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By J. B. Gwiazda

This work contains a description of a new design of lemniscate-guided powered roof support units, with a discussion of currently designed supports and their interaction with the roof. Conventional roof supports of this type present a serious reaction with the roof because of the horizontal displacement of the canopy in the course of decreasing support height as the support takes weight. The new design presented eliminates previous disadvantages of lemniscate supports and ensures proper performance of the latter in relation to the roof of a mine working.

Jan 1, 1986

By Arthur F. Daily

8.5-1. Bucket Line Dredge. INTRODUCTION. The modern bucket-line dredge for mining placer deposits is a combined floating excavating machine and a gravity concentration system. Material is excavated at the forward end by the buckets, elevated, washed and screened, concentrated for recovery of valuable mineral, and discharged off the stern as coarse and small fractions. The end product is a concentrate consisting either of the valuable mineral(s) and associated heavy minerals, which are taken to a treatment plant on shore; or in the case of gold, a clean amalgam ready for retorting. The bucket-line dredge may be adapted for purposes other than mining for valuable mineral by installing, in lieu of a concentrating system, machinery for pumping all of the solids ashore in a pulp, or for producing washed gravel in selected size fractions of which all or a selected fraction is pumped to disposal areas on shore or into side barges. The two principal types of bucket-line dredges designed for mining are: 1) those having an inclined belt conveyor stacker to discharge the coarse fraction of washed tailing well astern of the hull, and sluices to discharge the sand fraction (Fig. 8.5-1), and 2) those having only sluices for discharge of all tailing astern. Use of the latter type has been confined, with few exceptions, to mining cassiterite-bearing deposits in Malaysia and at sea off Indonesia. Small dredges having only flumes for washing and concentrating all material dug by the buckets have been used for mining shallow stream valley deposits. GENERAL ARRANGEMENT. The general arrangement of a modern bucket-line dredge for mining tin, with a primary concentrating plant on board, is shown in Fig. 8.5-2. The hull and superstructure of most dredges are fabricated of structural steel; wood also has been used for both small and large dredges with complete satisfaction. All-riveted or partially welded unit hulls have been widely used and are desirable for exceptionally severe service. The hull

Jan 1, 1968

By Michael Epstein, Daniel E. Rosner

Fume nucleation sufficiently close to vaporizing suvfaces can augment net vaporization rates into cooler environments. Environmental conditions favoring large vaporization rate enhancements are briefly discussed and a previous theoretical treatment of this nucleation phenomenon is generalized to account for the self-regulating effect of condensalion-heat release within the boundary layer. Despite kinetic limitations on homogeneous nucleation, and latent heat release, non-diffusive condensate removal processes appear to make possible large enhancements in steady-state vaporization rates, provided surface temperatures are well below the boiling point. When condensed phases vaporize (or dissolve) into cooler media, the diffusion-limited mass loss rate can be strongly influenced by the process of nucleation/con-densation (or precipitation) within the thermal boundary layer. This condensation process (which typically leads to mists or fumes in the case of evaporation into cooler gases) has the effect of steepening the vapor pressure profiles near the evaporating surface, since the condensation zone acts as a vapor sink. However. the resulting enhancement in the diffusion-limited evaporation rate can be estimated (as first done by Turkdogan1 for the case of molten iron/nickel alloys evaporating into helium) only if one has independent knowledge of the critical supersaturation, sCrit(T), required to homogeneously nucleate the vapor.* In a recent reformulation and generalization of the theoretical model of Ref. 1 it has been shown that, when log sCrit is approximately linear in reciprocal temperature, rather simple expressions can be derived4 for the ratio of the actual rate of vaporization j" to either the minimum (no condensation) rate j"min, or the maximum (equilibrium condensation) rate j"max In the present communication we wish to briefly report on further developments and implications of the formulation of Ref. 4, with emphasis on i) environmental conditions favoring large enhancements in vaporization rate, and ii) the self-regulatory influence of condensation heat release (neglected in Refs. 1 to 4) on predicted vaporization rates. Additionally, we take this opportunity to correct several misprints appear- ing in Ref. 4, and comment on Elenbaas's recent criticism5 of Ref. 1. MAXIMUM POSSIBLE VAPORIZATION RATE IN PRESENCE OF CONDENSATION A nonequilibrium theory is of interest because of the very large difference between the minimum (no condensation) and maximum (equilibrium condensation) vaporization rate. The magnitude of this maximum possible enhancement can be shown quite clearly by combining a result of Refs. 3 and 4 with the fact that for most liquids there is a simple relationship between the molar heat of evaporation, A, and its boiling point, i.e., A/(RTBp) = C, where the constant C, often called the Trouton ratio, takes on values not very different from 11.* More generally, for any substance (including The Trouton ratio (which for water is 13, for methane, 10, and so forth) will be recognized as the ratio of the molar entropy change upon vaporization (at TBP or Ttransf) to the unlversal gas constant R. Its near constancy reflects the fact that the change in atomic order upon vaporization depends only weakly on the kinds of molecules involved. those that sublime under ordinary conditions) we can define a characteristic transformation temperature. Ttransf, by a relation of the form Ttransf =A/(CR), and then examine the maximum possible evaporation rates as a function of how far removed from Ttransf are the surface temperature, Tw, and ambient temperature, T. Subject to the assumptions: 1) equilibrium vapor pressure, pv,eq, everywhere small compared to prevailing total pressure, p, and 2) negligible effect of condensation heat on temperature profile, the maximum enhancement ratio was found (Eq. [17], Ref. 4) to be: where, for most vapors, Nu/NuD (the ratio of heat transfer coefficient to mass transfer coefficient for the same configuration) is a number near unity.* Ex- *An alternative derivation of the Nu = NuD special case of this equation. revealing its validity for arbitrary velocity/temperature profiles in a laminar boundary layer, is given in Ref. 3. amining this result for a "Trouton substance", one obtains the results shown in Fig. 1, constructed for C = 11. Since we are concerned with vaporization enhancements (j'max/J"min > 1) at surface temperatures below Ttransf, this area of interest is shown unshaded. One notes that at a fixed ambient temperature (hence, T/TtranSf) there is a unique surface temperature, 2T , at which j"max/j"min attains its peak value; moreover, the peak enhancement ratio, see dashed locus. Fig. 1, is: (NuA/NuD)(C/4)(Ttransf/T,). Hence, if Nu = NuD, when the ambient temperature is less than 1/4 of TtranSf the peak enhancement exceeds the Trouton

