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By Adam L. Wesner, A. C. Richardson

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. DURING World War I1 the demand for electrode carbon was greater than could be met by the supply of petroleum coke available for this use. It was believed that coke made from an extremely low-ash coal might be a suitable substitute for petroleum coke. Requirements for electrode carbon to be used for the production of aluminum are that the silicon plus iron content shall be not more than 0.14 pct of the coke. However, during the period when petroleum coke was in short supply, it appeared that relaxations of the specifications might be allowed to permit 0.4 pct iron plus silicon in the coke. For this work, the specifications were that the coal should contain less than 1 pct of ash, and the combined iron and silicon content of the coal should not exceed 0.28 pct. Heavy-liquid separations made on a number of coals showed that if certain size fractions of some coals Were separated at low specific gravities the resulting float products would meet the required specifications. The objective of this work was to develop a continuous, commercially feasible process for producing low-ash coal. After consideration of various processes, the method chosen was a float-and-sink separation using a solution of calcium chloride as the separating medium. Eagle Seam coal was used for the experiments. The —10 +35 mesh fraction was found to be the most promising feed for producing the low-ash content coal. A batch of 30 tons of — --in. coal was screened at 10 and 35 mesh; the yield of —10 +35 mesh was 20 pct of the feed. This fraction comprised the feed for the separation tests. In spite of the fact that this portion was screened a second time at 35 mesh, the coal still contained 16 pct that was finer than 35 mesh. Three-fourths of the undersize was in the —35 +48 mesh range. A representative portion of the —10 +35 mesh fraction was subjected to batch float-and-sink separations to determine the best gravity to be used in the continuous tests to obtain a high recovery of low-ash content coal. The heavy liquid used was a mixture of carbon tetrachloride and benzene which assured complete wetting of all of the particles. As usual in batch separations, adequate time was allowed for each separation so that even the near gravity particles had ample time to separate. The separations were made at increments of 0.01 sp gr from 1.25 to 1.28 and also at 1.32 and 1.59. Each specific-gravity increment was assayed for ash content. From these data it was concluded that 1.27 sp gr was best for the continuous separation of the coal. The products from the batch separation were composited into float 1.27 and sink 1.27 sp-gr fractions; these were assayed for ash, iron, and silicon. Table I shows the results of the batch separations. The composite float 1.27 fraction contained 66.04 pct of the total weight, and had an analysis of 0.79 pct ash, 0.073 pct iron, and 0.112 pct silicon, or 0.185 pct iron and silicon combined. These data are indicative of the results possible if a perfect separation is made at 1.27 sp gr under static conditions. Inasmuch as the method under investigation is a dynamic system these results could not be obtained. It remained to be determined, however, just how close to the theoretical results the actual continuous separation would be.

Jan 1, 1953

By Adam L. Wesner, A. C. Richardson

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. DURING World War I1 the demand for electrode carbon was greater than could be met by the supply of petroleum coke available for this use. It was believed that coke made from an extremely low-ash coal might be a suitable substitute for petroleum coke. Requirements for electrode carbon to be used for the production of aluminum are that the silicon plus iron content shall be not more than 0.14 pct of the coke. However, during the period when petroleum coke was in short supply, it appeared that relaxations of the specifications might be allowed to permit 0.4 pct iron plus silicon in the coke. For this work, the specifications were that the coal should contain less than 1 pct of ash, and the combined iron and silicon content of the coal should not exceed 0.28 pct. Heavy-liquid separations made on a number of coals showed that if certain size fractions of some coals Were separated at low specific gravities the resulting float products would meet the required specifications. The objective of this work was to develop a continuous, commercially feasible process for producing low-ash coal. After consideration of various processes, the method chosen was a float-and-sink separation using a solution of calcium chloride as the separating medium. Eagle Seam coal was used for the experiments. The —10 +35 mesh fraction was found to be the most promising feed for producing the low-ash content coal. A batch of 30 tons of — --in. coal was screened at 10 and 35 mesh; the yield of —10 +35 mesh was 20 pct of the feed. This fraction comprised the feed for the separation tests. In spite of the fact that this portion was screened a second time at 35 mesh, the coal still contained 16 pct that was finer than 35 mesh. Three-fourths of the undersize was in the —35 +48 mesh range. A representative portion of the —10 +35 mesh fraction was subjected to batch float-and-sink separations to determine the best gravity to be used in the continuous tests to obtain a high recovery of low-ash content coal. The heavy liquid used was a mixture of carbon tetrachloride and benzene which assured complete wetting of all of the particles. As usual in batch separations, adequate time was allowed for each separation so that even the near gravity particles had ample time to separate. The separations were made at increments of 0.01 sp gr from 1.25 to 1.28 and also at 1.32 and 1.59. Each specific-gravity increment was assayed for ash content. From these data it was concluded that 1.27 sp gr was best for the continuous separation of the coal. The products from the batch separation were composited into float 1.27 and sink 1.27 sp-gr fractions; these were assayed for ash, iron, and silicon. Table I shows the results of the batch separations. The composite float 1.27 fraction contained 66.04 pct of the total weight, and had an analysis of 0.79 pct ash, 0.073 pct iron, and 0.112 pct silicon, or 0.185 pct iron and silicon combined. These data are indicative of the results possible if a perfect separation is made at 1.27 sp gr under static conditions. Inasmuch as the method under investigation is a dynamic system these results could not be obtained. It remained to be determined, however, just how close to the theoretical results the actual continuous separation would be.

Jan 1, 1953

By R. Q. Shotts

Nitric acid oxidation rate analyses of three coals, previously studied microscopically by the Bureau of Mines, revealed three components. Relative quantities agree with those found for the four components given by the Bureau and results are consistent with current ideas of coal constitution. Possible multi-component composition for bright coal and a reactivity-rank relation are suggested. THE physically dissimilar components of bituminous coals often are easily recognized mega-scopically. Under the microscope, reflected light or light transmitted through thin sections reveals the presence of the different components, even when these are intimately mixed. Optical methods for the quantitative estimation of the relative abundance of the various components, both by means of thin sections and by particle count, have been fully described.'. ' It has long been recognized that there are chemical and physical differences between the various petro-graphic components of bituminous coals, although analytical differences usually are small.:'. ' Only in the case of fusain have chemical differences been used for quantitative determination of a component. C. C. Hsiao and associates, at the Mineral Industries Experiment Station of the Pennsylvania State College, have described a method of analysis which is based upon the differences in the rate of nitric acid (8N) oxidation, fusain, and the other components of coal."," The reproducibility of their method and its applicability in checking microscopic determinations of fusain content have been supported by several independent investigations.'. " The writer has proposed the use of differences in oxidizability for the estimation of other components." "' The results of the oxidation of whole coals and of float-and-sink fractions of coals were reported. In most cases the plots of the logarithms of the percent dry, non-fusain, organic residue from oxidation, against time, revealed the presence of at least two distinct components. Both components appeared to oxidize according to a first order law, but the reaction constants for the components were distinctly different. One or more of the dull density fractions were found to contain but one component, and some of the lower rank coals oxidized in such a way as to suggest the presence of three components. A suitable way to check the identity and significance of the components delineated by oxidation would be to analyze a sample of coal both by the nitric acid oxidation procedure and by a microscopic method. The writer was wholly unfamiliar with either of the microscopic techniques commonly used, and to make such a comparison it was necessary to rely upon microscopic analyses made by someone else. It is hoped that some laboratory which is equipped to make both types of analyses will some day make them upon identical samples. During the past 20 years, four Alabama coals have been analyzed petrographically and the results published by the United States Bureau of Mines. They are: 1—Flat Top mine, Mary Lee bed; 2—Empire mine, Black Creek bed; 3—Wylam No. 8 mine, Pratt bed, all in the Warrior field; and 4—Soot Creek mine, Fairview bed, in the Coosa field."-" Of these, only the Flat Top mine is still operating. Because of the closing of these mines, it first appeared necessary to rely upon the indirect and unsatisfactory procedure of sampling the beds in other mines located as near to the closed mines as possible. Upon investigation, however, it was found that the Bureau of Mines still had, in storage, the very same samples which had been used in the published petrographic studies. The Bureau very generously furnished about 2000 g each of the Pratt, Mary Lee, and Fairview bed coals, largely lumps but with some fines. The blocks of coal, when received, still were covered by the paraffin coating which had been placed on the polished surface, in the case of the Mary Lee coal almost twenty years ago. Procedure The procedure for oxidizing the coal sample and removing the alkali-soluble humic acid has been described. In the present study, oxidation periods of 1/6, 1/3, 1/2, 3/4, 1, 2, 3, and 4 hr were used. All oxidations were made in triplicate. After the paraffin had been removed in boiling water and the coal washed carefully with cold benzene, the entire sample of approximately 2000 g, obtained from the Bureau of Mines, was crushed to pass a No. 4 sieve. About 200 g of this material was pulverized to pass

