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By T. D. Parris

At the Frood Mine of International Nickel, tests were begun in March 1966 to evaluate the use of low profile diesel Load- Haul-Dump equipment in cut-and-fill operations. In December 1966, this test was extended to drawpoint loading of caved ore at the Creighton Mine. The results of these tests conclusively demonstrated the effectiveness of this equipment in the operations at International Nickel. This paper describes the equipment used, the operation of the equipment, production data and operating details. Also discussed in the paper are other trackless equipment which works with the Load-Haul-Dump carrier. Production and Safety comparisons are made with the slusher method in bulk mining, giving the reasons for replacing the slushers at the mines of International Nickel Company of Canada, Limited, Ontario Division.

Jan 1, 1969

By Peter Zhang, Todd Minoski, Gabriel Esterhuizen, Jack Trackemas, Berk Tulu, Morgan Sears

Longwall mining is a highly productive and efficient coal mining method used in the USA. During longwall retreating, the belt entry must be maintained stable, and any roof fall in the belt entry would jeopardize mine safety and substantially interrupt the continuous production of coal in the longwall face. Past experience of longwall mining has shown that belt entry stability was the greatest challenge in longwall ground control, and the occurrence of roof falls in belt entries was largely associated with high horizontal stress. To properly support the roof in the belt entry, it is important to understand how longwall-induced horizontal stress changes affect roof stability in belt entries and to strategically install supplementary support to prevent any potential roof falls. Researchers from the National Institute for Occupational Safety and Health (NIOSH) performed horizontal stress measurements in the immediate roof of the belt entries for two Pittsburgh seam longwall panels oriented unfavorably to high horizontal stress. Hollow inclusion cells (HICells) were installed in the immediate roof at the intersection in each of the belt entries, and stress changes were monitored as the longwall face was approaching and passing the intersection. Numerical models were set up to calculate the longwall-induced horizontal stress changes in the roof at the monitoring sites. With the verified model, investigations were made of how horizontal stress is concentrated and relieved in the belt entry roof near the face under different panel orientations. Both measurements and modeling results showed that high horizontal stress is concentrated in the roof in the belt entry within about 15 m outby the face when a longwall panel is unfavorably oriented to the major horizontal stress. The study demonstrated that longwall-induced differential horizontal stress can be used as an indication of the degree of horizontal stress concentration. Roof support strategies are discussed on how to maintain the stability of belt entry under high horizontal stress concentration.

Jun 14, 2022

By F. N. Earll

4.1 – DEFINITIONS 4.1.1-MINERALS, ROCKS, ORES AND NONMETALLIC MINERALS F. N. EARLL A mineral is a naturally occurring inorganic element or compound. Most minerals are distinctly crystalline in habit, and many commonly occur in well-formed crystals or crystalline aggregates, but a few species are characteristically amorphous or noncrystalline. Nearly 2,000 mineral species are now recognized but, perhaps fortun¬ately, a rather large proportion are rarely encountered. Even so, the application of modern mineral-detection and identification techniques reveals the presence of surprising numbers of mineral species, even in deposits of outwardly simple mineralogy. A rock is an aggregate of grains or crystals of one or more minerals. Some rocks are essentially monomineralic. Thus, limestone consists largely of the mineral calcite, and many sandstones are dominantly quartz. Most rock varieties, however, contain a number of mineral species, and their classification or identification depends upon the minerals present, their relative abundance, and such factors as rock tex¬ture and origin. Ore is defined as an accumulation or deposit of one or more minerals that can be exploited at a profit. The profit criterion commonly is taken somewhat liberally, however, for, although always present, the specific mineral content that constitutes a viable product is transitory. The term ore, usually, but no longer consistently, is restricted to the description of mineral deposits that are of value for their metal content. Nonmetallic deposits, although similar to metallic ores in many respects, generally are valued for the minerals or compounds present, as with rock salt (halite), rather than for individual contained elements, though such minerals as elemental sulfur are of much value. 4.1.2-ROCK TYPES F. N. EARLL Rocks arc grouped tinder three broad headings: (1) igneous rocks formed by the crystallization of magma; (2) sedimentary rocks formed by surface processes of weathering and deposition; and (3) metamorphic rocks, which may have been derived from either of the foregoing, but have undergone important physical and/or chemical changes through the application of heat or pressure, or both. The three groups can he further subdivided for convenient reference. Thus, igneous rocks are described as being abyssal or plutonic if formed at great depth, hypabyssal if formed at intermediate depth, and extrusive or effusive if formed at or on the earth's surface. The sedimentary rock group is similarly subdivided into clastic sediments and chemical precipitates. The clastic sediments form by deposition of preexisting minerals that were released from their place of origin by weathering and were transported to their point of deposition as discrete grains

