Search Documents

By E. C. Olsen

This paper describes the unusual hazardous conditions encountered in mining thick, inclined coal beds under heavy cover in Carbon County, Utah. This description includes heavy roof conditions, sloughing ribs, steep pitches, and severe bouncing experience. Emphasis is placed on the top bench system of mining and development in the top half of a 16-ft seam, under the roof rock, which is bolted, and the recovery of the bottom coal on retreat; success of bolted roof and ribs; and shuttle car haulage on pitches up to 25%. The use of protective cages for operators of face equipment, the guarding of belt conveyor systems to prevent injury associated with deliberate entry, and the positive lockout of electrical equipment is described and illustrated. Most coal mining areas of the Western United States are characterized either by thick beds, steep pitches or heavy cover. Individually, each of these may present inherent safety hazards that in- fluence mining. When all are met at one location, the combined effect is to make the safe operation of the mining property a challenge to management. These combined conditions are met in the coal mines of the Columbia-Geneva Steel Div. of the United States Steel Corp., in southeastern Utah. This coal bed measures up to 16 ft in thickness, the pitch is as much as 25% and mining is under a cover approaching 3000 ft. The Sunnyside coal bed is found in the Book Cliff Mountains, a great escarpment extending from west of Castle Gate, Utah, some 180 miles to Grand Junction, Colo. West of Green River, Utah, the Book Cliffs swing sharply to the north because of the domelike uplift of the San Rafael Swell to the southwest. It is in this northerly extension of the Book Cliffs that the Sunnyside coal in the Blackhawk Formation of the Mesa Verde Group, Upper Cretaceous Age is found. The strata are transected by numerous faults with from a few inches to as much as a hundred feet of displacement. All of these faults have had some effect on the mining operation, either in roof control or, in the case of major faulting, the pattern of mine development. MINING PRACTICES Mining has generally been confined to the Lower Sunnyside Seam which measures from 7 to 16 ft in

Jan 1, 1963

By D. A. Dahlstrom

WITHIN the past decade, certain economic, legal, and technological changes have intensified the necessity for an efficient classifier or deslimer of particles at and below the 200-mesh point. First, mechanization of mining has increased the percentage of clays, silts, refuse, and high-ash material in the run-of-mine coal. Second, many washing plants must rely on smaller percentages of fresh make-up water because of shortages, economic factors, or legal restrictions on waste water disposal. Finally, the crushing of middlings and coarse refuse plus the trends toward the finer sizes in sales has further increased the amount of fines produced per ton of mine-run coal. A joint consequence of these circumstances has been the augmented percentage and build-up of —200-mesh material in the finer coal fractions and tipple circulating water. This results in the following disadvantages to the operator: (1) Higher ash percentages in the fine coal caused in large measure by the greater ash content of the —200-mesh material, (2) more difficult or expensive recovery of the fine coal from plant slurries, (3) lower filtration, drainage, and drying rates caused chiefly by lower permeability and the greater surface exposed per unit mass of the finer particles, (4) greater difficulty of beneficiatiing the fine coal because of lowered equipment efficiency and the partial "coating-out" of the —200-mesh fraction on the coarser coal particles. Obviously some cheap but effective fine-size classifier must be developed to aid in solving the above problems. To eliminate the generally unwanted —200-mesh fraction from coal slurries, several methods are in use. Probably the most popular with coal operators is the settling cone wherein the larger particles settle downward to the apex of an inverted cone while the slimes and fines are carried upward by vertical fluid currents caused by the overflow of the bulk of the water at the outer rim. The main advantage of the unit is the low initial and operating cost since no moving parts are involved. Disadvantages are primarily low capacity per square foot or floor space and the loss of good coal in the overflow above the 100-mesh size due to localized, high-velocity gradients and nonuniform, fluid-flow patterns, especially at high loadings. Also, when high slime concentrations are present, desliming efficiency is severely injured.', ' Other methods employed are the drag and spiral classifiers and the hydroseparator.' These involve mechanical separations achieved by moving parts and naturally result in higher initial costs. As they are all based on the settling velocity of the particles at the mesh of separation under operating conditions, area requirements are again large per gallon of throughput if they are to classify at the 200-mesh point. Usually the selection is made to operate at a higher throughput and thus sacrifice some coarser coal in the slime overflow. Cyclone Possibilities in Desliming To overcome the disadvantages of present deslirning methods, it is necessary to use a device that exhibits a very stable and uniform flow pattern while greatly increasing the settling velocity of

Jan 1, 1952

By D. A. Dahlstrom

D. R. Lyons (Republic Steel Corp., Cleveland)— Some tests using wet tables conducted by an operating company to determine the effect of table-water density on the quality of the washed coal should be of interest. The only variable during the testing period was the percentage of solids in the table water. Table operating data are shown in Table XV. The ash content of the washed coal remained constant or nearly so, as the percentage of solids varied while the sulphur content of the washed coal decreased as the percentage of solids increased.