Jan 1, 1969

Samuel T. Wellman, Cleveland pioneer steel man, who was often referred to as the "father of the open-hearth process of the United States," died suddenly on July 11, 1919, of heart disease, at Stratton, Me. For more than half a century, he was closely identified with the improvement of equipment used in the making of steel and contributed a number of inventions that virtually revolutionized the industry. Mr. Wellman was born Feb. 5, 1847, at Wareham, Mass., his father at the time being superintendent of the Nashua Iron Co., Nashua, N. H. After attending public school and a year at Norwich University, he served in the Civil War as a corporal of the first New Hampshire Heavy Artillery. He went to Pittsburgh in 1867 and assisted in starting the first crucible steel melting furnace built in America at the works of Anderson, Cook & Co. It melted a ton of steel with an average of 1000 Ib. of nut coal, which cost less than $1, while the melting of a ton of steel in crucibles in the old fashioned coke furnace took three tons of the very best coke, costing from $2 to $3 per ton. This furnace was a great success and in a very few years had driven the coke furnace out of use. From Anderson, Cook & Co., Mr. Wellman went to the Singer, Nimick & Co. works, Pittsburgh, where he built two crucible steel melting furnaces of the same type. After that he spent some time in the office of the Siemens agents in Boston and also at steel works in different parts of the country, starting crucible steel furnaces. He then went to the Bay State Iron Works in South Boston, Mass., where he built the first open-hearth furnace that was a commercial success in the United States. It made steel of a much better quality than had previously been made in this country. The principal use to which it was put in those days was the manufacture of locomotive fire-boxes. From South Boston, he went hack to the old works in New Hampshire where his father was still superintendent, and built for it an open-hearth furnace, a plate mill, and a bar mill. In 1873, Mr. Wellman went to Cleveland to design and build the plant of the Otis Steel Co. He remained with this company for 16 years as engineer and superintendent. During this time he developed the open-hearth charging machine and introduced the electromagnet for handling pig iron and scrap steel. The first charging machine was operated by hydraulic power, and although it was not entirely satisfactory it performed the work fairly well and encouraged its inventor to build another one, which was driven by electric motors. This proved

Jan 1, 1920

By W. J. McKenna

The International Smelting Co. concentrator at Tooele, Utah, first operated on a custom basis for the treatment of lead-zinc-iron ores on Nov. 1, 1924, with a capacity of 500 tons per day. On May 1, 1925, the capacity was increased to 1000 tons per day by the addition of another 500-ton section. The concentrator has since operated on the two-section basis, each treating 500 tons per day entirely by flotation. Laboratory and Test Mill In connection with the concentrator, a well-equipped laboratory and small continuous testing plant are maintained. The flotation equipment in the test plant is of both the Callow and Mineral Separations Sub-"A" types, so that actual concentrator conditions may be duplicated as closely as possible. Purchasing of Ore A new shipper usually sends to the plant a 25 to 50-lb. samplc, representative of the ore to be shipped. Batch flotation tests are made on this sample to determine its amenability to the concentrator treatment flowsheet and reagent combinations. The results of these tests, which are accurately correlated with those obtained at the concentrator, both as to grades of products and recoveries, determine the actual value of the ore as submitted in the ore purchasing department's treatment schedule. Hearty cooperation has existed at all times between the ore purchasing and operating departments in the selection of ores as conccntrator feed. Ores Treated The basic tonnages for mixtures on the two sections come from the Park City and Bingham mining districts. The Park City tonnages arc received by rail and consist of crude ores and gravity mill products. Those from the Bingham district arc received by rail and aerial tramway,

