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By Tore M. Larsen

The Rødsand Gruber mining company is a subsidiary of the Norwegian industrial corporation Elkem AS, and its mine is situated in the northwestern part of southern Norway. The ore from this mine is a relatively low grade, magnetite ore with a small amount of vanadium. The ore bodies are very irregular in shape, resulting in a high amount of waste rock intrusion in the stopes. The annual underground production in the mines is approximately 1 million t. Because of varying dimensions of the ore bodies, both transverse and longitudinal stopes are employed. The inclination of an ore body is usually 1.04 to 1.59 rad (60 to 90°), and the horizontal width varies up to 100 m. Sublevel stoping has been the traditional mining methods in this mine. Until a few years ago, the ore was primarily blasted using horizontal blasthole fans drilled from Alimak raises. Recently, however there has been an increasing demand for a les expensive mining method. Alternate systems were considered, and finally a method based upon large blasthole stoping was found to be the most suitable. GENERAL LAYOUT The general layout of a stope is shown in Fig. 1 As with all the figures in this chapter, only the prin¬ciples of the method are emphasized and the details may not be valid for any particular site in the mine. In transverse stopes, the horizontal width of the ore determines the length of the stopes. The average length is approximately 50 m, the width of the stopes is 20 m, and there is a 10-m wide vertical pillar between. The distance between the main levels is 60 m. Usually a horizontal (sill) pillar 10 m thick is left below the mucking level. Each stope is first undercut along the entire stope length. This undercut is made with 50.8-mm (2-in.) diam drill holes in a fan pattern. At the moment this is being done by a trackless Atlas Copco Promec jumbo equipped with two Cop 130 pneumatic rock drills. The drill drift is 4 X 4 m, and the crater is 20 m high. This leaves a 36-m stope height to mine with large blastholes. To achieve a fair portion of vertical drilling withlarge blastholes, there are two parallel drill drifts at the top of each stope. These drifts are 4 m high and 9 m wide, separated by a 4-m wide pillar in the middle of the stope. To ease the drilling of perimeter holes, the drifts extend 1 m into the vertical pillar on each side of the stope. The pillar will thus be only 8 m wide on the drill drift level, but 10 m wide down the rest of the stope.

Jan 1, 1982

By C. Anderson, B. Schug, C. Cristoffanini, S. Nazari

The Haile Gold Mine in South Carolina is increasing the automation level of their processing plant in order to improve operational stability. One step towards this goal is the application of a Digital Twin concept to provide Virtual Instruments. Haile has engaged ANDRITZ to implement Virtual Instruments at their grinding area hydrocyclones and at their tailing thickener underflow pump. These virtual instruments provide density values in a traditionally difficult location for reliable slurry density measurement. The Digital Twin is a first principles IDEAS1 ™ simulation model, connected in real time to the control system. Results have been promising, with the virtual instrument values matching the physical instrument under normal operating conditions, and showing better prediction under upset conditions. Haile plans to use the virtual instruments as optional process control parameters. INTRODUCTION OceanaGold’s Haile Gold Mine is located near Kershaw, South Carolina, about 50 miles southeast of Charlotte, North Carolina. This location was first mined in 1827, and has produced gold through several operating periods since then, including a heap leaching operation from 1985 through 1992. In October, 2015, OceanaGold acquired Romarco Minerals, the previous owner of the Haile Operation. The current operation began in 2017 as an open pit mine with additional underground resources. The nameplate throughput is 6,350 tpd ore, and 2017 production was 118k oz gold, with 140 to 155k oz gold expected in 2018. Proven and Probable reserves exceed 3 Moz. The process, as shown in Figure 1, includes a primary jaw crusher, comminution with a Semi-Autonomous Grinding (SAG) mill and ball mill circuit, flash flotation, rougher flotation, regrinding, pre-aeration, carbon in leach (CIL) leaching, carbon elution and regeneration, cyanide recovery, electrowinning and refining. The Haile Operation’s inaugural gold pour took place January 20, 2017. Mine life is expected through at least the year 2030. Based partially on prior experience of some key OceanaGold personnel with automation services from ANDRITZ on other projects, OceanaGold decided to include ANDRITZ on the Haile project. The scope of these services for Haile included design and supply of the process control application software, an ANDRITZ dynamic process simulator for operator training, and ANDRITZ Advanced Control Expert (ACE) ®2 systems for the SAG mill, ball mill, and rougher flotation areas. Since startup, ANDRITZ has continued to provide services for the site, particularly related to PLC and HMI equipment, programming and support, ACE system expansion, and the implementation of Virtual Instrumentation. The implementation of this new Virtual Instrumentation technology at Haile is consistent with Haile goals: increased automation, a change in the role of the operator from interactor to monitor, and improved stability by removing human error.

Jan 1, 2019

By Jasinder S. Jaspal

The United States has very large coal reserves in seams that are less than 1065 mm (42 in.) thick. These reserves are often not mined because of the difficulty of working in low heights and, in most cases, the seams are beyond the technical or economical limits. It is estimated that the remaining reserves in 355 to 710 mm (14 to 28 in.) thick seams that are not faulted or intruded, dip less than 15° and under a cover of 610 m (2000 ft) or less are nearly 910 Gt (1 trillion st). This is a very large percentage of coal reserves in the United States (Unrug, 1985; DuCarme et al., 1990). The US Bureau of Mines is attempting to develop new techniques for mining 150 to 1065 mm (6 to 42 in.) thick coal seams. Research has focused on existing thin seam mining techniques, testing of small- and full-scale physical models of the Bureau's slot mining concept and computer modeling of the full-scale process. In 1986, only 21 % of the total US coal production came from -1065 mm (-42 in.) thick seams. Of that, 355 to 710 mm (14 to 28 in.) thick seams share of production was minimal, probably 5% or less (Mayercheck et al., 1990). This shows that the thin seams, particularly the seams 355 to 710 mm (14 to 28 in.) thick, have not been mined to the extent of the available reserves. Thin seam coal, generally, is of premium quality. As thicker seams are depleted, the mining of thin seams could become more important. Thus, there is a need to develop technology to mine these seams (J. Farinelli et al., 1987). A successfully applied technology could change reserve categorization. In the Appalachian Plateau there are 417 Gt (460 billion st) in 355 to 1065 mm (14 to 42 in.) seams. This represents nearly 9.2% of the US remaining reserves (Unrug, 1985). At depths of 152 m (500 ft) or less, the Appalachian Plateau has about 214 Gt (236 billion st) of reserves in 355 to 1065 mm (14 to 42 in.) thick seams. Other mining areas also have substantial coal reserves in this seam thickness category. Those, however, are not as attractive for mining as the Appalachian Plateau. In the past, most of the reserves of the Appalachian Plateau region were partially worked by contour and highwall auger mining. It is estimated that these mining systems had left some 40,230 km (25,000 miles) (Readers Digest, 1987) of unreclaimed highwalls. The Bureau recognized this situation. In FY1987, the agency initiated a research project to develop anew technology for mining thin seams of the Appalachian Plateau.

