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By W. E. Bearce

The recent concern for the recovery of even the finest fractions of coal preparation plant feeds has created needs for new equipment. Thermal dryers currently available have difficulty reducing the filter cake moisture on down to the desired point. This difficulty involves 1) the physical problem of handling this wet, sticky material and making each particle accessible to the hot gases, and 2) removing the extremely fine dried particles from the cooled gases. The dryer described in this paper handles the first of these problems, and eliminates the second by the substitution of steel balls for air as the heat transfer medium. As a general statement, the cost of preparing coal for market is inversely proportional to the size of the lumps or particles. In the not too distant past the finer sizes were discarded as being uneconomical to treat and this practice might be still continuing if it were not for the insistence of the public that this material be kept out of our streams. This being the only cost-free outlet, the disposal of fine refuse became a significant item of expense in every preparation plant. When this happened it was only reasonable to look to the recovery of the values in this material, not only to defray this expense, but to reduce the tonnage which had to be wasted. There seemed to be no device available suited to reducing the moisture in this material to what the market requires. All of the fine coal dryers commonly employed make use of air to bring the heat of combustion into contact with the coal particles. Any device employing air as a heat transfer medium, has the following two strikes against it in attempting to handle 28 mesh by 0 filter cake: 1) It is practically impossible to bring the hot air into intimate contact with the enormous number of fine particles stuck together in the cake. 2) Even if this can be effected, it is impossible to remove the dried coal from the air. A wet scrubber must be employed, the discharge of which must go back to the filter and dryer as a circulating load of the most objectionable sort, being the extreme fines. Various compromises have been used to cope with the problem of drying this relatively small tonnage of high moisture coal. Often some of the dried coal is mixed back with the feed to the dryer in an attempt to make the mixture more handleable and more available to contact with the hot air than the filter cake alone would be. This may help to some extent with the first problem but it does nothing about the second and a dryer operated in this way has to be oversized. Sometimes the filter cake is mixed with a coarser fraction for the same reason. Here again, while this may make a mixture (the particles of which are more readily brought into contact with the hot air), the problem of getting the fine coal out of the air remains and a larger percentage of the total coal to be dried is handled by the most expensive means. Mechanical dewatering has accounted for more and more of total coal drying lately and in many cases would satisfactorily handle the whole load, down to 28 mesh, if there were a means available for doing a good job on the —28 mesh by itself. The dryer which is the subject of this paper was designed expressly for doing this job — drying fine material by itself regardless of how wet it may be to start with, how fine it may be, or how gummy and sticky these properties in combination may make it. The essential features are depicted in Fig. 1. A rotating cylinder is supported on a frame in such a way that one end may be raised or lowered by hy-

Jan 1, 1968

By W. B. Jamison

Man's material advance from one level of civilization to the next has involved the development of new, more useful tools and the utilization of energy greater than he alone could produce. These two essentials for material progress have developed at an ever increasing rate in the last century. While the first really successful use of mechanical energy derived from burning fuel was in James Watt's steam engine which pumped water from coal mines in Cornwall about 1770, there was still very little use of mechanical energy in mining by 1900 except for pumping and hoisting. Compressed air, having some of the characteristics of steam and which operated engines not unlike steam engines or steam hammers, powered the first attempts to mine coal with energy greater than human muscle could provide. When direct current-powered traction cars replaced horse-drawn transportation cars on city streets, it was natural that direct current locomotives should begin to replace animal and human power in mine haulage. Slowly, trolley and battery locomotives were developed and introduced. With d-c systems being established for haulage, it was natural that improved cutting machines should be developed and powered by d-c motors. Thus began the years of

Jan 1, 1961

By S. A. Braley

THE first commercial production of bituminous coal in the United States was in 1820, and formation of acid in the areas from which the coal was removed began at that time. Thus it is 130 years since the development of mine acid began. The coal operator has had to deal with the acid water produced in his mines, and his losses from corrosion of pumps and equipment have been tremendous. Other industries and the public have been forced to spend large amounts for their water requirements because of the acid discharge from the mines. There are many contradictions in the literature on this subject probably because of variations in the analytical procedures followed and differences in the methods of calculation and reporting. The opinions previously expressed may be summarized as follows: The production of mine acid has been looked upon as a simple chemical reaction wherein the oxygen of the air has reacted with pyrite to produce ferrous sulphate and free sulphuric acid. This reaction appears in many chemistry texts and generally has not been considered worthy of further study. Since the measurement of hydrogen ion concentration or activity was developed and expressed in terms of pH, indicator and instrumental methods for pH determination have become readily available, making it natural to think of it as a measure of acid content. Special purpose analytical methods applicable to potable waters, sewage effluents, and boiler waters, based in general upon the determination of what may be regarded as impurities, have been applied to more concentrated solutions constituting mine discharges, without reference to the fact that these specialized procedures had been developed to serve as operating guides for water-treatment plants. It has been assumed that the pounds of acid deliv- Table I. Showing pH and Titratable Acid of Varying Concentrations of Ferrous and Ferric Sulphates Ferric Ferrous ion Acldfty Ion Acidity (Fe+++) ppm CaCO3 We++) ppm CaCOa ppm PA equiv. PPm pH equiv. 11.200 1.80 30,000 11,200 3.98 20,000 5,600 2.01 15,000 5,600 4.15 10,000 2,800 2.21 7,500 2,800 4.41 5,000 1.120 2.48 3,000 1,400 4.68 2.500 560 2.60 1,500 140 5.00 250 280 2.65 750 28 5.00 50 112 2.81 300 56 3.07 150 28 3.38 75 ered by a mine was roughly a constant varying with the number of acres mined, and that the concentration of the acid in the drainage was inversely proportional to the volume of flow. Increased flow was thought to result in a dilution of the acid produced. When the Sanitary Water Board of the Commonwealth of Pennsylvania established a fellowship at Mellon Institute, the first approach to the problem was to review some of the chemical reactions involved and their relation to acid production. Since the sulphates of iron were among the principal constituents of coal-mine drainage, the first studies were based upon solutions of chemically pure ferrous and ferric sulphate. These are both neutral salts, that is, the iron and the sulphate are equivalent, but they produce acid reactions in water solution because as salts of strong acids with weak bases, they hydrolyze to form acid solutions. Solutions of each salt, of varying concentrations, were prepared and the pH, titrable acidity and iron and sulphate