Jan 1, 1952

By Mehmet Guney

At the Armatcuk mine in Turkey, various modifications of sublevel methods have been tried during the last decade in an effort to eliminate gob fires, dangers from roof falls, and surface subsidence, and also to increase extraction efficiency. From the experience thus gained, it was decided that the most likely system to give maximum benefits, considering the existing geological conditions and coa1-seam characteristics, was hydraulic stowing. This paper examines trends in the working techniques of hydraulic stowing and discusses the favorable results and technical problems that the method produced at Armutcuk. The Armutcuk mine, a subsidiary of Eregli Coal Mines (EKI), produces 1300 tons of run-of-mine coal daily from thick (about 10-90 ft) and highly dipping (about 50-90°) coal seams. The coal is highly volatile with good coking properties that make it highly suitable for industrial use. Most of this coal is used in the iron and steel industry at Eregli on the Black Sea coast. Complete recovery of the coal is of vital importance to industrial development and progress. Initial coal extraction was carried out by means of a sublevel mining system modified to accommodate local geological conditions and coal seam characteristics. This method is similar to that employed in metalliferous mines. Several problems arose from the application of the method, one of them being the prevalence of gob fires which caused periodic shutdowns in some sections. To achieve maximum coal extraction with minimum dilution, to eliminate gob fires and surface subsidence, and to reduce fatal accident rates, a decision was taken to change from this method to a hydraulic stowing system. The pres- ence of large deposits of conglomerate and sandstone on the surface contributed to the decision to employ this new method. Although hydraulic stowing practices have been carried out in America and Germany, Upper Silesia may, so far as coal is concerned, be considered the birthplace of the method. Here the process was adapted to thick coal seams on a large scale to reduce dangers from gob fires and damage from surface subsidence. Today, the mining system is not only widely used in the coal mines of France, the Saar, India and Poland, but has also been successfully applied to the iron ore and potassium salt mines in Russia and Germany. LOCATION AND HISTORY The Kandilli-Armutcuk coal field, the most western outcrop of the North Anatolian coal basin, extends for five miles along the Black Sea coast (Fig. 1). The Armutcuk Mine is situated on the southern boundary of the coal field about 40 miles from Zonguldak and about 12 miles from Eregli. The history of coal mining in the North Anatolian coal basin goes back more than a century. Mining commenced on a small scale in the Kandilli-Armutcuk coal field in the late 1880's, and from this time until nationalization in 1940 the mines were operated by private Turkish and foreign interests. In 1940, Eregli Coal Mines was established to takeover the management of these enterprises. The Armutcuk Mine was developed for the first time in the mid-1940's. Since then it has been in continuous production and it now produces an annual 440,000-ton run-of-mine output from which 320.000 tons of salable coal is obtained. MINE GEOLOGY The carboniferous inlier of the Kandilli-Armutcuk coal field crops out over an area about three miles long and one mile wide. The major part of the inlier consists of the Namurian formation with only a relatively small area of Westfalian-A visible on the surface (Fig. 2). The Namurian formation, which has no exploitable coal seams in the Armutcuk Mine, consists mostly of arkose sandstones and black shales often having

Jan 1, 1969

By A. P. Pipilen, M. Weintraub, W. F. Hosford, A. A. 230-000-000-006 Orning

In a study of the principal factors affecting the transport of coal-water mixtures through a centrifugal pump and a pipeline, the interrelation between solids concentration, velocity, and pressure drop was established for a minus 2-in. bituminous lump coal in concentrations up to 48% by weight. Optimum concentrations were found to exist for maximum capacity of a given pipeline and for minimum energy requirements per ton of coal. Coal size degradation by particle fracture took place in the pump; abrasion and attrition to form fines took place in both pump and line. A novel system of studying flow of solid-liquid mixtures in pipelines by the determination of relative sound intensities was developed. The hydraulic transport of minerals is in part an established art and in part a relatively new field of industrial technology. The established art developed especially where the mineral was under water, where water was used in hydraulic mining, or where large quantities of water had to be pumped from the mining site to expose the mineral. Power requirements and limited life of pumps and pipelines, under such conditions, were relatively unimportant. Examples may be found in dredging, and in hydraulic mining and transportation of phosphate rock.' In the design of a coal pipeline the most needed information relates to flow characteristics, power requirements, pressure drop, size degradation, and size segregation as a function of solids-water ratio and size distribution of the mineral. A considerable amount of information has become available concerning pressure drop and energy requirements.2'3 These topics were therefore not studied intensively. Coal degradation, however, has not been clearly delineated. Lammers and coworkers4 showed that coal may be severely degraded in size but did not distinguish between degradation produced in the pump and in the line. It is the primary purpose of this paper to outline techniques used to acquire some of the much needed design data. A full report of the investigation is to be published by the Bureau of Mines at a later date. APPARATUS AND PROCEDURE The apparatus for circulation tests (hereafter referred to as Type III tests) consisted essentially of a loop and a centrifugal solids handling pump, with attendant instrumentation, and coal weighing and feeding equipment. It also included an adequate water supply, as shown in Fig. 1. The selection of 6-in., schedule 40, carbon steel pipe was based on the top size of coal to be used. Assuming that the top size of coal should not exceed 1/3 the inside diameter of the pipe, a 6-in. pipe would accommodate 2-in. coal, a realistic size for transport from the mine face. An 872-ft loop was constructed first. Later a few tests were made with a lengthened circuit of 1361 ft by superimposing an additional loop on the original one. All bends used were of 3-ft radius. A noisemeter microphone, taped to the pipe at various points, but generally about 50 ft downstream of the pump, was used to sense segregation of the coal and the length of slugs of coal in water. To measure the stream velocity, radioactive sources were encapsulated in plastic cubes. The specific gravity of the blocks was adjusted to 1.3 and 1.8 to simulate low ash and high ash coal particles. Their passage through the pipeline was detected by a Geiger counter. Because the repeated circuits of the Type III tests did not permit coal degradation in the line to be distinguished from that produced by the pump, the two