Jan 1, 1928

By G. St. J. Perrott

ONE of the important hazards in coal mining is the danger of ignition of explosive mixtures of methane and air or coal dust and air, or both, by the explosives used in blasting the coal. It has long been recognized that certain explosives are more dangerous than others in this respect. In America and abroad, government testing stations have been established for officially passing on the suitability of explosives for use in coal mines. In this country the Federal government is given no power to enforce the use of the approved explosives but merely publishes from time to time, through the Bureau of Mines, those explosives which have passed the tests necessary to place them on the list for use in coal mines. In spite of this lack of police power, the increase in the use of permissible explosives since the promulgation of the first list in 1909 has been rapid. In 1912, only 8 per cent. of the explosives used in coal mines were permissible; in 1925, this figure had increased to 25 per cent. While the past record of permissible explosives has been most gratifying, it is very desirable that we pave the way for further improvement in explosives and blasting methods by learning more about the mechanism of ignition of gas and coal dust by explosives. To this end a considerable amount of investigative work already has been carried out and a large amount yet remains to be done. It is the purpose of this paper to discuss briefly the effect of the characteristics of explosives and method of loading in the borehole on the liability of a blown-out shot to ignite explosive mixtures of gas or dust. For an account of the methods by which an explosive is tested for permissibility for use in coal mines, the reader is referred to Schedule 17-A of the U. S. Bureau of Mines.

Jan 10, 1926

By H. F. Yancy

PooR results in coal washing and in ore concentration are sometimes attributed to the shape of the particles in the feed. It is well known that the shape of a particle influences its rate of fall in water. This is evidenced by the value of the constant C in the Rittinger equation for the velocity V of settling in water, V = C-D(d- z ), where C has the following values: 2.73 for roundish grains, 2.37 for long grains and 1.92 for flat grains. This report, which is based upon data obtained during the course of coal-washing investigations conducted by the U. S. Bureau of Mines in cooperation with the College of Mines of the University of Washington, gives the results of a method or procedure of screening in which the particles are separated into three shapes and shows the deportment of these shapes when coal is treated on a coal-washing table. The method consisted in first screening the coal with Tyler standard square-hole sieves, followed by screening tests of each of these products with screens of rectangular opening. For this purpose Tyler "Ton-Cap" screens were used; each square-hole product was subsequently sized with two Ton-Cap sieves so selected that their shortest dimensions were approximately 75 and 50 per cent., respectively, of the opening of the square-mesh sieves. For example: A product which is through 8 mesh (0.093 in.) and on 10 mesh (0.065 in.) is screened on slotted sieves which have openings 75 and 50 per cent. of the 0.065-in. opening in one direction and an opening of not less than 0.065 in. in the other direction. For convenience in discussion, the particles retained on the first slotted screen are termed "cubical"; those retained on the second slotted screen are termed "prismatic" and those passing the second screen are called "flat." (See Fig. 1.) The square-mesh screens separate particles according to two dimensions, length and breadth; the slotted screens divide them further according to the third dimension, or thickness. Prismatic particles have a mass of about two-thirds, and flat particles of about one-half, that of cubical particles of the same square-mesh size. The screen-

Jan 1, 1931

By B. Fagerberg, P. H. Fahlstrom

Crushing and grinding practice in Sweden follows closely the international pattern. Certain special circumstances, however, have had an influence on its development. Most mines are worked underground. These mines are either quite small or very large. Sulfide ores are commonly treated by custom milling and, apart from the cement industry, the rock species treated are hard and abrasive. Another important factor is the increasingly strict requirements placed on the mechanical condition of different products, especially lump ores and coarse iron-ore concentrates. Increasing requirements are being placed on the concentrate grade to suit the new steel processes. As a result, iron ore beneficiation has expanded considerably for the production of fine pelletizing concentrates. This applies especially to magnetite ores in the intermediate phosphorus range 0.1-1% P. In 1969, about 5.5 million tons of pellets will be manufactured. It should be pointed out, however, that fairly large quantities of coarse products are still being made at mills where the right conditions are available. Crushing and grinding is also of great importance in the production of rich concentrates at high recovery from sulfide ores. In fact, a large part of the development work on the concentration of sulfide ores has been devoted to the crushing and grinding processes. Boliden has played an important role in this connection.