Jan 1, 1952

By J. W. Leonard

This expanded and more detailed supplement to earlier published work1 is offered as a means to demonstrate the high empirical correlation which exists between the readily measured coal bench scale test parameters of volatile matter (d & a.f.), Hard-grove grindability index (HGI), Koppers cone bulk density (BD) and pulverization level (% passing a 1/8-in. screen) to the less readily measured full scale coke stability index. Thus, Fig. 1, 2 and 3 represent the correlated test data contained in Tables I, II and III respectively. Ninety-six coals are correlated so that 97% of the resultant cokes fall within two points and 100% fall within three points of the actual tested coke stability indices. The main generalizations which can be drawn from the correlation of these coals are as follows: 1) The coke stability indices for higher bulk density coals (46.0 BD to 50.6 BD) appear to be largely influenced by differing combinations of factors for coals above and below a pivotal point of approximately 36% to 37% volatile matter. Below this pivotal volatile matter coke stability index appears to be influenced largely as a result of changes to the coal volatile matter and pulverization level. While above this pivotal volatile matter coke stability indices appear to be dependent on the coal hardness, as measured by the HGI, and to some extent upon coal volatile matter. Pulverization appears to have little influence on the coke stabilities of the harder coals (52 HGI to 58 HGI) of greater than 36% to 37% volatile matter. 2) The coke stability indices for lower bulk density coals (41 BD to 45.9 BD) appears to be largely influenced by changes to coal volatile matter and pulverization level for softer coals above approximately 58 HGI. 3) The main empirical observation to be drawn from these results is that for coals above approximately 36% to 37% volatile matter the quantities of hard and soft constituents as measured by the HGI are important when defining coal rank for correlation to the coke stability index.

Jan 1, 1965

By J. Quigley

As the outcrop mines in the West developed into underground operations, systems of ground support were gradually evolved. In the early coal mines there was little need for support except near the dirt line in portals, where stone masonry was common. Where the top was shaley or broken, native pine props with light cross bars and legs furnished enough support even in Utah's 25-ft coal seams. As depth of workings increased. roofs and backs of the same general nature as those near the surface became more and more unstable and required more and more support. Some coal airways show this tendency very clearly. From the surface down the same type of roof shows deterioration which an experienced eye can translate into a measure of depth under surface rather than change in rock characteristics. Rock bolts, developed by various companies and by the U. S. Bureau of Mines, have become an effective substitute for timber in sections of some metal and nonmetal mines formerly requiring escessive timber support, and further use of war surplus landing mats, chain link fencing, and a new punched channel developed by one of the steel companies has enabled other mines to operate deposits where costs of timber and lack of clearance for timber support would have prohibited mining. The block caving mines have made extensive use of reinforced concrete underground to achieve similar ends under difficult conditions. Steel sets are standard in many Bureau of Reclamation projects, although these are usually covered in with concrete to make the permanent structures the Bureau's reclamation projects require. But the use of steel in mining operations is limited and has been confined principally to the iron ore mines of Michigan, Wisconsin, and Minnesota. Some mines have installed used rail as posts, caps, and crossbars, but a rail section is not suited for load carrying, and used rails are generally brittle. having a tendency to fail without warning when overloaded. European mines were the first to reach the size of worked out areas and depths of cover resulting in major roof problems. The Europeans resorted to pack walls and masonry walls, in conjunction with timber arched sets. rail arches, and combination timber and rail and steel arches. The give in these pack walls and wooden blocking was supplemented by a hinge in the center of the arch. This design is called an articulated arch Through various refinements of this principle of the support giving graduallv with the load. Toussaint-Heintzmann developed the yielding or sliding arches, in which yield is accomplished by friction in the overlapping joints of the arch. This type has gained widespread acceptance in the Ruhr and Lorraine Basin and is being manufactured by Bethlehem Steel for sale in this country. In North America the anthracite mines in Pennsylvania, followed by certain iron ore mines in upper Michigan and Canada, were the first to employ these arches to any extent. The practice was later adopted by Kennecott at Ruth, Nev., and by others. Despite high initial cost, the use of these arches is growing in many parts of the country because of their suitability in heavy ground. In its present form of manufacture the yield-able arch consists of open U-shaped rolled section with heavy beads on the edge. The open edge of the U is placed toward the wall. The section nests in another section of the same dimensions, and an arch can be built up from rolled radii and tangents of various weights and lengths. Sections are fastened together by U bolts and saddles. The lap on the joint varies from 12 to 24 in., and ordinarily the bolts are tightened with a 1-in. drive air wrench. The arches are spaced with channel struts held by J bolts and saddles. Sections can also be obtained that are composed of various combinations of radii and tangents and true circles. The joints can be placed to bear against anticipated loads and asymmetrical loads imposed by dipping strata. In the arches now being manufactured clearance widths up to 19 ft are obtainable in weights of sections from 9 to 30 lb per ft. The circular cross sections are available in the same weights ranging from 8 to 16 ft diam. At present most of the arches sold are supplied only in carload lots. It is hoped that demand will grow so that distributors can stock various weights and sections to give small operators a chance to try this new type of rock support under their own particular conditions. Several excellent papers have discussed the properties of various sections now manufactured, the dimensions of the sets obtainable, and their application under widely differing conditions. The present article will describe the methods and results of a special use of the arches at Kaiser Steel mine No. 3. Sunnyside, Utah. Problem at Mine No. 3 : In 1953 Kaiser Steel Corp. laid out Sunnyside mine No. 3 to recover coking coal left by the previous operator, Utah Fuel Co.. below workings that had been abandoned in 1928. Two seams had been worked, the upper and lower, separated by 30 to 42 ft of rock. Approximately 10 million tons of coal had been extracted from this area some 3000 ft down the itch from the outcrop to a 1500-ft depth of cover. The mine had been opened by slopes in both upper and lower seams. Sometime in the late 1920's the lower slope

