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By George S. Rice

The A.I.M.E. Ground Movement and Subsidence Committee, proposed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bituminous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Pitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O'Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1940

By John W. Buch

NeaRly 25 years ago operating officials in the northern anthracite field were confronted with the problem of profitably mining virgin beds of thin coal (those 48 in. and under) or destroying them by mining underlying thick beds. To solve this problem costly hand-mining methods were replaced by cheaper mechanical methods, which eliminated the need for the lifting of bottom or the brushing of top in excessive quantities, which was essential if proper height for hand loading1 was to be provided. During this period long-face mining was an outstanding development. In its favor in comparison with any other method were the ease with which undercutting machines could be used, a differential in powder cost of about fifteen cents per ton in favor of this method over mining from a solid chamber or pillar face, and a greater production per day from a single working place, resulting in lower capital expenditures for a given total production, closer supervision and lower transportation cost. The long faces were drilled by jackhammer and undercut by machines. The coal from the cut was loaded into mine cars placed on the gangway road by means of a double-drum mechanical scraper engine and a mechanical V-type scoop. A tugger hoist was used to spot empty cars at the loading point from a supply placed at the operation by an electric mine locomotive. During loading the roof was supported by vertical wooden props paralleling the face. After the loading was completed the scraper engine and the ropes that had been used to propel the scoop in the loading of coal transferred the props to and along the face. A row of unfilled cogs was then built parallel to the new face and about 6 ft. from it. These cogs were usually made 6 by 3 ft., the longer dimension paralleling the face; they were placed on an average at 12-ft. centers. These wooden cogs were left in place and were buried later by falling roof. Attempts to remove them were never successful. When a new long face was started, the cog spacing would be increased to about 20-ft. centers. As successive cuts were made, weight would begin on the cogs, and they

Jan 1, 1940

By W. L. Dennen, W. W. Wirth

Research looking toward the reduction of the cost of roof support by substitution of longer-life materials for wooden timber is fully justified by the fact that roof support is an important element of all production costs. Permanent or semipermanent roof support will reduce the expense of timber-maintenance crews during steady operations, but even more markedly during idle periods. The management, therefore, is faced with the important problem of timber renewals, which in times of regular operation are frequently obscured by more urgent problems of production. Increased use of timber in the Anthracite Region of Pennsylvania is caused mainly by increased percentage of coal obtained from robbing areas, which requires more timber for safety than development work, and maintenance of long haulageways through crushed ground despite the fact that less coal is transported over them. The timber reserve in the immediate vicinity of the mines has been depleted to such an extent that it is necessary to secure 70 per cent of the timber requirements from distant sources. Originally, mine timbering in this region was done by largely utilizing round props or legs with round collars. The whole cross section of the tree from which timber was cut was utilized by forming the joints between the legs and the collars with saw and adze, using plank or round saplings for lagging behind the legs and over the collars to prevent material falling between them. The Mine Inspectors' report of 1898 records an important digression from the use of round timber, when two tunnels at the Babylon colliery, near Pittston, Pa., were timbered with 12 by 12-in. square timber sets with mud sills, through quicksand for distances of 425 and 160 ft., respectively. The timbers were lagged with 3-in. plank to control sand. An early step toward the permanent type of roof support was the installation of "brick pillars or perfect arched brick rooms" instead of

Jan 1, 1940

By Reamy Joyce

Conditions in underground mining are so variable that in approaching the problem of the economies effected by the use of pressure-treated mine ties and mine timbers, it is necessary to secure specific data for each mine. Facts must first be dug out to show the service life of untreated timber and the reason for its failure. Generally speaking, decay is the usual cause of failure, but in some mines mechanical destruction is an important factor. Occasionally the two causes combine to produce fail-urrs, and it is difficult to determine positively which factor is responsible. In making comparisons, it is always necessary to start with the cost in place and for this reason there must be accurate information covering the labor cost of installation and the labor cost of renewals. When treated timber is used, it is necessary to know the specifications for the treating process, whether the timber was painted, dipped, treated open tank, or impregnated under pressure according to standard specifications. It is also necessary to know the preservative used and the amount of the preservative that was retained by the timber after treatment. It is also well to know whether the timber was thoroughly air-seasoned before treatment or was green at the time it was treated. The kind of timber available, both untreated and treated, must be known and all of the factors listed above must be taken into consideration in the study of any problem. Variable Conditions A study was made recently of the untreated mine ties used by a bituminous coal company in two of its mines, which are about 20 miles apart. The mine ties were secured from the same sources for both mines. A careful check was made over a 10-yr. period on the life of untreated main-haulage ties, and it was found that in the mine where the haulage roads were on intake air the average life was approximately 5 yr., while in the other mine, where the haulage roads were on return air, the average life was 2 yr. A similar variability in untreated timber

