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By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the syn-clines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no re-handling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor's bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1953

By E. O. Wagner, V. F. Parry, J. B. Goodman

The low-rank coals containing 10 to 50 pct natural bed moisture represent over half of the tonnage reserve of the available solid fuels of the United States, but only about 2 pct of United States coal production is derived from these fuels. Increased utilization of these large reserves will result if the bed moisture can be removed at the mine in high-speed low-cost processes. This paper presents a theoretical and experimental study of the factors affecting high-speed drying of coal dusts. The time required to heat particles of coal is proportional to the square of their diameter, and it is calculated that the probable time in minutes is equal to 6D2. In other words, a 1/16 in. coal particle should be heated to drying temperature in 1.5 to 2.0 sec. Likewise, 1/8 in. particles should be heated in about 6 sec. The net heat required to dry subbituminous coal and lignite varies from 240 to 375 Btu per pound, depending upon the moisture removed, and this heat can be obtained from 42 to 61 cu ft of hot gases at 600°F. The necessary heat also can be obtained from 7.5 to 11 cu ft of gas at 2000°F. A description of the design and operation of a pilot plant for flash drying coal dusts is presented. This plant dries 1/16 in. by 0 coal dust in suspension in 1.5 to 2 sec with 600°F gas at a rate of 1300 lb per hour per square foot of column area and operates continuously at an efficiency of about 85 pct. The design and operation of another pilot plant for drying 1/8 in. by 0 coal dusts Is also described. This unit dries coal in the fluidized state using the sensible heat of gases introduced at 2000°F and processes coal at a rate of about 1500 Ib per hour per square foot of grate area at 85 to 90 pct efficiency. Over 90 pct of the bed moisture of low-rank coal dusts can be removed by the methods described. The heating value of lignite can be increased 45 pet and that if subbituminous coal 25 pct in a few seconds. The physical and chemical properties of coal dusts before and after drying are presented in the report, and it is shown that the bulk density of the dried coal is equal to that of the raw coal. This phenomenon makes it possible to ship 25 to 45 pct more heat in a railroad car with consequent savings in freight costs. Introduction For several years the Subbituminous Coal and Lignite Section of the Coal Branch of the Bureau of Mines at Golden, Colo., has been concerned with upgrading low-rank coals by removal of natural bed moisture. These fuels contain 10 to 50 pct total moisture, which reduces their industrial value considerably. In 1938, investigations were initiated in cooperation with the University of North Dakota to study drying of lignite by the Fleissner process with heat derived from high-pressure saturated steam. Two pilot plants were built, and two reports were published giving the results of this work.l,2 It was demonstrated that lump lignite and subbituminous coal

Jan 1, 1950

By E. O. Wagner, V. F. Parry, J. B. Goodman

C. P. HEINER*—If you take out 35 pct of the total weight of the coal in the form of moisture, would that be about what it was in the case of North Dakota lignites ? V. F. PARRY (authors' reply)—That is about it. C. P. HEINER—That will be about a 35 pct loss and the final product will be about 11,000 Btu? V. F. PARRY—That is right. The heating value of the dried lignite is 10,000 to 11,000 Btu. C. P. HEINER—And the thought is not, then, to produce a commercial coal with that heavy loss, but to make other products—that is, synthetic fuel base, is that it? V. F. PARRY—The principal thought is to upgrade the coal, making it a more favorable raw material for industrial purposes, both for shipment and direct use. You see, the trouble in these lignites is that their moisture penalizes their

Jan 1, 1950

By R. L. Brown, A. C. Fieldner

Committee it supports pioneering research on the development of a coal-burning gas turbine and through the Mining Development Committee it promotes research on a new type of continuous mining machine that operates without using explosives. Other organizations have contributed major publications on fundamental and applied coal research in the last two decades, including the Minerals Industries Experiment Station of Pennsylvania State College, the Illinois Geological Survey and University of Illinois, the University of West Virginia, the University of Michigan, the University of Alabama, the University of North Dakota, the Massachusetts Institute of Technology, the Koppers Company, the U. S. Bureau of Mines, and, in recent years, the Institute of Gas Technology. Engineering development work, which is included in the overall estimate of $20,000,000, is conducted largely by manufacturers of equipment, and this portion of the estimate is difficult to gage. It represents a large proportion of the estimate and probably is too small rather than too large. For example, the development of mobile coal-loading machines and continuous mining machines, outstanding in recent mining research, undoubtedly involved a high cost. Coal Mining From an economic point of view, the greatest research development in the coal industry obviously has been the growing degree of mechanization in mining. The percentage of underground-mined bituminous-coal output mechanically loaded has risen from 3 pct in 1926 to 67 pct in 1949. Coal operators and equipment , manufacturers have united in designing machinery. Loading machines of large capacity and astonishing flexibility are in common use. Shaking conveyors transport coal from producing faces to transfer points from which it may be carried long distances underground and on the surface; and in many mines rubber-tired shuttle cars carry the coal from the loader to the