Jan 1, 1992

By John W. Rushton

Maintenance and inventory are usually the last items considered in mine development. They need to be addressed properly and early. The micro-computer brings sophistication within the reach of even the smallest mines. Maintenance and inventory programs, if properly installed, will reduce costs 15 to 25% and allow management time to address the mining and metallurgical problems at start up. Shops and warehouses are rarely designed properly. Large engineering firms spend 5 to 20 times more than is necessary for first class facilities. There is an optimum design, size, and layout. Take the time necessary to assure you get the right people, programs, and facilities in that order. Maintenance, inventory, and inventory control must be addressed early if a small mine is to have a chance at optimizing costs during the first two years of operation. This is hard work and is usually put on the back burner until the mine is in some degree of trouble. Hiring an experienced maintenance superintendent is important but it is not the complete solution. Maintenance systems, design of facilities, and use of computers are not skills that maintenance people usually have. In addition, the work of ordering spare parts and working with erection people is a full time job and any maintenance people on site during construction are quickly recruited. The end result is that few mines start up with an adequate Maintenance or Inventory System in place. The months, and in some cases years, that the mine attempts to operate without a system are both painful and expensive. Designs of successful integrated maintenance and inventory programs are virtually 100% supported with some outside assistance. Our system is the approach that we are most familiar with and this paper will isolate how Rushton International would go about setting up systems and designing support facilities. Micro-computers are the only way for a small mine to go. ~ini-computers cost more originally and operating costs are many times higher than even the most sophisticated micro system. Micro- computer software is developed further and recent advances in hardware makes medium priced units equal to most mini's in RAW COMPUTER POWER. Mines in general, and especially small mines, have no need for main frame power. MAINTENANCE A good maintenance program has: 1. Equipment Master List a. Equipment Detail Reports 2. Work Order and Backlog Control Program 3. Component Control Program 4. Tire Program 5. Wear Materials Program 6. PM Control Program 7. Maintenance History Program 8. Performance Program We also track miscellaneous items such as MSHA training and inspections, timekeeping, etc. Rather than going through the program step-by-step, it makes more sense to go through a typical Maintenance Day and show how the programs are used. The first thing maintenance supervision needs to know is "how we are doing." We

Jan 1, 1987

By John E. Frost

Duval's (now Battle Mountain Gold Co.) copper and gold deposits in the Battle Mountain District in Nevada are in two subdistricts - Copper Basin and Copper Canyon - about 18 km (11miles) apart. Location of the two subdistricts is shown in Fig. 1. The two centers of mineralization have distinctive geologic and ore characteristics that required different innovative approaches to exploration and development. Nearly all of the copper from Copper Basin has been recovered by heap leaching. Because of relatively high gold and silver values in the Copper Canyon ores, a significant part of the sulfide copper ores has been recovered by a conventional flotation mill at Copper Canyon. Recently, this mill has been converted to the recovery of gold that is mined from zones peripheral to the copper deposits. Prior to Duval's entry, small amounts of copper, gold, silver, lead, zinc, and antimony had been mined in the district over a nearly 100 year period, beginning in 1866. The bulk of the $13 million total production to 1962 was derived from copper and gold lode ores and gold placers. Duval's interest in Battle Mountain-arose directly from a submittal by Asarco, which controlled key claims in both the Copper Basin and Copper Canyon subdistricts. Asarco acquired its holdings in the district in 1959 and undertook limited exploration until mid-1961. At that time, the company decided to dispose of its interests. Duval was invited to evaluate Asarco's properties and accepted. On Dec. 6 and 7, 1961, J.E. Frost, R.C. James, R.W. Sayers, and H.A. Schmitt, representing Duval, and K. Richard and E.W. New- man, representing Asarco, visited Copper Basin and Copper Canyon. Frost concluded that the properties were of considerable interest and prepared, by Jan. 12, 1962, a voluminous report recommending that Duval acquire and explore Asarco's properties. Furthermore, he recommended a large initial exploration program that included not only geologic mapping, geochemistry, and geophysics, but also the extensive drilling that was to involve, as the first phase of exploration, 49 drill holes in Copper Basin and 21 drill holes in Copper Canyon. Following property agreements between Duval and Asarco, and Duval and several other claim owners, geologic mapping, geochemical surveys, and drilling began in August 1962. Exploration drilling was completed in July 1964 and engineering studies in April 1965. Development of the Battle Mountain properties was authorized in July 1965, and production commenced in the first quarter of 1967. Because of the unusually high oxide copper content of the ore bodies, particularly in Copper Basin, a much larger than typical proportion of the copper in the deposits has been recovered by heap leaching. The generalized geology and ore zones of the Copper Basin are as shown in Fig. 2 and of Copper Canyon in Fig. 3.

