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The October meeting of the Mining and Geological Society, Lehigh University, was held in the Eckley B. Coxe Mining Laboratory, and was called to order at 8:15 p. m., by President R. L. McCann. After the reading of the minutes of the previous meeting, a short business session was held, for the purpose of electing a Vice-President to succeed G. S. Borden, who has left college. Nominees for the office were: H. J. Sloman, P. S. Justice, L. Sargeant. Dr. F. F. Hintze, of the Geology Department, was elected a member of the Advisory Committee, which now consists of the President-

Jan 12, 1916

New York Meets first Wednesday after first Tuesday of each month. DAVID 11. BROWNE, Chairman. JOHN H. JANEWAY, Vice-Chairman. F. E. PIERCE, Secretary, 35 Nassau St., New York, N. Y. P. A. MOSMAN, Treasurer. LEWIS W. FRANCIS. BENJAMIN B. LAWRENCE. Boston Meets first Monday of each winter month. HENRY L. SMYTH, Chairman. ALFRED C. LANE, Vice-Chairman. HENRY A. WENTWORTH, Secretary-Treasurer, 60 India St., Boston, Mass. A. H. EUSTIS, F. G. STANTIAL. Columbia Holds four sessions during year. Annual meeting in September or October. STANLY A. EASTON, Chairman. FREDERIC KEFFER, Vice-Chairman. LYNDON K. ARMSTRONG, Secretary-Treasurer, P. O. Drawer 2154, Spokane. Wash. D. C. LIVINGSTON. FRANK A. ROSS.

Jan 4, 1916

New York Meets first Wednesday after first Tuesday of each month. L. W. FRANCIS, Chairman, WILLARD S. MORSE, Vice-Chairman,. THOMAS T. READ, Secretary, Woolworth Bldg., New York, N. Y. P. A. MOSMAN, Treasurer. LOUIS D. HUNTOON. - WILLIAM A. POMEROY. Boston Meets first Monday of each winter month. HENRY L. SMYTH, Chairman. ALFRED C. LANE, Vice-Chairman. HENRY A. WENTWORTH, Secretary-Treasurer, 60 India St., Boston, Mass. ROBERT H. RICHARDS, ALBERT SAUVEUR. Columbia Holds four sessions during year. Annual meeting in September or October. FRANK A. ROSS, Chairman. RUSH J. WHITE, Vice-Chairman. LYNDON K. ARMSTRONG, Secretary-Treasurer, P. O. Drawer 2154, Spokane, Wash. FREDERIC KEFFER, FRANCIS A. THOMSON.

Jan 5, 1915

By C. D. Stahl, S. M. Farouq Ali, W. E. Culham

One- and two-dimensional mathematical models have been developed that simulate transient, two-phase flow of hydrocarbon mixtures in porous media in a manner that accounts for interphase mass transfer. Numerical simulations of one-dimensional depletion-drive experiments using a two-component hydrucarbon fluid were used to establish the validity of the mathematical models. In addition, the experimental and numerical data were used to demonstrate that production rate had a relatively insignificant effect on the recovery of individual hydrocarbon components from the experimental system, and that attainment of equilibrium between phases is possible for a wide range of liquid and vapor velocities in reservoirs containing light hydrocarbon fluids. Results of some two-dimensional numerical simulations are also presented. INTRODUCTION This study was undertaken to develop a mathematical model that would simulate transient, two-phase flow of hydrocarbon mixtures in porous media under conditions that result in interphase mass transfer and to test the validity of the assumptions used to set up the model. In addition, the study was designed to determine if production rate influences the recovery of individual hydrocarbon components from reservoirs producing by depletion drive. Two-phase flow in porous media, with interchange of components between the two phases, is important in many petroleum recovery processes. Studiesl-5 conducted within the last 3 years have outlined methods of solving multiphase flow problems incorporating mass transfer. Some of these studies have also indicated the importance of accounting for mass transfer under various producing conditions. An earlier work6 first demonstrated the importance of combining relative permeability data with equilibrium ratios in compositional balance methods. The mathematical model presented in this paper is formulated so that a phase behavior package, as described in previous papers,5,17 is not required as an integral part of the routine employed to solve for the primary dependent variables. The finite difference formulation is designed so that all the primary variables can be solved for simultaneously. This is accomplished by utilizing one basic set of equations. These innovations, which are in contrast to other models2,3,5 but are similar in some respects to the approach used by Taylor,l render the total problem computationally simpler than any of the previously referenced formulations. The mathematical model was developed by combining Darcy's law with a continuity equation for each hydrocarbon component. The principal assumptions invoked in the formulation were (1) that capillary forces and diffusional effects are negligible, and (2) that thermodynamic equilibrium exists in the reservoir at all times. No assumption as to the type of vaporization process was made in formulating this model. Experimental data were required to complete this study. These were generated by conducting several depletion drive experiments. The experimental apparatus consisted of a sandstone core enclosed in a pressurized casing. The apparatus was designed in such a manner that the core could be charged with a liquid hydrocarbon mixture and depleted at different production rates. The experimental tests were designed to determine the effect of production rate on component recovery. In addition, direct comparison of experimental and mathematically predicted