Jan 1, 1973

By Gordon F. Revey

In underground construction and mining work, situations occasionally occur where hard rock or concrete must be excavated at locations near critical structures or utilities. Fearing damage from blasting, contractors and engineers often explore the viability of using mechanical methods or expansive concrete to facilitate the excavation work. However, for moderate to large excavations, the prolonged duration and high cost of using these methods is often impractical. In these situations, under the right conditions, controlled blasting and blast-management methods can be used to speedup the work and lessen costs. The process of determining whether or not blasting is practical is often complex, particularly when risks are high. In this paper, general methods of conducting blasting risk evaluations and outlined and considerations for blast design and risk management are reviewed. This paper also includes three blasting case histories demonstrating how these techniques were used at recent projects. In the first case, a cautious blasting program was developed to facilitate the removal of concrete and rock within a large section of an active mineshaft. The second case involved critical blasting to excavate a dry lake-tap connection between a tunnel and a shaft at the Southern Nevada Water Intake No. 2 project in Henderson, Nevada. The third case history describes how novel blasting methods were used to carefully drop a 2,721,500-kg (6,000,000-lb) concrete caisson into a1.2-m (4-ft) rock socket at the Detroit Outfall No. 2 project in Detroit, MI.

Jan 1, 2001

By S. G. Malghan

Steel consumption forms a significant part of the operating cost of a minerals processing plant. An estimated one-quarter billion kilograms of steel in the United States and over one-half billion kilograms in the world are consumed each year in wet grinding alone. The depletion of ore grades and the necessity of finer grinding of abrasive ores are expected to increase the steel consumption even higher. Through continued research and development by the mining industry and the suppliers of steel products several remedies to reduce steel consumption have evolved. This paper critically reviews the mechanisms involved in steel wear, and currently available methods, materials and operating conditions for reducing steel consumption. Particular attention is given to liner design modifications and substitution of steel liners with rubber liners.

Jan 1, 1981

By Raj K. Singhal

Three principal activities are involved in planning and developing a surface mine. These are 1) definition of economic and technical feasibility, 2) design, procurement and construction, and 3) mining and reclamation. The last is the only activity which begins in the feasibility study and continues throughout the life of a mine. It is also the activity which depends heavily on use of computers. This paper deals with a method of selecting mining software and dwells on the recent progress, made in this area. Discussed also is the status of computerized equipment management and information systems: a microprocessor based technology applied to dispatch systems, monitoring of equipment health and performance, wireless data transmission to a centralized data processing unit, and mine- wide information and reporting system. Reference has been made to upcoming developments which offer the potential- of transforming the surface mine of the year 2000 into a "manless pit". Canadian achievements in all of these areas have been highlighted at appropriate places in the text.