Jan 1, 1964

By H. L. Lovell, L. H. E. Weyher

Cyclones have been used successfully in the mineral industries to solve solid-liquid as well as solid-solid separation problems in the processing of fine particles. Substantial information is available about the operation of cyclones as thickeners, classifiers, and concentrators using a dense medium. However, very little is known about the operation of the cyclone as a fine coal concentrator not using dense medium. This application of the cyclone, namely to wash fine coal in water, requires that classification phenomena be minimized in order to arrive at an effective separation of the high density ash and sulfur-bearing mineral particles from the low density coal particles over a wide particle size range. The authors have investigated the role of feed pressure, orifice diameters, and vortex-finder length in the cyclone washing of fine coal (--1/4 in.) in water. The results of this investigation show that cyclone washing of fine coal in water is an effective unit operation particularly capable of reducing the inorganic pyritic sulfur content. The combination of hydrocyclones with complimentary equipment such as Sieve bends, shaking tables, and flotation units appears to be especially promising and may ultimately yield fine coal processing systems that could approximate the high recovery efficiencies of today's coarse coal cleaning circuits. In the future, fine coal beneficiation can be expected to receive even greater research emphasis as a result of increasing sulfur restrictions on all major coal markets. In a previous paper, work, conducted at the Pennsylvania State University, on the hydrocyclone washing of fine coal was discussed.' The initial evaluation based on that wo-rk demonstrated the feasibility of hydrocyclones as fine-coal cleaning devices, and in that previous paper selected test data were presented that showed in principle how product yield and assay respond to variations of the major cyclone operating variables. The present paper reports in greater detail on the findings and presents a comprehensive evaluation of the data as follows: (1) Hydrocyclone throughput rates were related to feed pressure and other variables by regression equations. These equations are compared with other capacity-vs.-pressure drop expressions taken from the literature on classification cyclones. (2) The flow ratio was found to be the most important variable for the separation of the solids. Control of this ratio through various cyclone operating parameters is discussed. (3) The data on the product yield as well as the ash and sulfur assays from the tests were evaluated by means of correlation and multiple regression analyses. Because size-classification phenomena cannot be suppressed completely in hydrocyclone washers, product yield and assay data are regressed against cyclone operating variables separately for different size classes. These equations then reveal how the interactions of the cyclone operating variables vary in terms of the separation for different size classes. (4) Density partitions from additional tests were analyzed by supplementing product-size analyses with sink-float analyses of the separated product size classes. The evaluation of these data show how classification phenomena affect separation by density. The results demonstrate in particular the feasibility of the hydrocyclone washer for removing pyritic sulfur from fine coal. In addition to these specific findings, the investigation points up the value, in general, of using hydrocyclone washers in combination with unit operations such as screening, table concentration, and flotation. In order to make optimum use of the hydrocyclone washers and work out the most effective combination of unit operations, it is important to have careful flowsheet analyses corresponding to the given feed and product requirements. Analyses of this kind are now possible by means of computer simulation of flowsheets and other operation research tools. INTRODUCTION Before proceeding with the detailed treatment and evaluation of our results, it will be useful to review the fundamental aspects of cyclone separators and also restate the experimental procedures and data handling employed in these studies. Review of Cyclone Fundamentals: AS shown in Fig. 1, the basic components of cyclones commonly used

Jan 1, 1970

By Shiou-Chuan Sun

THE fact that coals of different ranks and even of the same rank differ greatly in their amenability to iroth flotation is well known. In recognition of the need for an explanation of this phenomenon, two hypotheses have been suggested. Wilkinsl reported that the floatability of coals increased with an increase of the carbon content or rank. This postulate is handicapped by the fact that bituminous coals that possess moderate carbon contents are actually more floatable than anthracite coals that have high carbon contents, as shown in columns 6 and 9 of Table I. Taggart and his associates' implied that the difference of floatability between bituminous and anthracite coal was caused by the variation of carbon-hydrogen ratio. This is not applicable to the relative floatability of other coals and carbons. For example, column 11 of Table I shows that the carbon-hydrogen ratios of low-floating lignitic coal and non-floating animal charcoal are not only smaller than the moderate-floating anthracite coal, but are also similar to the high-floating bituminous coal. Furthermore, according to this hypothesis, high temperature coke-A (464), Ceylon graphite (1238), and lamp-black (357), all possessing extremely high carbon-hydrogen ratios, should be less floatable than other substances having much lower carbon-hydrogen ratios such as high volatile-B bituminous coal (11.9 to 22), anthracite coal (35.7 to 60.5), lignitic coal (15.6 to 33.6), and charcoal (13 to 26.2). However the former group is actually more floatable than the latter group. In this paper, a surface components hypothesis is Proposed to explain the different floatabilities of coals, carbons, and hydrocarbon minerals. The validity of the hypothesis is experimentally supported by the actual floatability, natural floatability, wettability, and adsorbability for neutral oils of coals, carbons, and hydrocarbon minerals tested. The combustible recovery of the flotation results, as used in this paper. was calculated from Eq. 1: P (100-Ep) 100 RWCP Rc= [1] F (100-E,) C, where R, is the percent combustible recovery; F and P are, respectively, the weight of feed and the weight of concentrate or product; E, and Ep are, respectively, the total percent of ash plus moisture in feed and in concentrate; Ru. is the percent weight recovery: and C, and C, are, respectively, the percent of combustible in feed and in concentrate. Except for ash and moisture content, all chemical components of a coal are assumed combustible. The experimental work included studies on flotation, ultimate and proximate analyses, contact angle tests, extractions of bitumen-A with benzene, adsorptions for liquid hydrocarbons, and wetting tests. Most of the flotation experiments were performed in a laboratory Fagergren machine; others were tested in a small Denver machine. The solid feed for the former was 300 g and for the latter was 30 g. The solid materials used for flotation were crushed to —48 mesh. After the mineral pulp in the flotation cell was agitated for 6 min and the pH was adjusted to 7.5 & 0.2 with sodium hydroxide or hydrochloric acid, a petroleum light oil having a viscosity of 5.73 centipoises at 77 °F was added and conditioned for 2 min. Finally, pine oil was introduced and the froth was collected for exactly 3 min. The weight ratio of petroleum light oil to pine oil was kept constant at 1.5 to 1. Tap water was used for all flotation tests. Contact angles were measured with a captive bubble machine. For each coal sample, three specimens were mounted in transoptic mounts and polished with levigated alumina, first on a sheet glass, then on a cloth-covered metal polishing wheel. The polished specimen was first washed with distilled water and wiped thoroughly on a cleaned linen pad, then transferred into the pyrex cell of the captive bubble machine and conditioned for 6 min., and finally measured for contact angles at three or more points. Except where otherwise stated, the induction time for each measurement was 1 min. The contact angle representing each material was obtained by averaging the measurements of three specimens. The linen pad was first washed with warm distilled water, then boiled 30 min in a 2N sodium hydroxide solution, and finally washed with distilled water until no trace of sodium hydroxide could be detected in the decanted solution. The cleaned linen pad was stored under distilled water. Immediately before using, the pad was rewashed and transferred into a clean pyrex petri dish partly filled with distilled water. The glassware and rubber gloves used were cleaned by soaking in sulphuric acid-potassium dichromate cleaning solution, followed by rinsing with distilled water. The polished specimens were handled only by glass forceps. The ultimate and proximate analyses were made in accordance with the ASTM standard procedures for coal and coke. The extractable bitumen-A was determined by weighing 1 g of —100 mesh sample and placing it in a desiccated and weighed ASTM aluminum-extraction thimble. The thimble was placed in condenser hooks and inserted into an extraction flask containing 100 cu cm of benzene. The flask was heated and the benzene vapor was condensed by water coils. At the end of 24 hr of percolation, the thimble was removed, desiccated, and weighed. Loss in weight of sample was taken as bitumen-A and calculated to dry and ash-free basis.