Jan 1, 1940

By L. I. Cothern

The introduction of mechanized mining has created a demand for long working faces. It has also prompted mining men to contest the old theory that longwall methods can be used only where roof conditions are ideal and where pack-wall material is plentiful. "Longwall" as used in this paper means "modified longwall," "semi-longwall," or any of the various ''wall" or "long-face" methods of mining. Many mines are now using long conveyor faces that differ only in their methods of development, types of roof support and method of loading coal onto the conveyor. Some types of longwall operations are being carried on successfully under conditions heretofore considered impossible by men experienced in longwall mining. With possibly one or two exceptions, long-face methods in Alabama are a direct result of the need for lower mining costs. The advantages over room-and-pillar systems were clearly expressed in an article by Messrs. Fies and Lacy of the DeBardeleben Coal Corporation1 as follows: (1) increased tonnage per miner; (2) increased percentage of lump coal; (3) reduction in cost of development; (4) elimination of the cost of room yardage; (5) reduction in the cost of haulage, from ability to use larger cars; (6) reduction in the cost of machine cutting; (7) reduction in cost for tracks; (8) reduction in the cost of deadwork; (9) reduction in the cost of supervision. An attempt on the part of the operator to gain these advantages by changing from room-and-pillar to longwall requires considerable courage, and credit should be given to those men who have done their bit to meet the competition of rival fuels. The initiatory problems to be solved are in reference to: (1) retreating or advancing system, (2) length of wall, (3) method of roof support, (4) method of entry development, (5) type of conveyor or loader, (6) hand vs. machine loading (onto conveyors), (7) hand vs. machine loading in entries, (8) control of roof pressure,

Jan 1, 1940

By R. V. Taborelli, P. B. Bucky

This investigation was carried on by means of models and the application of the principles of similitude to determine the effects of props, props and cribs and sand filling in longwall mining. The geologic structure and rate of face retreat were kept constant and the only change introduced was in the method of support. It is not felt necessary to repeat a description of apparatus, or to explain the fact that a scalar model made in all parts of the same material as the prototype will behave similarly to the prototype when rotated at speeds such that the centrifugal force is as many times the gravitational force as the prototype dimension is a similar linear model dimension. The reader is referred to earlier papers1,2 for further information regarding these statements. All models were made of sandstone and were as in Fig. 6. They are composed of a 3/16-in. ore body bed A, overlain by a 3/16-in. bed B, a ?-in. bed C and two ?-in. beds D and E. The model was rotated at + 2350 r.p.m. equivalent to a model ratio of 1268. It was worked back ¼ in. at a time before being replaced in the centrifuge and again rotated. Table 1 gives model and prototype dimensions occurring in this discussion, the prototype dimensions being 1268 times the model dimensions. Figs. 1, 2, 3 and 4 were taken while the model was rotating at 2350 r.p.m. One was taken for each face retreat of 26.4 ft., but for obvious reasons all are not presented. Fig. 1 (A-E) shows the behavior of this type of structure with face retreats of 26.4 ft. and no artificial support. Fig. 2(A-E) shows the behavior of this type of structure where breaker and face props were used. The props were exceptionally strong, being of steel ? in. square in cross section, and capped with 1/16 in. ± of soft wood and paper, to allow for roof deflection. The face props were

Jan 1, 1940

By David Sinclair, Philip B. Bucky

The dimensions and shapes of mine structures may at present be determined by (1) field experience, (2) structural calculations, and (3) barodynamic tests. § None of these, however, provide information as to the stresses, their direction and distribution in a mine structure. For the design of pillars, the placing of openings relative to working places such as gangways, haulage levels, etc., this information is desirable. Photoelastic methods provide this and, while relatively new, their application in all engineering fields is assuming increasing importance. This paper, therefore, will present the method with a few applications to mining, and with sufficient reference for further study by the reader. PART I The Photoelastic Method The photoelastic method of stress measurement determines the stress distribution in a two-dimensional model by optical means. It makes use of the fact that when a piece of isotropic transparent material, such as glass, celluloid, or Bakelite, is stressed and viewed in polarized** light a picture is seen which indicates by its color bands of light, or fringes, the magnitude of the stresses, their direction and their distribution. General Explanation The stress at any point 0 in a structure (Fig. I), may be resolved into two components, P and Q, each a pure tension or compression, at right angles to each other, called principal stresses.' To fully identify the stress, the values of P and Q and their directions or angle F must be known.