Jan 1, 1952

By Claude P. Heiner

W-hile the litle of this paper embraces the entire West Coast, the author, in the interest of simplification, has confined the discussion to California—particularly the central section. California's population has risen 45 pct since 1940. Its electric requirements have also increased, not only because of the growth in industry and population but also because of the widespread tendency of all types of electric consumers to use more electricity than ever before. The growth of California's electric utilities is shown by published data covering operations of three of its large systems and is given in Table 1. New Capacity Steam Driven In meeting these additional electric loads there has been a substantial increase in generation from fuel burning plants as shown by the published energy statistics given in Table 2. Increased Fuel Oil Consumption A study of construction plans of the three electric utility systems named in Table 2 clearly shows that expected new electric loads are to be met, to a large extent, by new steam generating capacity as shown by Table 3. D. D. Smalley, Vice President of the Pacific Gas and Electric Co., in an address before the Pacific Coast Electrical Association in June 1948, stated that California utilities plan to meet new electric loads through greater use of steal11 plank and that in 1951 approximately 45 pct of the area's electric production would be by steam as compared to 33.6 pct in 1946. He stated that if load growth develops as expected and if the year 1952 is dry, thereby requiring high load factor steam operation, the Pacific Gas and Electric Co. would need the equivalent of nearly 25 million barrels of fuel oil. Federal Power Commission reports show that the energy input to the Pacific Gas and Electric Co. system was approximately one billion kilowatt-hours greater in 1947 than in 1946. They also show that the steam plant operating capacity factor in 1946 was approximately 35 pct; in 1947 it, was 72 pct due, probably, to the extreme drought. able 3 shows that Pacific Gas and Electric Co. now has under construction 327,000 kw of hydro capacity which, by 1950, will bring its total hydro generating capacity to 1,313,899 kw. In addition to hydro capacity under construction that company also has 975,000 kw of steam generating capacity under construction, which by 1951 will bring its total steam capacity to 1,560,834 kw. If it is assumed that this total steam electric capacity is operated at a capacity factor of 45 pct, in 1951, over 6 billion kilowatt-hours would be generated. If the load growth on the Pacific Gas and Electric system continues at the current rate of approximately one billion kilowatt-hours per year and if it is assumed that 70 pct of this increase would be generated by steam (30 pct assumed to be generated by hydro facilities with average water conditions), the steam electric production in 1960 would amount to 12 billion kilowatt-hours, which would require fuel the equivalent of 24 million barrels of oil or 6 million tons of coal. Government Projects Inadequate Much has been written with respecl to the plans of the Department of Interior in regard to the installation of hydroelectric plants as a part of multipurpose dams relating to reclamation. flood control, and power generation. An examination of the effect of such plans on the power supply for central and northern California is of interest. The Depart,ment of Interior now has five 73,000 kw units at its Shasta hydro plant on the Sacramento River. It now appears that its three 25,000 kw unit Keswick hydro plant on the same river will be available during the winter of 1949 to 1950. Its estimates show that in a dry year there is sufficient water for operation of Shasta units at an average load factor of 36 pct and Keswick units at a load factor of 41 pct. It is also giving consideration to the development of the Pine Flat and Haas hydroelectric plants on Kings River for generating capacities of 45,000 kw each. The total installation of Government, units in Central Valley and Kings River in California is relat.ively small when compared with t.he

Jan 1, 1950

By Claude P. Heiner

mountain region to M tht Coast points for domestic consumption and for export are shown in Table 11. There is considerable disparity in rates from both Rock Springs, Wyo., and Castle Gate, Utah, to the four coast cities where the slack coal is to be used for purposes other than export. The rate on coal to be exported is the same from either starting point to any of the four coast cities even though there is a difference of as much as 341 miles in the shipping distance. It is interesting to note that the freight rate between Sunnyside, Utah, and Fontana, Calif., on coking coal is $5.05 per ton and that coal up to 8 in. can be moved on this rate if it is suitable for coking. This rate was published late in 1942 on a contemplated annual movement of more than 500,000 tons. There have been decreases in freight rates since 1923 on movements of slack coal from Utah into Seattle and Portland due to pressure on the railroads and to greater. quantity of coal shipped. It is the author,', opinion that a movement of slack coal in excess of 3 million tons of coal per year from Utah to any point in central California would justify a freight rate equal to that puhlished for Fontana, Calif.. or $5.05 per ton. The movement of 3 1/2 million tons of coal per year on the basis of 240 mine working days per year would require that 14,600 tons be handled each mine working day. If it is assumed that shipments could be arranged for a 6-day week, the average railway movement would he 11,200 tons, or approximately 3 trains containing fifty-five 70-ton coal cars per day. Such movements of coal would require railroad equipment represented by the investment amounts stated in Table 12 and entail the services of 250 men. Table 12 ... Railroad Equipment and Investment Required To Move 11,200 Tons of Coal a Day from Utah to California Railroad Equipment Cost 2800. 70-ton coal cara at $6.000 each $16,800.000 32 locomotives at $310.000 each. 9.920.000 Miscellaneous eauiument:.......... .5;000:000 Total........................ $31,720,000 It therefore appears that, under presently known mining methods, the lowest price at which coal could be sold f.0.b. the mine in amounts of 3 1/2 million tons per year for generation of power in central California would be $4.60. It also appears that the lowest freight rate that could be expecled between intermountain points and the central California area would be $5.05 per ton, making a total cost of coal delivered at a plant site of $9.65 per ton. Conelusions The following are the author's conclusions: 1. Coal mines in Utah and in the Kemmerer and Rock Springs districts of Wyoming could increase annual production by 6 1/2 million tons per year. 2. Under present conditions coal could probably be delivered to any steam electric plant in central California at a price not to exceed $9.65 per ton. 3. The use of coal at such a price, while higher than the equivalent present price of fuel oil, is entirely feasible. 4. There are adequate railroad facilities for movements of large quantities of coal from the intermountain region to the Rest Coast. 5. In a national emergency it probably would be extremely difficult, if not impossible, to obtain sufficient oil to meet requirements of the greatly expanded West Coast steam electric generatsing capacity. 6. The intermountain region contains ample coal reserves to supply all conceivable demands for West Coast power generation for a number of generations. 7. Increasing demands of labor threaten to lessen, if not eliminate, savings in cost of coal production through the use of new mining machinery. 8. Continuation of experiments in socialism by the Federal Government through construction of hydroelectric generating plants, particularly those unrelated to land-use reclamation, defies justification. Rates under this concept of a governmental function are subsidized through greater taxation of its people. Private capital is available to construct steam plants, or hydro plants where feasible, and should be permitted to continue in order to preserve the principles of our free enterprise system. DISCUSSION (L. C. McCabe and Robert P. Koenig, presiding) C. C;. BALl*—I was asked to lead off the discussion, but it is not with the thought that I might be able to add anything to the paper. The thoroughness with which it was prepared rather forestalls the asking of many questions. Your treatment, Mr. Heiner, is a very valuable contribution. I do want to suggest—that although you have limited your study to this specific question, with certain geographic limitations, many of the things in your paper apply just as well to the eastern coals. I want to agree 100 pct with your final conclusion concerning government-subsidized construction. Id. C. McCabe*—It is certainly worthwhile to take stock occasionally to see where we are going in problems of this nature. I agree with Mr. Heiner that ultimately the only reliable source of fuel that the West Coast has is coal hut the time factor is the difficult element to evaluate. Just before I came here I discussed this subject with N. B. Hinson, Vice President and Executixe Engineer of the Southern California Edison Company, and Chairman of the West Coast Inter-Power Exchange Committee. He has given much thought to utilities' fuel supply and it was very helpful to me in preparing a discussion of the paper to talk with him beforehand. Stock taking and forecasting of future development are essential to the continuing success of any enterprise. Mr. Heiner has called attention to the unprecedented growth of central and southern California and to the increased demands for fuel and power which have accompanied it. He discusses the increased fuel oil and natural gas requirements and the probable limits on the future use of these fuels and of hydroelectric power. In contrast to the calculable limits of these sources of electric energy, the author points to the availability of enormous reserves of coal in the Rocky Mountain States adjacent to the Pacific Coast which can be utilized for power generation. That there will be increased use of coal for power generation in the area under discussion is generally accepted hut it is in the timetable of such development that there is not complete agreement. In a recent report, Mr. Hinson reviewed the future power outlook for the Pacific southwest area. He pointed out that the use of steam plants in connection with water power plants in this region has made the maximum use of hydroelectric energy possible, and that the correct balance between hydro and steam generating plants produces the most economical overall system. Steam plants in the area which had been installed to protect against deficiency in hydro energy in dry years were used to carry war loads. Fortunately, no dry