Jan 1, 1960

By R. H. Swallow, George Hess

The purpose of presenting this paper is to show: 1. The applications of aerial photography to the map requirements of strip mining. 2. The methods and procedures for producing accurate aerial contour maps. 3. The advantages of aerial contour mapping over stadia or plane table methods. Introduction Aerial photography was once a crude, uncertain tool. 'Today it is a precision mapping instrument which saves important time and money for strip mining and other industry. Aerial photography began in the boxkite days of aviation. The Army first used it for rough reconnaissance photographs. Then, in the early Twenties, industry began to use aerial photographs of their plants and facilities for advertising or annual reports. The quality of these pictures was not very good by today's standards, but this work began the development of better cameras, better lenses, better photographic techniques. Then, some farsighted engineers and aerial photographers found that if the camera's axis were vertical to the earth's surface, these pictures could be the source of valuable mapping information. By viewing these pictures in stereoscopic pairs, accurate measurements could be made and contour maps produced from this aerial photographic base. So the aerial mapping business developed, using the plane and aerial camera to collect the data, working with skilled ground parties who established a network of ground control for this map information photographed from the air. Over the following years, a whole new science developed, known as photogrammetry. Precision cameras and lenses were perfected, and the narrow view of the surveyor's transit was broadened to the vast perspective of the mapping plane. Stereoscopic Coverage Obtained Today in making aerial topographic maps, aerial photographs are exposed so that full stereoscopic coverage for the survey area is secured. To obtain true elevations these photographs are processed through a series of precision measuring and contouring instruments which employ a principle similar to the military range finder. Here is how aerial mapping is being used by your industry. Topographic maps are but one of three products from aerial photographic surveys. First, there are the individual photographs. These are taken in flight strips, exposed with an overlap, just as shingle on a roof overlap. Exposures are overlapped from 55 to 60 pct along the line of flight so that a large part of the area in one picture also appears on the adjacent photograph. Successive strips are flown with a side-lap of between 20 and 25 pct. Each point on the ground thus will be somewhere near the center of at least one photograph in the series. They may be studied under a stereoscope and relief seen in a third dimensional effect. Geologic features that may go unobserved in ground exploration are often detected under stereoscopic inspection. Moreover, though field checking is not obviated with these aerial photographs, it is directed toward critical areas and toward worthwhile anomalies. Aerial Photographic Mosaics The second product of an aerial survey is the photographic mosaic. This is an assembly of carefully matched aerial photographs in mosaic form. Although not as accurate as the topographic map, it does reveal a wealth of cultural detail in true relative scale. Since the mapping plane is flown over the area at a fixed altitude, the peaks and valleys are pictured at different scales. During compilation these scale differences are compensated by adjusting one vertical picture to another according to plotted ground control positions. Only extreme variations in ground elevation cannot be

Jan 1, 1950

By R. H. Swallow, George Hess

C. G. BALL*—These maps are obvi-~,usly quite helpful in many types of mining engineering, but I want to find out if the prints which you obtain in the first step toward making any aerial contour map have already been corrected for tilt. Is that done afterwards only if you are going into the process of making a contour map ? GEORGE HESS (authors' reply)— The tilt correction is the first step in either compiling the mosaic or compiling the total map. They must be corrected. C. G. BALL—But are the contact prints themselves corrected for tilt? GEORGE HESS—They are not corrected. E. R. KAISER†—I enjoyed this paper very much, and am glad to learn that the methods have a high degree of accuracy and are in use in the coal industry. Accuracy in dimensions is understandable from the paper. Would you tell us how accuracy can be obtained in the vertical dimensions by means of the photographs so as to permit drawing contour lines? GEORGE HESS—To explain our process, even in a limited way, is quite a large order. However, the basic principle involved is the measurement of parallax by viewing two overlapping photographs stereoscopically. It would almost be necessary to have the various instruments here on view to adequately explain their operation. E. R. KAISER—Are the elevations for reference obtained at more than four points ? GEORGE HESS—The pattern of control is four points obtained in the corners. Primarily. that is to restitute the picture and horizontalize it. The vertical control points are read ill on the stereometer and a direct comparison is made while draw- ing the contour lines. You do have to have those four points. But one stereoscopic pair of 5 ft contours has some 340 acres on it, consequently for every four points picked up 340 acres are contoured. E. R. KAISER—How does your method compare with that used by the military reconnaissance during World War II? GEORGE HESS—The Army did not get into this refined mapping. Most of their work has been reconnaissance mapping by taking one vertical and two oblique views. Basically, the same principle is used. Our maps are for engineering purposes. The Army was primarily interested in reconnaissance. E. R. KAISER—You could take more than 48 pictures in one flight? GEORGE HESS—Yes, the Brock camera accomodates 48 pictures per magazine. We are limited by the hours of greatest light, though, 2 hr before high noon, and 2 hr after high noon. Shadows do affect the picture. The shadows can-

Jan 1, 1950

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines or phosphate ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pct. When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900" to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By H. Pew, J. H. Field