Jan 1, 1989

By Syd S. Peng

According to accident statistics (l, 2), roof fall accidents occur quite frequently at the intersections of underground? openings and account for up to 30% of the roof fall fatalities. But this figure will certainly go up considerably if roof falls not associated with any fatal accident are included in the analysis. Unfortunately, nonfatal roof falls are seldom documented. In the underground coal mines, there are two types of entry intersections. One is the three-way (V type) and the other is the four-way (T type) intersections. A three-way intersection occurs whenever the mainline track haulageway turns at an angle into a section haulageway. The four-way intersections by far the most common type in the U.S. under-ground coal mines are formed by the intersections of two entries which are oriented perpendicular to each other or by the intersections of entries and crosscuts. However, crosscuts are frequently driven at an angle to the entry to facilitate machine movements. The four-way intersections formed by this method of crosscut development are angled. At the intersection of two underground openings, the diagonal roof span is wider than the width of each individual opening. This is one of the major reasons why roof falls occur more likely at the intersections than in the entry between pillars. Consequently, special intersection support plans are usually employed (Figs. 1 and 2) for intersections. The most popular practices are to decrease roof bolt spacing 1 ft. less than or to increase roof bolt length 1 to 2 ft. longer than those employed in the entry between pillars (3). The increase in bolt length or decrease in bolt spacing can be applied to all of the bolts in the intersection or employed only for the strategically located bolts depending on the individual mine practices. These rather generalized practices seems to be adequate for some areas but in many cases inadequate for others. Occasionally, the neighboring two lines of chain pillars are staggered such that the crosscuts in one row of chain pillars face the centers of the chain pillars in the neigh-boring rows. This layout forms T-type three-way intersections which are also inadequate for most weak roofs (3, 4). A study was therefore initiated to develop special techniques for intersection supports. The first phase of the study was the analysis of the stress fields for the underground entry intersections. A three-dimensional stress analysis was necessary for this purpose. Unfortunately, the three-dimensional problems of complicated boundary conditions are beyond the existing analytical methods. The three-dimensional finite element method (FEM) was therefore employed in this study. In the analysis, square pillars of 40 feet wide and an entry height of 8 feet were assumed while entry widths varied at 14, 20 and 26 feet. This model was selected to simulate the mine where roof falls data were collected. The second phase of the study was the collection of field data. The sizes of the roof falls including heights, lengths and widths were collected and analyzed. In the third phase study, the results of the first and second phases were compared and discussed. The analysis of field data revealed an arching zone within which roof falls occurred. The arching zone above an entry obtained empirically was compared with the calculated stress contours. Assuming that the bolts must be anchored in the rock strata outside the arching zone, the preferable systems of bolting were proposed. Also the amount of pretension required to inhibit sliding between layers was discussed.

Jan 1, 1978

By J. R. Tonry

In the grinding of dry non-metallic materials such as rocks, ores, gypsum, lime, and cement clinker, it is recognized that only a relatively small portion of the energy is utilized for actual size reduction, and the major portion of the energy used is converted into heat. As a result, heat is generated inside grinding mills imparting to the ground product a higher temperature than that of the mill feed. In some cases, as for instance in grinding cement clinker and gypsum into finished cement, it may be desirable to keep this temperature increase within certain limits, to avoid any change in the properties of the product and, because consumers may specify a maximum temperature of the finished product for reasons associated, with the tulimate use of the cement. The actual proportions of heat and useful work are somewhat indefinite. Some tests have indicated that nearly all the energy is converted into heat, and. other investigators have found that they could account for "everything except the noise." Investigations sponsored by the American Institute of Mining and Metallurgical Engineers, and carried out with an adiabatic calorimeter, indicated that a surprisingly high 19 percent of the energy might actually be used for size reduc¬tion. However, for the purpose of the following discussion it will be assumed that: (Fig. 1) 10% is lost in motor and gears 10% is lost by radiation and convection from the mill body 10% is used for size reduction, and. that 70% of the energy input leaves the mill as heat in the ground. product. A typical example (Fig. 2) would be a 7x22-ft tube mill operating in open circuit, driven by a 400 hp motor, charged with 77,000 lb of grinding media and producing 70 bbl or 26,300 lb of cement per hour. Reducing this cement from a clinker size of 96 percent passing a No. 16 mesh sieve to a specific surface of 3,230 sq cm per gm (Blaine) will increase the temperature of the cement approximately 143F above the temperature of the feed, as explained later. If the tempera¬ture of the clinker therefore is 125F, the finished cement will be 268F, which in some cases may be too hot. : When cement temperatures approach a range during grinding where dehydration of the gypsum becomes a problem, a logical consideration would be some means of reducing the temperature before it attains the range conducive to gypsum dehydration. Regardless of the varying degrees of heat generated within the mill or mills during grindingto various finenesses, consideration of the maximum safe limit for any type of cement being ground, which may vary fromplant to plant, govern in refining the mill circuit to obtain a safe temperature range. All plants are not necessarily required to do the same cooling job, and may or may not use cement coolers following the grinding circuit. (Not discussed in this paper). In the past, various means have been employed for controlling this temperature increase, Such means have included water cooling of the outside of the mill shell; equipping the shell with cooling fins; injection of moisture inside the mill; ventilation through the mill; and., in cases of closed-circuit grinding, drawing cooling air through the classifier; the purpose in all such cases being to remove heat from the grinding system.