Jan 1, 1950

By Paul A. Taylor

MUCH information concerning pegmatites which was thought to be true a few years ago has been proved false, and what is now actually known about some pegmatites is not true of many others. The erratic and seemingly unpredictable structure and variable composition of this class of mineral deposits has been widely emphasized. Even parts of the same pegmatite body exhibit marked differences in texture, mineral composition, width, and attitude. Constructive geological thinking in respect to pegmatites now aims to establish general laws rather than to stress the confusing diversity of features having no special economic significance. Substantial progress has been made in classifying different types of these deposits according to general features, internal structure, mineralogy, and origin. In some cases it has even been possible to block out tonnage reserves in advance of mining. It is still easy, however, to make highly erroneous predictions after a preliminary examination of a pegmatite prospect. Pegmatites are important to the economic well being of the country and to its military security. They furnish virtually all the feldspar, strategic mica, beryl, columbium, tantalum, and caesium utilized in the United States, as well as sundry other minerals and significant amounts of lithium and rare earth minerals and gems. With the exception of vermiculite, occasional ilmenite-rutile, and perhaps soda-lime feldspar and garnet deposits, basic pegmatites are of little economic importance. Consequently in this paper, as in common parlance, the term pegmatite generally relates to coarsegrained acidic rocks or what is aptly called giant granite. Available data indicating the size and importance of the production and trade in specified pegmatite minerals are summarized in Table I. Geological Features Much of the latest thinking on the economic geology of pegmatites is now available in a 115-page monograph' by a group of experts who participated with geologists of the Federal Geological Survey in the widespread wartime investigations. Doubtless the most significant feature of the monograph is indicated by the title, The Internal Structure of Pegmatites, but it also contains a vast amount of other new information and includes the assimilated concepts of many earlier writers, whose works are given in a comprehensive list of references. The shape, size, attitude, and continuity of many pegmatite bodies is controlled by the structure of the older rocks in which they occur. If the older rocks are easily penetrated, e.g., biotite schist, most of the pegmatites in a given district will be found outside the parent granite mass as exterior pegmatites. Marginal pegmatites are more prevalent if the older rocks are massive, unsheared, and sparingly jointed. Networks of pegmatites are abundant in highly-jointed rocks. In strongly foliated schists the bodies are usually lenticular, whereas in highly-folded areas they assume tear drop, pipe or pod-like, bow-shaped, or sinuous forms. Jahns2 recognizes five major shape classes: l—dikes, sills, pipes, and elongate pods; 2— dikes, sills, pipes, and pods with bends, protuberances, or other irregularities; 3-—trough-or scoop-shaped bodies with or without complicating branches; 4—bodies with the form of an inverted trough or scoop; and 5—other bodies, including combinations of the above and miscellaneous shapes. Many pegmatite deposits are scarcely big enough to be recognizable as such. Most of them, in fact, are small tabular deposits less than 4 in. wide and usually without economic concentrations of minerals. On the other hand, some pegmatites are more than a mile long and over 500 ft wide. The ratios of length to breadth range from 1 : 1 to 1 : 100. Although the vertical dimension bears no invariable relationship to strike length, tabular deposits or large lenses are often symmetrical enough to show nearly as much continuity down dip as in their horizontal extension, and some pipes or pods are amazingly persistent in the vertical plane. Small pegmatites often string along a fairly definite trend line; in a given district major bodies may lie roughly parallel, and where only a few of them do not, the erratically disposed bodies generally differ in composition from those conforming to the regular pattern. This does not apply, however, in all districts. Characteristically, pegmatite veins pinch and swell or split into branches. When they pinch out entirely it is often possible to find a new body by prospecting the extension of the strike or dip, but the chances of finding a hidden deposit are ordinarily too uncertain to justify much subsurface prospecting. Diamond drilling may yield valuable information as to the continuity of known deposits whose upper portions are well-exposed. Some deposits, in fact, can be proved up for hundreds of feet by surface trenching and then intersected by drill holes at various depths like any other vein-like deposit. Others twist and branch, apparently defying all efforts to explore them short of actual mining.