Jan 1, 1954

By R. A. Falconer, H. L. Lovell

With increasing requirements for fine coal cleaning and the many advantages of the cyclone operating on a coal-water slurry, a series of studies designed to achieve an enhanced understanding of the cyclone parameters has been pursued. The purpose of this investigation has been to vary a series of dimensional parameters with an experimental 15-in. diam hydro-cyclone, and to determine the relative effects of parameter variations upon ash and sulfur removal. This study of the hydrocyclone is a continuation of a series of investigations, conducted in the Dept. of Mineral preparation at The Pennsylvania State University, which are directed toward the removal of ash and sulfur constituents from fine bituminous coal by size-density separations in cyclones, and toward a fuller understanding of the complex mechanisms involved in such separations. The purpose of this investigation has been to vary a series of dimensional parameters with an experimental 15-in. diam hydrocyclone, and to determine the relative effects of parameter variations upon ash and sulfur removal. Of special interest was the response to cone angle variation, since it is the high included cone angle that generally distinguishes the hydrocyclone as a gravity separator rather than a classifier. DESCRIPTION OF THE HYDROCYCLONE Fig. 1 is a dimensional drawing of the hydrocyclone used in the investigation. The unit was designed and fabricated by Heyl and Patterson, Inc., for experimental use. The dimensional parameters considered were: 1) Three cone sections having included angles of 45, 60, and 95,degrees, 2) Three 3 1/8 in. diam vortex finders having lengths of 15, 17, and 19 1/2 in., 3) Three 4 1/16 in. diam vortex finders having lengths of 15, 17, and 19 1/2 in. 4) Two apex sections having diameters of 2 and 2 1/4 in. The feed inlet orifice had a fixed diameter of 3 in. The hydrocyclone test installation was made at The Pennsylvania State University pilot plant at Barnes-boro, Pa., which operated at 50 tph. During the period that the hydrocyclone was tested, research was in progress at the Barnesboro site to reprocess refuse material from a near-by coal preparation plant in

Jan 1, 1968

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

One of the chief difficulties with which the operator of a coal washing plant has been forced to contend is the handling of the very fine coal. First he has the problem of separating the fine coal from washery water. This is usually accomplished by the use of settling cones or Dorr thickeners: in either event the separated fine coal contains a high amount (40 to 80 pct) of water and, somewhat loosely, dependent upon the size of the coal particles present, may be known as sludge or slurry. His second problem is the satisfactory further dewatering of such separated sludge and/or slurry. At the coal washing plant of the Colorado Fuel and Iron Corp., washed coal from the Deister Plat-O tables is sluiced to a stationary inclined de-watering screen provided with 3/16 by 1-in. slots, and the coarse coal passing over this screen is conveyed by a Baum-type elevator and a flight-type distributing conveyor to a battery of Carpenter centrifugal driers. The fine coal and water passing through the openings of the primary dewatering screens, plus the fine coal passing through the screens of the centrifugal driers (and no small portion of which is the result of the disintegrating action of these driers) is pumped to two 30-ft Link Belt settling cones. The sludge recovered as underflow from these cones was formerly dewatered on stationary screens having 1/16 by 1-in. slots and discharged on top of the coarse washed coal in the Baum drainage conveyor; but after a relatively few hours of op eration the circulating load of fine coal on this conveyor built up to a point where it became necessary to reduce the quantity of feed to the washing tables. The original arrangement of our coal washing and dewatering equipment is shown in Fig 1. In order to regain the loss in active throughput which had been caused by this circulating load, an installation of six vibrating filters was made during the year of 1936. This resulted in an increase in normal tonnage of washed coal from 1430 to 1750 tons per 8-hr shift, a gain of slightly over 22 pct. This increase in throughput has made a very substantial decrease in the " conversion cost" per ton of washed coal. Since the vibrators were installed, we have gained about 2,000,000 tons in production, compared to what would have been produced in the same number of shifts at the former rate. Putting it in slightly different form, we have saved in the past 12 yr the cost of 1400 operating shifts in the washery. Fig 2 shows the location of the vibrating filters as applied to the original flowsheet. It will be noted that the sludge from the bottom of the settling cones is passed through a distributing box and thence to the filters. The de-watered coal from the filters is de-livered to the conveyor receiving the dried coal from the Carpenter driers.

Jan 1, 1950

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

W. J. PARTON*—I have not had the opportunity to read this paper, and I do not have a written discussion. However, I thought it might be interesting for me to relate some of the experiences we had with equipment similar to the vibrating filter as described by the' authors. At the Tamaqua flotation plant of the Lebigh Navigation Coal Co. approximately 40 tons per hour of froth concentrate carrying 60 pct by weight moisture are produced. The major problem encountered at this plant is the dewatering of this coal froth so that a satisfactory product can be sent to market. In the original design of the plant a centrifuge of solid bowl type was included for de-watering this material. The centrifuge did not work out as well as we had hoped. High maintenance costs and moisture content in the cake were obtained. A Robbins dewatering screen was installed at a later date with the idea of using it in conjunction with the centrifuge. The froth concentrate from the flotation cells was fed directly to the Robbins dewatering screen. The cake from the screen carried approximately 55 pct of the feed solids. Moisture in the cake was approximately 24 pct by weight. The underflow from the screen carried 45 pct of the feed tonnage at about 80 pct moisture by weight. The underflow product was then pumped into the centrifuge with the idea of using the cen-trifuge for recovering the tonnage lost through the screen. This circuit did not operate as satisfactorily as we expected. The only benefit derived was in the reduction in the power consumed by the centrifuge. The maintenance on the centrifuge was approximately the same as previously. The next step in our experiments was to pump the underflow from the screen into a cyclone thickener which was mounted directly over the vibrating screen. This thickener increased the concentration of the solids to approximately 60 pct by weight and dropped the mate- rial back 011 the filter cake which had formed toward the discharge end of the screen. Unfortunately, the screen was not capable of handling this additional tonnage, and our experiments stopped at that point. We have been considering installing a second screen to make possible the complete mechanical dewatering of this product by the use of the dewatering screen and the cyclone thickener. Another possibility under study is to pump the underflow from this screen to a thickener which is available in the flotation plant, and to combine this thickeued underflow with the original feed going to the screen. Again, however, a second dewatering screen will be required to handle the total tonnage. 0. R. LYONS*—I had an opportunity to read this paper ahead of the meeting, and I did a little pencil engineering on it. As Mr. Bertholf said, it is very difficult to make a comparison and to carry the results of work at one plant over to what might be expected at another. What I did was to find information on filtering operations more or less comparable to the type of operation that Mr. Bertholf has with his vibrating filter. The only information that I was able to find was for drum type filters, and I found the operating characteristics of the vibrating filter and the drum type filters were very similar. The moisture contents of the cakes were almost identical. The output per square foot was about the only way that I could compare their capacities— using square foot of screen area against square foot of filter area—and I found the capacity of the vibrating filter to be slightly greater per unit area than the capacity of the drum-type filters. W. H. NEWTON†—Do I understand that the only escape for the solids is by overflowing the thickener? That is, does the filter have a chance to recover all the solids except that lost in the thickener overflow ? W. M. BERTHOLF (authors' reply)— Actually, the only escapc from that part of the circuit is over the top of the thickener. There are other places the fines could be lost in the washery. but once they get into that part of the circuit, they must go over the top to escape. W. H. NEWTON—I would like to ask Mr. Lyons if, in the study of rotary filters, he has any basis for comparison of operating costs? 0. R. LYONS—No, I had no information on costs. The only information I was able to find was on screen size, moisture content, and tonnage output per unit area. W. L. McMORRIS*—Are you wasting that overflow water or re-using it? W. M. BERTHOLF—Right now, we are not re-using it. D. R. MITCHELL† —What is the approximate per capita cost of one of these units? W. M. BERTHOLF—It appears to be somewhere in the neighborhood of $200, for the screen. W. H. NEWTON—The cost would be about $2500 for the complete unit including the vibrating power unit. G. A. VISSAC‡—I do not like to come on the floor after I have been talking so long, but I thought you might be interested in our experience in dewatering, as well as drying our very fine coals. We have used both centrifuge and vibrating screens. The type of vibrating screens we have used in Canada are called the Zimmer. That is a screen of German construction, and I guess it is along the same lines as the dewatering screens you are using now. We use wedge wires, and the minimum size opening is a quarter of a millimeter. In our experience, the cheapest way is still a dewatering bin. A dewatering bin takes 48 hr to do work that takes 20 min in a dewatering screen. We use old wedge wire from our driers which we cover with brattice cloth, and