Jan 1, 1970

AN excerpt from the Odessa Isvestia Vechernia, dated April 24, 1927, and translated by P. Routzitzky, summarizes a report by Professor Preobragensky ad-dressed to the Russian Government. Professor Preo-bragensky writes in regard to newly discovered potash deposits in eastern European Russia. Near Solikamsk in the Perm district, about 100 miles east of the Ural mountains and on the Kama river, large deposits have been located and the belt has been traced southward about 25 miles. Subsequently the search for potash salts has been transferred to the Berezovsky district, near Nikolsk, on the Kosva river, about 70 miles north-east-of the city of Perm. Toward the end of March, 1927, at a depth of 800 ft., drilling at the Levensk salt works revealed potash salts in, definite strata. The bed has a thickness which extends from a depth of- 245 to 307 meters (latest data in middle of April, 1927) and has not been bottomed yet. Professor Preobragensky goes on to say that these operations have definitely, developed very large de-posits, the limits of which are still unknown. There is at least an implication that the reserves far exceed those of Stassfurt, Germany. For European use they are, of course, much less accessible, for by through rail they lie 1200 miles east of Petrograd. Further, no particulars are given as to the potash content. It is to be supposed that the salt occurs in the Permian beds that extend north and south along the west flank of the Ural mountains, as the localities mentioned .all lie within the area so mapped on the new geologic map of Russia issued by the Scientific and Technical Di-vision of the All-Russian Council of Soviets in 1922. It is interesting to note that these salt beds are of the same geologic age as those of 'the Stassfurt region and of the west Texas-New Mexico district, in which the U. S. Bureau of Mines is at present drilling.

Jan 7, 1927

By E. T. Turkdogan, L. J. Martonik, R. J. Fruehan

Electrochemical measuretnents of the solid oxide electrolyte galvanic cells CY-Cr2O3 I ZrO2 (CaO) 1 O (in Fe alloy) CY-Cr2O3 I Tho2 (Y2O3)I O en Fe alloy) have been made at 1600°C (2912°F) in order to test the Performance of such cells at liquid steel temperatures. The oxygen pvobe (cell) consists of a disk of ZrO2 (CaO) or Tho2 (Y2O3) electrolyte fused at one end of a silica tube filled with a mixture of Cr-Cr2O3 which is the reference electrode. Upon immersion in liquid steel, the electromotive force readings achieve a steady value within a few seconds, and remain steady for 30 win or more. The perforwzance of the probes has been tested using Fe-O, Fe-Si-O, Fe-Cr-O, Fe-V-O, and Fe-Al-O alloys; the oxygen contents of liquid steel derived from the measured electromotive forces are in satisfactory agreement with those determined by arulysis. Use of the probe in the deoxi-datiorz of steel, in laboratory experiments, is discussed. The results indicate that there is insignificant electronic conductivity in ZrO2(CuO) at oxygen activities down to those corresponding to 10 ppm in steel. At lower oxygen activities, probes tipped with ThOn (Y2O3) disks perform satisfactorily at oxygen activities down to 1 ppm O or less. THE key to the control of deoxidation of steel is a sensing device to measure rapidly the concentration of oxygen in liquid steel in the furnace, ladle or tun-dish at any desired stage of deoxidation. The analysis of the cast steel by the neutron-activation or vacuum-fusion method gives total oxygen as oxide and silicate inclusions. This analysis is important for guidance to steel cleanliness; however, such a postmortem is of little value in the control of deoxidation of liquid steel. At the General Meeting of the American Iron and Steel Institute in New York, 1968, Turkdogan and Fruehan' presented a paper on the preliminary results of the work done in this laboratory on rapid determination of oxygen in steel by an oxygen probe. Details of the work done in this laboratory leading to the development of a galvanic cell for the determination of oxygen in liquid steel, and the results of the tests made are given in this paper. It was through Wagner's contributions, since the early Thirties, to the physical chemistry of semiconductors in general that it ultimately became possible to construct galvanic cells for application at high temperatures. In 1957, Kiukkola and wagner2 successfully demonstrated the use of several solid electrolytes in measuring the free energies of several chemical reactions, in particular, the use of lime-stabilized zir-conia in high-temperature oxidation reactions. Starting 7 years later, a number of papers appeared in the technical literature3-' demonstrating possible applicability of galvanic cells for the determination of oxygen in liquid steel. In the earliest work, Japanese investigators3j4 experimented with various types of reference electrodes, e.g., graphite-saturated liquid iron at 1 atm CO or Ni-NiO mixtures; the results obtained, though promising, were not of sufficient accuracy. Except for the work of Baker and West,6 all other investigators5,7,8 showed that ZrO2(CaO) electrolyte could be used for this purpose. The main part of the galvanic cell used by Fischer and ~ckermann' and by schwerdtfeger7 (the latter work was done in this laboratory), consisted of a ZrO2(CaO) tube, -1 cm ID, closed at one end, with a platinum contact wire fixed mechanically inside the closed end. The tube was flushed with a gas of known oxygen partial pressure, e.g., air, CO-CO2 or H2-CO2 mixtures; gas along with the platinum lead wire served as the reference electrode. The oxygen contents derived from measured electromotive forces agreed reasonably well with the oxygen contents determined by vacuum-fusion analysis. It is evident from recent investigations that the electromotive force technique using a solid oxide electrolyte is fundamentally well suited for the determination of oxygen in liquid steel. However, it is equally clear that the cell arrangement of the type as commonly used is in need of considerable improvement, as it exhibits several shortcomings for industrial and even laboratory use. 1) Because of its size, the zirconia tube, though stabilized, has a poor resistance to thermal shock. 2) Fine pores and microcracks, which are invariably present in zirconia tubes, are detrimental to the satisfactory operation of the cell, particularly when gas reference electrodes are used. 3) Air or carbon dioxide reference electrodes give rise to high electromotive force readings; as a result, the determination of oxygen in steel becomes less accurate. For higher accuracy, the oxygen partial pressure of the reference electrode should be in the range similar to that of oxygen in steel. 4) Even in laboratory experiments, difficulties are experienced when flushing the tube with gases and maintaining the proper gas flow rate. Fischer and Ackermann,' who used air as the reference electrode, reported that when the flow rate was too low, furnace gases would leak into the electrolyte tube, therefore lowering the oxygen potential and measured electromotive force. The required flow rate in order to avoid leakage depended on the tightness of the electrolyte tube which varied with different tubes, thus making it difficult to predict in advance the required flow rate. However, if the flow rate is too high the inside wall of the electrolyte tube would be cooler than the wall