To alleviate pollution more restrictive legislation is being enacted, either limiting emission of pollutants or the type of fuel that can be utilized. The nature and magnitude of air pollution problems affecting the mining, preparation, coking and combustion of coal are described. Methods for combatting particulate emissions by use of mechanical separators and electrostatic pre-cipitors are discussed. Proposed methods to meet the problem of gaseous emissions currently receiving considerable attention are described, with special emphasis on methods to decrease pollution by sulfur oxides. Concern about air pollution goes back several centuries, but until very recently most effort has been aimed at coal smoke and other visible pollutants. The classic example of a 'successful' campaign for smoke abatement and control is the fruitful combined effort of the city of Pittsburgh and its surrounding Allegheny County, which eventually led to the reconstruction of downtown Pittsburgh at an estimated cost of one billion dollars. Historically, the city's downtown Golden Triangle district had been afflicted by pollutants evolving from steel mills, from a variety of other industries, and from railroad locomotives. Efforts to alleviate the situation prior to 1943 were virtually ineffective. In 1945, however, a comprehensive redevelopment plan was prepared and backed by state authority. Within a few years a clean, modern metropolis has evolved where once stood America's famous 'smoky city.' But the victory in Pittsburgh, as in various other American cities, has not solved the national problem. Current estimates indicate that 133 million tons annually of air pollutants from all sources still are emitted annually into the atmosphere above the United States. About 10% of this annual effluent is particulate matter so that most of the remaining pollution problems will be solved only when other effluents are reduced. Essentially, these are sulfur oxides, nitro- gen oxides, hydrocarbons, and carbon monoxide. Over the years, both states and local communities have tended to increase the restrictions on smoke and fly ash — problems mostly of concern in the combustion of coal. Prior to the middle 1950's, ordinances sometimes permitted emissions of smoke equivalent to as much as No. 3 on the Ringlemann scale. Since 1956, no ordinance has been passed which allows smoke of greater than No. 2. Under today's conditions of improved fuels, equipment and practice, a few communities have passed laws prohibiting emission of smoke of any density darker than Ringlemann No. 1. The majority of existing laws on fly ash emission in the U.S. limit emissions equivalent to 0.85 lb of fly ash per 1000 Ib of flue gas. In recent years, however, regulations which have been adopted give cognizance to the higher level of performance now obtainable with improved equipment. A comparison of the restrictions of five codes adopted since 1960 is given in Table I. The most stringent of these is the one for New York City which provides for a maximum emission of 0.6 lb fly ash per million Btu heat release (equivalent to roughly 0.51 lb/1000 lb of flue gas). The first comprehensive effort to restrict the emission of SO2 resulted from the passage of a 1937 law in St. Louis. This regulation stipulated that coal containing in excess of 23% volatile matter and 2% sulfur must be washed, thereby presumably producing some effective reduction in the input sulfur content. This was followed in 1949 by a Los Angeles County law which prohibited the emission of SO2 in concentrations greater than 0.2%. Most SO2-restrictive legislation passed since that date has been based on this limiting 0.2% SO2 by volume, although modifications are occasionally permitted under selected conditions, sometimes based on the fact that certain limiting ground concentrations are not exceeded — such as in the rules adopted by the San Francisco Bay area. To date, no legislation has been passed in the U.S. to limit the generation of nitrogen oxides from the combustion of fossil fuels. However, such oxides are considered to be of potential importance in air pollution control because of their possible detrimental effects on health and their reported role in the formation of photochemical smog. Interest in reducing oxides of nitrogen from powerplant and auto exhausts is increasing and regulations limiting their quantity can be expected in the future.

Jan 1, 1968

By Louis C. McCabe

The control of dusts and fumes of submicron size is involved in many process industries. This paper presents in tabular form the quantitative data from a number of metallurgical operations and discusses allowable discharge in terms of effluent-weight ratio. Typical installations for recovery of stack effluents are cited. THERE are many sources of smoke, fumes, gases, and dusts that may cause undesirable contamination in an industrial area. Fumes and gases are the result of chemical action or volatilization, while dusts are produced in processes of disintegration and dispersion. Commonly, smokes are suspended particles formed by combustion or destructive distillation. Methods for controlling smoke are generally known, and great improvements have been made in equipment for removing fine dusts and fumes from gases. In most of the industrial centers of the country, the chief duty of the air-pollution abatement agency is the elimination or reduction of emissions resulting from improper combustion. Correct methods of firing, proper fuel selection, and installation of properly designed equipment may accomplish this. It is becoming more widely recognized that much of the contamination of an industrial atmosphere is caused by dusts, fumes, and gases that are not related to the process of combustion. As such emissions are usually a functional part of plant operation, they rarely can be prevented even by better design and careful operation. When dusts, fumes, and gases are an integral part of the operation, mechanical equipment must be designed to remove these emissions. Most of the dusts and fumes that should be controlled are less than 1 micron in size. Control of such small particles is largely a new and untried field in many industrial areas. However, where conditions have been critical, effective control equipment for dusts, fumes, and gases has been in operation for many years. The problem is not one of smoke abatement and requires an entirely different set of standards for effective solution. Identical standards need not be applied to all industrial areas because of differences in meteorology, physiography, plant location, plant density, and nature of stack effluents. In this paper, discussion of control of atmospheric contamination by industrial plants refers specifically to the Los Angeles area, but the method of approach is applicable to any area where similar problems are encountered. Quantities of particulate matter reported in analysis of stack effluents indicate the need for standards of emission. Many types of dusts and fumes that contribute materially to the air-pollution problem are discharged into the atmosphere at a lower rate than 0.40 grain per cu ft at standard conditions. On the other hand, some stacks discharge at a higher