Jan 1, 1950

By B. Trafton, H. L. Hartman

An earlier paper1 introduced a rapid solution for mine ventilation network problems, employing the digital computer. In this paper, refinements and additions to the computer program which expand its scope of application and a solution to a ventilation network having a coal-mine configuration are presented. The revised program-written in FORTRAN for the IBM 7074 computer in Penn State's Computation Center-is applicable to any mine ventilation network employing natural splitting and having its fan located externally. It accomplishes the following: 1) Distributes an assumed total quantity of air flow among all airways of the network, performing iterations until an acceptable level of error is attained and the flow in each airway is known. 2) Computes the pressure drop (head loss) in all airways and the overall network and mine head losses. 3) Determines the operating point for the mine and a given fan, adjusting all air flows and head losses to the actual values. The basis of analysis used is the Hardy Cross method of successive approximation for solution of any fluid network, explained in detail in the earlier article. In brief, corrections to assumed air flows are computed for all the loops of a network; following distribution to the airways of each loop, the process is repeated until acceptable accuracy is attained. When a hand or desk-calculator solution is used, there is advantage to close approximation of the actual flows in the original assumption of air-flow distribution in the network. Since publication of the previous paper, in which assumption of initial quantities was also necessary in the computer solution, the program has been modified to make the first distribution itself, assigning the total quantity to one specified path in a split and then proceeding to correct the flows. The number of iterations required for a high-accuracy solution is little affected, at least in networks of few loops, 10 to 15 being sufficient to reduce the error to 1 pct or less. The principal limitation of the original program was that it contained no provision for handling a fan. Further, it did not determine the overall head loss in the network and was not designed for multiple parallel openings, such as occur in coal mines. The revised program was written to overcome these disadvantages and also to print out the adjusted flows for all airways, corrected for the system operating point. At the present time, the program can handle only an external fan. The procedure employed in solving a ventilation network is now as follows: 1) Input data consisting of airway characteristics and loop designation along with assumed direction of flow, for each branch of the network and any trunk airways, are fed into the computer. 2) Employing the resistance relationship, where R is resistance, K is friction factor, P is perimeter, L is length, A is area, and n is number of airways (where multiple intakes and returns are employed); values are computed for all airways. 3) An arbitrary, continuous route of flow through the circuit—from beginning to end—is selected and the path identified by airway numbers and read in. 4) Based on a total assumed quantity of flow Q (usually 100,000 cfm.) the computer employs the procedure described in the previous article in distributing quantities of flow among all the airways until a satisfactory level of accuracy has been attained (in the current program, the machine is turned off when the cumulative corrective error in cfm, is reduced to five times the number of airways). 5) Airway head losses H are calculated by the relation H = R Q2 and cumulated along the original designated flow route to find the overall head loss for the network. Trunk airways are then included to obtain the mine head.

Jan 1, 1964

By H. L. Washburn, G. L. Judy

DURING recent years thermal drying of fine coal has increased at a tremendous rate, but very little discussion of the several excellent units has found its way into print. This is particularly true for operating and maintenance costs, and the operator who is faced with his first fine coal drying installation must depend chiefly on hearsay when trying to determine what his drying costs will be. This article presents the maintenance and operating costs for the seven drying units now installed in preparation plants of Pittsburgh Consolidation Coal Co. and lists other cost information that should be obtained before any particular drying unit is selected. Plant Description: The fine coal drying section of Plant A consists of four Flash driers operating on two furnaces equipped with spreader stokers. Plant B has one Flash drier operating (together with a screen-type coarse coal drier) on one furnace equipped with a spreader stoker. Plant C has two Cascade-type driers operating on a single furnace with a spreader stoker. Plant Costs: Tables I, 11, and III present the cost of operating these units as reported by the individ- ual plants. As can be seen, the labor rates vary, depending on the amount of overtime paid and the time interval the cost figures cover. There is considerable variation in thermal units used and in the value given to the thermal units at the particular mine. There is some difference in power costs at a given location and, finally, the power costs of conveyors leading to the drying plants were included in some cases. Maintenance supply costs are simply a tabulation of the cost of all supplies used on the driers, furnaces, and stokers. Depreciation is not included, and the tables omit items such as supervision, plant cleanup, oiling, general maintenance, taxes, and insurance. These tables in themselves do not offer a true comparison of costs. This information must be reduced to some common basis. But what is this common basis to be? A mine operator must calculate his overall mine and preparation plant costs in terms of tons of clean coal sold. Unit costs on pieces of equipment within a given plant are correctly calculated on the same basis. However, when unit costs are compared between one plant and another, some thought must be given to unit design. This reasoning is similar to the old argument concerning i solid bowl vs basket-type centrifuges—the units themselves are not directly comparable without their supporting equipment. So it is with thermal driers. Except in rare cases of low entering moisture i content, driers are sized on the basis of water that must be evaporated, while the wear each unit gets de-