Jan 1, 1970

By S. D. Thompson

An AIME transaction by P.L. Allsman (1960), "Analysis of Explosive Action in Breaking Rock," contains many useful concepts applicable to blast design. Of special interest was the derivation of a heave blasting burden formula, which is important in establishing the maximum burden. Unfortunately, its derivation contains a fundamental error and then a compensating one. This formula was: [bImpulse _P d At max =nwuwu where: bmax = maximum burden in m (ft). P= average detonation pressure in N/m2 (lb per ft2). A= duration of the average detonation pressure in sec. n= 3.1416. w= specific weight of rock in N/m3 (lb per ft3). u= minimum velocity that must be imparted to the rock in m/sec (ft per sec). d= borehole diameter in m (ft).] The derivation used the impulse-momentum principle, with a 0.3 m thick (1 ft thick) radius of rock mass being acted on by an explosive impulse. However, the rock weight was not converted to mass by dividing by the gravitational constant. This mistake can be verified by checking the units in the formula, which resolve into m1/2-sec instead of meters. Adding the gravitational constant corrects this: [P d Atbmax =gu g = gravitational constant equal to 9.81 m/sect (32.2 ft per sec2).] If the minimum velocity is now taken as 0.3 m/sec (1 ft per sec), as was recommended, excessively large burdens result. Any appropriate rock velocity should be capable of throwing the rock mass a minimum distance so that the swell needed for breakage can occur. Unless the burden is extremely small and the bench height is very high, it is unlikely that a velocity of only 0.3 m/sec (1 ft per sec) will be sufficient. A review of the literature indicates that average rock velocities observed by camera studies typically range from 15 to 30 m/sec (50 to 100 ft per sec) (duPont deNemours and Co., 1977; Tansey, 1979). On this basis, a minimum rock velocity of about 15 m/sec (49 ft per sec) would be expected and not 0.3 m/sec (1 ft per see). When a velocity of 15 m/sec (49 ft per sec) is inserted into the corrected formula, the results approximate those of the original formula with a velocity of 0.3 m/sec (1 ft per sec). The errors were approximately self-compensating [(9 = 1.24)]. ? References Allsman, P.L., 1960, "Analysis of Explosive Action in Breaking Rock," Transactions, AIME, Vol. 217, pp. 475-476. E. I. duPont deNemours and Co., 1977, Blasters' Handbook, 175th Anniversary Edition, Dupont, Wilmington, DE, pp. 233. Tansey, D.O., 1979, "The DuPont Sequential Blasting System," CIM Bulletin, Vol. 72, July, p. 80. Thompson, S.D., 1981, "Critical Review and Revision of Surface Blasting Practice," MS thesis, West Virginia University, pp. 308-316.