Jan 1, 1955

By J. D. Price, W. M. Bertholf

Making maximum possible use of available equipment and material, CF&I placed a high-efficiency, high-capacity washery unit in the existing buildings to gain simplified operation, reduced manpower requirements, raised efficiency, and boosted production. THE central washing plant of the Colorado Fuel & Iron Corp. was first operated in 1918 to furnish coal for two 60-oven batteries of Koppers design. Prior to that time the coke for the blast furnaces had been secured from a number of beehive operations, some with their own washeries. The original washery had four Elmore jigs, which did not do as well as had been expected. In 1923 the jigs were taken out and replaced by Deister Plat-O tables, with various changes in auxiliary equipment. In 1936 the throughput of the plant was considerably increased by the installation of vibrating filters for recovery and drying of fine coal, thereby reducing a large recirculating load of fine coal which passed through the screens of the Carpenter driers. In the late 1940's it became evident that something would have to be done to keep the washery abreast of progress at Pueblo and other washeries. Coal requirements were nearly two and one-half times the original requirements of the 120 small ovens, as additional ovens had been built in 1930, 1938, and 1945, whereas no additional coal washing capacity had been added since the 1936 modifications. Much of the equipment, moreover, had seen better days. One answer to the problem was to build a new and completely up-to-the-minute plant, with sep- arate sections for coarse, medium fine, fine, and very fine coal. On the other hand, it might be financially more profitable to place a high-efficiency, high-capacity unit in the existing buildings. The washing problem at Pueblo is not unique, but it is approached differently. There are no established specifications for coke ash and sulphur. The plant is operated on the principle that if it is more economical to wash it out than to slag it out, wash it out; otherwise, send it on to the blast furnaces. With coals at the Pueblo plant costing what they do, the benefits of washing at a low gravity are questionable. Hence the primary problem is to eliminate the high-ash components and retain all the low and medium-ash components. Sulphur is no problem at all, averaging 0.6 pct in the coke. Since a good proportion of the free impurity in the coals is as large as the coal or bone or larger, it appeared that it would be safe to go to jig washing and discontinue crushing the coal before washing, as had been the practice with the tables. This would boost the throughput. It might also decrease manpower requirements and washing costs considerably. It was assumed at the start and has been demonstrated in the first year of operation that a jig will do as well on the fines as the tables had been doing. It would not be necessary to discount any gains due to increased efficiency in washing the coarse material because of incapacity to wash the fines. By washing at —2+0 mesh it would be possible to dewater, by screening, a considerable tonnage satisfactorily, for example, at —2 + % or —2 + l/4, and then crush it in a hammer mill. The —?+20 could be dewatered in centrifugal driers and the remain-