Jan 1, 1958

By Vard H. Johnson

IN 1943 the U. S. Bureau of Mines undertook drilling in an effort to develop new reserves of coking coal in an area near Paonia, Colo., as a part of an attempt to alleviate the shortage of known coking coal of good quality in the western United States. Geologic mapping of the area was undertaken by the U. S. Geological Survey with the purpose of first furnishing guidance in location of drillholes and later aiding in interpreting the results of the drilling. The drilling program was under the general supervision of A. L. Toenges of the U. S. Bureau of Mines. J. J. Dowd and R. G. Travis were in charge of the work in the field. Geologic mapping was started by D. A. Andrews of the Geological Survey in the summer of 1943 and was continued from the spring of 1944 to 1949 by the writer. The first few holes drilled failed to locate coking coal, but in the summer of 1944 coking coal was discovered by drilling 6 miles east of Somerset, Colo., the site of present mining. In the succeeding years, 1945 to 1948, 100 to 150 million tons of coal suitable for coking were blocked out by drilling. The ensuing discussion of the geologic controls on the distribution of coking coal in the area is based on the geologic mapping as well as the drilling done in the Paonia area, more complete descriptions of which have appeared or are in process of publication."' In order that the possible geologic controls affecting the present distribution of coking coal may be considered, it is necessary to discuss briefly the indicators of coking quality coals. Coking Coal Coal that cokes has the property of softening to form a pastelike mass at high temperatures under reducing conditions in the coke oven. This softening is accompanied by the release of the volatile constituents as bubbles of gas. After release of the contained gases and upon cooling, a hard gray coherent but spongelike mass remains that is referred to as coke. This substance varies greatly in physical properties and, to be suitable for industrial use, must be sufficiently dense and strong to withstand the crushing pressure of heavy furnace loads. Western coals have a generally high volatile content and therefore form a satisfactory coke only when they attain a rather high fluidity during the process of heating arid distillation in the coke oven. When this high degree of fluidity is developed, the volatile constituents escape and leave a finely porous coke. On the other hand, when the degree of fluidity is low the product is an excessively porous and therefore physically weak mass that is called char." Small quantities of oxygen present in coal are believed to decrease the fluidity of the material during the coking process and to favor the development of char rather than coke. In consequence, coal chemists have for some time considered the possibility of developing an index to coking qualities by inspection of chemical analyses of coals.' A formula has now been developed that does permit a rough preliminary estimate of the cokability of coal on the basis of the analysis on an ash and moisture-free basis. Coals may be eliminated as possible coking fuels if the oxygen content is greater than 11 pct. Similarly the ratio of hydrogen to oxygen must be greater than 0.5 and the ratio of fixed carbon to volatile constituents must be greater than 1.3. If the coal, on the basis of these limiting factors, appears to have possible coking qualities, the following formula permits determination of the coking index: a+b+c+d Coking index = -------- 5 a equals 22/oxygen content on ash and moisture-free basis, b equals two times the hydrogen content divided by oxygen content on moisture and ash-free basis, c equals fixed carbon/l.3 x volatile matter, and d equals the heating value on moist, ash-free basis/13,600. Coking indices higher than 1.0 suggest that the coal will coke, and indices above' 1.1 indicate good coking tendencies. Although generally usable, this formula 'is not completely satisfactory because the percentage of oxygen shown in ultimate analyses is derived only by difference; i.e., by subtracting the sum of the percentages of the constituents determined analytically from 100 pct. Although the coking index indicates the coking tendencies of coal, it is necessary to make physical tests of coke before its industrial value can be determined. The U. S. Bureau of Mines has developed a standard procedure for determining the approximate strength of coke that would be formed from a given coal. In this test one part of ground coal, mixed with 15 parts of carborundum, is baked to form a standard briquette. The weight, in kilograms, necessary to crush the briquette is termed the agglutinating index. This test determines the relative fluidity attained in the coking process by measuring the cementing strength of the coal in the briquette. A

Jan 1, 1953

By Vard H. Johnson

IN 1943 the U. S. Bureau of Mines undertook drilling in an effort to develop new reserves of coking coal in an area near Paonia, Colo., as a part of an attempt to alleviate the shortage of known coking coal of good quality in the western United States. Geologic mapping of the area was undertaken by the U. S. Geological Survey with the purpose of first furnishing guidance in location of drillholes and later aiding in interpreting the results of the drilling. The drilling program was under the general supervision of A. L. Toenges of the U. S. Bureau of Mines. J. J. Dowd and R. G. Travis were in charge of the work in the field. Geologic mapping was started by D. A. Andrews of the Geological Survey in the summer of 1943 and was continued from the spring of 1944 to 1949 by the writer. The first few holes drilled failed to locate coking coal, but in the summer of 1944 coking coal was discovered by drilling 6 miles east of Somerset, Colo., the site of present mining. In the succeeding years, 1945 to 1948, 100 to 150 million tons of coal suitable for coking were blocked out by drilling. The ensuing discussion of the geologic controls on the distribution of coking coal in the area is based on the geologic mapping as well as the drilling done in the Paonia area, more complete descriptions of which have appeared or are in process of publication."' In order that the possible geologic controls affecting the present distribution of coking coal may be considered, it is necessary to discuss briefly the indicators of coking quality coals. Coking Coal Coal that cokes has the property of softening to form a pastelike mass at high temperatures under reducing conditions in the coke oven. This softening is accompanied by the release of the volatile constituents as bubbles of gas. After release of the contained gases and upon cooling, a hard gray coherent but spongelike mass remains that is referred to as coke. This substance varies greatly in physical properties and, to be suitable for industrial use, must be sufficiently dense and strong to withstand the crushing pressure of heavy furnace loads. Western coals have a generally high volatile content and therefore form a satisfactory coke only when they attain a rather high fluidity during the process of heating arid distillation in the coke oven. When this high degree of fluidity is developed, the volatile constituents escape and leave a finely porous coke. On the other hand, when the degree of fluidity is low the product is an excessively porous and therefore physically weak mass that is called char." Small quantities of oxygen present in coal are believed to decrease the fluidity of the material during the coking process and to favor the development of char rather than coke. In consequence, coal chemists have for some time considered the possibility of developing an index to coking qualities by inspection of chemical analyses of coals.' A formula has now been developed that does permit a rough preliminary estimate of the cokability of coal on the basis of the analysis on an ash and moisture-free basis. Coals may be eliminated as possible coking fuels if the oxygen content is greater than 11 pct. Similarly the ratio of hydrogen to oxygen must be greater than 0.5 and the ratio of fixed carbon to volatile constituents must be greater than 1.3. If the coal, on the basis of these limiting factors, appears to have possible coking qualities, the following formula permits determination of the coking index: a+b+c+d Coking index = -------- 5 a equals 22/oxygen content on ash and moisture-free basis, b equals two times the hydrogen content divided by oxygen content on moisture and ash-free basis, c equals fixed carbon/l.3 x volatile matter, and d equals the heating value on moist, ash-free basis/13,600. Coking indices higher than 1.0 suggest that the coal will coke, and indices above' 1.1 indicate good coking tendencies. Although generally usable, this formula 'is not completely satisfactory because the percentage of oxygen shown in ultimate analyses is derived only by difference; i.e., by subtracting the sum of the percentages of the constituents determined analytically from 100 pct. Although the coking index indicates the coking tendencies of coal, it is necessary to make physical tests of coke before its industrial value can be determined. The U. S. Bureau of Mines has developed a standard procedure for determining the approximate strength of coke that would be formed from a given coal. In this test one part of ground coal, mixed with 15 parts of carborundum, is baked to form a standard briquette. The weight, in kilograms, necessary to crush the briquette is termed the agglutinating index. This test determines the relative fluidity attained in the coking process by measuring the cementing strength of the coal in the briquette. A

Jan 1, 1953

By H. F. McCullough

The success of a coal-mining venture as relates to operations at the gob or break-line, such as the drawing of pillars or the working of long-faces, depends upon the control of face conditions. The measures required to control the latter are dependent upon the action of the coal seam, roof and floor at and in the vicinity of the faces, but the manner in which these will act is dependent upon the rate of face movement as well as upon the facc characteristics and upon the forces exerted. Yet the merits of a method of working—the practicability of controlling face conditions and the means required for such control when applied to any particular case—are usually judged only upon a consideration of the characteristics of coal seam, roof and floor, and the manner in which they have been observed to act. Practically every coal-mining man knows that face conditions are influenced by the rate of face movement, but few have appreciated that this is as much of a factor in the control of face conditions as are the face characteristics and the forces exerted. Fewer yet have recognized or taken advantage of faster rates of face movement as one of the principal means of establishing control of face conditions or of securing more favorable face conditions. Seemingly inherent difficulties, particularly in the control of roof conditions, but also in the control of conditions of the coal seam and floor, are frequently mitigated or disappear when the rate of movement of the faces is sufficiently increased. The characteristics of the coal seam, roof and floor and the forces exerted upon them do determine the general trend or tendency of face action, but the manner in which the coal seam, roof and floor will act and be aeted upon and the magnitude of the action during the time the site of mining operations remains at any particular point are very largely determined by the rate of movement of the faces. Progressive Nature of Forces Exerted Coal seam, roof and floor are composed of materials having definite properties or characteristics. The potential or ultimate values of the