Jan 10, 1969

By John W. Cabot

THE following series of analyses, which may, be of interest to Bessemer-steel makers, is submitted as a contribution to the much-vexed discussion of manganese in Bessemer steel. They were made by the writer in the spring of 1881, at the works of the Pittsburgh Bessemer Steel Company Limited, during a period of about eight days, part of, which time the rail-steel showed great variation in rolling qualities. During the first part of the time in question, it rolled well; later on it began to roll worse until, finally, it cracked very badly, reaching its worst stage. After this it rapidly improved until it again rolled smoothly. During this .time of transition 'from. good rolling to bad, and back to good again, the course of the manganese was followed' quite closely. It is the striking correspondence between these varying manganese contents and the differences in rolling which is here presented as worthy of notice. The other ordinary elements were determined occasionally, but showed no unusual variation Tim steel, in question, was ordinary rail-steel of about "the following composition : C=.35.P=.08.Mn=.98. Si=.08. S=.05. The samples for manganese determinations were taken at first from an average of two days' work, and later on, when the steel

Jan 1, 1882

By W. J. McKenna

THE International Smelting Co. concentrator at Tooele, Utah, first operated on a custom basis for the treatment of lead-zinc-iron ores on Nov.' 1, 1924, with a capacity of 500 tons per day. On May 1, 1925, the capacity was increased to 1000 tons per day by the addition of another 500-ton section. The concentrator has since operated on the two-section basis, each treating 500 tons per day entirely by flotation. LABORATORY AND TEST MILL In connection with the concentrator, a well-equipped laboratory and small continuous testing plant are maintained. The flotation equipment in the test plant is of both the Callow and Mineral Separations Sub-"A" types, so that actual concentrator conditions may be duplicated as closely as possible. PURCHASING OF ORE A new shipper usually sends to the plant a 25 to 50-lb. sample, representative of the ore to be shipped. Batch flotation tests are made on this sample to determine its amenability to the concentrator treatment flowsheet and reagent combinations. The results of these tests, which are accurately correlated with those obtained at the concentrator, both as to grades of products and recoveries, determine the actual value of the ore as submitted in the ore purchasing department's treatment schedule. Hearty cooperation has existed at all times between the ore purchasing and operating departments in the selection of ores as concentrator feed. ORE TREATED The basic tonnages for mixtures on the two sections come from the Park City and Bingham mining districts. The Park City tonnages are received by rail and consist of crude ores and gravity mill products. Those from the Bingham district are received by rail and aerial tramway,

Jan 1, 1928

By N. J. Themelis, A. W. Last

Autogenous grinding, broadly defined as the self-comminution of ore in a cylindrical tumbling mill without using auxiliary grinding media, was first applied in the early 1930s in the so-called Hadsel mills (Hardinge, 1955) as seen in [Fig. 1], mainly used in gold mines on the US West Coast. A number of these mills were built in sizes of 6.1 to 7.3 m (20 to 24 ft) in diameter and 1.5 to 1.8 m (5 to 6 ft) in length but were later abandoned because of operating and maintenance problems. Further development of autogenous grinding led to the Hardinge wet grinding mill, the Aerofall dry grinding mill, and others. A tabulation of all known autogenous and semiautogenous installations around the world was sent recently to the authors by G. J. Kushner of Koppers Metal Processing Division. In the size range of 7.3 to 9.8 m (24 to 32 ft) in diameter, it lists a total of thirty mills used in the processing of copper ores. An additional 120 units are listed in the same size range for the processing of iron ores and other minerals. The largest mills in existence are 11.1-m-diam X 4.6-m-long (36-ft-diam X 15-ft-long) mills in Minnesota, powered by 8950 kW (12,000 hp) motors and used in taconite beneficiation. It is indicative of the phenomenal growth of this technology that none of the above 150 mills was in operation prior to 1964. Grinding rates in autogenous mills are very sensitive to variations in ore breakage characteristics and even with fairly constant ore type can exhibit large fluctuations in throughput. Autogenous grinding also depends on availability of a fraction of the ore to be ground in the form of "competent" rock which can withstand to a certain degree the tumbling and impact action within the mill. If the ore is uniformly "soft", it will soon be crushed into small size particles and further grinding, which depends on the existence of large size rocks in the mill, becomes impossible. This apparently is the case with most copper ores, as discussed by Bachman, Last and Nabbs (1970) who described pilot testing of a large number of ore samples from Kennecott 's Chino Mines. Under circumstances where autogenous grinding is impractical, it is possible to realize most of its benefits by utilizing large grinding balls in place of the "competent" rock and practice what has become known in the mining industry as semiautogenous grinding. In semiautogenous grinding (SAG), a grate discharge tumbling mill is fed the product of a primary crusher or run-of-mine rock. Forged steel or alloy grinding balls, in proportions ranging from 3.10% of the total volume in the mill, are used to complement the self-grinding ability of the ore. Semiautogenous circuits can be either designed for single-stage closed-circuit grinding which produces a final product for recovery processes, or for two-stage operation in which the SAG mill product is fed to a second stage of grinding in a conventional ball mill.