Jan 1, 1991

By J. B. Gwiazda

Introduction Thick seams constitute a considerable portion of Polish coal reserves. In producing mines and those under construction, the reserves where the height is more than 4.5 m (15 ft) total about 24v0 of all the hard coal reserves. The reserves that are under urban and industrial areas can be extracted using stowing methods for surface protection. About 50v0 of thick seam reserves can be extracted by caving methods. Thick seam mining methods Thick seams are usually extracted by multilift methods. There are some methods of thick seam mining that generally can be divided into caving and stowing methods. The stowing method is more expensive, but must be used when mining to protect surface features. The more known methods of thick seam mining include: • extraction of thick seam in ascending order with stowing; • thick seam mining in descending order with consolidated stowing; • thick seam mining with gob consolidation; • thick seam mining in descending order with roof caving; • sublevel caving thick seam mining system; and • thick seam mining system using an artificial roof. • Thick seam mining with introduction of an artificial roof The most advantageous method of thick seam mining is the method using an artificial roof. The artificial roof usually consists of wire mesh. Such wire mesh must be durable and must prevent small rock fragments from passing through. The mesh is usually made of steel properly protected against corrosion. The artificial roof is placed on the coal floor of the top slice (Fig. 1). Caving rock falls on the artificial roof and is thus separated from the coal. The next slice is extracted under the artificial roof. The producing face of the next slice must progress within 10 m (33 ft) behind the front of the first slice. As the second producing face progresses, the wire mesh moves down on the coal floor and may be used as an artificial roof for the next slice. Thus, one artificial roof allows the extraction of several consecutive slices. Application of the artificial roof allows for clean extraction and reduces coal losses. It also eliminates the hazard of spontaneous combustion. Introduction of the artificial roof has been limited due to a lack of artificial roof laying equipment. Manual laying of wire mesh is a labor-intensive operation, and it delays face progress, resulting in poor productivity. To eliminate these problems, artificial roof laying equipment was designed at the Polish Central Mining Institute. The support and the method of laying operation is covered by a patent. Artificial roof laying powered support To extract thick seams using an artificial roof; four-leg shield supports with lemniscate linkage were developed (Fig. 2). This shield is used in the first top

Jan 1, 1988

By George McArthur

GEOLOGY One of the largest copper resources in North America occurs in the Bingham mining district, located 40 km (25 miles) southwest of Salt Lake City, UT. Anaconda's present contribution to the development of this resource is the Carr Fork mine. (Fig. 1, Bingham district plan map.) The Carr Fork story began about 300 million years ago when a thick sequence of sedimentary quartzite and limestone was deposited in the Oquirrh Basin. Subsequent earth stresses formed a series of west-northwest-trending overturned anticlines and synclines that were later offset by northwest-striking thrust faults and northeast-trending normal faults. (Fig. 2, Bingham district cross section.) About 38 million years ago the Bingham stock was intruded into some of the folds, resulting in the formation of porphyry ore in the intrusive, and fissure, replacement, and skarn ore deposits in the sediments. The Carr Fork skarns were formed preferentially in two limestone units, called the Yampa and Highland Boy, which average about 46 m (150 ft) in thickness, separated by about 76 m (250 ft) of quartzite. Copper mineralization consists almost entirely of chalcopyrite associated with pyrite and molybdenite in a gangue of garnet, diopside, magnetite, quartz, and carbonate. Molybdenum, gold, and silver are economic byproducts. (Figs. 3 and 4, indicated ore in plan and section.) EXPLORATION HISTORY The West Mountain mining district, of which Carr Fork is a part, was organized in 1863, after leaf-silver ore was discovered in the Carr Fork of Bingham Canyon. However, it was not until 1873, when the Bingham and Camp Floyd railroad to the town of Bingham was completed, that the mining camp flourished. Many small mining companies in the district were reduced, often through consolidation. By 1937 only the National Tunnel and Mines Co. was active in the Carr Fork area. In 1948 it declared bankruptcy, and was later purchased by The Anaconda Co. Anaconda conducted geologic investigations in the Bingham district until 1968, when a concerted diamond drilling program was initiated to fully evaluate the geologic potential of the Carr Fork area. The exploration program commenced with one diamond drill, but was enlarged to three drills following discovery of the first major ore intercept. Near completion of the exploration effort, large oil well rotary drills were used to drill to depths of 762 to 914 m (2500 to 3000 ft), after which the holes were completed by standard diamond drilling techniques. All holes were collared vertically, but considerable deviation occurred because of structure and rock conditions. Wedging and directional drilling were used to keep the holes on target and to provide multiple ore intercepts in the deeper holes. Geophysical surveys, in-

Jan 1, 1982

By Robert J. Hite, James P. Searls, Sherilyn C. Williams-Stroud

Potash is a generic term that includes potassium chloride, potassium magnesium sulfate, potassium sulfate, potassium nitrate, and sodium-potassium nitrate mixtures. In the ceramics industry, potash is also used to refer to potassium oxide. Potash, primarily in the form of potassium carbonate, was the first industrial mineral produced in the United States, and the first US patent issued was for an apparatus and process developed in 1790 for its production (Paynter, 1990). Prior to the 1860s, potash was primarily sold as an impure form of potassium carbonate produced by burning hard- wood trees and leaching the potassium salts from the ashes. The major early uses of potash include soap and glass making, dyeing fabrics, baking, and saltpeter for gunpowder. In 1859, the development of a purification process to remove the sodium and magnesium chlorides was developed for the carnallite found at Stassfurt, Germany, and mined potash became available. With the appearance of mined potash and the earlier (1840) discovery in Germany by Justus von Liebig that potash was a nutrient for crops, potash started to be used for high valued crops such as cotton and vegetables. The German potash companies quickly developed a manufacturing process for producing potassium sulfate for tobacco fertilization. German potash supplied nearly all American needs until the embargo of the First World War when imports from Germany were interrupted (Bateman, 1918). With the discovery of potash deposits in New Mexico in 1931, the United States became self-sufficient in potash. In 1962, the United States began importing potash from Canada, and two years later domestic apparent consumption began to exceed domestic production. Along with nitrogen and phosphorus, potassium is one of the three essential plant nutrients, the "K" of NPK terminology. As a result, 95% of potash production is used as plant fertilizer. In all plants, inadequate potassium diminishes growth, causes increased disease, stalk and stem breakage, and susceptibility to other stress conditions. Plants take up large quantities of potassium from the soil, and potash fertilization replaces this loss so that each new crop can be grown with the same vigor and productivity as the previous year's crop. The potassium depletion of the soil from growing repeated cotton and tobacco crops is well known in the history of southern agriculture in America. George Washington was known to have studied alternative crops that could be grown on soil that had been depleted by repeated tobacco crops. Most of the remaining 5% of potash consumption is by the chemical industry, as potassium hydroxide to produce soaps and detergents, glass and ceramic products, dyes, explosives, alkaline batteries, and medicines. Potash as chemical is used in oil field drilling mud, the aluminum recycling industry, and the electroplating industry. Additional minor uses for potassium chloride include water softener regeneration, sidewalk deicing, and salt substitution for human consumption. Potash is used in the food industry as potassium phosphate, and in production of glass products as potassium carbonate or nitrate. GEOLOGY Potassium is the seventh most abundant element in the earth's crust and the sixth most abundant element in seawater. It is found in silicate minerals of igneous, metamorphic, and sedimentary rocks and is also a major constituent of many surface and subsurface brines. The majority of world potash resources are found in subsurface bedded salt deposits which yield high grade, large tonnage ore bodies and are amenable to low cost mining and beneficiation. Because of the relatively high solubility of potassium minerals, potash from salt deposits is ideal for use as fertilizers. Some potash production is from evaporation of naturally occurring brines, but the vast majority of current domestic and international production is from bedded salt deposits. Sylvite, carnallite, kainite, and langbeinite are some of the more important potassium minerals (Table 1). Sylvinite, a mixture of KC1 and NaCl is the highest grade potash ore. Carnallite can be considered a potash ore when removal of magnesium chloride is included in the beneficiation, but it can also be considered a contaminant when mining for sylvite. Potassium sulfate and potassium nitrate are typically manufactured products. Potassium sulfate is produced from mined minerals through conversion processes in Italy, Germany, and Carlsbad, NM, and from brines in southern California and at the Great Salt Lake in Utah. Natural deposits of potassium nitrate occur only in small amounts in Chile. The majority of potash-bearing bedded salt deposits are believed to have originated from the evaporation of seawater or mixtures of seawater and other brines in restricted marine basins (Schmalz, 1969). The reflux depositional model for evaporite deposition was first described in the literature in 1888 by Ochsenius. A shallow bar, or sill, across the mouth of a basin lets in a restricted flow of seawater which evaporates into a salt-precipitating brine (Fig. 1). The density of the brine at the distal end increases with increased salinity, sinks to the bottom, and sets up a reflux current of higher density brine back toward the ocean. The sill, which restricts the inflow of seawater, allows inhibited flow of evaporation-concentrated brines back to the ocean. The least soluble salts are precipitated nearer the sill, and the most soluble components come out of solution in the deeper parts of the basin. The result is a lateral facies change in a tabular-shaped deposit that is due to the salinity gradients in the brine (Fig. 2A). The asymmetrical facies distribution of the Paradox Formation (Middle Pennsylvanian) Utah (Hite, 1970), the Prairie Formation (Middle Devonian) in Saskatchewan (Holter, 1972), and the Salado Formation (Upper Permian) in New Mexico (Lowenstein, 1988), might prompt explanation by such a model. Other deposits, such as the Salina Formation (Upper Silurian) in Michigan (Matthews and Egleson, 1974), show a facies distribution that could be described as a bull's eye pattern. Although some small subbasins of high grade sylvite are found near the margins, the potash is generally located in a central part of the basin surrounded by successively less soluble facies (Fig. 2B). The sparse