Jan 1, 1954

By F. G. Miller

Earlier studies of the mechanism of flotation showed that coal flotation products are contaminated by composite coal and mineral particles. The mineral matter of these particles has the dual effect of increasing the specific gravity and decreasing the coal surface of the particle available for bubble attachment. The present study showed that these particles are more loosely held in the froth column than the high grade coal particles, especially in the coarser sizes. Consequently, it is possible in the laboratory to produce a secondary concentration of the product by sprinkling the froth product with water to wash these loosely held particles from the froth during formation and recovery. To improve coal flotation, Bethlehem Steel Corp.'s research department has been working on a number of related problems based on both physical and chemical factors. The ideal is of course full understanding of the interrelated effects of both sets of factors. However, detailed studies1 of each physical factor have produced immediately usable results in commercial coal flotation and are providing the basis for our current work on the effect of chemical factors. Since these detailed studies provided the basis for the present investigation of froth sprinkling, their key results are highlighted here. For example, for a highly floatable bituminous coal, particle floatability was shown to be directly related to the percentage of coal surface available for bubble attachment and inversely related to particle mass. The data also underscored that in a given size range the addition of mineral matter to a particle makes flotation more difficult both by adding to the mass of the particle and by decreasing the amount of coal surface available for bubble attachment. Size also affects particle floatability in that as the size of particles of a given specific gravity increases, mass increases and the ratio between coal surface area and mass of the particle decreases. Flotation tests showed that any recovery of the coarsest particles in a coal feed is accompanied by progressively higher recovery of finer and finer particles. Consequently, fine high-specific-gravity particles are recovered with the coarse good coal. As a result, when the entire 14 mesh x 0 size range is processed by froth flotation, the good coarse coal is contaminated with high-specific-gravity fines. It is the small quantities of these fines that must be rejected from the froth product if flotation efficiency is to be increased above its present level. These principles and findings, together with work by Professor V. I. Klassen,' led to the froth sprinkling investigation reported in this paper. Working with Russian coals, Klassen had shown that froth sprinkling washes loosely held refuse particles from the froth and compensates for the water lost in the upper froth layers and increases its stability. As a result, the coal particles are retained in the froth and the loss of coarse coal from the froth is prevented. Decreases in ash of 0.1% to 1.5% were reported by Klassen. Working with a 28 mesh x 0 American coal, the author of the present paper found that single-stage flotation coupled with froth sprinkling produced a product comparable to that obtained by the more expensive method of roughing and cleaning.3 The investigation reported in the present paper expanded study of the effect of froth sprinkling to include each of the standard size fractions of a bituminous coal. The main findings are as follows: (1) For those size fractions between 48 mesh and 150 mesh, froth sprinkling produces a marked reduction in both sulfur and ash. In the case of those sizes coarser than 48 mesh, while this reduction is not as significant, the additional benefit of shorter retention time is still a plus for froth sprinkling. Furthermore, when froth sprinkling is coupled with normal fast flotation, the product purity approaches that obtained by starvation reagent feeding, i.e., gradual feeding of flotation reagent to float the recoverable coal from the pulp slowly in several increments. The specific results for these two size groups are: (a) 14 x 48 Mesh. By using froth sprinkling with fast flotation, product sulfur can be reduced to a level

Jan 1, 1970

By Shiou-Chuan Sun

THE effects of oxidation on the flotation behavior of sulphide minerals have been extensively studied 1-3 but no similar study has been made of coals. Coals of bituminous and lower rank undergo atmospheric oxidation during mining and storage, and their degree of oxidizing increases with temperature, 4,5 and time of exposure." The problem of recovering fine oxidized coal from culm piles, rivers, and streams by froth flotation has become increasingly important to the coal industry as well as to those concerned with air and stream pollution. This paper describes the flotation behavior of oxidized and unoxidized coals and carbons in the presence of oily collectors and presents the reagents found effective in the flotation of oxidized coals and carbons. Except where otherwise stated, flotation tests were performed in a 50-g Denver flotation cell. Solid feed for each test was 30 g, and the pulp dilution was kept at 8.5. The —65 mesh coal sample was wetted by agitation with water in the flotation cell for 6 min, then conditioned with a predetermined amount of collector for 2 min, and finally floated with pine oil as frother for 3 min. The oxidation of coals and carbons was performed in a vertical, electrically heated glass tube, shown in Fig. 1. One hundred grams of —65 mesh coal were fed into the glass tube and heated to 10° to 20 °C below the predetermined temperature of oxidation. Compressed air was admitted to the tube at a rate of 0.05 cu ft per min and the temperature was controlled by a variable voltage regulator. The oxi-