Jan 1, 1981

By Allen D. Gorrison

EFFORT to develop fundamental quantitative information and improved technique in the production of petroleum has long been faced with difficulties of a particularly evasive nature, owing to a combination of circumstances. In the first place, petroleum comes from far below the ground, out of sight, so that information derived from indirect observations must be extrapolated to reservoir conditions. If one does not force the analogy too far, the petroleum engineer may be said to be somewhat in the position of the astronomer. The actual phenomena which he wishes to observe in detail appear at a place so deep in the earth?s crust that personal observations and direct experiments are impossible.

Jan 1, 1942

By T. Koulomzine

The theory of the Askania mag-netometer, as well as a complete discussion of all factors influencing magnetometer readings, is very ably described by J. Wallace Joyce.1 We will assume that the reader is thoroughly familiar with this theory and we will use whenever possible the same notations for the physical values involved in the theory. On the other hand, we believe that the principal formulas of the magnetometer can be deduced much more easily by the use of differential calculus than by the methods employed by Joyce. Consequently, we are presenting herewith a resumé of the theory of the vertical magnetometer in order to determine the correct formula of the temperature coefficient which we propose to compensate. Short Theory of the Vertical Magnetometers We shall accept the following notations: m = total mass of the magnet system, in grams. g = acceleration of gravity, in dynes. a = projection on the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. h = projection on the perpendicular to the axis of the magnet system of the distance between the axis of support and the center of gravity of the magnet system, in centimeters. M = 2pL = M, magnetic moment of the magnet system. Z = vertical component of the earth's magnetic field. ? = deflection angle of the magnet system. = focal length of objective lens in the optical system S = scale reading. In a vertical magnetometer the axis of the magnet system is kept horizontal within 2". Therefore, referring to Fig 1, we see that the magnetic torque of the magnet system influenced by Z must be counterbalanced by the gravity torque. MZ cos ? = mga cos ? + mgh sin ? or dividing by cos ? MZ = mga + mgh tan ? If we now move to another magnetic station and the temperature conditions change, the values of all the variables will change also: M, a, and h under the influence of temperature. 2, both with time and space. ? and S, as a result of all the above changes. Taking increments in Eq 2 we obtain: (Mo + ?M)(Z + ?Z) = mgao + mg?a + mg(h + ?h) tan (? +??) but tan (? +??), because ? is small, can be replaced by tan ? + tan ??; in fact tan (?+ ??) = sin ?.cos ?? + cos ?.sin ??/ cos ?.cos ?? — sin ?sin ?? but sin ? sin ???0 tan (8 + ??) ? tan ? + tan ??. Fur-thermore, neglecting all terms of the equation containing two differentials becomes: or a differential and a tan ??, Eq 3

Jan 1, 1950

By A. T. Beckwith

THE Lehigh Navigation Coal Company Inc. operates steep-pitch, relatively deep mines in the Panther Creek Valley, at the eastern end of the southern anthracite coal field. Commercially minable coal beds are the Mammoth,† Primrose, Orchard, and Buck Mountain,‡ their respective thicknesses average 50, 15, 1o and 12 ft. Sand slates immediately overlie and underlie these beds, though in some locations sandstone or conglomerate beds occur. Underground operations extend to a depth of 1200 ft. below the surface, and comprise a bulk of virgin work on the lower levels, and a small amount of remining on the upper levels. With increase in depth, ventilation has become a more important factor. DEVELOPMENT In early days slopes or drifts were driven in the outcropping coal on the hillsides. Subsequent development was by water-level tunnels, with portals above high water in Panther Creek. These tunnels have been maintained because they intercept and drain much water that enters the mined-out portion of the bed and otherwise would follow down to the deep workings. Below these tunnels, the coal was customarily developed by slopes. Muleways, which are zigzag chutes easy to travel, were also driven to connect lower openings with the water levels. When it became necessary to go still deeper, the slopes were abandoned, and vertical multicompartment shafts were sunk from the surface. Levels have been at vertical intervals of from 195 to 250 ft. Recent practice has been toward the latter spacing; splitting the total lift, however, by starting to mine from chutes extending halfway up to the level above. Chutes customarily are spaced at 50, 60 or 70-ft. centers along the gangways. At some mines, additional shafts have been sunk from the surface to the level containing the main pumping station, and these carry the discharge lines from the pumps. As the mines are ventilated by the exhausting system, both the hoisting and so-called water shafts serve as intakes. The main fan is always on the surface, and is connected with the return airways from the various levels by an air hole or a vertical shaft. For economy, when the present deepest levels were developed, main return-air shafts or air holes were not extended to the new level, but connections were made from the new level to that above by two or more holes. It is doubtful that this method will be continued, as complications in the system, increased resistance to flow, and other difficulties undoubtedly will counterbalance any economies this provision may afford.