Jan 1, 1951

By Louis C. McCabe

The control of dusts and fumes of submicron size is involved in many process industries. This paper presents in tabular form the quantitative data from a number of metallurgical operations and discusses allowable discharge in terms of effluent-weight ratio. Typical installations for recovery of stack effluents are cited. THERE are many sources of smoke, fumes, gases, and dusts that may cause undesirable contamination in an industrial area. Fumes and gases are the result of chemical action or volatilization, while dusts are produced in processes of disintegration and dispersion. Commonly, smokes are suspended particles formed by combustion or destructive distillation. Methods for controlling smoke are generally known, and great improvements have been made in equipment for removing fine dusts and fumes from gases. In most of the industrial centers of the country, the chief duty of the air-pollution abatement agency is the elimination or reduction of emissions resulting from improper combustion. Correct methods of firing, proper fuel selection, and installation of properly designed equipment may accomplish this. It is becoming more widely recognized that much of the contamination of an industrial atmosphere is caused by dusts, fumes, and gases that are not related to the process of combustion. As such emissions are usually a functional part of plant operation, they rarely can be prevented even by better design and careful operation. When dusts, fumes, and gases are an integral part of the operation, mechanical equipment must be designed to remove these emissions. Most of the dusts and fumes that should be controlled are less than 1 micron in size. Control of such small particles is largely a new and untried field in many industrial areas. However, where conditions have been critical, effective control equipment for dusts, fumes, and gases has been in operation for many years. The problem is not one of smoke abatement and requires an entirely different set of standards for effective solution. Identical standards need not be applied to all industrial areas because of differences in meteorology, physiography, plant location, plant density, and nature of stack effluents. In this paper, discussion of control of atmospheric contamination by industrial plants refers specifically to the Los Angeles area, but the method of approach is applicable to any area where similar problems are encountered. Quantities of particulate matter reported in analysis of stack effluents indicate the need for standards of emission. Many types of dusts and fumes that contribute materially to the air-pollution problem are discharged into the atmosphere at a lower rate than 0.40 grain per cu ft at standard conditions. On the other hand, some stacks discharge at a higher

Jan 1, 1951

By Orville R. Lyons

CENTRIFUGAL force has been utilized for the dewatering of fine coal for over 50 years by means of machines commonly called centrifugal dryers. In any centrifuge the coal and water are subjected to a spinning action which usually inctedtocreases in intensity as the coal progresses through the machine. This spinning action, or the centrifugal force that it induces, tears water away from coal particles and produces a dewatered coal. The effectiveness of the dewatering action for any particular machine is governed by the size-consist of the coal and the centrifugal force imparted to the water on the coal. Four makes of centrifuge are currently being used to dewater fine bituminous coal in the United States. These include the vertical, basket-screen C.M.I. (Centrifugal and Mechanical Industries) Carpenter, Reineveld, and the horizontal solid-bowl Bird. The C.M.I. and Reineveld centrifuges are similar in construction and are variations of the Elmore centrifuge, varying principally in the slope of the basket, operating speed, size of perforations in the basket, and diameter of the basket. The same operating description suffices for both. These centrifuges are made up of two rotating elements, an outside conical screen frame and an inside solid cone which carries spiral pusher blades. The screen frame is supported by a hollow shaft, and the solid cone by a spindle shaft passing through the hollow shaft. Both the screen frame and the solid cone rotate at high speeds; the solid cone, with its scraping flights, rotating somewhat more slowly than the screen. The wet coal enters the machine at the top, falls on the cone, and is thrown against the screen by centrifugal force. The coal slides down the screen until it meets the upper end of the flights, where it is moved slowly along the flights until discharged at the bottom. The elfluent passes through the sieve basket or screen and discharges into a launder. The Carpenter centrifuge consists of a conical rotating element with a vertical axis, built up of three rows of perforated stainless-steel plates or stainless-steel wedge-wire or round-wire sections cut and rolled to conform to the surface of the cone. Wet material is delivered to a conical hopper inside the centrifuge casing and is distributed in a thin, even layer over the inner surface of the cone by a disk mounted on the rotor shaft. Impacting of the coal against the screen and the centrifugal force exerted as the material moves downwards creates a dewatering effect. The moisture is forced through the coal bed and then through the perforated plates of the basket, collecting in a circular trough built into the centrifuge. The effluent is conducted from the trough through the centrifuge casing by outlet pipes. The dewatered coal is discharged over the inside bottom edge of the rotor. The Bird centrifuge consists of a tank or truncated conical shell, which is revolved at the desired speed by means of a drive sheave. A screw conveyor rotates inside the cone or bowl at a slightly lower speed in the same direction of rotation. The feed entrance, in the center of the large end of the truncated cone, is high enough to allow formation of a pool of slurry. Adjustable effluent-discharge ports are located in the large end of the bowl so the level of the liquid can be regulated. The solids are moved forward by the screw conveyor as fast as deposited and carried above the level of the pool for an interval before leaving the bowl. In the past these four makes of centrifuge could be compared only by indirect means. The author published a rough comparison method in 1949' but