Jan 1, 1955

By R. Sporcic, P. J. Mudra

During the past year personnel of the Roof Control Research Group of the Bureau of Mines designed and developed encapsulated hydraulic cells for measuring pressure changes in coal in situ. Preliminary results from feasibility tests in progress in the laboratory and field indicate that the cells respond favorably to changes in pressures associated with a coal-mining environment. The ultimate objective is to use these cells to obtain engineering information relating to coal bump or coal outburst phenomena. Mining research has experienced a remarkable growth during the past decade and from all indications should surge forward at even a greater pace in the future. This increased interest in mining research has been brought about by the ever-multiplying number of technical problems being encountered by the mining industry in its effort to produce pay-dirt tonnages safely and economically from greater underground and surface pit depths. Among the various divisions in the field of mining research, the study of ground stress has received a large part of the overall attention. Numerous technical papers have been written with regard to ground stress conditions around mine openings and yet the total usable knowledge concerning this subject is extremely limited. Research teams around the world are endeavoring to bridge the gap between theoretical and laboratory approaches and the solution to design problems in the field. The problem of ground stress is not a new one for the coal mining industry. Operators in both eastern and western coal fields of the United States as well as in many foreign countries have long been plagued with a number of ground stress problems, one of which is commonly referred to as the coal-mine bump. This term or phenomenon has been aptly defined by earlier investigators as the violent and sudden failure of coal and adjacent rock. Direct results of such failures have been the loss of human lives and thou- sands upon thousands of dollars of sustained mine damages. In an effort to alleviate the magnitude and frequency of these failures, mine operators have employed and are presently employing various corrective techniques. These include: Inducing failure in heavily stressed areas by augering large-diameter holes into stressed pillars, using extensively yield-able steel arches in both long-term haulage and short-term face entries, backfilling unused or abandoned entries with mine waste to gain additional support, and using systematic pillar coal panel extraction techniques including mechanized longwall mining. The Health and Safety Activity of the Bureau of Mines has for a number of years contributed considerable time, funds, and talent in an effort to define better the ground stress problems associated with coal-mine bumps. As a part of this investigative program, in situ ground pressures are presently being studied in coal mines located in West Virginia and Utah. The description and development of the device being used to measure these pressures is the specific topic of this paper. CHOICE OF INSTRUMENTATION After consideration and evaluation of the various instruments and techniques that are being used by researchers in the field of rock mechanics to detetmine in situ stresses, a simple hydraulic system was selected for detailed investigation. Aside from simplicity, a hydraulic system is relatively inexpensive and also eliminates the need for electricity in its installation and operation. The hydraulic device referred to is, in essence, a simple unidirectional pressure-measuring cell. Hydraulic cells, which had been used for measurement of pressures in hard rock, were investigated but found not entirely suitable for use in coal, particularly since they involved (1) preparation and emplacement of cement grout at the test site, and (2) a waiting period before the cells became operational. At the outset, it was decided that certain design parameters be established. The cell should be simple and inexpensive, constructed of materials readily obtainable and easily built in the laboratory or field by unskilled labor. It was further decided that it was desirable to encapsulate the cell in a suitable mate-

Jan 1, 1963

By J. Visman

A simplified method of assessing the characteristics of a cleaning unit, including washability curves, yield figures, ash error, separating gravities, and error curve. FOR more than 25 years evaluations of coal cleaning units have appeared in many publications, and during this time considerable progress has been made in developing the theory and techniques of calculating washery results. A variety of characteristics have been defined over the years, many receiving general recognition during the first post-war Coal Preparation Conference in Paris in 1950. Coal preparation has become a science, and the approach to preparation problems is being revolutionized. Formerly, for example, if a coal operator wanted to know the performance of a given cleaning unit for a particular grade of his run-of-mine, he either shipped a carload to the manufacturer for a test run or relied on a report of a test carried out on a coal similar to the one he intended to clean. In modern practice the expensive, time-consuming test work generally can be avoided if the washability curves of the coal in question and the error curve of the cleaner are known. Normally a guarantee can be given on the strength of this information. General acceptance of modern standards has thus reduced the need for pilot-scale models of existing types of coal cleaners. Needless to say, the method requires exhaustive and accurate information on the cleaning unit and on all the additional circumstances affecting its results, information that ultimately can be secured only from actual tests made in the laboratory or in the washery. It is not the purpose of the present discussion to compare the merits of various characteristics, but to describe the general method of calculation, including a method of obtaining data with a minimum of laboratory work. All data presented here were derived from tests made at the plant with portable field equipment, using small quantities of varsol, carbon tetrachloride, and bromoform. It will be explained later that the use of standard fractional ash contents designated normalized ash contents aids considerably in reducing the routine testing of coal cleaners. Like any other statistical procedure, this method has limitations, and it is not recommended for any work in connection with guarantees. Even a varying commodity such as coal, however, will follow the law of large numbers, which is the basic law of statistics, thus allowing simplification of test work without appreciable loss of accuracy. The characteristics needed for evaluating cleaner performance are normally derived from a float-sink analysis of samples of the clean coal and the refuse, followed by ash determinations of the various fractions. Usually a sample of feed is also collected to find its ash content and to make a single specific gravity cut, both in connection with determination

Jan 1, 1955

[Abstmct of remarks by R. II. Sweetser arid suhscqueut discussion at the New York Meeting, February, 1926. The Transactions mill ultimately contain a report of work resulting from these discussions. Need for a standard of ualuration.—There should be an accepted method for evaluating the different kinds of bituminous coals based on their carbon just as that for iron ores is based on iron. Variation in the classes of coals and their desirability fol the different industrial uses is analagous to the case of iron ores. The iron people have developed a standard. In the past coals have not been intelligently distributed because the public has not been informed as to its uses. Many operators have known nothing about the final disposal of the coal shipped from their mines. Sales agents and retailers have had no means of evaluating coals except to make the price as high as the market will stand. Producers and distributors need to know how much more valuable the coals would be if prepared so as to contain less objectionable material. Mr. Swectser cnlletl on the Goal and Coke Committee of the Institnte to organize a study through producers and users to develop a method scientifically and comnereially sound E. A holbrook said that the information exists but the public is unwilling to investigate. He praised Ashley's1 nomenclature of the different kind of coal which has been accepted by the Coal Mining Institute of America and recommended to the American Enginecring Standards Commitee. G. S. Rice pointed out the wider variety and distribution of consumption of coal compared to Iron ore. C'. M. Young approved Mr. Sweetser's remarks and hoped lor a research institution, probably endowed, to study utilization of coal after it comes from the ground. Mr. Sweetser cited incidents to prove his former statements. He said that a reduction of 1 per cent. ash in the coal would cut down the cost of pig iron at a blast furnace 30 c. a ton, that the coal men should tell what the iron men would have to pay to get better coal. H. J. Prryn endorsed the need of blast-furnace operators to know the kind of coal from which their- coke is made and the latler's ash content, as this affects coke consumption. A. C. Fieldncr explained that the difficulty of evaluating coal on the basis of chermical analyses for calorimetric tests is: (1) representativr