Jan 1, 1940

By C.P. Proctor, J. T. Crawford, J.A. Younkins

Coal-cleaning plants using the launder process generally wash the fine coal (minus ? or minus 5/16-in.) separately in a plant consisting of washing launders or troughs placed one below another and sometimes supplemented by additional washing equipment such as tables and jigs. The coal and its associated impurities enter the feed end of the top launder, with a current of push water. As the material moves along the impurities, being heavier than coal, settle toward the bottom of the launder and are drawn off, with some coal, through "free discharge" boxes to the launder below. These boxes were given their name because their discharge falls freely; that is, it does not discharge into a water seal. Thus the impurities are progressively concentrated in the lower launders until box products are sufficiently free of coal to be discharged to refuse. The overflows of the upper launders go to clean coal while the overflow of the lowest launder, and perhaps that of the one immediately above, is usually recirculated to the plant feed. Thus the bulk of the clean coal is removed rapidly and the impurities are concentrated more slowly. The chief purpose of this paper is to present the cleaning data of such a plant under different operating conditions. OpeRating Equipment The washing units considered here are part of a complete plant cleaning minus 4-in. coal. The primary cleaned product is a high-quality gas and metallurgical coal and another portion of the output may be shipped as a steam coal. The units consist of two parallel Rheolaveur free-discharge units each containing five launders, together with three 7 by 14-ft. Plato tables. The launders are designated by the letters A-B-C-D-E (Fig. I), launder A being the top and longest; E being the bottom and shortest. Launder details are given in Table 1. The launders are equipped with a series of free-discharge Rheolaveur boxes, which are integral parts of the launder (Figs. 2 and 3). Each box with-

Jan 1, 1940

By Kenneth B. Ray, George R. Gardner

The application of wet cleaning processes for the beneficiation of bituminous coal has created in some localities a problem in the recovery and disposal of fine solids in the washery water. The maximum allowable concentration of these solids in the washery circuit that is consistent with satisfactory operation of the cleaning units remains a controversial point. It has, however, been generally recognized for a number of years that some degree of control should be exerted over the concentration of such material. Early methods of control involved intermittent or continuous discharge of a portion of the plant slurry into near-by streams or settling ponds, and this persisted even after the adoption of settling tanks or cones. The practice is not always economical, for two reasons: first, because of the waste of coal, and, second, because it is sometimes difficult and expensive to obtain an adequate supply of water. The latter condition often is aggravated by chemical treatment of the make-up water that is necessary to prevent excessive corrosion of plant equipment. The modern approach to this problem has been the adoption of continuous thickening devices. However, because such equipment occupies considerable plant space and is somewhat expensive to construct, the practice of loading these thickeners beyond their rated capacity is widespread, therefore often the performance of such units is not altogether satisfactory. In recent years, European washeries have increased the rate of settling of the fine coal from the slurry by the addition of specially prepared starch solutions. These reagents cause flocculation of the solid particles and thereby increase the rate of sedimentation. The first important work on the flocculation and clarification of coal slurry was done by Henry, who developed the Henry process and used it successfully in Belgian preparation plants. Soon after the initiation in Belgium, the process was introduced into England, where subsequent investigation developed similar processes, which were successfully used in

Jan 1, 1940

By Charles C. Russell

Pressures developed by the coal during the coking process have been responsible for serious trouble to many companies that operate or build by-product coke ovens. The insidious nature of this trouble is indicated by the fact that oven operators, unless they have ascertained the characteristics of the coals they use, cannot determine the existence of dangerous pressures until after the damage has been done. A rather popular misconception is that ease of pushing of the charge at the completion of the coking period indicates that no dangerous pressure has been developed, but this is far from the truth; pressures developed during the coking period are often completely dissipated by the shrinkage of the charge in the completion of coking, and so any criterion based on the ease of pushing is futile. However, methods of testing coals are available that will definitely distinguish the dangerous coals from those that are safe, and such methods have been brought to a high degree of certainty. The operator should take no risk that involves either a lack of knowledge of the behavior of coals during coking or the failure to believe that certain of the coals or coal mixtures can cause damage to the oven structure. This merely tends to minimize the problem and leads to claims that coking coals of all kinds may be carbonized with no danger of damage to the oven structure provided some particular kind of byproduct oven is used. This lack of appreciation of the true situation is in itself one of the main factors leading to damage of coke-oven structures. Coals that develop more than a certain safe pressure cannot be used in by-product coke ovens without causing damage to the side flue walls. The following paper presents a highly certain method for the measurement of the pressures developed during the carbonization of coal—a method that is the culmination of many years of work on the problem in which a number of methods for the purpose were investigated and used. While actual pressures found may appear to be small when expressed in pounds per square inch, the number of square inches in each oven wall is so large that the total pressure on the oven wall becomes an enormous load.