Jan 1, 1958

By M. Robb, S. J. Bate, M. Boisvert

The never-ending quest for gold-bearing ore deposits is a primary industry in Canada, employing people of widely differing expertise. The exploration models for gold deposits are generally geologically complex and the targets have relatively small cross-sectional areas. These problems are compounded in Canada by the minimal outcrop exposure in most mineral provinces and the high percentage of water-covered areas. Naturally the chances of successfully delineating an economic gold deposit are very much greater if all parties involved can act as a team of explorationists constantly confemng amongst themselves particularly as the applicable disciplines generally overlap and complement each other. Three case histories (see Figure 1) are presented to illustrate this philosophy with emphasis placed on the role of the geophysicist in translating the exploration model as defined by the geologist into physically measurable parameters and identifying target areas with similar characteristics. The Conwest-Miminiska property in Ontario is a grassroots exploration play. Additional ore reserves needed to be delineated on the Salmita property in the Northwest Territories. At Laura Lake, near Chapais, Quebec geophysical methods were employed to trace thin high-grade mineral intersections under deep overburden and severe cultural interference.

Jan 1, 1989

By A. W. Fletcher

A nickel-copper sulphide concentrate was treated in a pilot plant at Warren Spring Laboratory during 1961, by a hydrometallurgical roast-leach- solvent extraction process devised to cleanly separate the metal values, and produce acid sulphate liquors containing the metals. It is shown that 86 percent of the nickel and 95 percent of the copper could be leached from the calcine, after a sulphatizing roast on finely ground flotation concentrate. Iron was removed from the leach liquor by limestone neutralization with a small loss of values. Copper and nickel were separated by solvent extraction with naphthenic acid in kerosene after partial neutralization of the iron-free liquors with ammonium hydroxide. Hydrogen sulphide and chlorine were used to remove zinc (and residual copper) and cobalt impurities respectively. The recovery of pure copper and pure nickel, by electrowinning, from the acidic solvent extract strip liquors finally produced, appears technically feasible.

Jan 1, 1964

By Daniel Willis, Ryan O’Connell

An 8-foot tall horseshoe adit tunnel located within an abutment of the Mount Morris Dam in Mount Morris, NY required repair of a failed reinforced concrete liner. The adit tunnel is entirely founded within highly weathered shale bedrock units, which induced unloading stress on the original tunnel liner. An epoxy-impregnated carbon fiber fabric wrap manufactured by Sika was chosen to repair the failed adit tunnel liner. Kilduff Underground Engineering completed the design of the tunnel repair utilizing both empirical and numerical methods. Merco Engineering Contractors completed the injection of chemical grout and repair liner installation.