Jan 1, 1955

By H. L. Riley, B. W. Gandrud

AS far as the present writers know, this system of flotation has not been used elsewhere in this country, but in the last couple of years it has been introduced, with minor variations, at one plant in England and one in Wales.' The system has been described and discussed in a number of publications.2-5 The following is quoted from an abstract of the latest of these,5 a paper presented at an International Conference on Industrial Combustion in 1952. On the basis of experience to date with the commercial plants, it is believed that the kerosene-flotation process incorporates all the necessary elements to make it greatly superior to anything else now available for treating of fines in wet processes of coal preparation. Additional study and investigation are still needed, however, to determine if certain phases of the process can be improved to such an extent as to make it generally satisfactory and acceptable to the industry. Further improvements will be needed with respect to the capacities of the flotation cells and the reagent consumption. The situation referred to above explains why an investigation is being made of the possibilities of achieving better cell capacities. Results obtained from this investigation, which is still in progress, are believed significant with regard to both cell capacity in general and the relation of cell design to cell capacity in particular. Commercial equipment now being used in a laboratory-type investigation should have performance characteristics similar to those of the larger machines. Equipment and Procedures: All flotation tests have been made in a standard Denver sub-A 24x24-in. unit cell of 12-cu ft volume. Cell modifications to make it more suitable for the tests were an adjustable front-wall section for varying cell depth and a perforated scraper-drag assembly for removal of the float product. There is also an apron dry-coal feeder, a gravity-feed water supply, reagent feeders, and a centrifugal pump that feeds the mixture of coal, water, and reagents into the flotation cell. A wattmeter connected into the drive-motor circuit records the power requirements of the impeller throughout each run. Dry coal, water, and reagents are all fed through a pan-type intake to the feed pump. A Sturtevant blower was set up to furnish air for supercharging. A centrifugal pump with a garbage-can intake provides for disposal of refuse flow to an outside settling tank. Figs. 1 and 2 show the flotation cell; Fig. 2 also illustrates the blower for supercharging. For purposes of this investigation, the percentage by weight of the feed coal recovered in the float product under a standard set of conditions has been considered as the criterion of cell capacity. The authors realize that such a criterion may be somewhat unorthodox, as the term cell capacity is usually understood to refer to feed input and ordinarily takes into account the ash analyses of the float product and refuse. However, the word capacity is flexible enough so that Webster gives one definition as maximum output, a definition which seems to justify, at least partly, acceptance of the above criterion. It has been the authors' experience in the Birmingham district that the ash-reduction efficiency of the coal-flotation process is generally satisfactory and that the only real problem is to increase the rate of float recovery so that the feed rate to any given bank of cells can be increased without undue loss of coal in the refuse. Originally it was planned to operate the flotation cell to simulate continuous operation during sampling periods. It was assumed that operating for reasonable time with feed coal, water, and reagents turned on would stabilize conditions so that the weight of float coal discharged during a fixed time interval would be an accurate measure of the rate at which the coal was being floated. It developed, however, that this supposition was erroneous. The float coal, caught for fixed time intervals and weighed, gave widely varying results in duplicate runs. Efforts to correct this trouble failed, and it was decided to try to operate on a batch-test basis, whereby all the float coal produced during a run on a known weight of feed coal would be caught in tubs, dewatered, and weighed. This method gives consistent and reproducible results, with total float product weight rarely varying by more than 3 or 4 pct on duplicate runs. The standard test procedure is as follows: A 132-lb sample of dry feed coal is weighed and placed in the feed hopper. The feeder is adjusted for a rate of 800 lb per hr. Feed water and reagents are turned on, and the feed and refuse pumps are started. One minute later the impeller is started. Six minutes are allowed for the cell to fill up with the water-reagent mixture. The feed of dry coal is started at the end of this 6-min period. One minute later the float-coal removal drag is started. The float coal is caught in one tub for the first 6 min after the flow of feed coal starts. Tubs are then changed, and the float coal is caught in a second tub until the feed coal runs out, when the tubs are again interchanged to catch the float coal for the remainder of the run in the first tub. The cell is kept running for 3 min with the water and reagents on after the feed stops to allow residual float coal to be removed. At the end of a test the wet float coal in both tubs is weighed and the total weight recorded. The product in the second tub is used for moisture determination and screen-size analyses. When the

Jan 1, 1956

By Thomas L. Kesler

I all of the 14 million tons of kaolin produced in Georgia through 1949 had been mined from a single deposit 20 ft thick, it would represent a mined-out area of less than 1 sq mile. This measure of depletion is in such strong contrast with the large size of the area shown in Fig. 1, from which most of the output has come, that there has developed an optimistic outlook concerning future supplies of kaolin equal in grade to that already mined. Dozens of mines are widely but not evenly distributed within this area, which consists mostly of parts of Twiggs, Wilkinson, and Washington counties. There are scarcely half a dozen kaolin mines elsewhere in the state that have been appreciably productive, although the areas in which they occur may have good possibilities. Most of the efforts to find new deposits, therefore, will be confined to the area shown in Fig. 1, for it is here that the industry has established its facilities, and prospecting is encouraged by relatively numerous outcrops and opened deposits of kaolin. The Ocmulgee and Oconee rivers drain the area. They and their main tributaries have flat, swampy floodplains covered with alluvium and for the most part thickly wooded. The divides between these streams are intricately dissected by the minor tributaries, which are spring-fed in steep-walled, cirque-like heads. Gullying and under-cutting are active along the streams but are partly controlled by unusually dense vegetation. The mean altitude is about 350 ft above sea level, and locally the relief ranges up to 200 ft. The area is near the inner margin of the Coastal Plain, and is underlain by Cretaceous and Tertiary formations. Its general location is shown in black in Fig. 1. The division between the Coastal Plain and the Piedmont plateau is actually a very sinuous line, the fall line, marking the extent to which erosion has stripped the Coastal Plain sediments from the southeastward-sloping surface of the crystalline rocks of the Piedmont. Cretaceous Host Formation The kaolin deposits, which are described in the succeeding section, occur sporadically in lenslike bodies in a thick series of poorly sorted clayey sands constituting the lowermost of the Upper Cretaceous formations, the Tuscaloosa. The basis for this correlation has been reviewed by Cooke.1 The same formation has been termed the Potomac group by Ladd2 and the Middendorf formation by Smith' in reports on the kaolin deposits. In another report, Veatch4 called the formation the Tuscaloosa but correlated it with the Lower Cretaceous. The Tuscaloosa in this area consists of fine to very coarse sands, commonly cross-bedded, which contain widely different proportions of white, pink, and yellow kaolin and considerable white mica. A few of the coarser beds contain quartz grains up to 3 in. in length, with bodies of pure white kaolin of equal size that appear to have been formed by the weathering of feldspar grains in place. In the kaolin-producing area, gravel and cobbles occur mainly at the base of the formation. Fig. 2 shows the relations of the Tuscaloosa and its kaolin deposits to the adjacent formations. Its base, on the eroded surface of the crystalline rocks, dips southeastward 50 to 60 ft to the mile, according to LaMoreaux.5 The upper surface of the Tuscaloosa is also an unconformity,