Jan 1, 1942

By T. E. Ockershauser

The Southwest Antioch Oil Field located in T2&3N - R2&3W, Garvin County, Okla., was discovered in February, 1946, by The Globe Oil & Refining Co. and The Vickers Petroleum Co. at their Melinda Gibson No. 1. Oil production was encountered at a depth of 6,525 ft in 14 ft of lower Pennsyl-vanian sandstone. The producing sand, locally called the ."Gibson" or "Third Deese," is a sand bar deposit strung along the west flank of the Pauls Valley uplift in a northwest-southeast direction. The oil reservoir is defined to the east by the edge of sand deposition along the uplift and to the west by the gradation of sand into shale. Within the Antioch Unit, the sand dips from 150 to 400 ft per mile southwesterly and has a pay thickness varying from 10 to 55 ft and averaging 28 ft. Several areas of Gibson production have been developed, including the Katie pool on the south, the Elmore Area of the Southwest Antioch Gibson Sand pool. the Southwest Antioch Gibson Sand pool proper, the Southwest Maysville pool and the New Hope pool on the north. Production extends over a belt 22 miles long by about two miles wide and is continuous except into the Katie and New Hope pools. It appears that drilling will soon prove these intervening areas productive. To March 1, 1950, approximately 18,000 acres have been developed with 450 Gibson wells. It is likely this count will ultimately reach 32.000 acres with 800 wells. RESERVOIR The natural production of oil from the Gibson reservoir is primarily due to solution gas expansion though gravity also plays a part. There is no indication of water drive or evidence that an initial gas cap existed. Gibson oil has a stock tank gravity of 43" API and was saturated at the original reservoir pressure of 2.925 psi. It contained 1,3440 cu ft of gas per bbl in solution and had a high shrinkage of 70 per cent. Connate water amounts to 15 per cent. Reservoir temperature is 130°F. Porosity averages 15.5 per cent and permeability ranges from less than one md to over 1.000. Based on meacurement of 400 core plugs from 15 wells, 78 per cent of the permeabilities fell within the 10 to 1,000 md bracket and averaged l67. Primary oil recovery is estimated at 140 bbl per acre-ft or 22 million bbl for the present 137-tract unit. Due to unitization

Jan 1, 1951

By T. E. Ockershauser

The Southwest Antioch Oil Field located in T2&3N - R2&3W, Garvin County, Okla., was discovered in February, 1946, by The Globe Oil & Refining Co. and The Vickers Petroleum Co. at their Melinda Gibson No. 1. Oil production was encountered at a depth of 6,525 ft in 14 ft of lower Pennsyl-vanian sandstone. The producing sand, locally called the ."Gibson" or "Third Deese," is a sand bar deposit strung along the west flank of the Pauls Valley uplift in a northwest-southeast direction. The oil reservoir is defined to the east by the edge of sand deposition along the uplift and to the west by the gradation of sand into shale. Within the Antioch Unit, the sand dips from 150 to 400 ft per mile southwesterly and has a pay thickness varying from 10 to 55 ft and averaging 28 ft. Several areas of Gibson production have been developed, including the Katie pool on the south, the Elmore Area of the Southwest Antioch Gibson Sand pool. the Southwest Antioch Gibson Sand pool proper, the Southwest Maysville pool and the New Hope pool on the north. Production extends over a belt 22 miles long by about two miles wide and is continuous except into the Katie and New Hope pools. It appears that drilling will soon prove these intervening areas productive. To March 1, 1950, approximately 18,000 acres have been developed with 450 Gibson wells. It is likely this count will ultimately reach 32.000 acres with 800 wells. RESERVOIR The natural production of oil from the Gibson reservoir is primarily due to solution gas expansion though gravity also plays a part. There is no indication of water drive or evidence that an initial gas cap existed. Gibson oil has a stock tank gravity of 43" API and was saturated at the original reservoir pressure of 2.925 psi. It contained 1,3440 cu ft of gas per bbl in solution and had a high shrinkage of 70 per cent. Connate water amounts to 15 per cent. Reservoir temperature is 130°F. Porosity averages 15.5 per cent and permeability ranges from less than one md to over 1.000. Based on meacurement of 400 core plugs from 15 wells, 78 per cent of the permeabilities fell within the 10 to 1,000 md bracket and averaged l67. Primary oil recovery is estimated at 140 bbl per acre-ft or 22 million bbl for the present 137-tract unit. Due to unitization