Jan 1, 1927

By Carroll F. Hardy

The day by day loss of industrial plants to gas and oil is chiefly by default. The coal industry is not selling its superior economy, safety, and other advantages to its customers. THE position of the coal industry has been affected by a wide variety of developments in the production and use of energy. The tempo of development and change has been increasing and the end is not in sight. Legislation is currently being proposed for commercial use of atomic power, and the employment of atomic energy in significant quantity will probably occur about the same time as the decline in production of petroleum and natural gas. But these developments are in the future and have little immediate effect on utilization and marketing of coal. While no one should try to suppress or retard the development of a new and economical source of energy, both the coal and private utility industry should be allowed to question how the nuclear power is to be used, who is to use it, and who is going to pay for it. The taxpayers have a monopoly on fissionable material and the knowledge to employ it. Any commercial use must stem from this source. It is not hard to visualize either taxpayer-subsidized private utility atomic power plants on one hand and super TVA's on the other. In view of the gains of gas and oil in the home heating field, it is interesting to compare the 1940 and 1950 census reports on the kind of fuel used for heating in occupied dwelling units. Table I shows that whereas coal provided 77 pct of the fuel for central heating (furnaces and boilers) in 1940, it was down to 45.4 pct in 1950. However, only about 1 1½ million units were lost in this 10-year period. In the non-central heating category, which principally includes stoves, the percentage declined from 39.2 to 25.6, but the units declined about 2½ million in number. The big increase was in heating units designed to burn gas and oil. Use of wood for central heating declined about one-third. Data on amount of fuels used for residential heating are not available, but information is on hand for residential and comnlercial space heating, see Table 11. Commercial space heating includes office buildings, churches, schools. and similar structures. The annual use of bituminous coal in these two categories declined about 1 million tons in the 10-year period. Other forms of solid fuel showed greater losses, except wood, which remained the same. Domestic stokers reached their high point in 1948 with about 1,200,000 in use. At the end of 1951 there were approximately 1,116,790 stokers in use. Conversions to gas and oil have been from hand-fired heating plants in the ratio of about 7 to 1 compared to stokers. In other words, for every one stoker which has been converted to gas or oil, seven hand-fired units have been converted to gas or oil. A bare recital of these data would indicate that the coal industry is holding its own reasonably well. However, 93.4 pct of the new homes built in 1951 were heated by gas or oil. Oil-burning equipment was installed in 37.8 pct and gas equipment in 55.6 pct of the new homes. This indicates that the public prefers gas when it is available, and that oil is second choice, with all forms of solid fuel apparently used when it is unavoidable. It must be pointed out, however, that during the period of rapid expansion of gas pipelines gas has been sold for house heating at prices that are in some cases actually lower than coal prices, or very nearly on a par. Gas has been sold at wells at far below the comparable price for oil produced from the same wells, and far below its actual worth. This situation is being remedied at the present time by increases in gas prices at the wells. For example, the wellhead price of gas in Texas averaged 7.494 per Mcf in 1952. In 1949 it was 4.59c per Mcf. This increase in price is being reflected in pipeline gas prices, and in most of the markets served by the pipelines the tendency is to get it out of the bargain basement type of sales. The American Gas Association estimates that at the end of 1952 there were in the United States about 11 million customers for gas house-heating, and the Association expects additional gains each year until around 18 million homes will be heated by gas in 1975. By 1975 there should be 60 million dwelling units to be heated in the United States, if dwelling units increase at the same rate as the population. If the gas industry heats 18 million dwelling units by that time, this still leaves 42 million units to be heated by some other fuel. If oil is used to heat 18 million dwelling units in that same year, 24 million would of necessity be heated by coal, coke, wood, electricity, or another fuel. The total number of dwelling units using coal listed in the 1950 Census was 18,776,000, so it would appear that coal has a chance at least to stand still in the tonnage sold for domestic use. In the first quarter of 1953, 2044 domestic stokers

Jan 1, 1955

By G. E. Keller, G. D. Coe

Several methods have been devised for rapid determination of sulfur in coal using a high-temperature combustion furnace. The fundamental principles of the various methods are similar but the techniques vary widely. The method described herein is no different in principle from some of its predecessors but varies considerably in technique. Originally devised to overcome the low-accuracy criticism leveled against the others, the tube-furnace method has proved successful for limited applications. Work is being done to widen the applicability of the technique. Subsequent to the work of A. Vita1 in 1920, a number of investigators have published data dealing with the rapid determination of sulfur in coal using a high-temperature combustion furnace.2-5 These, together with several unpublished methods, involve wide differences in technique although they are all similar in principle. Investigation of typical details of procedure showed that poor precision or accuracy could be attributed to one or more of the following: 1) inadequate quantity of sample to be representative, 2) a delivery tube temperature too low to confine condensation to a washable area of the tube, 3) no provision for back washing the delivery tube, 4) incomplete absorption of the sulfur compounds due to inadequate bubbler, and 5) the false assumption that all of the sulfur is evolved as sulfur dixoide. The method herein discussed is no different in principle from some of its predecessors; the differences are in technique. Although the method is rapid—a determination can be completed in 5 min— accuracy has riot been sacrificed in order to achieve speed. MATERIALS AND EXPERIMENTAL WORK Fig. 1 shows the arrangement of the equipment. A metered volume of oxygen is passed through calcium chloride for removal of moisture, then through Ascarite for removal of carbon dioxide, and thence through a glass tee to the feed end of the combustion tube. Connected to the side arm of the glass tee is a rubber tube that is closed with a pinch clamp during combustion of a sample. The products of combustion are expelled through a Vycor bubbler, the end of which is submerged 3 cm below the surface of a 1 pct hydrogen peroxide solution contained in a 300 ml high-form beaker. At the completion of the 4-min combustion period the contents of the beaker are titrated to a yellow end point, using a methyl red indicator, with 0.05 N sodium hydroxide (or sodium tetraborate). The pinch clamp on the tube leading from the tee in the oxygen line is then opened and mouth suction applied to draw the hydrogen peroxide up into the bubbler past the ell, almost to the rubber tube connecting the bubbler to the combustion tube. This washing of the bubbler is repeated until no further acid reaction with the indicator is obtained. The hydrogen peroxide-sul-furic acid solution is then titrated to the final yellow end point. Test work conducted at the Robena Laboratory disclosed the following critical conditions that had important bearing on the precision of the method: Weight of Sample: Several series of tests wherein the sample weight ranged from 0.1 to 0.5 g of -60 mesh coal showed that the lowest variance (standard deviation squared) resulted from the use of 0.5-g samples; the variance was slightly greater when 0.3-g samples were used, and with less than a 0.3-g sample weight the variance increased rapidly with decrease in weight. Although the use of a 0.5-g sample invariably resulted in a very slight explosion as soon as the sample entered the hot zone of the combustion tube, this did not detract from the precision of the method. Temperatures: Mott and Wilkinson5 found that a combination of low temperature (1250°C) and low oxygen flow rate (less than 700 ml per min) tended to give high results due to formation of oxides of nitrogen. Although the method herein employs a much higher oxygen rate than 700 ml per min, as an additional precaution to assure accurate results, 1315°C (2400° F) is recommended as the most desirable furnace temperature. The temperature of the elbow of the bubbler should be at least 75°C to minimize condensation