Jan 1, 1986

By C. P. Besse, G. W. Osborne

Drilling operations on the Gulf of Mexico continental shelf are following two general plans. The first plan placed all buildings. equipment, and materials on a large platform while the second copies Lake Maracaibo drilling methods by placing a minimum of equipment on the platform and using a floating drilling tender or drill barge anchored adjacent to the platform in an attempt to take maximum advantage of flotation in an effort to reduce costs. A complete power rig, mud pits, pipe racks, and small auxiliaries have been placed on the platform as shown in Figure 1, to assure proper handling of drilling operations under the most adverse weather. This has the economic advantages of eliminating the fabrication of a large structure, a material reduction of rigging up time on each structure. and elimination of delay time on equipment while quarters and equipment placed on the tender are being rigged up. These factors have certainly proved their economic value in drill barge operations in the Louisiana and Texas coastal area. Naturally the heart of such an operation is the anchorage. This anchorage, to permit taking advantage of maximum economies. must hold the ship in a relatively fixed position under the following weather conditions: (1) Winter northwesters (2) Southeasterly gales (3) Severe rain squalls In hurricane weather all operations must be suspended regardless of method of operation and the floating tender will, if possible, be moved to a protected location. However, in case of an emergency the floating tender should be prepared to swing with the wind and seas on a single anchor as there is considerable doubt that either ship or anchorage could withstand the severe strains imposed by fixed anchorage in hurricane weather. Northwesters are reported to have developed wind gust velocities as high as sixty miles per hour and steady winds of approximately forty-five miles per hour. Fortunately these winds do not travel a great distance over water before arriving at our location, and, while they can possibly develop seas approximately eight feet high, the waves are short, and if taken on the bow or stern of a ship 250 feet or more in length they should not present a difficult problem. On the other hand southeast gales having a long fetch and maximum steady wind velocities of approximately thirty miles per hour can possibly develop long swells approximately twelve feet in height. These should develop the most severe anchorage problem. To combat these waves any workable anchorage plan should permit maneuvering the ship in such a manner as to head the bow or the stern into the seas and still maintain contact with the structure. Rain squalls develop wind gust velocities as great as sixty miles per hour. However, they are local and the seas have not been a source of great trouble. To permit continuous operation the anchorage must be able to withstand these winds broadside to the ship as the wind direction often changes radically in a very sliort time. All of these weather conditions are dangerous to small craft, deck cargoes,

Jan 1, 1949

By C. P. Besse, G. W. Osborne

Drilling operations on the Gulf of Mexico continental shelf are following two general plans. The first plan placed all buildings. equipment, and materials on a large platform while the second copies Lake Maracaibo drilling methods by placing a minimum of equipment on the platform and using a floating drilling tender or drill barge anchored adjacent to the platform in an attempt to take maximum advantage of flotation in an effort to reduce costs. A complete power rig, mud pits, pipe racks, and small auxiliaries have been placed on the platform as shown in Figure 1, to assure proper handling of drilling operations under the most adverse weather. This has the economic advantages of eliminating the fabrication of a large structure, a material reduction of rigging up time on each structure. and elimination of delay time on equipment while quarters and equipment placed on the tender are being rigged up. These factors have certainly proved their economic value in drill barge operations in the Louisiana and Texas coastal area. Naturally the heart of such an operation is the anchorage. This anchorage, to permit taking advantage of maximum economies. must hold the ship in a relatively fixed position under the following weather conditions: (1) Winter northwesters (2) Southeasterly gales (3) Severe rain squalls In hurricane weather all operations must be suspended regardless of method of operation and the floating tender will, if possible, be moved to a protected location. However, in case of an emergency the floating tender should be prepared to swing with the wind and seas on a single anchor as there is considerable doubt that either ship or anchorage could withstand the severe strains imposed by fixed anchorage in hurricane weather. Northwesters are reported to have developed wind gust velocities as high as sixty miles per hour and steady winds of approximately forty-five miles per hour. Fortunately these winds do not travel a great distance over water before arriving at our location, and, while they can possibly develop seas approximately eight feet high, the waves are short, and if taken on the bow or stern of a ship 250 feet or more in length they should not present a difficult problem. On the other hand southeast gales having a long fetch and maximum steady wind velocities of approximately thirty miles per hour can possibly develop long swells approximately twelve feet in height. These should develop the most severe anchorage problem. To combat these waves any workable anchorage plan should permit maneuvering the ship in such a manner as to head the bow or the stern into the seas and still maintain contact with the structure. Rain squalls develop wind gust velocities as great as sixty miles per hour. However, they are local and the seas have not been a source of great trouble. To permit continuous operation the anchorage must be able to withstand these winds broadside to the ship as the wind direction often changes radically in a very sliort time. All of these weather conditions are dangerous to small craft, deck cargoes,