Jan 1, 1994

By G. Swartz, R. J. Brison, R. A. Wyman, J. P. Vandenhoeck, S. E. Khalafalla

Magnetic fluids are commonly composed of ultrafine submicron particles of magnetite (Fe3O4) dispersed in either a kerosene or a water carrier medium with oleic acid as a protective colloid. The unique property of these fluids is that when the magnetic particles are attracted to a magnet the carrier medium moves along with the particles so that the liquid also appears magnetic. The availability of a fluid whose gross behavior changes sharply in a magnetic field without affecting its rheologic characteristics is becoming critical for many technologic applications. About two gener¬ations ago suspensions of relatively large, micron-sized ferromagnetic particles in oil proved useful in clutches, brakes, and dashpots. These were called magnetic clutch materials, and their viscosities were highly dependent upon the applied magnetic field, in contrast to true mag¬netic fluids which retain their fluidic characteristics under all applied fields and field gradients. While the term "ferrofluid" was used by Rosensweig and Kaiser&apos; to designate a magnetic colloid composed of a dispersed magnetic ferrous material, the more general term, "mag¬netic fluid," is preferred. This is because these fluids may contain ferro- or ferrimagnetic substances, such as cobalt, nickel, gadolinium, dysprosium, or magnetic ferrites in carrier liquids, such as hydrocar¬bons (kerosene), silicones, water, fluorocarbons, etc. Preparation of Magnetic Fluids Magnetic fluids were first prepared by Papell and Faber, at the NASA Lewis Research Center, Cleveland, Ohio, by tumbling ferrite materials for several weeks in a carrier medium (such as kerosene) and a dispersing agent (such as oleic acid) with steel balls. Rosensweig et al.1 compared the stability of their ferrofluids obtained from both metallic iron and the magnetic oxides of iron, Fe3O4 and yFe2O3. They found that, although the saturation magnetization of iron is 3.5 times greater than that of magnetite, its colloidal solutions deterio¬rated much more rapidly. The iron particles were undoubtedly rapidly oxidized to a nonmagnetic state due to the enormous surface area of these colloidal particles. Because the properties of these fluids are highly dependent on particle size and may acquire opposite characteristics by small incre¬ments in that size, more controllable methods of preparation have been developed recently by Khalafalla and Reimers.6 In one method metastabilized pyrophoric wiistite is ground to colloidal size in a kerosene oleic acid mixture, and then it is disproportionated to colloi¬dal magnetite and iron at temperatures just below the eutectoidal temperature of wüstite stability (570°C); thus, t < 570°C 4 FeO t > 570°C Fe + Fe30, (1) Because of its antiferromagnetic characteristics, the precursor wüstite is much easier to grind than magnetite. Another method3 avoids grinding entirely and utilizes a peptiza¬tion technique from ferrous materials in the molecular state of aggrega¬tion. This latter procedure has resulted in preparation of strong fluids at considerable savings in cost.

Jan 1, 1985

By C. William Parisi

Since the introduction of mechanical mining, some 35 years ago, a great deal of effort has been expended by the coal industry to solve the coal dust problem. The original concern was the possibility of a dust explosion or the propagation of a gas ignition; however, in recent years it has come to be recognized as a possible &apos;health hazard as well. In the middle thirties, the Pittsburgh Coal Company began an intensive study of the dust problem to determine the dust levels created by the various cycles of operation in mechanical mining and to make a comparison with those of hand loading operations. Numerous dust samples were collected with a midget impinger and it was found that dust counts varied from 38 to 268 million particles per cubic foot during hand loading and from 96 to 227 million particles per cubic foot during mechanical loading. The study indicated that the source of underground dust in coal mining could be classified as primary or secondary. A primary source is one associated with appreciable degradation of size, that is, where most of the fines capable of becoming air-borne are produced. A secondary source is one in which most of the fines have been produced at some previous operation and are in such condition that they can become air-borne. Cutting, drilling and blasting are the operations which produce the primary dust. Secondary dust is formed by operations which include; loading, transportation and discharge at transfer points. The study showed clearly that the coal cutting operation was responsible for the greatest amount of dust and that the amount of dust produced varied directly with the speed of cutting; type and dullness of cutting bits; and the distance the cuttings had to fall to the floor. The information obtained from the study enabled our company to decide on a course of action and the direction to follow. With the cooperation of mining machine manufacturers, water tanks with a 200 gallon capacity were built into the machine trucks and were provided with a pump driven by a small motor. A length of 1/2" water hose carried the water from the pump to the two water spray nozzles located at the front of the machine. Water was delivered to the machine tanks by water cars which were filled on the surface or from an underground water supply. It was during these early days that wetting agents were introduced into the water system. Our studies revealed that wetting agents did little better than plain water in the abatement of the dust problem. It was found that when using a wetting agent less water was required to obtain the same results but we could not take advantage of this fact because it was almost impossible to keep the smaller nozzles open to reduce the amount of water being used.