Jan 1, 1927

By F. J. Chernosky

Most previous studies of coal flotation utilized chemically pure reagents. Since such reagents are not available in quantity, a study of various wmmercially available reagents as frothers was undertaken. Flotation procedures similar to those used in a large number of commercial coal flotation plants were employed. The reagents used were compared on yield-selectivity-wst basis. The selectivity was found to be relatively constant with yields above 70% but the reagent cost varied greatly. The types of reagents available were compared. In the past, a number of papers have been written on the use of various reagents as coal flotation frothers. For the most part, these have been an attempt to correlate the intrinsic chemical and physical proper- ties of the reagents or reagent groups to the floata-bility of various fine coals. Because of the type or nature of evaluations made, chemically pure reagents were normally used. Generally, these reagents are unavailable in commercial quantities. Therefore, the results obtained with these reagents were usually limited in usefulness to the plant operator or metallurgist. This paper presents the results of a study in which commercially available forms of various reagents were used as flotation frothers. Flotation procedures which would conform in general to those employed in a large number of commercial coal flotation plants were utilized. The objective of the test was to compare these reagents on yield-selectivity-cost basis, that is, 1) the yields obtained, 2) the ash levels obtained and 3) the reagent cost relating to these results. REAGENTS The reagents employed have been grouped into the following chemical classes: 1) aliphatic alcohols, 2) terpines, 3) glycols (synthetic water-soluble

Jan 1, 1963

By S. A. Braley

DRAINAGE water from coal mines is probably the most serious water pollution problem today, varying in importance according to location of the mines and geological structure. Drainage may be either acid or alkaline in character. Acid discharge, the most severely detrimental to a stream, is caused by natural oxidation of the sulfuritic material (FeS2) in the strata associated with the coal seam. Since the acid is the result of a natural reaction the acid water differs because it does not cease with abandonment of the mining operations. There is no known economical method of neutralizing acid mine water or any practical method to prevent oxidation of exposed pyrite. Since production of acid from a mine does not stop when mining stops, the total quantity produced depends entirely upon the excavated areas. The increasing volume of acid water in manv mines has greatly increased operating costs. Pumping is expensive and acid mine waters are destructive of all equipment, especially metals in pumps and piping, and necessitate the use of corrosion-resistant materials. Discharge of the acid mine drainage into streams neutralizes their normal alkalinity, causes them to become acid, and produces an environment unfavorable for aquatic life and unsuited for industrial or domestic use without costly treatment. Mine drainages vary in percentage composition over wide limits, although the usual dissolved substances are ferrous and ferric iron, aluminum, calcium and magnesium sulfates, and lesser amounts of sodium, potassium and manganese sulfates and chlorides. Some alkaline discharges may contain heavy concentrations of iron as iron bicarbonate. These waters may produce iron hydroxide deposits in the receiving stream but do not cause it to become acid. In the extensive literature on acid mine water there appears to be a great deal of confusion about the importance of various components and the methods for their determination. In many instances faulty conclusions have been drawn from use of unsuitable methods of analysis. It is desirable that the factors and terms used in evaluation of analyses of mine waters should be so clearly defined that any interested person could properly appraise any analytical report. Some analysts report complete chemical analyses of mine waters and neglect to record drainage volumes. Others report only a minimum of analytical data after taking no precautions to preserve the original composition of the water during the time elapsing between collection and analysis. Some use methods of analysis intended for so-called pure waters of the potable and boiler water classes. These methods are not applicable to highly buffered waters such as mine water. Probably the most common criteria for evaluation are pH, free acidity, or acidity or alkalinity to methyl orange or methyl red, total acidity or acidity to hot phenol-phthalein, and the sulfate content. If these determinations are made on carelessly collected samples after a few days to weeks standing in warm rooms, they do not in any way represent the character of the water flowing from the mines. It is hoped that a brief discussion of the fundamental value of some of these factors may lead to a bettqr understanding of the need for more careful evaluation of mine water discharges. The term pH is one used by chemists to express relative acidity or alkalinity in terms of concentration of effective hydrogen ion in a solution. It is defined as the negative logarithm of the hydrogen ion concentration or activity in equivalents per liter. pH = logarithm A neutral solution, which is one containing the same number of hydrogen and hydroxyl ions, has a pH of 7. As the hydrogen ions increase and the solution becomes more acid, the pH decreases toward zero; and as the hydroxyl ions increase and the solution becomes more alkaline, the pH approaches 14. When dissolved in water to a dilute solution acids like sulfuric and hydrochloric, commonly known as strong acids, ionize completely, and the pH or hydrogen ion concentration varies with molar concentration of the dissolved acid. However, in high concentrations of such acids, the pH or hydrogen ion concentration is less than the acid concentration because the acid does not completely ionize. In only very dilute solutions does the pH represent the total amount of acid that can be neutralized by an alkali. All ionization reactions are equilibrium reactions. If other chemicals are added to the solution of an acid and the added chemical produces an ion that is the same as one of the ions of the acid, the degree of ionization of the acid is altered and the pH changes to some value that represents the active hydrogen ion of the new solution. Thus if iron sulfate is added to a solution of sulfuric acid the pH increases, since the common sulfate ion suppresses the degree of ionization of the sulfuric acid and decreases the effective hydrogen ion. However, the total acidity of the solution is increased. There are two salts composed of iron and sulfate— ferrous and ferric sulfate. In these salts there is no hydrogen ion that can ionize to give an acid solution, but when they are dissolved in water, the pH is less than 7 and the solution becomes acid. This is caused by a reaction known as hydrolysis and is represented by the equations FeSO4 + 2H2O ? Fe(OH)2 + 2H + SO,: or Fe,(SO,)3 + 6H2O ? 2Fe(OH)3 + 6H+ + 3504 A solution with a total acidity of 5000 ppm according to the first equation will have a pH of 4.40 but one with an acidity of 75 ppm, according to the sec-