Jan 1, 1949

By B. Ehgartner, M. Karmis, C. Haycocks

INTRODUCTION Interaction effects between mining operations developed on different levels of contiguously placed bedded deposits are a major problem in many parts of the United States. Efficient design criteria, derived from careful analysis of field studies as well as theoretical and model analysis, have shown that many problems that may result from seam interaction can be avoided, minimized, or on occasion used with beneficial results. Selection of seam mining sequence is rarely based only on ground control considerations, but rather on factors such as seam ownership, characteristics and accessibility. Therefore all possible mining sequences and methods must be considered in the development of multi-seam design criteria. Increased utilization of the longwall method of mining creates special problems in multi-seam mining due to the inevitability of caving and, therefore, subsidence. Considerable valuable experience has been gained in this direction from British mining operations where interaction problems are extremely common and the longwall method is the primary method used. However, limitations persist in the application of much of this information to U.S. conditions due to variation in layouts between advancing and retreating longwall systems and the inherent difference in the geologic environment. Considerable field data is available for U.S. room and pillar operations, and where successful pillaring operations have been carried out close parallels may be drawn with ground control mechanisms in longwall operations. VARIABLES AFFECTING INTERACTION Analysis of ground control multi-seam problems shows that factors contributing to interaction may be classified into variables that are fixed by the geo- logic environment and those that depend on engineering design. These may be described as follows (Haycocks et al, 1982): Fixed: (i) Depth*, (ii) Innerburden thickness*, (iii) Innerburden physical characteristics*, (iv) Stress field, (v) Seam thickness, (vi) Coal characteristics, (vii) Completed mining operations either above or below the seam to ve mined, and (viii) Age of workings. Mining: (i) Height, (ii) Dimensions and geometry*, (iii) Method*, and (iv) Spatial location of entries* *These variables are regarded as being critical effects. Few attempts have been made to quantify the variables controlling interaction. Dunham's (1978) linear regression analysis, combined with the National Coal Board's MRDE Roadway Monogram, is probably the most noteworthy. This study involved eighteen multi- seam longwall mines in the United Kingdom. The following variables were taken into account: seam depth, pillar width, opening shape, floor and roof conditions, rate of face advance, vertical seam interval, and age and geometry of overlying workings. A multi-linear regression analysis of these variables fitted by appropriate field values developed a prediction equation for lower seam stability. Although the study included many variables, it neglected the nature or geology of the innerburden. The prediction equation was found unreliable, but it did show the significance of geometry as a factor controlling lower seam stability. MINING SEQUENCES Irrespective of the type of mining method selected only four possible mining sequences are possible for two seams. Finlay and Winstanley wrote in 1934. .. "Experience has shown that whether the upper or lower seam was worked first, the subsequent working of the other seam in the same area seriously affected the roads in the seam first worked." Those effects depended upon the four basic mining sequences: 1) The upper seam is mined out prior to mining the underlying seam; i.e., the seams are mined sequentially downward and mining of the underlying seam is not commenced until operations above are completed and have ceased.

Jan 1, 1982

By Arthur L. Crawford

PLAGER gold is perhaps the most difficult of the common mineral deposits to evaluate. Not only are the erratic pay streaks a source of never-ending uncertainty, but the spotty distribution of the gold in iso¬lated nuggets is a perpetual source of error and disappointment to the sampler. Experienced engineers have long realized this difficulty and have been wary in predicting assay values except in the most general terms, and then only after extensive and usually expensive sampling. During the past year I have had the opportunity of checking a number of gold placers, and from the results obtained I believe that the microscope may be employed to advantage in placer sampling.

Jan 1, 1933

By G. F. Loughlin

Nearly 55 years have elapsed since the U. S. Geological Survey was organized. During this period the mineral industries have grown from infancy or early childhood to well developed maturity, and some mining industries appear to be declining. Decline in mining is even more conspicuous in certain districts, or even states, than in any mining industry as a whole, and discoveries or developments of important new districts have been very few in recent years. Our reserves of most metals and nonmetals are adequate for many years of normal production, subject to changes in rate of exploitation that may be caused by foreign competition and by changes in technology and market conditions; but the declining production of many districts and the scarcity of important new discoveries to compensate for them is cause for some apprehension. It is likely that future production will depend to a considerable extent on deep exploration and on the thorough examination of places in and adjoining existing districts where significant outcrops are lacking and only apparently insignificant geologic features offer encouraging clues—more so than on the discovery of entirely new districts. Even the possible discovery of new districts, which have escaped the intensive search by the nontechnical prospector, must be dependent on geologic information. There is therefore no need for argument as to the value of geology to the mining industry; and the need of continuing to advance the science. The U. S. Geological Survey's past contributions to mining geology, as expressed in its official reports and other publications, are recognized by Ransome in his historical review of mining geology in the Lindgren volume.' These contributions have helped to form the basis of modern mining geology, which has grown rapidly in the last 25 years. This growth has witnessed the organization and maintenance of geologic staffs by mining companies in several districts, and their detailed work, together with the fact that nearly all of the larger mining districts have