Jan 1, 1952

By W. J. Skewes, C. T. Holland

Properly constructed stopes are necessary underground for proper ventilation and safe, healthful working conditions. Mining companies are concerned with providing the best possible stoppings at an economic cost. It is important to consider 1) proper materials, 2) construction that ensures maximum performance from the material used, and 3) keeping costs within reasonable limits for the unit operation—stopping—involved. The authors investigated all aspects of the problem and have attempted to layout a study program to provide information on all phases. Mining companies every year purchase huge amounts of material for the construction of ventilation stoppings. This material by the time it is in place in the stoppings will cost the company many thousands of dollars. In addition and more importantly these stoppings are supposed to confine air travel along certain well defined paths into areas where it is needed to provide safe and healthful atmospheres for workmen. The cost of forcing this air through the mine is large and in general increases as the total quantity increases. Consequently, if the stoppings leak, increased cost results and in addition unsatisfactory ventilation may prevail in the mine. Therefore, it is essentia1 to provide the best possible stoppings at an economic cost. When attempting to provide the best stoppings possible at an economic cost, three aspects of the problem must be considered. These are 1) getting the most suitable material for stopping constructions, 2) constructing the stopping so as to get the best performance from the material used, and 3) keeping the cost of stopping construction to a minimum for a suitable stopping. This investigation is concerned with all phases of the problem and has attempted to lay out a study to provide information on all of them. In order to develop information on the best stopping material available it seemed wise, after con- sidering the job a stopping must do, to investigate the following aspects of material used for stoppings: 1) strength in compression, 2) compressibility or yield under load, 3) porosity, and 4) permeability. To provide information on the best manner of construction the following aspects of air leakage through the stopping were investigated: 1) when the stopping is built of dry stacked materials—i.e., the joints are not mortar-laid (See Fig. 1); 2) when the stopping is built of dry stacked material and then plastered on one side (see Fig. 2); 3) when the

Jan 1, 1961

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By G. H. Matheson, J. M. W. Mackenzie

The flotation rate of coal has been studied using a continuous laboratory flotation cell and a multiple exposure photographic technique. The effects of particle size, reagent additions and cell turbulence on the flotation rate of coal have been investigated. The results have been analyzed in terms of current theories of particle bubble encounter and adhesion. The rupture forces which occur during particle bubble encounter have been shown to have an important effect on the flotation rate of coarse coal particles. The flotation of coal is a time dependent process, and in order to obtain a complete understanding of the operation it is necessary to determine the kinetics of the process. In coal flotation, the flotation rate and the associated values of cell productivity in terms of tons per kilowatt-hour and tons per cubic foot of cell volume are of greater importance than in most other mineral flotation operations. The reasons for this are associated with the low unit value of coal and the relatively high cost of coal flotation compared with other coal cleaning processes. An important aspect of coal flotation kinetics is the design of coal flotation cells. The cells used in many coal cleaning operations are often replicas of those used for the treatment of sulfide ores and are not necessarily suitable for the treatment of coal washing pulps, which usually contain a higher percentage of coarse and floatable material than the sulfide pulps for which the cells were designed. This disadvantage of many coal flotation cells has been recognized by Russian engineers who, by the addition of baffles to their subaeration coal flotation cells have increased the cell capacity by 100%. Investigation of the variables which influence flotation kinetics and cell productivity are, therefore, important to the coal industry where flotation is the only accepted method of cleaning the fine (-1/2mm) coal. The kinetics of the flotation process have been investigated in detail by Schumann10 who developed a continuous flotation cell which permitted the study of flotation under steady state conditions. This technique has been used by Gaudin et a1.,6 Brown and Smith and recently by Jowett and Savfi.7 The latter authors have developed a refined mathematical treatment for interpreting the results of continuous laboratory tests. Attempts have also been made to study flotation kinetics using conventional batch testing techniques. The experimental difficulties, however, and the problems of determining the order of the flotation rate process have made this technique of dubious value. A summary of the proposed rate equations for batch testing and the difficulties associated with this technique is given by Brown and Smith. 2 The experimental techniques described above are concerned with the rate of production of froth product. However, several authors, notably Sutherland, 11 Bogdanov1 and Philippoff, 9 have analyzed flotation

Jan 1, 1963

By J. T. Peters, N. Schapiro, R. J. Gray

Petrographic studies of coal have resulted in a better understanding of the origin and composition of coal and have added to the knowledge of how and why coals differ in their response to various prepamtion and utilization processes. Petrographic methods can now be used in scientific economic exploitation of coals. Because the United States Steel Corp. is one of the largest consumers of coal and coke, there is a continuing interest in the corporation in perfecting a better understanding of coal so that it can be utilized more effectively. Coal is a heterogeneous substance composed principally of plant material that has been altered by a complex geochemical process to form rock. The science of coal is anthracology. The ultimate aim of anthracology is determining the relationships and differences among coals. The fundamental aim is finding out why coals differ. The practical application is determining how the differences in coals affect their utilization properties. Coal, like other sedimentary rocks, consists of more than one material. Fig. 1, a photograph of a natural fracture and a polished surface of a lump of coal, displays heterogeneity in terms of the banded ingredients present. The earliest attempts to describe coal were based on its hardness and appearance. ' Thus, coal has been described as hard or soft, bright or dull, and banded or nonbanded. The variations in the hardness and physical appearance of the coal reflect differences in materials that constitute the coal. The banded nature, lustre, and texture depend upon the relative proportions of the various constituent materials. Even the earliest petrographic investigations of coal produced evidence of of the heterogeneity of the coal substance and shoul have destroyed the concept of the unit coal molecule. Coals (Fig. 2) that appear the same frequently consist of different microscopic materials and may behave differently during mining, pulverization, washing, or coking. In Fig. 3, the upper photograph shows a bright at-trital medium-volatile coal, and the micrograph at the right shows some of the materials that constitute the coal. The lower photograph shows a bright attrital low-volatile coal and the corresponding micrograph shows material of similar origin to those from the medium-volatile coal. Any petrographic description of a coal that has meaning in terms of its utilization begins with a microscopic examination. Numerous attempts have been made to describe coal on the basis of the materials that constitute the coal.2*3 Initial petrographic studies were directed towards the question of the origin of coals without any true basis for studying the metamorphic changes in the