Jun 23, 2024

By P. A. Korpi

The angular spiral lining system reduces energy consumption and grinding media consumption in grate discharge ball mills. This paper discusses the effect of this lining system on mill product size distribution. Performance comparisons of a grinding circuit outfitted with angular spiral liners in the ball mills are made with respect to standard double-wave liner installations. The angular spiral lining system was developed by Waagner-Biro of Austria. The installation of this lining system at Magma Copper Company, San Manuel, Arizona, is the first of its kind in a grate discharge ball mill used for wet grinding.

Jan 1, 1979

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

INTRODUCTION This manuscript is based on site inspections of five existing mines and the detailed analysis of two of the five. The detailed analysis and mathematical models of the two mines are presented herein in order to demonstrate a method of predicting mine water inflow using existing mines as a data base. Water management procedures were studied at all mines. Our original criteria for mine selection were that the mines be less than five years old and have an inflow rate in excess of 2000 gpm (7.6 m3/min). Background hydrologic and hydrogeologic data were required so that two of the mines could be used for our computer modeling experiments. Our original criteria proved to be difficult to meet. We contacted numerous mines via various sources. Many mines were unable to meet the cri¬teria stated or were not interested in cooperating with us. Therefore, we allowed some flexibility in the application of these criteria to satisfy our orig¬inal requirements. Although the visitations at all the mines are not discussed herein the characteristics of the two mines simulated are presented. These simulated mines consist of a uranium mine in New Mexico and a lead mine in Missouri. The mines and owners will remain anonymous by request. URANIUM MINE-CASE HISTORY NUMBER 1 Mine Location The uranium mine that we simulated is located in northwestern New Mexico. The ground surface at the uranium mine is at an elevation of approx¬imately 7000 ft (2134 m) above mean sea level. The climate is arid to semi-arid with an average annual precipitation of approximately 10 in. (25.4 cm) (John and West, 1963). Geologic Setting It is essential that a conceptual model of the stratigraphy, hydrostratigraphy, and structural geology of a mine site be developed prior to simulation by a mathematical model. If such a conceptual model is not developed prior to simulation major errors are difficult to avoid. The uranium mine we have simulated is located in the San Juan Basin (Fig. 68), which is a structural basin located in northwestern New Mexico and adjacent parts of Colorado, Arizona, and Utah. The sedimentary rocks present in the deepest part of the basin are nearly 15,000 ft (4572 m) thick (John and West, 1963). The regional dip of the sedimen¬tary rocks in the Grants area is generally northward toward the San Juan Basin; the regional attitude is modified locally by normal faults and folds (Hilbert, 1963). The rocks of interest in this area are sedimentary in origin and range in age from Penn¬sylvanian to Cretaceous. They rest on the Precam¬brian core of the Zuni Uplift. Associated intrusive and extrusive rocks are of Tertiary and Quaternary ages (Hilbert, 1963). The deepest sedimentary unit of interest at this mine site is the Morrison Formation as described by Hilbert (1963). The Morrison Formation con¬sists of three members in the general area of the mine site; the deepest member is the Recapture. The Recapture Member consists of alternating beds of gray sandstone and grayish red siltstone or mud¬stone with strata up to several feet in thickness. Only a few small deposits of uranium are contained within the Recapture Member. The Westwater Canyon Member overlies the Recapture Member of the Morrison Formation. The Westwater Can¬yon Member is approximately 150 ft (46 m) thick and consists of light brown to gray, poorly sorted, arkosic sandstone, and some interbedded gray mud¬stone. The Poison Canyon Formation denoted in subsequent discussions, tables, and figures was de¬termined to be a part of the Westwater Canyon Member. The separate connotation is used for the description of mining operations and hydraulic properties but a separate physical description is not included. This member of the Morrison Formation contains many large deposits of uranium. This is (some pages in this article are damaged)