Jan 1, 1951

By T. E. Tietz, J. E. Dorn

Activation energies for creep of copper at intermediate temperatures, where crystal recovery was negligible, were determined by the simple technique of rapidly alternating the test temperature between and T2 (T2 = TI + about 10°K) throughout a constant stress creep test. The activation energy for creep H was found to be 37,000± 3,000 cal per mol, independent of stress and strain. The same creep laws as have been previously established for high temperature creep were found to be valid for creep at intermediate temperatures. But the AH was found to be lower than that for self-diffusion in the intermediate temperature range whereas it is known to be equal to that for self-diffusion at high temperatures. IT is now well established that creep at high temperatures can be correlated by the functional relationship'-" e = f(?) s = constant [I] where E is the total plastic strain, f is the function that depends on the stress, ? = f e -- dt, t is the duration of test, e is the base of natural logarithms, AH is the activation energy for creep, R is the gas constant,T is the absolute temperature, and s is the stress. Thus, identical values of are achieved during creep under a given stress at two alternate temperatures at the same total strain. At the same strain therefore where subscripts refer to the two alternate temperature conditions of test. Consequently, the activation energy for high temperature creep is readily obtained with the aid of Eq. 2. Activation energies so obtained have been observed to be insensitive to the creep strain at which they are evaluated, the applied stress, grain size, minor alloying additions,'-' cold worked state, and the presence of thermally stable dispersions of intermetallic compounds in an a solid solution matrix Furthermore, the activation energy for high temperature creep is very nearly equal to that for self-diffusion in all cases where adequate data are available for comparison." " The apparent identity of the activation energies for high temperature creep and self-diffusion strongly suggest that high temperature creep arises from a dislocation climb process. This hypothesis is further strengthened by the fact that the energy of activation for high temperature creep is insensitive to the applied stress," and by the observations that the correlation suggested by Eq. 1 cannot be ex- tended to temperatures below those for rapid crystal recovery. When using the activation energy for self-diffusion, Eq. 1 fails to give the required correlations between creep curves at various intermediate temperatures.' This suggests that the dislocation climb mechanism for creep might be superseded by some alternate mechanism in the intermediate range of temperatures. This investigation was initiated in a preliminary attempt to ascertain the type of creep law that might be valid in the intermediate range of temperatures. Special consideration, however, was given to the determination of the activation energy for creep in this range because the Becker-Orowan', " and the Mott-Nabarro10 theories for transient creep suggest that the activation energy should increase

Jan 1, 1957

By Reuben B. Ellestad, Fremont F. Clarke

In the early days of the lithium industry most of the production was from lepidolite, zinnwaldite, and amblygonite. Nearly all the early extraction processes described in the literature involve heating the finely ground mineral with sulphuric or hydrochloric acid. On subsequent water leaching most of the bases present in the mineral (especially aluminum) are dissolved as sulphates. As a result, the leach solution required extensive chemical purification before the lithium could be precipitated as carbonate. Following the remarkable growth of the lithium industry to its present size, zinnwaldite and amblygonite ores must be considered of minor importance only. Attention is now focused on spodumene, abundant enough in North America to be a major source of supply, and there are important supplies of lepidolite and petalite in Africa. The extraction processes described below all apply to spodumene, although several will also operate on other lithium minerals, such as petalite. Base Exchange with Alkali Sulphates: A distinct advance was made with the disclosures of Wadman and von Girsewalt. In these methods the finely ground silicate ore (spodumene or lepidolite) is intimately mixed with an excess of alkali sulphate (usually K2S04) in at least a 1 to 1 proportion, and the mixture was heated to a relatively high temperature. Base exchange results, with the formation of lithium sulphate. A water leach dissolves the lithium sulphate, together with the excess potassium sulphate. Successful operation of this type of process requires very thorough grinding and mixing, as well as careful temperature control. The use of K,SO, is objectionable from cost considerations since purification of lithium carbonate re- quires the use of potassium carbonate, if the K2S04 is to be recovered and recycled. The lower solubility of K,SO, as compared with Na,SO, is also objectionable, since it limits the concentration of the Li,SO, solution to be precipitated by K,CO,. Early laboratory-scale investigation of this process by Lithium Corp. was not encouraging