Jan 1, 1935

By W. E. Bearce

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

Jan 1, 1968

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By Orville R. Lyons

DISCUSSION Judson S. Hubbard (The Humphreys Investment Co., Denver)—In this very interesting paper several brief references are made to the Humphreys spiral, a device used for cleaning fine coal. In Table I, Plants 76 and 77, data are given on spiral performance treating Raton and Trinidad coal. The fine coal, as fed to the spiral in these instances, is actually a table middling, hence the more easily treated material was previously removed and a large number of particles were present which were difficult to clean. Mr. W. M. Bertholf of the Colorado Fuel and Iron Co. presented a paper in February, 1946, Cleaning Table Middlings from a Coal Washery with the Humphreys Spiral Concentrator, from which I quote: "In considering the results of our tests it should be noted that the feed was our table middling, and that any real separation is a 'moral victory' as there is little material that could properly be called coal and practically no heavy rock, the consequences being that previous attempts to clean the middling have not at all been successful." Referring again to Table I, Plants 72, 73, 74, and 75, these data were obtained by Yancey and Geer and others and presented at the February 1950 Meeting, AIME, in a paper entitled Laboratory Performance Tests of the Humphreys Spiral as a Cleaner of Fine Coal. Results shown for those tests involve all particles from 8 mesh through the colloids, which admittedly is not an ideal situation for spiral feed if much refuse is contained in the —80 mesh or —100 mesh size range. As an illustration of the effect of treating too broad a size range, let us consider Plant No. 75, Kentucky No. 9 seam. Spiral feed was 8 mesh x 0. Now had this been 8x100 mesh the percentage of misplaced material would have been 8.0 pct instead of the reported 15.26 pct. Similar comparisons can be made on the other data presented with respect to the spiral. Other types of equipment show a similar trend in that whenever too fine a size is treated in a given unit process the percentage of misplaced material increases. Since the spiral is working near the finer end of the size range, it will sometimes be advantageous to treat the entire range of —8 mesh material rather than to deslime and make a fair showing on, say, the +80 or + I00 mesh. Desliming is subsequently done in any case in the dewatering or thickening operation. Results obtained by spiraling any given coal depend on factors too numerous and complex to discuss here, but there are strong indications that proper preparation of feed to the spiral can improve results obtained on some of the raw coals tested. This is clearly pointed out at the end of the aforementioned paper by Yancey and Geer. "The spiral is an extremely simple device which involves no moving parts and is constructed almost entirely of unmachined castings. Since it is such an uncomplicated mechanism, operation is simple and virtually foolproof. These characteristics, which go far toward insuring low cost operation, are attractive attributes in any coal cleaning unit." Certain equipment used in conjunction with the spiral has resulted in a decrease in the percentage of misplaced material, notably in actual practice the launder screen which is used to remove objectionable high ash fines from a spiral-washed coal product. Private correspondence with the U. S. Bureau of Mines has intimated that an additional yield of coal is possible by flotation of the spiral middling. Possible future improvements and developments may result from other methods now under consideration. Finally, some compromise must be made between the best metallurgical performance and the best practical or economical results. Mr. Lyons emphasizes in his summary this objective of overall economy in selection of equipment. G. B. Walker (American Cyanamid Co., Stamford, Conn.)—I had the pleasure of reviewing the draft of this paper and my curiosity was aroused by the data given for Tromp plants, in that all of the examples shown appeared to be 2-product separations, whereas all the Tromp plants with which I am familiar have been 3-product units. The data given for plants No. 101 and 102 appear to be taken from Tromp's brochure on his process and represent the results obtained at the Dominale plant in Holland which was operated for many years by Mr. Tromp. The plant, which was designed to treat 58 tons per hr, was sampled while treating 35 tons per hr of 3Y4x-in. coal. Example 101 appears to conform to what would result if the middling product were calculated into the refuse product, while Example 102 represents the calculation of the middling into the coal product. It is believed that Examples 103 and 104 represent the operation of the Willem-Sophie Mine in Holland recalculated on the same basis. In checking the English examples given by numbers 14, 15, 16, and 17, the same procedure seems to have been followed. These results have, apparently, been taken from an article in Colliery Engineering in August 1941, describing the initial operation of the Williamthorpe Colliery of the Hardwick Colliery Co. Two vessels are employed in this plant, one to treat soft coal and one to treat hard coal. Example No. 14 presents the results that could be obtained from the soft coal bath if the middling were calculated into the refuse, and Example 15 the results when the middling is calculated into the coal. Examples 16 and 17 represent the same expedient in the case of the hard coal bath. Of interest to this discussion is the fact that during the past year the Simon-Carves Engineering Co. in England has installed in the Williamthorpe plant their new "Sim-Car" medium cleaning system which is based on magnetic extraction and control and which is licensed under the Heavy-Media Separation Processes patents by the American Zinc, Lead and Smelting Co. This system has been described in the December 1951, issue of Colliery Engineering. It is reported that since the Williamthorpe Colliery was changed from the Tromp system of medium cleaning to the Magneto-Motive method of medium control the opera-