Jan 1, 1952

By Harlan W. Nelson, James B. Purdy

The great majority of stokers used in residential heating installations are of the clinkering type. Because of inherent characteristics of the underfeed combustion process as it occurs in these small stokers, and because of the wide range of operating conditions encountered in residential heating schedules, not all bituminous coals can be used successfully in residential underfeed units. The problem of providing consumer satisfaction has been attacked by the joint efforts of the stoker manufacturers and the coal industry; the stoker manufacturers have been concerned with improving their product to permit the use of a wider range of coals, and coal operators have made every effort to market selected and better prepared coals. In view of the wide range of physical characteristics and chemical properties of bituminous coals, the problems are real and are receiving increased attention. Some twelve years ago a laboratory method was developed at Battelle Memorial Institute for the relative evaluation of coals for use in residential underfeed stokers. The method was developed in conjunction with work being done at that time for Bituminous Coal Research, Inc., and for individual coal producers, to obtain information on the characteristics of coal that determined the successful performance of the small stoker. Details of the equipment and test procedure were given by R. A. Sherman1 in a paper presented at the University of Illinois Fifth Short Course in Coal Utilization. During the past several years, the evaluation of over forty different coals has been determined by investigations sponsored at Battelle by coal operators in the eastern producing districts. The purpose of this paper is to compare the results and conclusions derived from the various tests with marketing experiences of the operators, and to discuss briefly the relationship of test results to standard laboratory determinations. Equipment Used in Tests Fig 1 shows schematically the equipment used in the evaluation tests. The stoker is a commercial model of conventional design, having a nominal feed rate of 30 Ib per hr. A special laboratory furnace is used, which is tightly sealed to prevent the entrance of secondary air. The complete stoker-furnace assembly is mounted on platform scales which permits the measurement of combustible weight losses occurring during the test. The furnace stack, in which are a thermocouple, flue-gas sampling tube, and smoke meter, is connected to the laboratory exhaust breeching through a friction-less connection. Through use of an auxiliary blower system, provision is made for the measurement of the air delivered by the stoker fan to the windbox. A water seal in the duct work eliminates inter-

Jan 1, 1950

By Harlan W. Nelson, James B. Purdy

A study of data obtained during laboratory tests to determine the suitability of bituminous coals for use in residential underfeed stokers of the clinkering type has led to the following general conclusions: 1. Performance data obtained during the test furnish useful and practical information upon which to judge the performance of stoker coals. 2. Conclusions derived from the results of evaluation tests are in general agreement with use experience of the coals in the field. The method is of practical use in this regard. 3. Few direct relationships were found between data furnished by the evaluation test and determinations by standard laboratory tests. This is particularly true with regard to coke formation and clinker formation, two of the most important factors in determining the degree of satisfaction to be obtained in applications of the small underfeed stoker. It has not been the intention in this paper to prescribe this evaluation test as the best method for the evaluation of coals for use in residential stokers. That there are other test methods having similar objectives serves to point out the necessity for full-scale laboratory tests. At the present time, a standard test procedure for the evaluation of coals for use in residential stokers is not available. For the past several years a joint committee, appointed from the membership of the Residential Stoker Committee of Bitulninous Coal Research, Inc., and from the Engineer- ing and Research Committee of The Stoker Manufacturers Association, has been working to develop a standard procedure for testing and evaluating bituminous stoker coals. This procedure has been completed to the extent that it has been drawn up in tentative form. It is now planned to install the equipment in several laboratories and to run check tests, using portions of identical lots of coal, to determine the degree of reproducibility offered by the method. Completion of this standard test procedure should greatly facilitate the evaluation of coals for use in residential stokers. Acknowledgments The cooperation of coal operators sponsoring the coal evaluation tests, in granting permission to use data and results from their tests, is gratefully acknowledged. Acknowledgment is also made to Carroll F. Hardy, Chief Engineer, Appalachian Coals, Inc., for his interest and timely suggestions on the preparation of this paper. Special thanks are due to Ralph Sherman, Assistant Director, Battelle Memorial Institute, who was largely responsible for development of the evaluation procedure, for helpful suggestions and advice. References I. H. A. Sherman: The Evaluation of Coal for Use in Domestic Stokers. Univ. of Ill., Eng. Experiment Station, Circular No. 39 (1939). 2. . Q. Shotts: The Relation of Free- swelling Indexes to Other Characteristics of Some Alabama Domestic Stoker Coals. TP 2314, Coal Tech., (Feb. 1948); Trans. AIME (1948) 177, 502. 3. . J. Helfinstine and C. C. Boley: Correlation of Domestic Stoker Combustion with Laboratory Tests and Types of Fuels 11. Combustion Tests and Preparation Studies of Representative Illinois Coals. Ill. State Geol. Survey, Re1. of Zntlestigations No. 120 (19467 ) p. 35. DISCUSSION E. R. Kaiser*—AS a former fuel engineer assigned to residential stokers at Battelle, the writer has read with considerable interest the paper by Messrs. I'urdy and Nelson. It is gratifying to note that earlier test techniques have been developed and that criteria found important in the years 1935 to 1938 are still important in judging the performance of stoker coals for residential heating. The authors have plotted a number of the primary coal data singly against performance data of the stoker in an effort to establish relationships. Unfortunately, wide scattering of points usually resulted, which did not illustrate what experience generally indicates to be the case. For example, Fig 3 indicates almost no trend in the percentage of released ash converted to clinker with change in ash-softening temperature. Other factors, such as fuel bed temperatures and zones of heat release must have influenced the results. In the present state of our knowledge, we cannot explain all of the reasons why suitable stoker coals perform satisfactorily, nor why one satisfactory coal may be better than another. It would therefore