Jan 1, 1962

By Richard A. Mullins, James J. Merle

Systematic sampling and analysis in coal-washing plants results in product control and economical operation. A well-organized laboratory system reduces operating costs and increases sales if the operating and sales companies make full use of information it supplies. The operating cost of a control laboratory is less than 1 pct per ton of coal produced. THE system of laboratory control as practiced by the Ayrshire group of mines was begun with the construction and organization of the first laboratory in conjunction with the completion of its first wash-ery in 1941. The primary purpose of this laboratory was to maintain a check on the quality of the coal produced from the washing plant. The laboratory was equipped to determine proximate analyses, float-and-sink tests, and screen analyses, and was staffed by a chemist. an assistant. and a technician. Members AIME, are Assistant Preparation Manager and Chief Chemist, respectively, Ayrshire Collieries Corp. and Subsidiaries, Danville, Ill. Joint AIME-ASME Fuels Conference. French Lick. October 1949. TP 2855 F. Discussion (2 copies) may be sent to Transactions AIME before June 30, 1950. Manuscript received Aug. 22, 1949. Soon it became apparent (1) that product control could be accomplished, (2) that a substantial reduction in operating costs could be effected by close control of coal losses in the rejects, and (3) that information furnished by the laboratory was of continuous value to the operating and sales companies. Consequently the laboratory function was expanded to supply them with additional data as required. As the parent company expanded and new mines were added, control laboratories were established at each mine equipped with washeries, and a central laboratory was maintained at one of the mines that was located equally convenient to all of them. A well-planned control laboratory assures production of a uniform, high-quality product. Whenever coal of substandard grade is loaded inadvertently, the sales company can be informed and the proper channel of distribution made. Conversely only constant and strict laboratory supervision can insure proper production of the various grades. Knowing the product, the producer can keep out of many customers' arguments that may be both awkward and expensive. Double checking between customers and mine laboratories have solved many disputes' Some nager insist On buying coal On the analysis basis only, with penalties for substandard coal and premiums for coal above normal quality. A substantial percentage of the coal sold by the Ayrshire group is of this nature. The usual procedure is for the customer to sample and analyze the coal as it is unloaded at the plant, and the mine laboratory samples it as it is loaded at the mine for subsequent analysis by the central laboratory. Only the unloaded moisture is considered, rather than the average of mine and plant moisture, and the buyer sends duplicate samples to the central mine laboratory for this determination. The average of the two analyses determines the price of coal at the mine. In this particular case, the laboratory plays an important part in increasing realization for two reasons: (1) every attempt is made to maintain plant efficiency and product quality as high as possible, (2) check on analyses and sampling by the buyer or seller.

Jan 1, 1951

By Richard A. Mullins, James J. Merle

Systematic sampling and analysis in coal-washing plants results in product control and economical operation. A well-organized laboratory system reduces operating costs and increases sales if the operating and sales companies make full use of information it supplies. The operating cost of a control laboratory is less than 1 pct per ton of coal produced. THE system of laboratory control as practiced by the Ayrshire group of mines was begun with the construction and organization of the first laboratory in conjunction with the completion of its first wash-ery in 1941. The primary purpose of this laboratory was to maintain a check on the quality of the coal produced from the washing plant. The laboratory was equipped to determine proximate analyses, float-and-sink tests, and screen analyses, and was staffed by a chemist. an assistant. and a technician. Members AIME, are Assistant Preparation Manager and Chief Chemist, respectively, Ayrshire Collieries Corp. and Subsidiaries, Danville, Ill. Joint AIME-ASME Fuels Conference. French Lick. October 1949. TP 2855 F. Discussion (2 copies) may be sent to Transactions AIME before June 30, 1950. Manuscript received Aug. 22, 1949. Soon it became apparent (1) that product control could be accomplished, (2) that a substantial reduction in operating costs could be effected by close control of coal losses in the rejects, and (3) that information furnished by the laboratory was of continuous value to the operating and sales companies. Consequently the laboratory function was expanded to supply them with additional data as required. As the parent company expanded and new mines were added, control laboratories were established at each mine equipped with washeries, and a central laboratory was maintained at one of the mines that was located equally convenient to all of them. A well-planned control laboratory assures production of a uniform, high-quality product. Whenever coal of substandard grade is loaded inadvertently, the sales company can be informed and the proper channel of distribution made. Conversely only constant and strict laboratory supervision can insure proper production of the various grades. Knowing the product, the producer can keep out of many customers' arguments that may be both awkward and expensive. Double checking between customers and mine laboratories have solved many disputes' Some nager insist On buying coal On the analysis basis only, with penalties for substandard coal and premiums for coal above normal quality. A substantial percentage of the coal sold by the Ayrshire group is of this nature. The usual procedure is for the customer to sample and analyze the coal as it is unloaded at the plant, and the mine laboratory samples it as it is loaded at the mine for subsequent analysis by the central laboratory. Only the unloaded moisture is considered, rather than the average of mine and plant moisture, and the buyer sends duplicate samples to the central mine laboratory for this determination. The average of the two analyses determines the price of coal at the mine. In this particular case, the laboratory plays an important part in increasing realization for two reasons: (1) every attempt is made to maintain plant efficiency and product quality as high as possible, (2) check on analyses and sampling by the buyer or seller.