Jan 1, 1967

By Meng Wang

"In this study, multiple methods that include geological surveying, laboratory studies, and numerical modeling have been used to study the deformation characteristics of a “deep inclined rock roadway with weak planes” (DIRRWP). Failure of the DIRRWP was found to be triggered initially by the shearing of the rock near the weak plane. The movement of the failed rock resulted in reduced confinement of the deeper rocks, and causing the failure of the supporting units and the severe deformation of the surrounding rocks. Hence, it is concluded that DIRRWP stability is affected predominantly by rocks near the weak plane. Corresponding control strategies for DIRRWP deformation were proposed: (1) greater support intensity is necessary to maintain the stability of the whole DIRRWP; (2) additional support is needed for rocks near the weak plane to reduce the potential of failure; (3) to increase the bearing capacity of the floor strata, as the floor needs to be supported; and (4) a supplementary support technique should be considered for cases of large deformation. The case study presented here indicates that the proposed control strategies can significantly reduce the potential of failure in rocks near the weak plane, thus preventing severe deformation of the DIRRWP.INTRODUCTIONAs shallow seams are mined out, the cover depths in most Chinese coalmines have become increasingly greater. Deep cover in a coalmine is considered to occur when the cover depth exceeds 800 m (Sun, Chen, and Miao, 2018). The loading and failure behavior of rocks in a deep cover setting are quite different from those in a shallow cover setting owing to the increase in in-situ stress, temperature, and hyperosmosis (Lobanova, 2008; Zhang et al., 2013). For instance, experience shows that bulking and dilating deformation in deep roadways is much greater than that which occurs in shallow roadways (Hou and Yang, 2018; Zhang et al., 2016). In addition, anisotropic behavior and thickness deviation of coal measure rocks, combined with the appearance of weak planes (e.g., faults, joints, and coal cleats) in coal and/or rocks, can significantly affect the stability of an underground roadway (Ghazvinian et al., 2013; Sun, Chen, and Miao, 2018). The presence of weak planes can significantly complicate displacement and stress transmission in nearby rocks. Additionally, tensile, compressive, and/or shearing failure in rocks near weak planes can be easily caused by high in-situ stresses in deep roadways; such failures may lead to further severe deformation of the surrounding rocks, resulting in roof falls and rib spalling (Kamata and Mashimo 2001; Wu, Ohnishi, and Nishiyama, 2004). A typical failure of the so-called “deep inclined rock roadway with weak planes” (DIRRWP) is shown in Figure 1. Severe deformation can be found near the weak plane. In this study, a plane or thin layer of rock/coal parting that can significantly reduce the strength of the surrounding rocks—and thus affect roadway stability—is defined as the weak plane."

Jan 1, 2019

By Donald Whittaker

Severe climatic conditions in mines are normally associated with working at depth. More precisely, it is a matter of working in regions of high strata temperature, since, for example, some coal mines in the Ruhr have strata temperatures at 1000 m which are matched only at 3 times that depth in the South African gold mines. In Britain we are perhaps fortunate in encountering strata temperatures in our deeper coal mines some 10-15 K lower than at the same depth in the hot mines of the Ruhr. Nevertheless the problems of high temperatures and humidities in UK mining operations, which in the past have been confined to a few specific sites, have in recent years become more widespread, and there is every reason to suppose that the trend will continue. Clearly strata temperature is not the only .causative factor. Long ventilation circuits serving districts remote from the shaft must also play a part, while the heavy use of machinery and the rapid, rates of excavation of modern mining methods are obviously important. It is necessary to understand the mechanisms of all such factors, and to measure their magnitudes and fluctuations, if remedial measures are to be properly designed or predictions made which are accurate enough to be of use in ventilation planning. This paper presents the results of a long-term (and continuing) investigation carried out in several mines to measure the major sources of heat and moisture on longwall districts and trunk conveyor roadways. The variation with time of these sources is reported, and the resultant effect on the climate is considered. In their investigations of the heat exchanges occurring in a whole mine Oakes and Hinsley (1955-56)1 listed and estimated the magnitudes of a large number of heat sources. Some of these, namely autocompression of the air in the shaft, diesel engines and compressed air mains, do not apply to the present investigations, which refer to longwall districts with conveyor transport and no compressed air supply. Others, such as oxidation of coal and, desorption of methane, have not been taken into account in the measurements, but are discussed later. One other source, the importance of which was pointed out by Gracie and Matthews (1975)2, is the coal on the conveyor, a form of strata heat emission, but one which is strongly dependent on the rate of coal production. The heat emissions measured in the present investigations were the energy consumed by electrical machinery, the heat conducted from the strata and the heat lost by the conveyed coal. Most of the emitted heat was removed from the district by the ventilation air, and was measured by continuously recording the enthalpy of the air at fixed stations. The heat removed by service water, supplied for machine cooling and dust suppression, was also measured. Account was of course taken of the relatively minor changes in level around the districts, in terms of the potential energy gained or lost by the coal and by the ventilation air (autocompression).