Jan 1, 1986

By W. C. Goldberg

This paper presents pit slope monitoring and back-analysis case studies con-ducted to evaluate and, if possible, stabilize slopes at the Berkeley Pit in Butte, MT. Two specific slopes were studied in detail. The first slope was determined to be near limiting equilibrium, as evidenced by its response to the initiation and cessation of mining, while the second failed on three occasions. Back-analysis of the failed slope verified the laboratory-derived geomechanical properties of the earth materials. Early detection of slope movement can show a need to take action, including modifying slope geometry, dewatering if the slope were wet, and/or investigating and implementing artificial supports. Monitoring slope displacements in open-pit mines can supply data for predicting collapses and preventing serious damage, injuries, and/or loss of life.

Jan 1, 1989

By G. Akdogan, S. M. Bradshaw, A. P. Van Wyk, T. Punt

This study investigated the performance of citric acid as lixiviant for cathode material from end-of-life lithium-ion batteries (LIBs). Black mass containing 84.2 weight percent MNC (LiNi0.45Mn0.4Co0.15O2) and 15.8 weight percent LCO (LiCoO2) material was leached at solid-to-liquid ratios of 20, 50 and 100 g/L. Leaching with 1.5 M citric acid, 2 volume percent H2O2 and a solid-to-liquid ratio of 20 g/L at 95 °C extracted 90 percent Co, 95 percent Li, 94 percent Mn and 94 percent Ni within 20 min. At the highest solid-to-liquid ratio of 100 g/L with 1.5 M citric acid, 10 volume percent H2O2 and 95 °C, 84 percent Al, 84 percent Co, 87 percent Li, 86 percent Mn and 96 percent Ni were leached after 40 min, producing a leach solution containing 0.68 g/L Al, 20.0 g/L Co, 6.0 g/L Li, 13.6 g/L Mn and 19.3 g/L Ni. Using stepwise addition of H2O2, instead of initial bulk addition, did not significantly improve the leaching efficiency but did reduce the required leaching time. Temperature control of the reactor was also more manageable with stepwise addition of H2O2, and evolution of gas was less vigorous. It was observed that the leaching efficiencies of Co, Li and Mn decreased slightly at solid-to-liquid ratios of 50 and 100 g/L, while the leaching of Ni increased slightly. The solution leached at 20 g/L could be stored for six weeks without any spontaneous precipitation. However, solutions leached at 50 and 100 g/L showed changes in concentration after one month of storage, suggesting that processes that have a large material throughput and use high solid-to-liquid ratios around 100 g/L must ensure that the solutions are not stored for extended periods of time.

Apr 1, 2025

By Fabio Mielli

The mining industry is a multifaceted industry with special characteristics that make efficient and intelligent operations uniquely challenging, and subsequently call for distinctive and tailored solutions. Not only does mining call for processes that are complicated and energy intensive, but complex material movements and an intricate logistical chain add to the industry?s obstacles in achieving sustainability and performance. This article explores three major challenges and the resulting issues currently being faced in the industry, as well as potential solutions to these impediments by use of energy management, automation and control strategies.