Nov 1, 1955

By W. Weissberger, Frankiewicz, L. Pommier

Introduction In the U.S., about one-quarter of the fuel oil and natural gas consumption is associated with power production in utility and industrial boilers and process heat needs in industrial furnaces. Coal has been an attractive candidate for replacing these premium fuels because of its low cost, but there are penalties associated with the solid fuel form. In many cases pulverized coal in unacceptable as a premium fuel replacement because of the extensive cost of retrofitting an existing boiler designed to burn oil or gas. In the cases of synthetic fuels from coal, research and development still have a long way to go and costs are very high. Another option, which appears very attractive, is the use of solid coal in a liquid fuel form - coal slurry fuels. Occidental Research Corp. has been developing coal slurry fuels in conjunction with Island Creek Coal (ICC), a wholly-owned subsidiary. Both coal-oil mixtures and coalwater mixtures are under development. ICC is a large eastern coal producer, engaged in the production and marketing of bituminous coal, both utility steam and high quality metallurgical coals. There are a number of incentives for potential users of coal slurry fuels and in particular for coal-water mixtures (CWMs). First, CWM represents an assured supply of fuel at a price predictable into future years. Second, CWM is available in the near term; there are no substantial advances in technology needed to provide coal slurry fuels commercially. Third, there is minimal new equipment required to accommodate CWM in the end-user's facility. Fourth, CWM is nearly as convenient to handle, store, and combust as is fuel oil. Several variants of CWM technology could be developed for different end-users in the future. One concept is to formulate slurry at the mine mouth in association with an integrated beneficiation process. This slurry fuel may be delivered to the end-user by any number of known conveyances such as barge, tank truck, and rail. Slurry fuel would then be stored on-site and used on demand in utility boilers, industrial boilers, and potentially for process heat needs or residential and commercial heating. An alternative approach is to formulate a low viscosity pre-slurry at the mine mouth and to pipeline it for a considerable distance, finishing up slurry formulation near the end-user's plant. Finally, at the other extreme of manufacturing alternatives, washed coal would be shipped to a CWM manufacturing plant just outside the end-user's gate. Depending on fuel specifications and locations of the mine and end-user facility, any of these alternatives may make economic sense. They are all achievable in the near term using existing technology or variants thereof. The Coal-Water Mixture CWMs contain a nominal 70 wt. % coal ground somewhat finer than the standard pulverized ("utility grind") coal grind suspended in water; a complex chemical additive system gives the desired CWM properties, making the suspension pumpable and preventing sedimentation and hardening over time. Figure 1 shows the difference between a sample of pulverized coal containing 30 wt. % moisture and a CWM of identical coal/water ratio. The coal sample behaves like sticky coal, while the CWM flows readily. The combustion energy of a CWM is 96-97% of that associated with the coal present, due to the penalty for vaporizing water in the CWM. Potentially any coal can be incorporated in the CWM, depending on the combustion performance required and the allowable cost. CWMs are usually formulated using high quality steam coals containing around 6% ash, 34% volatile matter, 0.8% sulfur, 1500°C (2730°F) initial deformation temperatures, and energy content of 25 GJ/t (21.5 million Btu per st). Additional beneficiation to the 3% ash level can be accomplished in an integrated process. There are a number of minimum requirements which a satisfactory CWM must meet: pumpability, stability, combustibility, and affordability. In addition, a CWM should be: resistant to extended shear, generally applicable to a wide variety of coals, forgiving/flexible, and compatible with the least expensive processing. It was found that a complex chemical additive package and control of particle size distribution are necessary to achieve these attributes simultaneously, while maximizing coal content in the slurry fuel. Formulation of Coal-Water Mixtures A major consideration in the manufacture, transportation, and utilization of a slurry fuel is its pumpability, or effective viscosity. Most CWM formulations are nonNewtonian, i.e., viscosity depends on the rate and/or duration of shear applied. Viscosities reported in this paper were obtained using a Brookfield viscometer fitted with a T-spindel and rotated at 30 rev/min, thus they are apparent viscosities measured at a shear rate of approximately 10 sec-1. The instrument does reproducibly generate a shear field if spindle size and rotation rate are held fixed. By observing the apparent viscosities of several slurries at fixed conditions it is possible to obtain a relative measure of their viscosities for comparison purposes. A true shear stress-shear rate relationship at the shear rates at which the CWM will be subjected in industry may be obtained using the Haake type and a capillary viscometer. These viscometers are used for specific applications. However, for comparison purposes, apparent viscosities are reported.

Jan 1, 1985

By R. H. Goodrich

Development in this district during the year 1929 was little different from that of any other year despite the somewhat depressed condition of the oil business in general. The year was marked by: (1) An intensive geophysical campaign in the search of deep-seated salt domes; (2) the rather successful exploration and development of lateral sand production on the flanks of some of the older domes. Geophysical Exploration Geophysical work is confined almost exclusively to the torsion balance and seismograph. At the beginning of the year there were approximately 185 individual instruments of all types engaged in the search of structures in the Gulf Coast of Texas and Louisiana. This total includes 3 magnetometers, 83 torsion balances and 100 seismographs. The number of parties operating the torsion balances was 47; seismographs, 20; magnetometers, 3. It is estimated that 4,800,000 lb. of dynamite or other explosives were used in seismic exploration. During the year the torsion balance and seismograph were credited with having discovered the following salt domes: Texas, Danbury dome, Brazoria County, H. & T. B. R. Co., and Clodine dome, Fort Bend County, John Fredricks and Wm. Stanley, Satowsky Survey; Louisiana, Iowa dome, Calcasieu and Jefferson Davis Parishes, T. 9 S., R. 6 and 7 W., and Cameron-Meadow dome, Cameron Parish, T. 14 S., R. 13 W. Because of the secrecy maintained in these developments, it is somewhat hazardous to venture opinion as to the number of salt domes found during one year. The discovery of gas at Mykawa district and at Genoa district, in Harris County, Texas, is considered of more than passing importance, because there is indication in the performance of these two districts that commercial accumulation of oil and gas may be found in the Gulf Coast on structures other than the wore or less definite salt dome type. While it is generally conceded that the accompanying structure in these instances are deeply buried salt domes, it is a fact that neither the seismograph nor the drill have indicated the presence of salt.

Jan 1, 1930