Jan 1, 1961

By C. W. Thompson

The two-way belt conveyor for coal mine service simultaneously carries coal from faces and transports men and supplies into the mine, largely eliminating the necessity for rubber-tired supply and man haulage equipment and the necessity of maintaining additional headings. SINCE 1920 use of belt conveyor transportation has increased steadily, and capacity in tons per hour per unit has increased at about the same ratio. Belt conveyor systems for simultaneous two-way material transportation have been widely discussed, but heretofore have proved impractical for specific operations. Studies of a proposed means of disposing of wash-ery waste on the return trip of a main haulage conveyor, as yet uninstalled, led to the application of two-way belt haulage to Weirton's underground supply and mantrip problems. Except for a single-tracked main haulage entry, Weirton mine is essentially a full-belt haulage operation with attendant problems of supply and man-trip haulage that could not be handled with conventional belt conveyors during the regular work shift. It was necessary, therefore, to maintain rubber-tired haulage equipment and haulageways serving multiple-shift continuous operations. The conventional underground belt conveyor is primarily intended to transport coal from working faces to surface. Belts are usually reversed to haul men to face areas at the start of the working shift or to transport supplies. The two-way belt conveyor for coal mine service is not yet offered for sale as a unit by any manufacturer and must be built on order. This conveyor, with return run traveling parallel but opposite in direction to the outgoing or coal haulage run, permits not only haulage of coal from faces, but simultaneously permits transportation of men and supplies into the mine at any time, see Figs. 1 and 2. This eliminates, to a great extent, the necessity for rubber-tired supply and man haulage equipment and the maintenance of additional headings. Unlike conventional installations the return belt is in the open for inspection and maintenance. It is less subject to edge wear, since it runs on regular troughing idlers, see Fig. 3. The only points where the edge of the belt can rub are at the drive unit and the tail units where improved design should eventually eliminate edge wear entirely. Individual conveyor sections are lighter, since all that is needed is a tubular stringer on each side to carry troughing idlers only, no cover plates being necessary. The return run is carried on an assembly identical to that carrying the coal haulage run. These assemblies are definitely easier to handle and are cheaper than installations for the same length of conventional belt conveyor. The four pipes required for the complete conveyor frame may serve only as supports, or, with proper couplings, may be used for mine water discharge line, pressure line for face spraying, telephone cable conduit, or any other purpose requiring a pipe line, see Fig. 4. Additional expense, if any, for the increased number of pulleys needed to turn the belt over for the return service is more than offset by reduced expense for rubber-tired supply and mantrip haulage. The two-way belt haulage system is confined to a single heading on intake air, isolated from all other headings by stoppings on each side, and open for air circulation only at the main haulage way and at the faces. One additional heading, also on intake air, is maintained as an emergency rubber-tired haulage way, all other headings being center-posted and used as airways only until retreat work begins. This method, of course, can and should be used with any belt conveyor, since it provides two independent intake airways for escape in case of fire on the belt or elsewhere. Belt heading is heavily timbered even under good top conditions and space between the two runs permits center posting for the full length

Jan 1, 1954

By C. W. Thompson

Paul D. Suloff (Goodyear Tire and Rubber Co., Inc., Akron, Ohio)—I would like first to comment on problems of the conveyor belt discussed in Mr. Thompson's excellent paper, since that is what we hope we know most about. Twists in relatively wide conveyor belt unavoidably produce a lateral maldistribution of tension, raising tension at belt edges and reducing it at the center. They also produce a lateral collapsing force on the belt at the center of the twist owing to the inherent tendency of all the longitudinal elements of the belt to try to pass through a point at the twist center. Calculation of the twist geometry by the methods shown in Mr. Thompson's paper keeps these extraordinary forces within limits which the belt designer can tolerate. No reduction in belt life due to twisting need be contemplated when this geometry is maintained. There is a minor exception that belts of extreme lateral flexibility will tend to curl laterally at the center of the twist. However, any ordinary fabric construction will perform satisfactorily in this respect. These twists are always made in regions of low tension in the conveyor so that even in the edges of the twist, belt tension does not exceed the average tension found in highly stressed regions of the conveyor. Offsetting these out-of-ordinary belt stresses is the advantage that Mr. Thompson has brought out of getting the return run up out of the dirt where it can be seen. This not only makes it easier to train, but also, in the event that it is not properly trained, frees it of the normal return run edge wear hazard. It is well known that return run edge wear is a prominent cause of belt mortality underground. An interesting aspect of this two-way conveyor is that the belt may be made what is known as a Mobius Strip. A Mobius Strip is obtained by splicing a belt after turning one end of it 180" about its longitudinal axis. In other words, one end is turned upside down before splicing. A belt spliced in this fashion turns itself upside down every time it comes around, but the twist which has been put in the splicing, of course, stays at one location on the conveyor, in this case one of the twist sections at the end. Turning the belt over every revolution might have advantages in some cases. Belts could be made with equal covers and the two sides worn equally and simultaneously. In this case there would be no problem of getting belts on upside down by mistake. However, the two-way conveyor does not have to be a Mobius Strip. It can be twisted in such fashion that the same side is up on both runs. It is simply a question of which way the final 90" twist is made before joining the ends. Another interesting aspect of the two-way conveyor is the problem of operating two-way conveyors in series. Here the sequencing of starting brings up some new problems. It will be recognized, although not always at first glance, that if the starting sequence is planned for one run of the conveyor the reverse will result on the other run. With the two runs carrying bulk material in both directions a reverse sequence on one run would be intolerable. In this situation the only solution appears to be a simultaneous starting of all conveyors in the series. However, with the coal in one direction and intermittent supplies in the other it would be entirely practical to sequence the conveyors for the coal run and accept a reverse sequence on the supply run. The two-way conveyor also lends itself to new driving possibilities. First, it is quite possible to drive at the head end of each run, which of course, means a drive at each end of the two-way conveyor. Driving in this way a given belt can be extended to substantially greater lengths than a conventional conveyor with drive at one end only. In addition to this, under certain conditions the conveyor can be extended to extreme length by driving at one end and at some intermediate point on the most heavily loaded run. As a particular case, a belt carrying coal downgrade and supplies back upgrade could be extended to extreme lengths by driving at the head of the coal run and at an intermediate point of the supply run. Mr. Thompson has been a pioneer in belt conveyor transportation underground and his accomplishment here with the first two-way conveyor of any consequence is another notable addition to the art.

Jan 1, 1954

By E. F. Thomas

THE late George S. Rice was active in the inves--I- tigation of bumps, particularly in the last ten years of his career as chief mining engineer of the U. S. Bureau of Mines. Since most of his investigation was carried out in Great Britain, continental Europe, and—to a lesser extent—Canada, his thinking on prevention was influenced considerably by the experience of those countries. It is not surprising, therefore, that when he was called upon a few years before his retirement to investigate bumps in the U. S. and suggest ways to prevent them, he turned to longwall mining. A longwall method had been most successful in combating the bump hazard in mining coal under deep cover, especially in Great Britain, but the prevailing method there at the time was advancing longwall mining, which he knew was uneconomical under U. S. mining conditions. For this reason he proposed a modified retreating longwall system that he believed included the best features of the advancing method. As brought out by Rice,' if the cover is 2000 ft and 50 pct of the coal is extracted, the static load on the remaining pillars will be about 4000 psi, which exceeds the ultimate crushing strength in most instances. If the pillar coal is overloaded before a pillar line is established, then the abutment zone preceding a line of extraction is no place to split pillars or extract them by any method other than an open-end system. Rice therefore advocated open-end mining, preferably by longwall, but he was willing to compromise with long-face mining if the longwall method was not acceptable. Rice's system was put into operation in a mine in Harlan County, Kentucky,3 but subsequent experience has shown that it did not take into account two important factors—avoidance of pillar-line points and maintenance of adequate development in advance of the pillar-line abutment area. For ten years after Rice's retirement the USBM did little investigation and research on bumps, chiefly because so few were occurring that there was not much cause for alarm. But in 1951 there were three occurrences involving fatal injuries, and the Bureau began a statistical survey in that year. C. T. Holland, head of the department of mines at Virginia Polytechnic Institute, was retained as a consultant. The resulting study' of 117 case histories brought out these important conclusions: 1) Almost invariably the bump occurred in a locality affected by the abutment zones of one or more pillar lines. 2) In most cases the locality of the bump was influenced by the abutment zones of more than one pillar line. The term pillar-line point has been used for many years in the Appalachian region for such a situation. Point is used in the geographical rather than the mathematical sense. 3) In pillar-line extraction the following practices are safest in preventing bumps: a. The mine layout should provide for pillars of uniform size and shape along the extraction line. b. The mine layout should be planned so that no development need be done in the abutment zone of a pillar line. c. The layout should permit open-end extraction of pillar lines from the next goaf, so that it will not be necessary to resort to pocket mining, splitting pillars, or any practice that will involve driving in the direction of the goaf within the abutment zone. d. Pillars should be large enough to support area without undue roof and floor convergence before establishment of a pillar line. These are, of course, generalities, and while they are useful in laying out areas where bumps can be expected, they are of limited help in many mines that were committed to a system of mining before it was realized that they were subject to bumps. Under such conditions it becomes necessary to choose between the following alternatives: 1) Abandon the territory, except for pillars that offer no extraction problems. 2) Through experience select the pillars that are most heavily loaded, and, by augering, induce bumps from a safe vantage point so that impinged loads are relieved. This method was first developed at the Gary, W. Va. mines of U. S. Steel Corp. and later adapted to mining thick coal beds at Kaiser Steel's Sunnyside mine in Utah. No scientific method is available to determine where to drill within a loaded pillar. Although this method of unloading has worked very successfully at Gary—with one exception—