Jan 1, 1980

By Richard Dunmore

Most coal seams in the United Kingdom contain firedamp, some of which is released into the ventilating airways as the strata around a working relax. The close sequence of seams in carboniferous strata is usually such that seams in adjacent roof and floor strata are also disturbed and potentially large quantities of firedamp can enter the workings from these sources. Since emission depends on the disturbance rate of the seams, higher coal production rates associated with increased mechanisation in the colliery accentuate the problem. Firedamp concentrations in UK underground workings are subject to statutory regulations1. To enable production to continue uninterrupted and in safety the problem is tackled by ventilation engineers by the provision of adequate ventilation in the airway and by the technique of firedamp drainage. In the UK the latter usually takes the form of cross-measure boreholes in roof strata, and increasingly also in floor strata, drilled from the gateroads, especially from return airways. In this way firedamp emitted from disturbed seams in adjacent strata can be extracted directly to the surface or to some safe position outbye, and the airways of the working district are not contaminated. Despite the use of firedamp drainage, considerable quantities of firedamp enter the ventilation air- stream and the concentration levels may be seen to be related to the coal production activites in the district. Figure 1 shows a typical recording of methane concentration obtained by a BM1 methanometer at a position outbye in a return gate where firedamp drainage was operating. The general rise in level throughout the working week may be seen, together with peaks occurring during production shifts and the gradual decay of emission over the weekend. In view of the fact that ventilation and firedamp drainage contribute a considerable proportion to the operating costs of the colliery, it is essential that the ventilation and firedamp drainage requirements be calculated as accurately as possible at the planning stage of the working. In the past the ventilation engineer has had to rely heavily on experience in neighbouring workings but with the development of mines in previously unworked areas there is clearly a need for a methodical approach to meet the requirements of planning. The work described in this paper represents such an approach and it is hoped that successful development and application of the method of prediction will lead to the widespread adoption of a routine procedure. The current development of a method of prediction of firedamp emission into longwall workings is based on the theoretical treatment of gas emission from coal seams proposed by Airey2 in 1971, which was based on his study of gas emission from broken coal 3 which was published in 1968. In common with other experimental methods of prediction, the method enables a calculation to be made of the total make of gas emitted from all sources of gas in the strata per tonne of mined coal.

Jan 1, 1980

By Thomas J. Peters

At the request of the U.S. Bureau of Land Management, the U.S. Bureau of Mines and U.S. Geological Survey conducted a mineral investigation of about 85,470 acres of the Diablo Mountain Wilderness Study Area, to determine mineral resources and mineral resource potential. The study area is in Lake County Oregon, about 45 miles northwest of Lakeview. Fieldwork was conducted in 1986, 1988, and 1989. Geologically, the study area lies on the northwest edge of the Basin and Range physiographic province, a region of fault block-formed mountains and basins characterized by interior drainage. Consolidated rocks are mostly Tertiary basalt and tuffaceous sedimentary rocks. Low-lying areas are covered by Quaternary alluvial-fan, sand-dune, playa, lacustrine, and landslide deposits. The principal structural features are normal faults that have large vertical offsets; these faults are typically concentrated at the margins of large horst and graben. The 34,310-acre "additional," or western part of the area was mainly studied in 1988 and borders the east shore of saline Summer Lake. In this portion of the study area there is potential for undiscovered resources of soda ash, boron compounds, and sodium sulfate as well as by-products, which include magnesium compounds, salts, potash, bromine, lithium, and tungsten associated with brines. No energy or mineral resources were identified in the study area, but brines within lake and playa sediments contain concentrations of chemical components suitable for the production of soda ash, boron compounds, and sodium sulfate. Possible by-products include potash, salts, bromine, lithium, magnesium compounds, and tungsten. Limestone from the study area could be used in the recovery process of brine components or agricultural uses. The study area includes thermal springs and lies within 2 miles of the Summer Lake Known Geothermal Resource Area. Local production of geothermal power by the binary process may be economically feasible. Soda ash, soda ash products, and boric acid are widely used in fluxing metals, and have important applications and markets in the aluminum industry in the northwest and the developing Chinese counterpart. Evaporite commodities are essential to many "backbone" industries, and to many new applications and advanced materials. Markets for brine mineral products appear to be undergoing steady and strong growth, especially in the Pacific Northwest and in Pacific Rim countries. Soda ash produced near the study area would be 55 percent closer by rail to marine export at Portland, Oregon, than trona deposits of the Green River District (Wyoming). Soda ash and a daughter product, caustic soda, and sodium borohydride will receive increased application in the bleaching of paper pulp for environmental reasons. Caustic soda, used in many industrial processes, is more environmentally friendly when derived from soda ash than when produced from sodium chloride salt, because of the absence of a chlorine by-product. Producing soda ash from brines, a type of in situ" mining, can be done with minimal environmental degradation, and mining natural soda ash is environmentally as well as economically preferable to synthetic soda ash production.

Jan 1, 1995

By N. I. Fedunets

The development of variousworks on automatising of operational processes at coal¬extractive and coal-mining enterprises introduces an ever growing necessity to carry out investigations in the wide range of problems concerning the possibitity of joint functioning of different levels, which seems impossible without the use of integrated systems of passing on and prossessing the information on the basis of modern calculating machines and other means of automation. This view on the problem of automation of the enterprise operation presupposes that the structure of automated operational dispatcher control (AODC) should be very closely connected with that of integrated data processing system! In this case the AODC structure from the point of view of automation of its functions (or operations), i.e. from the point of view of informational techniques coinsides in general with the structure of dispersed information and calculating network (ICN), realizing the passing on and processing of the information as well as solving the functional tasks on the basis of local calculating networks of electrical calculating machines with correspond to different levels of AODC and located with regard to productional hierarchy of operational dispatcher control (ODC). In this approach however the use of dispersed, ICN gives opportunity to. realize the possibility of action only with realisation on the CM of algorithmic&apos; task of ODC, when a dispatcher &apos;is the lastto make a final decision.. At the mining enterprises it&apos;s rather difficult for a dispatcher to make a final decision because of the great variaty of productional situations, the presence of contradictions and a number of other subjective factors which often lead to the dispatcher&apos;s mistaks causing great material losses of the enterprise. Hence the necessity arises to automize and formalize making a decision and including the expert systems (ES) into the ontour of AODC, which enables the dispatcher 2to get a neeessary advice in different situations, especially in case ofemergency on the basis of establishing the mechanism of logical layout of non-formalized data base (DB) procedures relying upon the background knowledge and experience of experts. The presently presented in this paper theory of establishing the. ISAODC on the basis of dispersed ICN and expert systems is aimed at trasformation ofoperational&apos; tasks of the higher level into the technological ones, at solving the tasks of diagnostics and forseeing of the process of production, with a dispatcher getting an opportunity of making qualitative and effective decision concerrning the choice of operational impacts in actual time on the basis of generalization the realisation of presented conception must, as has been already mentioned, be accomplished only on the basis of the foundatoin of the united dispersed informational - calculating network of electrical calculating machines.. This includes ICN which would be operating in . different operational contours. . From what has just been said it followsthat the ISAODC Structure will conside with that of DICN (in our case) and hence all the problems of ISAODC structure would be regarded as the questions of DICN.