Jan 1, 1951

By M. R. Geer, H. F. Yancey, P. S. Jacobsen

Two growing problems confront the preparation engineer—still further restrictions on stream pollution and a greater proportion of fine coal as more and more continuous miners come into use. The de-watering screens, centrifuges, and settling ponds that sufficed a few years ago must often be supplemented by more effective equipment, and in some instances the finest solids are now being recovered by vacuum filtration, once considered too costly in many coal washeries. Unfortunately not all slurries can be readily filtered. Some are virtually unfilter-able, and others do not permit the high cake rates needed to hold costs within reason. Often, however, these difficult slurries can be rendered filterable by flocculation. Starch and lime have long increased settling rate in thickeners and can also aid filtration. More recently there have been other flocculants, primarily synthetic polymers or gums. Reports of their great effectiveness prompted the U. S. Bureau of Mines to test their use in filtration as part of its program on recovering and cleaning fine coal. The object of the present work was to compare some of the newer flocculants with starch and lime and to test the reaction of different types of slurries to flocculation. Three natural slurries, three synthetic coal-clay mixtures, and five flocculants were tested, all with a laboratory filter leaf. Slurries and Flocculants Tested: First of the slurries was the thickener feed from the washed coal section of the Michel colliery of Crow's Nest Pass Co. Ltd., Michel, B.C. As shown by the screen and ash analyses in Table I, practically all this slurry was finer than 28 mesh, and about half the solids were finer than 200. Substantially all material coarser than 200 mesh was clean coal, but the ash content of 20.7 pct in the finest size indicates a moderate amount of impurity. This coal is medium-volatile bituminous in rank. Slurry was also obtained from the Black Diamond washery of Palmer Coking Coal Co., King County, Wash. This underflow of a 1/2-mm vibrating slurry screen is a waste product discharged to a settling pond. As shown by the data in Table I it contained about 20 pct material finer than 200 mesh, which analyzed 77.5 pct ash, indicating a high proportion of clay. The coarser sizes were also high in ash content because of the presence of bone and shale. Black Diamond coal is on the dividing line between bituminous and subbituminous rank. The third slurry, from the washery of Roslyn-Cascade Coal Co. in Kittitas County, Wash., was the underflow of a battery of 8-in. cyclones that is wasted with the coarser refuse from the plant. In both size and ash content this slurry was similar to the Black Diamond, but since it was a cyclone-underflow product, the fraction finer than 200 mesh undoubtedly contained a much smaller proportion of finely divided clay than was present in the Black Diamond slurry. The Roslyn-Cascade coal is high-volatile A bituminous in rank. Initially the slurry samples were stored in the laboratory at a concentration of 40 pct solids. A progressive change in filtration and flocculation characteristics during storage was noted, however, presumably because of progressive disintegration of the clay. After careful testing demonstrated that the slurry solids could be dried and then repulped without changing the original filtration characteristics, this procedure was adopted. The flocculants used were Separan 2610, Kylo 27, Jaguar MD-A, Idaho potato starch, and lime. The lime was used as a slurry, the Separan 2610 as a 0.10 pct aqueous solution, and the other flocculants as 0.50 pct aqueous solutions. The starch was causti-cized with sodium hydroxide before use. As some of the flocculants are reported to deteriorate on storage, new batches of flocculants were prepared every two days. Test Procedure: Fig. 1 diagrams the filter leaf apparatus employed. This consists of a test leaf having an effective area of 0.10 sq ft., a 60°, 8 1/2-in. diam cone to hold the slurry sample, a variable-speed electric stirrer, vacuum gage, bleed line for controlling vacuum, filtrate flask, moisture trap, and vacuum pump. Two filter cloths were selected—Saran SA 603, a monofilament, coarse-weave cloth, and polyethylene PO-801 HF, a monofilament, fine-weave cloth. The Saran was selected to provide maximum cake rate without regard to clarity of filtrate and the polyethylene to provide the best quality of filtrate without regard to cake rate. The test charge for the filtering vessel consisted of 800 g of solids (previously soaked for 24 hr) and

Jan 1, 1960

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal. ribs and movement of the roof and pavement. Output from the three operating mines in the Cumberland field is 3000 tpd. The field is highly inclined, inclination varying from 12° to 32°. Overlying beds consist of shales and massive sandstone lenses of extreme toughness and are responsible for bumps when the stresses are relieved by extraction. At greatest depth these are among the deepest coal workings in the world. Depth of cover ranges from 2300 to 4000 ft. This prohibits room-and-pillar working and necessitates longwall operation. Working of contiguous seams concurrently to maintain output increases the already difficult conditions. The Sydney field, with a frontage of about 30 miles, is the most important of the Nova Scotia coal fields. With the exception of one small area it is now wholly submarine. Output is approximately 21,000 tpd. Seams are 21/2 to 8 ft thick, and cover in the areas varies from 600 to 2300 ft, with an average of 1500 ft from sea bottom to the top of the seam. The seams dip in a seaward direction, pitches ranging from 6" to 4.0". The shoreline is the last place of entry to the seams and distance from the bank to the working faces is generally over 3½ miles, in some cases as much as 6½ miles. Ventilation is a problem and requires the construction of large permanent airways. Getting and Loading Coal: In 1925, in view of heavy pressures exerted by thickness of cover overlying the seams, roadways and pillars of the room-and-pillar system being worked began to break up and coal was lost. It was decided that a change in the method of extraction in areas with heavy cover was necessary, and experiments were made with many short walls and longwalls varying between 90 and 250 ft. Trial and error proved that the best operating length was between 400 and 500 ft, delivering all coal to the dip, with roadways to the face following a level course. The change-over was gradual, and the technique of roof control developed with the system, so that falls to the face are now very infrequent. An advantage of longwall mining is that it yields 95 pct extraction, especially important in coal seams of high quality or in seams where faults or disturbances restrict the workable areas. This percentage of extraction is based on the fact that the longwall advancing system takes development faces where pitch permits instead of driving headings and leaving roadway pillars. This system yields high tonnage during development and limits loss in extraction to duff left during operations. Accompanying disadvantages, on the other hand, are the heavy construction cost of main roadways and the necessity of driving all new flank face roadways through the gob. If the main roadways are driven through the solid, and large enough pillars are left on each side for protection against flank face weights, then the width of solid coal is approximately 1700 ft. This represents 10.6 pct of the coal available, or 89.4 pct extraction of the whole. These pillars, however, are of such size that they provide a useful pillar drawing area as a final operation of the mine. It may be that although the seam cannot be generally mined by the room-and-pillar method, it can be adapted to

Jan 1, 1955

By Frank Doxey

Louis A. Turnbull (U.S. Bureau of Mines, Washington, D. C.)—This paper is a well-organized concise discussion of progress in mechanization at coal mines in Nova Scotia during the last 50 years. It shows the perseverance and resourcefulness of those responsible for the production of high-quality coal essential to steel production in the area. The development of the Dosco miner and its use in longwall mining at the Sydney

Jan 1, 1955