Jan 1, 1959

By W. A. White

UNDERCLAY squeeze is the plastic flowing of underclay below coal pillars into mined-out entries and rooms. Squeezes may be caused either by wet mine conditions where the moisture is taken up by the clay—making it more plastic—or by the presence of certain clay minerals, such as mont-morillonite, which are plastic when there is enough natural moisture. Under either condition, sufficient stress will cause the underclay to flow. Squeezes may fill an entry or room within several hours, or it may be several months before movement stops. Squeeze areas commonly cover hundreds or thousands of square yards. The Coal Div. of the Illinois State Geological Survey has collected samples of underclays from various mines and from cores drilled by coal companies in the state. From these samples natural moisture content, particle size and particle size distribution, and clay mineralogy are determined. Data from samples collected from the underclay below Herrin (No. 6) coal in the Lumaghi Coal Co. mine near Collinsville, Ill., are used in this report to clarify points in the discussion. Data from samples taken from other mines lead to the same conclusions. A generalized geological section for the area is given in Table I. The underclay below coal No. 6 ranges from a few inches to 3 or 4 ft thick. The upper 6 to 18 in. are noncalcareous; the lower part becomes more calcareous with depth. Limestone nodules become larger with depth, almost grading into the limestone below. In areas where the clay is nearly absent the limestone may thicken to take its place; thus they may complement each other in thickness. There are two types of squeeze areas in a mine. In one type up to 3 or 4 ft of underclay lie in what appears to be a depression in the limestone underneath. This is suggested by a rim surrounding the squeeze area where the underclay is thin and the limestone lies within a few inches of the coal. In the second type the underclay of the squeeze area is surrounded by underclay of similar thicknesses which does not squeeze. Test Methods: Natural moisture content, particle size and particle size distribution, and clay mineralogy were determined by the following methods: Natural moisture is the moisture content a clay contains in its natural environment, expressed in percentage of the weight of ovendried clay (110°C or 230°F). To obtain the natural moisture content, given in Table 11, the clay samples were sealed in containers as soon as the samples were taken in the mines. Particle size and particle size distribution were determined by the pipette method,' and the data were plotted as cumulative curves, Fig. 1, on semi-logarithmic paper. Particle size was plotted on the logarithmic scale and percentage on the arithmetic scale. The quartiles (diameters that correspond to frequencies of 25 and 75 pct) and medians (diameters that correspond to the frequency of 50 pct) of the samples were determined from these curves. The clay minerals were identified from the data obtained by differential thermal' and X-ray powder" techniques. The differential thermal analyses were made on the whole samples, whereas the X-ray analyses were made on fractions of less than 2p. Data and Discussion Moisture Content: After the sealed jars had stood for 24 hr, those that contained squeeze samples had perspired, so that droplets of water had accumulated on the insides of the jars. Table II shows the moisture in the clay after it had perspired for 24 hr. (This perspiring was not noted in samples from nonsqueeze areas.) After perspiration one of the squeeze samples still had a higher moisture content than the samples that did not squeeze. Clay Mineralogy: The thermal analysis, Figs. 2 and 3, indicates that the clay minerals are chiefly micaceous in type; the low-temperature endothermal

Jan 1, 1957

By Cloyd M. Smith

A few years ago, the Ventilation Committee established the practice of resenting one topic each year for discussion at the annual meeting. The practice has met good response on the part of committee members and I suggest that it be continued. The topic chosen for this year, "Underground Anemometry," is a topic which has bothered me for more than 20 years. It seems to me that the coal industry is content to rely on slipshod methods for measuring the rate of flow of air underground, so I prefaced my discussional charge to committee members with the statement that I regard air measnrements made in the usual way, with hand held anemometer, as no good. Agreements and disagreements came in from more than a dozen engineers, some of whom are with operating companies, coal and metal; some with manufacturers; others with government agencies. The statement was accompanied by a questionnaire on the use of the rotating vane anemometer and by one describing two methods of using a mechanically held anemometer. The questionnaire will be considered first. The questionnaire and statement are as shown on pages 5 and 6, the committee members and respondents are given on page 4, and the general comments of the latter on page 5. Questionnaire 1. Has your company or agency issued written instructions for care and use of anemometers? If so, please enclose a copy with reply. Only one answer, McElroy's, was affirmative. It gave reference to Bureau of Mines publications1'5 which recom- mend the hand held anemometer for rough measurements and indicate that am accuracy of 5 pct can be had if calibration and method factors are used. Mathews said that instructions are principally oral while Maize reported that state inspectors of his department are well trained in use and care of anemometers. 2. Are your anemometers calibrated periodically? If so, by their manufacturer? or by? Are calibralion corrections applied to all observed mean velocity readings? Only one respondent, Lee, answered negatively as to calibration. This means that anemometers are generally calibrated but the questionnaire failed to ask how often this is done. As no one volunteered the information, we have no data on this point. In six cases the instruments are sent to their manufacturers for calibration. but Krickovic reports that his company limits manu-facturers' calibrations to anemometers which are used by operating personnel; those used by the engineering department being calibrated by U. S. Bureau of Standards. The Anaconda Copper Mining Co. has its ventilation engineers calibrate its anemometers. Most of the respondents say that a calibration correction is applied to each mean velocity reading, but Krickovic limits this to surveys made by the engineering department. Since Lee does not calibrate, he has no correction to apply. Maize reports that his department has its anemometers calibrated but does not apply corrections. 3. Do your men hold the anemometer by hand in measuring air flow? for 1 min? or traverse the section? for 1 min? or at? points for 5 sec each? Of 10 replies, 6 were "yes," 3 were "no," and one was "seldom" with respect to holding by hand. Among the six hand holders, four hold in a central position in the measuring section for a minute, except that two of them, Krickovic and Matthews, traverse the section by hand for survey or fan test. Their operating personnel hold by hand, centrally, for routine measurements. McElroy sometimes traverses with hand-held anemometer in rapid survey work. 4. If the anemometer is not held by hand, how is it supported? Augustadt supports the anemometer on an adjustable rod, Condon on "a rod of sufficient length to reach all points with observer standing in one position throughout traverse and at arm's length from plane of traverse." I presume that arm's length must be interpreted liberally enough to allow for arm movement, otherwise it would be impossible to manipulate the anemometer throughout the traverse section. Mancha upholds Condon in this method of traversing. Glanville hangs the anemometer on the end of a 4-ft staff by the hasp at the top of the anemometer frame. McElroy mounts it on the end of a rigid square shaft, 12 in. long, the staff being at right angles to the axis of the instrument. He traverses the section in two halves, holding the anemometer 3 feet from his body.