Jan 1, 1992

By James H. Nulty, Joseph S. Roti

INTRODUCTION During a recent panel discussion on the topic of millwater antiscalants, the question was posed: "What would you like to see in the future?" Answer: &apos;We need an antiscalant that: can be fed without a pump (no maintenance), functions on residual basis (no fault), can be tested for colorimetrically in sub-part-per- mi 11 ion concentrations (no guessing) and will recycle back to the mill from tailings (no cost).” Impossible? Based upon today&apos;s level of technology, such a product does not exist. However, a few of the qualifications listed above are certainly food for future thought. Meanwhile, progress is being made in upgrading three essential elements in this critical area of ore processing. They are: 1. Antiscalant reagent research and development. 2. On-line monitoring programs. 3. Reagent storage and feed systems. The objective of this paper is to out- line the current direction being taken in each of these areas and to present case histories as examples of how theory has been applied in field situations to better improve operations. LABORATORY SCREENING TESTS Recent technical papers have explained the various generic chemical raw materials which are available in the marketplace. While the merits of each chemical component are open to debate, the processes used by most major reagent suppliers to evaluate products in the laboratory are fairly standard. Common techniques incorporate both static and dynamic conditions in the search for the most cost effective product. Static Test Perhaps the most common test procedure used is based upon NACE Standard Test Method 03-74, which is "a laboratory screening test to determine the ability of scale inhibitors to prevent precipitation of calcium sulfate and calcium carbonate from solution." The objective of this test is to measure the change in calcium ion concentration in a container of saturated stock solution prepared from calcium chloride and sodium carbonate which is heated in a water bath. After 12 hours at constant temperature (160°), the mother liquor is pipetted into a beaker and a titrated for gm/l of unprecipitated calcium ion. A simultaneous test is performed on stock solution with antiscalant added. The results are then charted to determine the most effective antiscalant (with respect to preventing precipitation of calcite). It should be noted that this procedure examines only the ability of a chemical additive to retain calcite in a saturated solution where thermal stress is applied. This is commonly referred to as sequestration or stabilization. It has been documented that products which act in this manner (i.e., inorganic and organic phosphates) are good scale inhibitors. However, low molecular weight anionic polymers function in a much different manner by distorting calcite crystal size, shape and charge density, thus preventing scale formation. Because of this different mode of operation, static tests may not prove conclusive where polymers and sequestrants are being compared. The simplicity of this test makes it easily adaptable for field use, although modifications are often made.

Jan 1, 1986

By Donald E. Ralston, William B. Seward, Washington Bernard, William J. Halvorsen, Richard P. Killmeyer, Kelvin K. Wu, Peter Phillips, John N. Murphy, W. H. Mullins

Despite a recession that&apos;s crippling many US industries, and a UMW strike that wasn&apos;t supposed to happen, 1981 was a fairly good year for the coal industry. The big news was record-high exports. Queues of colliers waited for weeks at Hampton Roads, VA, and Baltimore, MD, to load coal destined for Europe and the Far East. Private industry and city governments began scrambling for a share of the export boom, and port developments were planned in nearly every coastal state. Coal exports jumped from about 82 Mt (90 million st) in 1980 to a record 83 Mt (92 million st) in 1981, according to the National Coal Association. NCA projects exports to remain at this level 1982. The following statistics illustrate industry performance in other areas: • Demand by domestic utilities and overseas customers was so strong that total US consumption reached 760 Mt (838 million st) in 1981. DOE predicts another record year in 1982, with consumption forecast at 804 Mt (886 million st). • 1981 production of 730 Mt (805 million st) was just slightly below the record high of 748 Mt (824 million st) in 1980, "a remarkable figure considering that UMW production was shut off for 72 days during 1981," says NCA President Carl Bagge. Production in 1982 should set a new record, 800 Mt (880 million st). • Eastern coal production fell 7.4%, while western production-largely unaffected by the UMW strike-rose almost 10%. NCA says eastern output will rise almost 12% to 538 Mt (593 million st) in 1982, and western production will increase by 6.6% to 266 Mt (293 million st). • Production per manhour is improving. After the strike, weekly production hit the 16 Mt (18 million st) mark and has approached the 18 Mt (20 million st) per week level. •Coal&apos;s share of the US electric generation market increased from 51% in 1980 to 53% in 1981. Electric utilities are coal&apos;s biggest customer, and their financial situation is precarious. Inflation has thwarted necessary expansion plans, while shifting environmental regulations have played havoc with construction costs. Utilities increased power sales by only 2.5% in 1981; NCA predicts that utility coal use will increase 4% in 1982. The Association&apos;s long-range forecast indicates that growth in electricity demand in the immediate future will be much lower than in the 1960s and 70s. Administration Impact Industry optimism following the election of Ronald Reagan and a Republican-dominated Senate didn&apos;t survive the year. One bright spot was the appointment of James Watt as Interior Secretary, and his promises to review surface mining regulations and resume leasing. The administration, however, has declined to support federal eminent domain for slurry pipelines and expects port users to foot most of the bill for expansions and improvements. It has also drastically reduced government financial support for synthetic fuels. The outlook for synfuels-one of three major markets for coal-has