Jan 1, 1953

By R. Bainbridge

THE Consolidated Mining and Smelting CO. of Canada Ltd. has operated a lead smelter at Trail, B. C., for many years. In order to take advantage of metallurgical advances, as well as to improve materials handling methods, this company, commonly known as "Cominco," commenced planning a program of smelter revision and modernization some years ago. The first stage of this program involved the design and construction of a new blast furnace gas cleaning system. The selection of equipment, the design of facilities, and preliminary operating details of this system will be dealt with in this paper. The essential problem was to clean and collect 100 tons of dust daily from 153,000 cfm* (12,225 lb per min) of lead blast furnace gas which varied in temperature from 350º to 1100°F. Because it was desired to collect the dust dry, either a Cottrell or a baghouse cleaning plant was to be selected. Comin-co's many years of experience with both systems provided a background for choosing the most satisfactory installation. All information pertinent to the two methods of dust recovery was carefully investigated, and it was decided to replace the existing equipment with a baghouse. Very briefly, the reasons for this decision were as follows: 1—A baghouse installation would be practical because the SO2 content of the gas was low and corrosion would not be a problem if the baghouse operating temperatures were held sufficiently above the dew point. 2—Variations in the physical characteristics of fume and dust, which are inherent in this blast furnace operation, should not substantially affect the operating efficiency of a baghouse. 3—For the same capital cost, metal losses (stack and water losses) would be appreciably less in a baghouse. 4—A baghouse would be easier to operate, and would not require the use of highly skilled labor. 5—Operating and maintenance costs of a bag-house would be lower. 6—The only available space for reconstruction was relatively small, and not suited to a Cottrell installation. Once the baghouse system was decided upon, detailed design of the installation was begun. Baghouse Design Gas Cooling: Before the required capacity of the baghouse could be determined, the method of cooling the gas to the temperature necessary for bag-house operation had to be chosen. The problem confronting the design engineers was how best to cool 153,000 cfm of gas from a temperature ranging from 350°F to brief peaks of 1100°F, down to 210°F, the maximum safe baghouse inlet temperature. A survey of existing blast furnace gas temperatures in the outlet flue showed that the normal range was as given in Table I. The obvious choices of cooling method were: 1— cool completely by the addition of tempering air; 2—utilize a heat exchanger; 3—cool by radiation; and 4—cool with water spray in conjunction with the admission of tempering air. The advantages and disadvantages of the various cooling methods were: Air Addition: To cool completely by the admission of tempering air involved tremendous volumes, Fig. 1. For example, to cool 1 lb of blast furnace gas at 450°F requires 1.84 lb of air at 80°F or 1.60 lb at 60°F. As it is necessary to design for peak conditions, it can readily be seen that volumes of tempering air in the order of 1,500,000 cfm would have to be handled. Using the normal design figure of 2.5 cu ft per sq ft of bag area, a baghouse installation comprising some 600,000 sq ft of filter cloth would be necessary. Such design requirements would be prohibitive, not only from a standpoint of capital expenditure, but also because of space limitations. Heat Exchanger: The utilization of a heat exchanger was given serious consideration. A horizontal tube unit using air as the medium to cool the required volume of blast furnace gas from 400" to 250°F was investigated. Cooling above 400°F would be done by water spray, and below 250°F by admission of tempering air. The estimated capital cost of such a unit was found to be prohibitive. From an operating standpoint, there was considerable doubt as to whether the soot blowing equipment provided would effectively keep the dust from building up on the tube surface. The performance of heat exchangers operating on dusty gas in other company operations had not been too favorable. Radiation Cooling: Although somewhat cumbersome, gas cooling by radiation from 'trombone' tubes or other similar equipment (cyclones) is employed in many metallurgical operations. Such an installation was also considered. However, calculations showed that an installation much larger than the space available would be required to handle the gas volume involved. For example, to cool 153,000 cfm of blast furnace gas from, say, 600' to 250°F (i.e., remove in the order of 58,500,000 Btu per hr with heat transfer rates varying from 1.1 Btu per sq ft per hr per OF for the higher temperature ranges to 0.88 Btu per sq ft per hr per OF for the lower ranges) would need a cooling area of some 175,000 sq ft.

Jan 1, 1953

By Charles T. Holland, David A. Olsen

Considerable work has resulted in the development of a formula by which the strength of coal pillars for an L to T ratio of up to 12 may be estimated. One of the factors involved in this formula is the coefficient of friction between the coal pillar and the adjacent rock with which it is in contact. This paper presents the results of studies on sandstone, limestone, shale and coal and the coefficient of friction between coal and the other three under wet and dry conditions. With the coefficients of friction thus determined and the use of results obtained earlier by Meikle and Holland, it seems probable that coal pillar strength may be more closely estimated. Several investigators1-3 have found that the com-pressive strength of cubical coal specimens can be correlated by an equation such as

Jan 1, 1969

By Joseph P. Hodgson

In 1880, mining operations were commenced at the Copper Queen mine. The famous Queen ore body, which extended to the surface, was first quarried from a large open cut in the outcrop. The ore body was followed down to the 300 level with the Queen incline, and stoped. This, we believe, marks the introduction of square setting in the Bisbee district, which is still the method most generally used. Orebodies The orebodies in the Copper Queen mine occur in the limestones, and the main portion of the ore has been mined from the Abrigo, Martin, and Escabrosa limestones. mile the orebodies outcrop in the extreme western end of the mine, the general dip is to the east and south, at an angle of about 20°, although this dip is not by any means regular, being, in fact, very irregular locally. The ore is remarkable for its variableness in character, some of it being very soft, requiring a large amount of timber, and other portions consisting of extremely hard sulphides. The orebodies, in general, while remarkable for their continuity, are very irregular, both as to shape and size. The mine having been in operation so many years, and mining having been prosecuted over such a large area, the operations have caused the shrinkage of the overburden, and there is a constant movement, which is shown by large cracks and interstices in the surface rocks, which largely contribute to the very high deadwork or maintenance cost of the mine. It is surprising that for year 1913 it cost almost $400,000 just for the one item of repairs and dead work in this property. Haulage and Hoisting The underground openings have been extended in the neighborhood of 200 miles. The ore hoisting is performed at a centrally located shaft, the Sacramento, to which all the ores from different parts of the mine are conveyed by electric haulage, of which there was slightly over 9 miles in operation in 1913. The hoisting levels at the Sacramento shaft are placed 200 ft. apart, commencing at the 400 level and continuing down to the 1,600 level, the average hoisting distance being 1,000 ft. The