Jan 1, 1940

By C. E. Lesher

Low-temPeRatUre carbonization needs no introduction to the literature on coal. This paper will attempt no review of that literature; it tells the story of the commercial development of one of the processes for carbonizing coal at so-called low temperatures, dealing with the technology and economics of the process rather than with the design of equipment. The process used is that covered by the patents of C. B. Wisner, with improvements and changes made by the Pittsburgh Coal Carbonization Co. The important and novel feature of the process is the phenomenon of making the product in ball form (Fig. 1). Melting coal dust into smokeless solid fuel in an apparatus as simple as a revolving cylinder of ordinary mild steel has been proved to be practical. Furthermore, it has been found that the process has a wide application, for by applying a knowledge of why and how this balling action is accomplished, the apparatus can be designed and adapted to treat any coal that has a certain minimum agglutinating property. The operating plant is 20 miles west of Pittsburgh, adjacent to the Champion preparation plant of the Pittsburgh Coal Co. The coal used is cleaned minus ?-in. or finer sizes. The products are low-tem-perature coke, tar and tar products. Present capacity is 7000 tons per month of low-temperature coke and 140,000 gal. of tar. There is no surplus gas for sale, and none of the light oils are stripped from the gas. The coke is marketed in two sizes, 1 to 2 in. and 1 to 6 in., through retail coal dealers for domestic consumption in hand-fired furnaces, and for fireplace fuel. The smaller size is bagged by dealers. Crude tar and many distillates and road tars are sold. Acknowledgments During the 10 years in which the work that is summarized in this paper has been under way, many have contributed time and effort, faithful service and fruitful thinking. Whatever success has attended

Jan 1, 1940

By H. L. Brunjes, G. B. Gould

The softening temperature of coal ash, as determined in the laboratory, has been used for years as an indication of the tendency of coal to form clinker and slag. It has not, however, provided an index of unfailing and complete reliability for forecasting the relative performance of different coals, even in the same plant and under the same conditions of operation. A large amount of investigation has been done in attempts to account for this, but practically all of this has been in the direction of studying the chemical constitution of coal ash. Much has been learned, but without much progress toward a better practical application of knowledge to the selection of coal. There has always seemed to be an important factor that eluded discovery. It has long been recognized that the ash-softening temperature as determined in the laboratory had one weakness, in that it represented the temperature at which a thorough mixture of the whole ash of the coal would melt, the mixing having taken place before the heal was applied. This does not, of course, simulate what takes place in practice. With certain exceptions, which will be noted later, almost all of the investigations that have been made, especially the chemical studies of coal ash, have likewise dealt with a mixture of the whole ash in the coal. The present investigation differs from these fundamentally in its approach to the problem. It started from the observation that some clinkering and slagging obviously takes place in a boiler furnace at temperatures below the laboratory-determined ash-softening temperature of the coal. Knowing that coal as it is shipped to market is a heterogeneous mixture of diverse mineral substances, each of which may act independently of the others in the initial stages of clinker and slag formation, it seemed logical to investigate separately the softening temperature of the ash in physically separable portions of the coal, and to determine the relative quantities of ash contributed by each portion. While the data are far from complete, covering as they do only a few coals, and while the conclusions to be reached from the data at this stage arc, in some measure, speculative, they are presented at this time because

Jan 1, 1940

By Joseph P. Kittredge

At the time this matter first came up in 1912, the National Malleable and Steel Castings Co. had seven basic-bottom open-hearth furnaces in its plant at Sharon, Fa., using fuel oil, then costing about 2½¢ per gallon. The fuel was satisfactory but its cost fluctuated widely, therefore the use of pulverized coal, which had been successful in the cement industry and in some metallurgical furnaces, was suggested. After a thorough investigation, and an actual trial on one of the open-hearth furnaces, to be certain that the siliceous ash in the coal would not seriously affect the basic slags, the management of this company approved its adoption and a complete installation was started. The conclusion had been reached that the success of the operation depended on pulverizing the coal to greater fineness than had thus far been accomplished, and all equipment was chosen with this in mind. A coal drier was deemed essential, and one of ample capacity and efficiency was selected, which has done all that was expected of it, working without failure and with very small maintenance cost, and is still in operation. Pulverizers The company demanded a pulverization of at least 90 per cent through a 240-mesh screen, with consistent results, and a machine was chosen that was grinding other materials than coal to the fineness desired. The control of fineness was obtained through a vacuum separator, and was governed by the size of the separator and the number and construction of the valves necessary to expand and direct the air current. These valves can be set to obtain any degree of fineness desired, and the result is consistent. Two machines were installed, each capable of producing 2½ tons per hour, 95 per cent through 240 mesh. These machines have done everything expected of them, have run continuously, and are still in operation. An examination of some of our pulverized coal by elutriation and microscopic measurement disclosed that a 1-lb. lump of bituminous coal,

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