Jan 1, 2011

By Kenneth P. Katen

Though the use of continuous mining machines consolidated the operations of cutting, drilling, blasting, and loading in one machine that would theoretically provide uninterrupted production, it was soon realized that the face haulage and roof support operations remained as major impediments to continuous production. The ability to provide not only continuous cutting and loading, but also continuous haulage and roof support (see Fig. I), was the basis for the introduction of mechanized longwall systems in this country from Europe in 1951. The early systems employed friction props for roof support, but were to become obsolete shortly after introduction of the first hydraulically powered, self-advancing supports in 1960. Since then, use of the longwall technique has grown to the point where it is currently providing more than 5% of the annual output of underground coal mines in the United States. The shortwall system had its origin in Australia, but was first introduced into this country in 1962 following the development of powered roof supports. It, too, provides for continuous cutting, loading, and roof support (see Fig. 2), but has not experienced the widespread popularity of longwall due to its inherent inability to easily provide for continuous haulage. While there are in the neighborhood of 100 longwall systems currently in operation in this country, there are only a handful of short- wall operations providing production of slightly more than 907 184.7 t/a (1,000,000 stpy) Both of these systems provide uninterrupted roof support for face operations by means of hydraulically powered steel roof supports. In each case, the coal removal consists of open-end cuts made from the narrow end of a large rectangular pillar (panel) which has been blocked out by continuous miners developing entries around its perimeter. However, this is the extent of the similarities between the two systems. It is hoped [ ] that the information presented in this chapter will enable the reader to better understand the two mechanical systems and the mining concepts involved in their employment.

Jan 1, 1981

By Arthur Miller

Most of the production of copper in the United States comes from a few large companies. Often overlooked, however, are the more numerous small mining operations, which collectively produce a significant portion of the nation's copper. The objective of this paper is to describe the operations of a small mining company - Ranchers Exploration and Development Corporation - how it views its place in this industry of giants, how it exploits the advantages offered by its relatively small size, and how it protects itself from the hazards being small introduces. Environment is, of course, a critical factor for such a company because it does not have the resources which help large companies weather periods of financial stress. Ideally, this environment is characterized by a heal thy economy, strong product demand, relatively stable prices , adequate grades and tonnages of ore commensurate with depth and over-burden, ready access to utilities, a dependable labor pool, and suitable technical "know how". When the environment is right, properties which were previously uneconomical to operate attract the attention of companies whose primary interest is a return on investment commensurate with risk over a reasonable period of time. Naturally, the amount of the investment and the duration of operation determines the size of the company which would be interested in such a venture. Such is the case with Ranchers Bluebird Mine at Miami, Arizona - a small property operated by a small company in a favorable economic environment.

Jan 1, 1967

By R. V. Ramani, C. J. Bise

INTRODUCTION Classification of coal seams as thin, medium, and thick is both subjective and relative and is more a reflection of the statistical mode of the thicknesses en- countered in a mining region. A geological factor more precisely understood as defined than thickness is the inclination of the seams to the horizontal, commonly referred to as the pitch or the dip of the seam. Thick and steeply pitching seams present far more challenges to the ingenuity of the mining men to devise methods for their extraction than the thick and less steeply inclined ones. In a marginal case, where a thick coal seam may be minable by surface or underground methods, the decision bears a close relationship to the technological developments in the mining scene. Near surface thick coal seams are a special gift of nature and are not as wide- spread as the ones that call for the application of under- ground methods. Recent advancements in surface mining equipment have made it possible to surface mine to great depths. Stripping ratios of 30: 1 (cu yd of over- burden removed per ton of coal) have been achieved. The extraction of thick seams which preclude the use of surface methods is the concern of this chapter. Specifically, the objective here is to outline the special technological factors in underground mining of thick coal seams and to illustrate through several case histories the methods that are practiced in the North American continent. For the purposes of this chapter, a thick seam is defined as one which is not readily minable with currently available equipment without modifications to the extraction method or the equipment. At the present time, a thickness of 4.8 m (16 ft) can be considered as the lower limit, although developments in equipment and methods can be expected to push this limit upwards. The data for the case histories were collected from two extended mine visits in 1975 and 1978, and from other published accounts of the mines visited. Reference is drawn particularly to reports by Bise, Ramani, and Stefanko (1976) and Rarnani and Kenzy (1979). THICK SEAM MINING METHODS The various methods of mining thick seams can generally be grouped into three major categories: full face methods, where the entire thickness of the seam is re- covered in a single pass; slicing methods, where the seam is extracted to its full thickness by horizontal or inclined slices either in an ascending or descending order; and caving methods, where the seam is undercut and the en- tire coal seam is recovered by caving the top coal into the excavated cut, or some variations of the basic sub- level caving method. As shown in Table 1, in each of these methods, the coal is removed by either room-and-pillar or longwall methods. Hydraulic mining and transport of the coal in open flumes or in pipelines are also practiced. Addition- ally, the void created by the removal of coal may be filled with a packing material (stowing) except in the caving methods. The pitch (dip) of the seam, particularly when it is greater than 0.26 rad (150), imposes severe operational limitations on mobile equipment. A more detailed general discussion of the methods can be found in Bise et al. (1976) and Vorobjev and Deshmukh (1964). Specific discussion on the many variations of the three basic methods can be found in the papers presented at the international symposia that were held in