Jan 1, 1915

By Robert T. MacMillan

Of all the mineral substances utilized by man, salt or sodium chloride has one of the longest and most varied histories. Because all animal life is descended from marine organisms, sodium and chlorine became necessary to life, and as sodium chloride meets requirements most readily, common salt has been essential to human and animal diets from earliest times. As civilization progressed, man learned to evaporate sea water and brines of saline lakes and springs to obtain salt, and gradually ceased to depend on the flesh of animals and the occasional salt outcrops or "salt licks" they frequented. He also learned that salt had a multitude of uses, particularly as a preservative of meat and other foods. The importance of salt in the minds of early men may be readily seen from many allusions in the Bible and other ancient writings. Expressions such as "not worth his salt," and "the salt of the earth" are deeply embedded in the English language. Salt has symbolized friendship, fidelity, purity, and hospitality. It has been used as currency; had religious significance; caused wars; and has been a favorite commodity for taxation. The English word "salary" stems from the Latin word for salt, in which Roman soldiers received part of their pay. Although salt today retains its dietary importance, it has thousands of other uses. Of each 100 pounds produced in the United States less than three pounds is for table use. The other 97 pounds goes for agricultural, industrial, and chemical uses. Fortunately, reserves of salt are enormous and reasonably well distributed on the face of the earth. The sea is an inexhaustible source of salt which may be had at prices which vary with climate and other factors. In addition, large underground salt deposits and many salt lakes, springs and playas are scattered throughout the world. Most countries have access to a salt supply, although some are more favored than others. Properties Crystallized sodium chloride or common salt is known as halite. Its molecular weight is 58.454 and it contains, by weight, 39.34 pct sodium and 60.66 pct chlorine. Alternate positions in the cubic lattice are occupied, respectively, by negative chloride and positive sodium ions; the crystal form is isometric. Crystallization from water produces concentric shells, and when broken, a crystal shows conchoidal fracture. Salt has a hardness of 2.5; its specific gravity ranges from 2.1 to 2.6 for rock salt and is 2.165 for the pure crystal. The refractive index is 1.554; it melts at 800.8 ± 0.5°C and boils at 1413°C. Its heat of fusion is 123.59 cal/ gram. Salt is transparent or translucent and highly diathermous. It is colorless when pure and has a vitreous luster. Salt containing traces of iron oxides has varying shades of yellow, orange, brown, pink, or red. The peculiar shade of red sometimes found in that produced by solar evaporation of sea water has been attributed to tiny organisms. Blue mottlings sometimes seen in halite crystals have been attributed to colloidal sodium. Salt is soluble in water to the extent of 35.7 parts per 100 parts of water at 0°C and 39.8 parts per 100 of water at 100°C. The small difference in solubility with increasing temperature is important in the economics of producing salt by evaporating brine. Salt has a negative heat of solution which is 1.281 kg-cal/ mole at 18°C in 200 moles of water.

Jan 1, 1960

By W. A. Ver Wiebe

Spurred on by the very high demands for new crude supplies resulting from World War 11, the oil men of Kansas strained every effort in 1943 to find as many new pools as possible. In all, 1761 test wells were drilled, ,of which 895 were successful in finding new oil supplies and 32 in finding new gas supplies. The total number of wildcat wells drilled in Kansas was 577, including the so-called "extension wildcats," which are located less than 2 miles from production. As a result of the very determined effort to find new oil reserves, 49 of the wildcat test wells found new oil pools in western Kansas and 7 found new oil pools in eastern Kansas. There were 289 rank wildcat tests, and among these 28 were successful ventures. They are responsible for 26 of the new oil pools recorded for the year and 2 of the new gas pools recorded. The ratio of dry holes to producing wells among such rank wildcats is thus approximately 10 to I (289 dry and 28 good). This compares with a ratio of II: I in 1942 and 6:1 in 1941; the same in 1940 and 5:Iin 1939. With nearly all wells flowing to their maximum economic capacity, the total production for Kansas during 1943 reached a new high; that is, 107,700,000 bbl., compared with 97,000,000 during the previous year. Similarly, the amount of gas sold during 1943 made a new high point on the production curve. The total amount was 122 billion cubic feet, compared with about 98 1/2 billion cubic feet for the year previous. As usual, the large Hugoton field of south- western Kansas accounted for approximately half of the total gas production. Highlights of the Year Despite the fact that 56 new pools were found in Kansas during 1943, it must be admitted that the future reserves of the state have not been increased greatly. Most of the new pools seem to be of minor value, although it is admittedly impossible to gauge the future production of any pool in Kansas with any exactitude. Judging by the later performance of some of the pools found in 1942 and in 1941, less than five of the new pools are likely to make any great contribution to the state's future oil reserves. Taking up the counties of Kansas in alphabetical order, the following statements summarize the highlights of the year. In Graham County, the Morel pool was extended to the southeast and this pool now promises to become one of the important pools of western Kansas. Several of the older wells have already contributed more than I000,000 bbl. of oil, and such a figure is considered very satisfactory in Kansas. In McPherson County, important new reserves were found by persistent drilling in township 17 S., R. I. W., the most northeastern township of the county. Many new wells in older pools and discovery wells in new pools make this region look promising. The Lindsborg pool, which had a sensational revival during 1942, was also extended in a number of directions and has now become a rather large producing area with a fairly high future potential capacity.

Jan 1, 1944