Jan 1, 1958

By J. R. Lucas, H. E. Rutherford

The primary objective of this research was to evaluate the feasibility of ultrasonic testing for the purpose of analyzing the structure and properties of rock material. A secondary objective of this study was to determine experimentally the elastic wnstants of various rock specimens with the use of ultrasonic equipment. Elastic constants can definitely be determined quickly and accurately by ultrasonic investigation. With proper equipment there are indications that ultrasonic waves could be utilized in the detection of flaws, such as fractures and planes of weakness, measurement of thickness, and the study of structure of rock material. Ultrasonic principles have been applied in many industries for the purpose of determining the structural characteristics of various materials. The principal applications of interest here are detecting flaws, measuring thicknesses, and determining elastic constants. The main advantage of ultrasonic analysis is that it is nondestructive. It also has extreme sensitivity, high penetrability, good accuracy, rapid response, and the necessity to have access to only one side of the specimen. These various applications and advantages lead one to think that ultrasonic principles might be applied to rock material in situ for purposes of analyzing its structural characteristics. Therefore, the primary objective of this research was to evaluate the feasibility of ultrasonic testing for the purpose of analyzing the structure and properties of rock material. A secondary objective of this research was to determine experimentally the elastic constants of various rock specimens with the use of ultrasonic equipment. The term ultrasonics is given to that branch of acoustics dealing with periodic sounds whose fre- quencies are above the audible range or above approximately 15,.000 to 20,000 cps. The first practical production of this high frequency sound was made possible after World War I due to developments of electronically controlled sound wave sources. At the present, the maximum practically attainable ultrasonic frequency is approximately 1000 megacycles per sec. Sound waves of sufficiently high frequency can be made to travel in a beam with little spreading. Therefore, such a beam can explore a material and detect nonhomogeneities in it by reflection. This is the principle of the ultrasonic reflectoscope invented by F. A. Firestone' in 1942. An ultrasonic beam, on pulse-penetratink a solid material medium, produces an echo when it strikes a flaw which has different acoustical properties than the surrounding medium. This echo can be detected by the same transducer which sends out the original pulse. This principle is the basis of the important industrial techniques of nondestructive ultrasonic testing of metals and other solid material. Later developments in the field of ultrasonic testing of solid material have been mostly improvements in equipment and instrumentation. To date, most of the industrial applications of ultrasonic analysis of solid bodies have been in the aircraft, electrical, and nuclear energy fields on the nondestructive testing of metal. The one solid material similar somewhat to rock material, upon which some work has been done to apply ultrasonic testing techniques to study structural integrity, is concrete. Jones and Gatfield3 briefly summarize this work on concrete. Their basic conclusion was that the strength of concrete was a function of the change of the longitudinal sound wave velocity in the concrete. Only recently has any extensive study been made to apply ultrasonic testing techniques to rock material. Lutsch studied rock material by an ultrasonic pulse method to determine the extent and degree of fracture of rock faces. Lutsch states that the number of observed reflections is a measure of the degree of fracture and that only ultrasonic waves are able to detect fractured zones subjected to high pressure. Suzuki, Sasaki, and Siohara5 studied the characteristics of the sonic wave in rock for propagation over a wide range of wave lengths. They indicated that, in rock, the propagation constant and attenuation of the

Jan 1, 1962

By L. Valentik

Many attempts have been made during the last 40 years to evaluate the performance of gravity separation equipement, that is, the effectiveness with which light and heavy particles are separated. The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris in 1950. The methods suggested by the Conference were accepted and very widely used in the last two decades. This paper discusses an improved method of evaluation in the light of the now-accepted standard presentation. The float-and-sink analysis of the product is presented on a Gaussian distribution curve, resulting in an easier visualization of the inherent difficulties of separation. The ogives of the distribution curve me then plotted, giving a quantitative measure of the deviation from perfect separation as an error distance instead of an error area. Illustrations of the new method are given both for gravel and for coal preparation, but the content is valid and applicable to other types of minerals which are separated by gravity methods. Many attempts had been made during the last forty years to evaluate the performance of heavy-media separation (HMS) equipment, that is, the effectiveness with which floats and sinks are separated.'-' The most comprehensive treatment of the subject was made by Cerchar at the 1st International Conference on Coal Preparation held in Paris. 6 The primary aim was the thorough understanding of the mechanism of separation and the unified presentation of data on gravity separation so that the evaluation and comparison of washery performance could be made from all over the world. No strict overall standardization has been achieved, but after the conference a more or less uniform presentation of performance was accepted, which, during the last two decades, has been very widely used. In this paper, illustration of the old methods and an improved method of evaluation will be given. HEAVY-MEDIA SEPARATION (HMS) PERFORMANCE CRITERIA In the ideal HMS process, all material lower in density than the specific gravity of separation (SGS) would be recovered as floats and all material of higher density would appear as sinks. In order to evaluate the misplaced material, the washery products are tested at the density at which the washing unit is operated. The original type of plot1,7, 8 is shown in Fig 1; this was developed primarily for coal cleaning units. The curve for raw coal represents the cumulative percentages of sink material. The refuse curve is also plotted as a cumulative sink, the percentages being expressed in terms of raw coal. This diagrammatic representation of the results of washing units has the merit of easy visual observance of the degree of separation obtained. The error areas (cross-hatched) are a measure of the amount of misplaced material and therefore they can be used to characterize the quality of separation. The ideal and actual separating performance between floats and sinks can be best seen from the partition curve developed by Tromp,2 where the ordinate is the percentage recovery of the sinks, and the abscissa is the specific gravity (Fig. 2). It can be seen from the shape of the curve that as the SGS is approached, the proportion of material reporting to the improper product increases rapidly. In fact, the SGS can be defined as the density of the material in the feed that is distributed equally between float-and-sink products. When the upper half of the curve is inverted, a shape similar to that of a Gaussian error distribution curve is obtained and therefore the analysis of gravity separation may be carried out by using the law of probability. The shape of the curve in Fig. 2 is determined partly by the density composition of the feed, and partly by the sharpness with which the unit separates floats from the sinks.9, l0