Jan 1, 1950

By Cloyd M. Smith

B. F. TiLLson*— The manifold difficulties of accurate anemometry in irregular sections of mine passageways, the irregular distributions of velocities in cross sections of the same, and the disturbing influence of the observer upon the air flow, all indicate an undesirable inaccuracy of results obtainable by any standardized method of traversing the section by an anemometer. It seems obvious that another, and simpler, method should be used to determine the volume of air flow in mine passages, namely: 1. At appropriate locations cement or calked framework rings should be installed permanently to equalize the irregularities of sectional contour and provide a place and means of attachment for a temporary cloth brattice which bears a rigid orifice. 2. The measurement of the velocity of air flow through the orifice may then be by anemometer or, preferably, by Pitot tube measurements of the differential pressures on both sides of the orifice in accordance with the standard practices available in engineering 1iterature. † The constants may be determined for various measuring positions in relation to the resulting "vena contracta." 3. The position of the person who makes the measurements is behind the brattice out of the air stream. The Pitot tube does not offer as disturbing an obstruction as the anemometer. A recording gauge may be employed to integrate fluctuations in air flow through that portion of the mine. No traverses are required because the reading may be at a single central point. An anemometer can be used with an orifice flow. The orifice will increase the air velocity at the measuring point, with correspondingly more accurate measurements where the normal air velocity through the passageway is low. Portable brattices might be devised with the cushioning rims which would seal against irregular rock surfaces where permanent rings were not available or feasible. The development by the Ventilation Committee of standard procedures and devices for the orifice measurement of the flow of mine ventilating air might be a desirable project for this coming year. C. M. Smith (author's reply)—Thank you for your discussion of my paper on underground anemometry. Your suggested method of measuring underground air flow is a novel one which might be applicable in some situations. It should be tested along with other suggested methods in any investigation of this subject. G. E. McElroy*—In spite of the adverse publicity that vane-anemometer methods of air measurement have had in the past and that contributed by the present paper, I endorse Mr. Erickovic's statement that anemometer traversing "has proved to be widely applicable, expeditious and simple" and add that available methods are accurate enough for the purposes for which they may be used. The fact that the great majority of minor mine officials assess relative changes in rates of air flow by comparison of crude vane-anemometer measurements, known to average 20 to 30 pct high, has no important bearing on this subject, because state inspection standards were based originally on such methods of air measurement. Federal inspection standards are based on actual rates of flow as determined by traversing, and interest in traversing methods is rapidly increasing. In considering traversing methods, three aspects are of major importance: (1) the absolute accuracy of calibrations; (2) the degree of interference with normal flow conditions introduced by traversing methods designed for accurate measurement by shaft-mounted instruments; and (3) the proper "method" factor to use for approximate measurements by hand-held instruments. With respect to absolute accuracy of calibrations, we have always placed reliance on calibrations made by the National Bureau of Standards, with which manufacturers' calibrations have usually agreed very closely. It is therefore particularly disturbing to find7 that calibrations made previous to June 1947 are presumably about 5 pct in error because of excessive registry caused by the thin flat plates on which anemometers were mounted for calibration. Velocities corrected for calibration have therefore averaged about 5 pct low in all probability. In this connection, it is interesting to note that an anemometer calibrated against Pitot-tube measurement by a single-point method in the Bureau of Mines experimental coal nine in 1923 indicated this same difference of about 5 pct and that the same instrument calibrated by a traversing method in a metal mine some months later indicated a difference in the same direction of about 4 pct. These results are reported by the Bureau of Mines.8 Regarding the degree of interference, or changes in velocity distribution, caused by the position of the observer's body in traversing operations, misconceptions seem to be especially prevalent, resulting in increasing advocacy of methods, such as the "clear section" method outlined in this paper, that cause just the type of interferences that they are designed to avoid. The degree of interference for any method may be gauged easily by a few experiments with a velocity-pressure gauge connected to a Pitot tube or with an indicating velocity meter such as the Velometer. In an experiment cited by McElroy and Richardson,# a decrease of 5 pct was noted at ten widths upstream from a 6-in. plank, whereas an observer's body at about the maximum practical distance of 6 ft downstream from the instrument is only about four widths away. In the Bureau of Standards paper previously mentioned, it is recommended that supports used in calibrations be at least 16 widths downstream. In practice, therefore, a downstream position of the observer is ruled out as far as accurate measurement is concerned. Operation of the anemometer by rigid shaft support from a point outside the section is seldom practicable; however, accurate results can be obtained, with the anemometer rigidly attached to a short shaft and held at arm's length, by an observer advancing across the traversed section while he faces the opposite wall and stands sideways to the current, provided that he keeps the instrument at least 3 ft away from his body at all times and traverses the entire section with it. If the traverse can be started with the observer in a side recess, the entire section can be covered in one operation. Normally, it would be covered in two half-sections. The presence of the observer's body does not, as is commonly supposed, increase the average velocity throughout the remaining part of the section. Rather, the velocities 1 to 2 ft on either side of his body are increased, but the distribution of velocities throughout the rest of the cross section remains normal, and a traverse made as stated gives a true average velocity for normal-flow conditions. Regarding the proper "method" factor for accounting for interference in the approximate methods of traversing with hand-held instruments, here again confusion prevails, for which the writer must assume some of the blame. Comparison of consecutive traverses made by shaft-held and hand-held 4-in. anemometers in field work after the tests reported by McElroy and Richardson' gave method factors

Jan 1, 1950

By J. D. Forrester, Erich Sarapuu, T. C. Cheasley

Electrocarbonization of coal will produce cheap fuel for electric power plants. The complete electrification of industry and domestic power consumption of gas generated as fuel in situ is a worthwhile goal for the coal industry. Production of synthetic oils, gasoline, and coal chemicals from coal gas will create additional new markets for the coal industry. RANKING high in the nation's wealth of natural resources, readily available liquid and gaseous hydrocarbons are nevertheless falling steadily in supply, while industrial requirements for hydrocarbon compounds are greater today than during the peak period of World War 11. Theoretically, several processes will increase the future supply of liquid and gaseous fuels, but the practical application of these methods depends on two controlling factors: 1—that enough carbon-bearing material be available to warrant industrial exploitation, and 2—that the given process be physically, chemically, and economically applicable. With regard to the first factor, the Western Hemisphere is fortunately well supplied, see Table I. Table I not only shows the extensive occurrence of many fuel sources, but also the wide discrepancy among quantities of fuel substances in various materials. Great quantities of potentially available fuel have not been recovered. Many coal seams and other hydrocarbon-bearing deposits are so situated that production is uneconomical by present-day mining or oilwell pumping methods, and even in many exploitable deposits where the most efficient production methods are used a large proportion of material is never recovered. Industrial production of hydrocarbon liquids and gases must be better balanced, and processes must be developed which will convert coals, oil sand, and oil shales into fabricated commodities. Among the techniques employed by scientists and engineers to secure fuel masses not economically available at present are several processes of underground gasification. These are all concerned with converting raw materials into gaseous products by thermal treatment of source substances while they are still in the ground and therefore require that a subsurface heat (fire) zone be induced to affect pervasively the buried raw fuel deposit. This is possible only where sufficiently high temperature exists and where the fuel material is satisfactorily opened or channelized. Also, to compete successfully with common mining or oilfield practices, a given method must be controllable and must not involve greater operational expense than regular methods. Preferably, it should be less expensive. Since the close of World War 11, England has continued field studies of underground gasification methods. Experiments have also been conducted in Sweden, Estonia, France, Belgium, Italy, and Germany. The U. S. Government has sponsored field studies at Gorgas, Ala., and because gasification in situ is vitally important to establishment of the nation's fuel resources, the mining engineering department at the Missouri School of Mines has been actively interested in developing the process of underground electrocarbonization. Research was conceived by Erich Sarapuu, and preliminary experiments in treating bituminous coal were begun at the school in September 1947. The work consisted of laboratory experimentation until June 1948, at which time, through a cooperative agreement with the Sinclair Coal Co. of Missouri, pilot plant studies were instituted at the Tiger mine of the Hume-Sinclair Coal Co. near Hume, Mo., to test the encouraging laboratory findings. This comprehensive field program was pursued continuously until the beginning of 1952."'" The broad term underground electrocarbonization describes a unique gasification process that releases by electricity the desired economic constituents from a raw carbon-bearing material, converting the substance of such deposits into deliverable form and greatly reducing the cost-of-mining factor. Underground electrocarbonization is not subject to the major limitations of many other underground gasifi-

Jan 1, 1955