Jan 1, 1950

By J Price, M. R. Geer, H. F. Yancey

COAL mine operators recognize coal as an abrasive material, because the wear of drilling, cutting, and conveying equipment is reflected as a cost item for replacement of parts. Similarly, industrial consumers of coal experience abrasive wear on all coal-handling equipment. Operators of pulverized fuel plants are doubtless most keenly aware of the abrasiveness of coal, because under the high contact pressures developed between coal and metal in pulverizers, abrasive wear is increased many fold. Moreover, experience in operating pulverized fuel plants has demonstrated that some coals are much more abrasive than others. Hardgrove' stated that maintenance costs entailed by the wear of grinding elements is often a more important variable than the cost of the power required to pulverize different coals. Craig2 also reports that one coal may cause pulverizer parts to wear several times faster than another. It is apparent, therefore, that those concerned with pulverizing coal could profitably employ a method for estimating the abrasiveness of different coals, just as they utilize standard tests for thermal value, grindability, and ash-fusion temperature to assist in selecting the most suitable and economical coal to use in a particular plant. The objective of this investigation was to develop a test procedure that would be suitable for general use in estimating the abrasiveness of coals. However, few, if any, of the standard tests now used for evaluating the properties of coal are the product of a single investigation or the result of a single investigator's efforts. Rather, in each case, a testing procedure was devised by one investigator, used by others on a wider variety of coals, and finally refined completely as the result of the joint efforts of a number of interested people. Thus, the test procedure for estimating abrasiveness developed in the course of this work may not be refined sufficiently in its present form for general use, but it may serve as the starting point from which an acceptable test procedure can be developed. The method has been used thus far on only about a dozen coals, and there has been no opportunity to attempt a correlation between experimental results and actual plant experience. Only wider use of the procedure by other investigators and correlation with plant experience can determine to what extent the method will have to be modified to render it suitable for general application. Test Method Although the literature contains no record of an attempt to devise a method for estimating the abrasiveness of coal that could be used industrially, several investigators have tested properties of coal that are closely related to its abrasiveness. The abrasiveness of a material generally is considered to be related to its hardness, and hardness tests for coal have been employed by Heywood,' O'Neill," and Mathes. Also, the resistance of coal to abrasion, a property that presumably is related to the abrasiveness of coal, was measured by Heywooda and by Simek, Pulkrabek, and Coufalik.2 11 these investigators tested only individual pieces of coal. Since coal is a heterogeneous material having components of varying properties, tests of this type can yield results having little more than academic interest. Only a test method that utilizes a representative sample of coal can give results that are useful industrially. The abrasion tests used for various other materials have been considered for adaptation to testing the abrasiveness of coal. The tests used for metals,7-9 paving and flooring,'" and rubber," cannot be used because coal is not sufficiently abrasive.~ The present experimental work was begun before World War II and was conducted by three research fellows"'" working under a joint agreement between the University of Washington and the Bureau of Mines. After a great deal of preliminary work with a variety of apparatus and materials, a test procedure was developed which consisted of rotating a test disk 2Yz in. diam in a steel mortar containing the coal sample. The shaft carrying the test disk at the lower end and a 100-lb load on the upper end was free to move vertically. The bed of coal in the mortar was kept fluid by low-pressure air admitted through a port near the bottom of the mortar. Measurable wear on an Armco iron disk could be obtained in this test procedure, but, despite extensive efforts to eliminate them, several major disadvantages remained in this test method. First, with most coals the amount of wear on the iron disk did not exceed a few milligrams. Second, a single type of disk was not applicable for all coals. A smooth iron disk gave satisfactory results with both bituminous and sub-bituminous coals, but hardly any wear with anthracite or coke. A disk having studs or projections gave more satisfactory abrasion losses with anthracite and coke and presented no operating difficulties with free-burning bituminous and sub-bituminous coals. It could not, however, be used with caking coals because these coals formed a

Jan 1, 1952

By D. S. Choi, R. Stefanko

There are a number of complex factors that influence the effectiveness of anchorage to maintain tension in rock bolts. However, a plastic analysis of the anchorage site employing certain simplifying assumptions with application of the Mohr-Coulomb criterion appears to explain the observed phenomena. Such an analysis has been made and a correlation sought with field and laboratory tests. Field tests were made in an anthracite mine in eastern Pennsylvania and included pull tests and long-term tests of a variety of anchorage devices in two basic lengths, 30 and 42 in. in two widely differing seams. Performance is reviewed for wedge, expansion shell, and resin anchorage. Laboratory tests duplicated many of the field conditions but in addition compared the performance of shells with normal and reversed serrations. This performance was compared with the predicted results from the plastic analysis. One of the major problems in conducting long-term underground tests is the selection of suitable instrumentation. All installed bolts were fitted with spherical and hardened washers to insure the best possible torque wrench readings. In addition, commercially available load cells were used. Finally, the performance of a specially developed strain-gage-equipped ring cell is viewed. Rock bolting as a method of support continues to increase with applications in many other industries in addition to mining. Nevertheless, with nearly 55,000,000 roof bolts installed in coal mines alone last year, this remains as the single greatest use. While bolts have frequently supported ground where conventional timbering could not, there are relatively few design criteria; and trial-and-error procedures prevail. Furthermore, there has been a lag in development of suitable instrumentation that is simple to install and read out, sensitive, durable, reliable, safe, and economical in evaluating the effectiveness of a bolt over long periods of time. Therefore, the pull test continues to be the most popular method of evaluating the applicability of a certain type of roof bolt under specific installation conditions. At The Pennsylvania State University in the Dept. of Mining, research has been conducted for a number of years to measure bleed off in carefully controlled laboratory experiments as well as in underground investigations."-' Unfortunately, most of the instrumentation developed has been primarily suitable only for research purposes, not possessing all of the aforementioned characteristics desirable for routine underground use. Other groups also have met with restricted success. Therefore, while relatively crude, the torque wrench continues to remain as the most widely used load measuring device. While both field and laboratory tests continue to be con- ducted, analytical analyses are attempted to discover the more important design parameters in order that more efficient anchorage might be devised. Bolts are being used for a greater variety of purposes in mining. Suspending wire sideframe belt conveyors from roof bolts is a common application. The suspension of a monorail transportation system presents yet another. One such system has just been installed in a recently reopened anthracite mine and is presently being evaluated under production conditions. Preliminary studies revealed that a considerable cost reduction was possible by suspending the monorail on bolts anchored in the top. The monorail was to be installed under two widely differing conditions—a competent sandstone above the Buck Mountain seam and a softer shale top above the Skidmore. The type of anchorage device, length of bolt, and long-term performance, consistent with economy and safety, had to be established for the installation once the decision was made to suspend the system on rock bolts. This paper describes some of the testing procedures leading to a final selection. Theoretical Analysis of Expansion Shell Anchorage A detailed look at an expansion shell assembly might shed some light on the factors involved in the design of a suitable shell, Fig. 1. When a bolt is rotated, the tapered plug is forced downward, expanding the leaves laterally to grip the sides of the hole. Two friction surfaces are present: (1) the interface of the plug and leaf and (2) the interface between leaf and rock. The relationships of these friction planes, geometry of expansion shell, and properties of the rock are important in the design of an expansion shell. Therefore, an analysis assuming the rock to behave as a rigid plastic material with its yield governed by the Mohr-Coulomb criterion was made." Furthermore, the effect of friction between the leaf and rock produced by serrations was analyzed.