Jan 1, 1970

By T. C. Shelton, W. M. Huang

One of the problems arising from the use of modern machines and systems in coal mining is increased emission of explosive gases into the mine passageways. Interest in degasification of coal seams has grown. The authors summarize the results of coal permeability tests done at Virginia Polytechnic Inst. on one of the fundamental aspects of the problem. There are undoubtedly many factors that affect the emission of gas from coal but the permeability of coal to gas flow would seem to be fundamental. However, there is very little information on coal permeability to gas flow available in the U.S. at the present time, despite the fact that such knowledge would appear to be significant in any degasification effort. Studies made recently in the mining engineering laboratories at Virginia Polytechnic Inst. have determined values of permeability on 1-in. cubes of selected coals. COAL PERMEABILITY The permeability values given in this paper are based on adaptations of Darcy's law formulated by Henry Darcy in I856 from experiments with fluid flow in porous media.' Using air as the gaseous fluid, they were obtained by employing similar methods as are used to dete1:mine the permeability of petroleum and gas reservoir rock specimens. Permeability is. a measure of the ease with which fluids flow through a porous substance, and it is governed by the connected pores within a porous medium. If a few large-sized pores, fractures or channels are present within the medium, the medium can have great permcnability without significant porosity. It is reasonable to assume, therefore, that the permeability of coal, due to its structure, is based on fractures connecting the pore space left as the result of some degree of devolatilization or other changes of the substance from which coal was formed. The results of this irvestigation tend to validate this assumption in that the values obtained indicate a close relationship between the rank and permeability of coal. COALS TESTED AND PROCEDURE The coals whose permeability values were determined in this investigation are given in Table I.

Jan 1, 1962

By W. N. Poundstone, G. R. Spindler

A comprehensive report on extensive experimentation in degasification of the Pittsburgh Coal seam is presented. Detailed accounts of the procedures and results are given for permeability tests, vertical bore hole tests, measurement of gas flow from sealed drainage areas, horizontal rib hole tests, and water infusion tests. The removal of methane gas from coal seams in advance of mining is not a new idea with mining men—much attention has been given to the possibility for many years. In other parts of the world, a principal concern has been in the recovery and utilization of the gas and a number of successful projects have been developed in Europe and elsewhere. In this country, the principal motivation has been related to the safety aspect with some possible side benefits concerning costs and efficiency in mining as secondary objectives. To date, experimental work in this country has been largely or entirely exploratory to determine bed characteristics with respect to the liberation of combustible gases to investigate or establish the possibility of degasification in advance of mining and as an aid in projecting mine ventilation systems to provide for safe, effective and economical handling of combustible gases liberated during mining. With these objectives in mind, the School of Mines and Engineering Experiment Station at West Virginia University, in cooperation with the Christopher Coal Co. of Osage, W. Va., initiated some experimental work in May 1952 and the project has been more or less continuous since that time. The work has been centered entirely in mines of the cooperating company in Monongalia County, W. Va., and Greene County, Pa. LOCATION AND GENERAL CONDITIONS OF THE AREA The principal mining in the area involved is in the Pittsburgh coal seam which lies at the bottom of the Monongahela series of the Pennsylvanian. The Pittsburgh coal in the area is generally uniform in its structure and physical characteristics. The bed thickness averages about 7 to 8 ft and the coal has a characteristic blocky structure as the result of pronounced face cleats and definite, but somewhat less pronounced, butt cleavage. It is a medium volatile coal, averaging about 37 pct, and is marketed principally as a steam fuel and for metallurgical purposes. The Pittsburgh coal outcrops along the Monongahela River and dips gradually, but not uniformly, slightly north of west (about N75°W). The westward extensions of the bed, from the areas contiguous to the Monongahela River which have been heavily mined, constitute the principal reserves and are all well below drainage with the depth of cover increasing with the distance from the river. In general, with deeper cover, it may be expected that methane liberation will be more of a problem in future mining as compared to areas mined previously; therein lies the interest in possible de-gasification in advance of mining. Fig. 1 shows the Pittsburgh coal area in the vicinity of Morgantown in Monongalia County, W. Va., and Greene County, Pa., in which the experimental work discussed is located. Fig. 2 shows the contours on the Pittsburgh coal in the test area. It will be noted that the first two bore holes for the experiments were at the edge of a bench or terrace with considerable change in the gradient in the immediate vicinity of the test holes and that the same general condition applies to later test areas. Mining is intensively mechanized in the area and in the mines involved in the experimental work with corresponding rapid advance of working faces a typical characteristic. In the areas now mined the total methane liberation ranges from moderate to relatively heavy, but the actual face liberation is extremely variable and may range from practically

Jan 1, 1961

By A. C. Watts

For several years, a mine in Colorado experienced considerable trouble from small fires caused by the blasting of coal. Although a well-known make of permissible powder was used, it was first thought that the powder ignited gas that might be present and that this, in turn, set fire to line brattices or to the coal. Experiments were, therefore, conducted with another make of powder, but the results convinced the superintendent of the mine, Robt. Williams, Jr., that the fires were not caused by the powder but by the high-voltage electric current used for detonating the shots. The mine "makes" gas very freely, giving off from three-fourths to one million cubic feet of methane every 24 hours. The coal is a good grade bituminous of a very free burning quality. Electric shot-firing from the outside, after all men and animals were outside, had been practiced for a number of years. The shooting circuit was connected with the main power circuit at the mouth of each entry. Shooting circuit switches were mounted in boxes, which were kept locked with the handle of the switch, in open position, showing through the bottom of the box. The door of the box could not be shut and locked when the shooting switch was closed. At the entrance to each room, or working place, a small switch, called the miner's switch, was placed on the shooting line. It was the duty of the miner, when entering his place at the beginning of the shift, to see that this switch was open and that it was kept open until he left the placc at the end of the shift to go out of the mine. Grounded return was used for main power lines. The current was 500 volts, direct current, with 100 to 300 amp. The nearest working face was 9600 ft, from the shot-firers' switch outside the mine and the most remote working face was 2½ miles. The various working faces were scattered over a wide area between these two points. A description of the main power line would be valueless; the conductors were of sufficient size, the drop in voltage being not more than

Jan 1, 1927

By J. P. Newell, R. W Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953