Jan 5, 1982

By David M. Parkes, Mike Fisher

In 1947, coal production in Western Canada was 10 Mt (11 million st) which predominantly came from 120 underground mines. By 1968 production was down to 1 Mt (1.1 million st). Production in 1981 was back up to 29 Mt (32 million st), but only 1.5 Mt (1.7 million st) of the total was from underground mines. In fact, because of competition from open-pit mines, only two underground mines were operating in 1981-the McIntyre and B.C. Coal mines. While several other underground mines are being planned, this article will deal with only the two in production. Geology The geology in Western Canada is frequently complicated; hence only more regular deposits are mined. Coal is generally thick, often steeply pitched and usually soft or well cleated. The 5000-km (3,100-mile) coal belt that runs from the Yukon to Mexico is well developed in Western Canada. It is divided into two distinct regions-the Rocky Mountain Lower Cretaceous coals and the mostly uniform level Upper Cretaceous coals of the plains area to the east. Isolated pockets of coal occur at Vancouver Island and Hat Creek. Hat Creek&apos;s massive 300-m (984-ft) thick geologically recent lignite deposit is 250 km (155 miles) inland. Geography The infrastructure of Western Canada is shown in Fig. 1 along with the locations of the two mines. Both rely on exports of premium low volatile metallurgical coal mostly to the Pacific Rim. The rail distance from both mines to port is 1100 km (684 miles)-accounting for a significant portion of the sales cost of the coal. Mining Methods Two main mining methods are used-room and pillar and hydraulic. Longwall has been attempted and is being considered again. In the past, a simple hand method used gravity to help mine most of the coal on steep pitches. Levels were driven in the seam, rising slightly to assist in draining. On the advance, mining would start at the upper level, usually from adits driven in the side of the mountain. Rooms or angles were driven up 100 m (328 ft) at a suitable angle for the coal to flow down. On retreating out of the level, the pillars would be taken, starting from the highest point, using batteries to stop coal or rock flushing down the room. Batteries consisted of three to six stout wooden posts set at an angle to resist material flow. Coal loading was directly from chutes at the bottom of raises into 1- to 2-t mine cars hauled by diesel or battery locomotives. Often output was restricted by the supply of empty cars. Mechanization began in the 1960s but was difficult in these conditions. Mines such as Canmore and Coleman continued operation into the 1970s but closed because steep seams, high methane emissions, and faulting made mining difficult. McIntyre Mines The McIntyre mines are located 360 km (224 miles) west of Edmonton, Alberta, in the foothills of the

Jan 4, 1983

By Michael L. Cramer, Kunsoo Kim

INTRODUCTION The Columbia River Basalt beneath the Hanford Site is one of several rock types being considered by the U.S. Department of Energy for the first commercial high-level nuclear waste repository. The Basalt Waste Isolation Project (BWIP) has been carrying out a comprehensive investigation to assess the feasibility of constructing a repository in basalt and to develop technology required for repository engineering and site characterization. Development of a nuclear waste repository requires a number of design considerations not usually necessary in the planning of conventional underground structures. Principal among these is the thermally induced stress field resulting from waste emplacement. This thermal stress, super- imposed on the preexisting in situ- and excavation- induced stress fields, needs to be evaluated in the design process due to the sensitivity of waste emplacement density to this factor. The nature of this thermally induced stress field is dependent on several factors, including the temperature distribution around the repository openings, the thermal properties of the rock mass, and the deformability of the rock mass. Re1iable performance of the repository structure is largely determined by the ability to under- stand the above rock mass characteristics and predict the underground opening response to waste emplacement based on this understanding. To gain an understanding of this rock mass behavior in a repository condition, the BWIP has conducted a large-scale heated block test in basalt at the Near-Surf ace Test Facility (NSTF) on the Hanford Site. This paper presents the results from the elevated temperature phase of the block test in which the thermal, mechanical, and thermo- mechanical response of the closely jointed basalt rock mass was evaluated as a function of stress and loading orientation at varying temperatures up to 200 oC. SITE GEOLOGY The NSTF test areas, including the block test site, are located in the entablature zone of the Pomona basalt f low. The Pomona flow entablature is structurally typical of most flows in the Columbia River Basalt Group in that it is characterized by long, undulating, polygonal columns up to 2.4 m in length. Columns vary from 15 to 30 cm in diameter, averaging about 20 cm. Over their length, columns can vary in dip by 150 to 200. The column structures are dissected by subhorizontal cross joints that are generally discontinuous and seldom cross columnar joint boundaries. Spacing of the cross joints also averages about 20 cm. A representative joint pattern taken from mapping of the tunnel wall near the block test site is shown in figure 1. Application of the Rock Mass Rating System (Bieniawski, l979) to the block test site yielded an overall value of 62. The corresponding value from application of the Q-system (Barton et al., 1974) is 8.7 (Cramer and Kim,1985). The NSTF is approximately 50 m below the surf ace. However, this depth is sti1l above the water table and conditions are generally dry. Interstitial water exists as the result of percolation of meteoric water from the surface.

Jan 1, 1986

By Richard B. Muter

In determining the layout or combination of unit operations to be utilized in a modem preparation plant, several factors must be taken into account. Among these are the washability data, size consist data, the particular market or markets for the coal, the mining method used, the degree to which the coal must be prepared, expected fluctuations in raw coal and markets, and the overall economics of the process. In addition, there are the health and safety aspects of coal utilization and preparation. Chapter 1 of this book discusses the characteristics of coal and coal markets, while Chapter 2 deals with the design of preparation plants. Once the unit operations for a particular plant have been decided upon, these arrangement may be depicted by using a flowsheet. A coal preparation plant flowsheet is a two-dimensional graphical representation of the sequence of operations, the arrangement of machinery, and the flow of materials for a particular coal cleaning operation. The flowsheet is usually qualitative in nature, becoming more quantitative as material balances, water balances, numbers and sizes of various pieces of equipment, etc. are added. The usual convention in preparing a flowsheet is to show the feed flowing from left to right, starting at the top left and working down as in an all-gravity process. Various symbols are used for each unit operation as depicted on a flowsheet, and Fig. 14-37 is a compilation of the commonly accepted international standard symbols for each unit operation. Fig. 14-38 through Fig. 14-40 are flowsheets of various coal cleaning processes as published in the January 1976 issue of Coal Age. Each process has its own unique features which, among others, depend upon the coal being cleaned, the coal market, and the mining operation. Fig. 14-38, for example, is a flow diagram for a jig plant preparing steam coal from two strip mines located a few miles apart. In addition to the jig washer, cyclones are used to treat the finer size coal. Fig. 14-39 is a flowsheet of a heavy media plant using a heavy media washer for + %-in. (+6.3-mm) coal, froth flotation for -28 Tyler mesh (-0.6 mm) material, and dewatering phase units (dwp) for Y, in. x 28 Tyler mesh (6.3 x 0.6-mm) cleaning of steam coal. In addition, thermal drying, centrifuge drying, and the use of vacuum disk filters and thickeners for dewatering are depicted. Fig. 14-40 is a flowsheet of a plant producing metallurgical coal. In addition to the heavy media and froth flotation circuits, a tabling circuit is also employed. These flowsheets are intended to give the reader an example of how the various unit operations for coal cleaning may be linked together. Figs. 14-41 and 14-42 illustrate the conversion of block diagrams into an IS0 prescribed set of symbols.

Jan 1, 1991