Jan 1, 1982

By D. A. Norrgran, J. A. Wernham

Introduction In view of the continuing depletion of natural resources, requirements for energy conservation, plant optimization and hazardous waste management, and environmental consciousness, recycling and secondary recovery of materials are being developed with increasing interest Many secondary recovery processes are being developed that apply mineral processing technology. Gravity separation, magnetic separation and flotation have all emerged as explicit methods for recovering residual values from various process streams and hazardous constituents from waste streams. Recently, a technique for concentrating metallics has been successfully reintroduced. The concept of the alternating magnetic field eddy-current separator was developed over a century ago using rudimentary electromagnets (Patent 400317 issued to Thomas Edison on March 26, 1889). Widespread acceptance was, however, hindered by high costs, marginal results and poor economic incentives for secondary recovery. With the recent increased interest in secondary recovery and the development of high-energy rare earth permanent magnets, the eddy-current separator has now gained prominent acceptance. The separation forces imparted by an eddy current separator are the function of the magnetic field intensity and the alternating frequency of the magnetic field. The evolution of permanent magnets has provided a cost effective alternative for the generation of high-intensity magnetic fields. High-energy rare earth permanent magnets have substantially reduced the capital and operating costs over the electromagnetic circuits employed in antecedent eddy-current separators. Specifically, in recent years, the strength of permanent magnets has increased several fold with neodymium-boron-iron rare earth magnets now providing an energy product of 35 million gauss-oersted. Figure 1 shows the evolution in the strength of permanent magnets. The development of these rare earth magnets has led to the design of circuits possessing a magnetic force an order of magnitude greater than that of conventional permanent magnetic circuits. The eddy-current separator has been successfully applied in several metal sorting and recovery operations. Most common is the sorting of metal from shredded automobile scrap and municipal waste. With the advent of high-frequency, high-intensity magnetic fields, the separator has advanced to the point where it has direct application in the beneficiation of fine sized metals. Relatively fine sized (-1/4 in.) metal bearing slags and spent foundry casting sands as well as precious metal hearing electronic scrap demonstrate excellent metallurgical response to the eddy-current separator, and applications have progressed to finer materials and more selective separations. Theory Electrical currents are induced in all conductors when exposed to an alternating magnetic field. The induced current generates a magnetic field in the conductor that opposes that of the alternating magnetic field. Figure 2, is a schematic illustration of generated eddy currents. In this particular example, the alternating magnetic field is produced by a series of permanent magnets mounted on the circumference of a rotor. The permanent magnets have alternating polarity. As the rotor revolves, an alternating magnetic field is produced, and the rate of revolution determines the frequency of the alternating magnetic field. When a conductor, such as a metallic disc in the illustration, is placed in the alternating magnetic field, a closed loop current flow occurs. The current loop in the conductor produces a magnetic field that is a mirror image of the alternating magnetic field. At any point in time, the induced magnetic field in the conductor directly opposes the alternating magnetic field. This opposition of magnetic fields produces an instantaneous repulsion in the conductor. The conductor is consequently repelled from the rotor. The alternating magnetic field has no effect on nonmetallic materials as they pass through the magnetic field and discharge the rotor in a natural trajectory. [ ]

Jan 1, 1992