd cleaner concentrate is fed to the 3rd cleaners which produce a final bulk concentrate. Table 1 lists cell types and specifications important to the circuit. DEVELOPMENT OF MODELS To develop flotation cell models a systematic procedure was employed (4). This involves the selection of responses and factors, proposing models to be fitted, the acquisition of raw data, the statistical adjustment of raw data, fitting models to adjusted data and testing the adequacy of fitted functions. Responses and Factors Responses of interest included recoveries, concentrate grades and concentrate solids/water flow rates for each stage. Factors that influence these responses, such as reagent addition rates, air flow rates to cells, solids and water flow rates to stages and feed grades were not manipulated during the sampling period. Thus, the investigation was limited,

Jan 1, 1986

By R. Woods

INTRODUCTION Gaudin (1), considered that "the mechanism of mineral collection is the central problem of flotation theory". From their work on adsorption phenomena in flotation for more than fifty years (2), Gaudin and his co-workers have made major contributions to the elucidation of this mechanism, a fact acknowledged by their contemporaries in the field (e.g. (3)). Their investigations in a number of areas, e.g. demonstrating that submonolayer quantities of adsorbed collector species can induce flotation, or identifying the influence of substances which can activate or depress (1), have helped to provide the basis for and the development of a theory to account for the interaction between mineral and collector. It was established in the late fifties (4) that the presence of oxygen is a prerequisite for the flotation of metal sulfides with thiol collectors. This fact rendered inadequate the concept of simple ion adsorption being responsible for rendering these minerals hydrophobic. The requirement for oxygen has been explained (1) in terms of this species being a component in chemical reactions between the sulfide surface and the collector. Gaudin has been instrumental in identifying the products of chemical reactions at sulfide minerals (1.5), a line of work which has been continued particularly by Finkelstein and co-workers in South Africa (6). In presenting a detailed mechanism of the mineral-collector- oxygen interaction, it has often been considered that the mineral surface - must first be oxidized to a metal oxy-sulfur species which

Jan 1, 1976

By M. A. Ouederni, G. C. Chorosz, E. P. Déliac

This paper deals with the heat flow problems in deep workings (depth greater than 1000 m in France). The approach benefitted from experience gathered in France and South Africa. Deep mines are located in the East of France (coal in Lorraine and potash in Alsace), where the geothermal gradient can reach 4 'C per 100 m (Virgin Rock Temperatures, VRT, around 50 'C). This paper is concerned with the following topics: method of studying the heat flow problem on a mine-wide basis, specific models for heat and moisture pick-up, interfacing of heat and ventilation calculations and a case history in the Lorraine collieries. The heat flow problem on a mine-wide basis is studied in a fashion similar to ventilation networks. Each branch is considered as a "black box" where the air picks up heat and moisture (or is cooled, in the case of an air-cooler). Each box is therefore modeled as a heat transfer unit, hence the calculation of outlet temperatures (wet- and dry-bulb) from the intake ones. The whole network can thus be computed as a sequence of boxes from the air intakes towards the outlets. The second section reviews a number of models used for the assessment of heat and moisture pick-up in the "boxes", with emphasis on stopes and development ends, which are particular to the mining method in use. Two types of stopes are considered: longwall (horizontal and inclined) and cut-and-fill (subvertical). Mechanized rock breaking and blasting can be accounted for. A particular attention is given to development ends, especially complex from the heat flow point of view (rock characteristics, wetness factors, etc.) and usually quite hot. These workings, which are critical in order to secure new production areas, are studied on an average basis as well as in extreme conditions. The calculation takes into account the history of rock temperatures. The interaction between ventilation and thermodynamic calculations is still under development. The idea consists of coupling ventilation calculations with the "box" models: the first step determines the direction of air flow, hence the sequence of the boxes; thereafter the calculations yield the effect of temperature and presssures on air flows. The final section reports on the validation of the whole approach against field data. It concerns several heat flow surveys aimed at checking the models for shafts, stopes and development ends. The accuracy of the method is found very satisfactory.

Jan 1, 1987

By Douglas B. Silver

The key to unlocking value is to find opportunity where others have missed it. How does a company find value in a thoroughly explored asset? This article explores the approach taken by one company and the results it achieved from working this problem. The town of Shafter is located on a small bend in the road in west Texas' Big Bend Country. Its past emulates the glory days of so many other mining camps. A discovery was made, fortunes were built and World War II shut down the operation. The Shafter deposit continues to be one of the larg¬est undeveloped primary silver deposits in North America. Gold Fields Mining Co. recognized the area's potential in 1977 and embarked on studies to reopen and develop the district. Gold Fields spent $16 million on the Shafter project. It drilled more than 86,240 m (282,940 ft) in 475 holes, sunk two shafts and developed 1,550 m (5,100 ft) of underground workings. Today, the mothballed 447-kW (600-hp), single-drum Nordberg hoist and permanent production-type, single-compartment headframe continue to crown the shaft. Other buildings onsite house compressors, mine pumps, spare parts, drill core and other supplies. All these items look as though they arrived yesterday. By 1982, Gold Fields had completed an extensive feasibility study with ambitions to. construct a 1.36-kt/d (1,500-stpd) agitated cyanidation mill and base metal gravity circuit. Reserves at Shafter stand at 4.23 Mt (4.67 million st) grading 193.5 g/t (5.65 oz/st) silver. Gold Fields planned to produce 71.5 t/a (2.3 million oz/year) of silver at a head grade of 196.2 g/t (5.73 oz/st) silver with a 77% recovery rate. Cash costs were projected to be $5.54/oz with an anticipated initial and total capital cost of $46.2 million and $51.3 million (1984 dollars), respectively. The study assumed a silver price of $8.60/oz that has not been seen since the Hunt brothers quit buying silver. When silver prices declined, Gold Fields put the property up for sale.