Jan 1, 1971

By S. C. Sun, John A. L. Campbell

Objective of the ittvestigation was to determine the electrokinetic differences, if ally, of anthracite lithotypes and thus establish the feasibility of making a coqlstituent separation by froth flotation or other surface-dependent processes. The elec-trokinetic properties were determined by streaming potential methods. In addition, chemical and physical analyses of the lithotypes were made to establish nonsubjec-tive parameters for their diflerentiation. Electrokinetic results indicate a difference in the characteristics of the lithotypes. The isoelectric points of 4 anthracites and 12 lithotypes were obtained. It was suggested that the relative flotation rates of the lithotypes were pH dependent and that the maximum flotation rate of anthracite occurs in an acidic circuit. Although there has been an increased interest in the electrokinetic properties of coal,' research on the subject is still in an early stage of development."' To the authors' knowledge, no electrokinetic studies have been made on coal lithotypes, particularly anthracite litho-types. This paper presents the results of an investigation of the double layer properties displayed at the anthracite/water interface (whole coal and lithotypes) as determined by streaming potential methods. It is hoped that this investigation and related future investigations will provide correlations from which the surface properties of any coal can be predicted and then related to surface-dependent processes such as froth flotation and flocculation. Another aspect of this research, undertaken to determine the physical and chemical parameters necessary for anthracite lithotype differentiation, was previously published.? The number of parameters involved, lithologic, organic and inorganic, and the wide divergence of these parameters within an individual coal limits the practical utilization of the results; however, the result? are of value and of interest in guiding theoretical developments. Keeping the limitations in mind, the results are analyzed in terms of their application to froth flotation processes and are used to indicate the feasibility of making a lithotype separation by flotation. Experimental Procedure Prior to discussing the experimental work per se, it would be of value to establish an understanding of anthracite lithology. Anthracite Lithology: Coal seams display two different modes of variation, rank and type. A coal's position in the continuous series ranging from peats through anthracites determines its rank. The rank of the material under consideration is anthracite. The different bands of coal which constitute a coal seam are distinctly contrasted in regard to physical properties and chemical composition. These megascopic bands are the coal types and are termed lithotypes., A megascopic examination of 11 anthracites revealed three distinct, homogeneous, lithotypes. The classic description of bituminous coal lithotypes as proposed by Stope" was found to be quite adequate for recognizing the three lithotypes in anthracite, namely vitrain, durain, and fusain. The term vitrain describes the brilliant constituent of coal; durain describes dull coal; and fusain is the term applied to fossil charcoal. The anthracites studied were generally found to contain 60-80% vitrain, 10-30% durain, and 1-10% fusain. A complete description of the physical characteristics and chemical composition of the studied lithotypes can be found in a previous publication.' Sample Preparation: Nine of the anthracites used in the investigation were from Pennsylvania, one was from Russia, and one from Wales. Table 1 shows the various designations and sources of the employed anthracites. The letter V, D, or F added to the whole coal designation indicates respectively the vitrain, durain or fusain lithotype of that coal. The samples were carefully collected to avoid contamination. To insure against oxidation, the samples were obtained as fresh as possible and stored under nitrogen or water. Generally a representative sample of the seam and samples containing high concentrations of one or another of the petrographic constituents were secured for each coal. Pure lithotypes were prepared by manual separation and selection. The pure lithotypes and representative samples were ground to —35 mesh and screened to 35 x 48 mesh. The 35 x 48-mesh fraction was then passed over a magnet to remove iron contamination, washed with doubly

Jan 1, 1971

By J. A. Harrison

The physical composition of coal and the distribution of its physical constituents will partially determine the efficiency of the process chosen for coal preparation. Distribution and concentration of coal macerals will, to a degree, be controlled by preparation processes. A number of preparation processes have been evaluated to determine this interrelation. These evaluations will help provide information needed for the production of coal that meets definite petrographic composition specifications. etrographic composition and degree of coalifica-tion of coal can determine the choice of cleaning processes and equipment used in coal preparation. The purpose of this paper is to present data on studies of the preparation processes of breakage, sizing, cleaning, and blending and to show the effect coal's petrographic composition has on these processes. Herrin (No. 6) and Danville (No. 7) Coals from Illinois were used in these tests. Nomenclature used herein is that approved in the Glossary of the International Committee for Coal Petrology,' based primarily on the Stopes-Heerlen system of classification2 and used in coal petrography laboratories throughout the world. BREAKAGE TESTS Breakage tests of Illinois coals related to petrographic composition have been conducted at various times at the Illinois State Geological Survey. These tests have been classified under two headings -impact tests and compression tests. Impact tests, as the name implies, consisted of breaking the coal by a sharp, sudden force. Impact tests were further divided into fall tests and block tests. Compression tests consisted of applying a constant pressure on a piece of coal until the coal fractured.

Jan 1, 1963

By J. R. Lewis, J. H. Hamilton, V. D. Hanson, K. W. Heinlein, J. M.194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1'/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953

By J. H. Hamilton, J. R. Lewis, K. W. Heinlein, V. D. Hanson, J. M. 194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1'/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953