Jan 1, 1999

By William A. Vogely, Lawrence J. MacDonnell

This section begins with an introduction to several key environmental laws. This is followed by a discussion of some of the issues raised by such environmental regulation. Finally the subject of regulatory reform is addressed. Overview of Selected Major Environmental Laws The number, type and complexity of environmental laws and regulations affecting mineral resources development have grown at an astonishing rate since the 1960s. The most significant laws and regulations are those addressing the use of the public lands; those concerned with air quality, water quality, and surface reclamation; and those defining government responsibility for its actions having important environmental consequences. The laws and regulations governing activities on the public lands will be taken up in the section 5.15C, "Minerals and Land Management." In this section we will concentrate on five primary laws: the Clean Air Act, the Clean Water Act, the Surface Mining Control and Reclamation Act, the Endangered Species Act, and the National Environmental Policy Act. The Clean Air Act: The legislative and regulatory framework for controlling air pollution has developed in a series of steps. Most important for present purposes are the Clean Air Act of 1970 (P.L. 9 1-604) and the 1977 Amendments (P.L. 95-95). The rather complex air quality control scheme includes four key elements-national ambient air quality standards, state implementation plans, the prevention of significant deterioration program, and the non- attainment program. National Air Quality Standards. The 1970 Clean Air Act is a landmark law because it contains the first explicit recognition that pollution is a nationwide problem requiring a national control program. The federal government through the Environmental Protection Agency (EPA) was to establish national air quality standards for at least the six pollutants named in the Act (sulfur dioxide, carbon monoxide, total suspended particulates, photochemical oxidants, hydrocarbons, and nitrogen dioxide). These standards were to be established so as to protect public health, "allowing an adequate margin of safety." The cost of achieving these standards is not to be considered. The 1970 Act unrealistically required that these standards be achieved in all areas of the country by 1975. As the program has developed, the air quality control regions are now to be designated as "attainment" or "nonattainment" with respect to each of the national standards. State Implementation Plans. Under the Act the states are given the primary role in achieving the ambient standards. States are to develop and implement control plans governing emissions from all existing and new sources. These plans can include emission limitations for existing stationary sources, transportation control plans to reduce emissions from mobile sources, and pre- construction review of new stationary sources to assure compliance with new source performance standards. The EPA must approve all state implementation plans. Sanctions are given to the EPA to be applied if states fail to submit acceptable

Jan 1, 1985

By Donald O. Rausch, William E. Hawkes, Odin Jr. Strandberg

INTRODUCTION Kennecott Copper Corp.'s Tintic Div. Burgin mine is a 725 t/d (800 stpd) lead-zinc producer, located in the east Tintic mining district, 5.6 km (3.5 miles) east of Eureka, and 130 km (80 miles) south of Salt Lake City, UT. The Burgin mining complex includes three shafts, shop, dry, and office facilities, and a 454 t/d (500 stpd) flotation concentrator. Production and ser¬vice entry to the mine is through a 4.26-m (14-ft) diam concrete-lined production shaft, and a three-compart¬ment rectangular shaft. A connection with the Apex No. 2 shaft and the Burgin mine provides a second exhaust airway. Bear Creek Mining Co., Kennecott's exploration sub¬sidiary, began exploration of the Burgin mine area in 1957 with the sinking of the 335 m (1100 ft) deep No. 1 shaft. A station was cut at the 320-m (1050-ft) level of the shaft and extensive drifting, diamond drilling, and some winzing and raising confirmed the presence of a major lead-zinc ore body, the main portion of which oc¬curred below a hot saline water table. Surface facilities and the 4.26-m (14-ft) diam concrete-lined production shaft were completed in 1966. Ore production has been steadily increased to 725 t/ d (800 stpd) since 1966. The collar elevation of the No. 2 production shaft is 1731 m (5678 ft) above sea level. The mine's three levels, the 1050, 1200, and 1300 are at elevations of 1414.6, 1356.4, and 1338.1 m (4587, 4450, and 4390 ft), respectively. The approximate elevation of the hot saline water table is 1382.6 m (4536 ft). The earliest production on the east Tintic mining district was from the Eureka Lilly mine in 1909. Later, in 1916, the Tintic Standard mine was discovered, and pro¬duced ores with a gross value of $80,000,000 through 1956. The Eureka Standard mine located 1.6 km (1 mile) southwest of the Burgin mine produced ore worth $12,000,000 from fissures in the Tintic quartzite along the Eureka Standard fault. Ores produced from the east Tintic mining district through 1971 are valued at $151.3 million. The production from the entire Tintic district is approximately $500 million. GEOLOGY The Burgin ore deposit occurs as replacement bodies in limestone, dolomite, and shale. Numerous fault struc¬tures and rock formations make the mine geology com¬plex. The mineralization is strongly associated with sev¬eral major geological features, which include the Ophir formation, consisting of limestone, dolomite, and shale as the favorable host rock, and the structural influence of the east Tintic thrust, the Eureka Standard and sev¬eral other significant faults. Fig. 1 depicts the Burgin 1300 level workings, a generalized configuration of the ore body, and the main geological features. Covering the quartzite, limestone, dolomite, and shale facies enclosing the Burgin ore deposit are 183 to 274 m (600 to 900 ft) of andesite. The Burgin ore consists of massive sulfide and oxide lead-zinc ores with significant silver and cadmium val¬ues. The main ore minerals are galena, cerrusite, and spalerite. Extensive manganese oxide deposits are associated with the lead-zinc ore occurrence. A limited

Jan 1, 1982