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By Robert S. Williams

On several occasions, when 60-40 brass is first obtained in the beta condition by quenching at about 825" C. and is then reheated, the writers have noticed that reerystallization will take place in the form of twin crystals at the grain boundaries of the original beta crystals. It was, therefore, thought that interesting results would be obtained if this twinning process was studied further, especially as to the temperature at which it first occurs. The material used in the investigation was a bar of extruded muntz metal, 1/2 in. in diameter, and of the following composition: Copper, 60.21 per cent.; zinc, 39.73 per cent.; lead, trace; iron, trace; tin, nil. Specimens 1 in. in length were sawed from this bar and a hole was drilled into the side of each specimen until it reached the center, and the tip of a platinum-platinum rhodium thermocouple was inserted. A small Hoskins' resistance-type electric furnace was used for heating the specimens. The specimens were first heated at 840' C. for 2 hr. and then quenched ' in cold water, in order to get the material entirely in the beta state. In every case, in the reheating process, the furnace was brought to the desired temperature before the specimen was introduced. The time occupied in preheating was from 20 to 30 min. By sawing back 1/4 in. on each side of the hole, two parts were obtained from each specimen and duplicate observations were possible. The specimens were held at the reheating temperatures for periods of 15 min., 30 min., 1, 2, and 4 hr. All specimens were etched with a freshly prepared mixture of equal volumes of concentrated NH4IOH and H2O2.

Jan 1, 1924

By H. J. Wasson, E. B. Hopkins

Throughout 1928, production of oil from Venezuela steadily increased, and at the close of the year, the output was at the rate of nearly 400,000 bbl. a day. The total for the year was approximately 106,500,000 bbl., representing an increase of about 66 per cent. over the 63,000,000 bbl. produced in 1927. Venezuela now stands next to the United States as the second largest oil-producing country in the world; and there is abundant evidence to suggest that this relationship will endure for a number of years. Reserves were notably increased in the Lake shore fields, and the potential output of the Maracaibo Basin is much greater than the current production. In spite of this, the percentage of production increase was comparatively moderate, and the rate of this increase continued to fall off slightly, as has been the case for the last three years. This does not suggest that further annual increases in Venezuelan output will not materialize; only that the rate of these increases appears to be on the decline, and that with every succeeding year the future tendency will be in the direction of a stabilized level of output. The following table shows the production trend of the country since 1925: 1925 increase over 1924, 132 per cent. 1926 increase over 1925, 80 per cent. 1927 increase over 1926, 75 per cent. 1928 increase over 1927, 66 per cent. Many factors will work to prevent a reversal of this trend, but the principal ones are strong ownership in large-acreage units, and the probable indisposition of the world market to absorb a more rapidly increasing Venezuelan output at prices profitable to the producers. The distress oil element is not a factor in Venezuela; also, any unforeseen jump, such as results in the United States from the discovery of a Seminole or a deep sand at Long Beach, or Santa Fe Springs, is impossible in Venezuela, on account of lack of means for transporting a sudden surplus of oil, and the slowness with which additional transportation equipment in the form of shallow-draft tankers can be obtained.

Jan 1, 1929

By N. Schapiro, R. J. Gray, G. D. Coe

A petrographic and chemical investigation was conducted to determine the occurrence and distribution of the organic and inorganic forms of sulfur in various size and specific gravity fractions of a high volatile, A-rank, Pittsburgh seam coal. The investigation was conducted to aid in the resolution of the problem of how metallurgical coke of lower sulfur content can be produced from high sulfur coals. The Pittsburgh raw coal contains about 2.4% S in two dominant forms: pyritic and organic. Pyritic sulfur constitutes about 57% of the total sulfur in the raw coal and occurs in macroscopic and microscopic-size particles. A plant cleaning process mechanically separates between 50 and 60% of the pyritic sulfur, mainly the macroscopic particles. The organic sulfur constitutes 41% of the total sulfur in the raw coal and 55% of the sulfur in the washed coal. Thus, organic sulfur is not reduced by physical methods of separation. Organic sulfur occurs in all the coal entities, but in general, increases as the exinoid content of the coal increases. In this study, it was found that carbonizing certain selected low gravity coal fractions containing an average of 0.45% pyritic sulfur and 1.01% organic sulfur produced cokes containing an average of 0.22% pyritic sulfur and 0.91% organic sulfur. However, the organic sulfur in the coke is probably derived in part from the heating of pyrite in the presence of carbon. Thus, the total sulfur in the coal carbonized is the most important factor affecting the sulfur content of the coke produced. Concentration of microscopic pyritic and organic sulfur in the middling-type coal indicates the necessity for rejecting as much of this material as is economically feasible. Most of the sulfur introduced into the blast furnace originates from the coke. Although the sulfur content of the hot metal can be controlled by the addition of flux, the flux used restricts the productivity of the furnace by displacing ore and requires first to be smelted; therefore, low sulfur coke is highly desirable. Since the sulfur content of coke is dependent upon the sulfur content of the coals from which the coke is produced, low sulfur coals are required to produce low sulfur coke. In the western Pennsylvania area, large quantities of good coking coal are available from the Pittsburgh seam, but this coal unfortunately contains relatively large quantities of sulfur. To provide information that might be useful in reducing the sulfur content of this coal in the cleaning process, an investigation was conducted to determine the physical and chemical forms in which this sulfur occurs and with what petrographic components it might be associated. The Pittsburgh seam coal, like most other coals, contains two dominant forms of sulfur: one occurs as the minerals pyrite and/or marcasite and is referred to as pyritic sulfur; the other is combined with the coal substance and is called organic sulfur. Pyritic sulfur is amenable to physical methods of separation, whereas organic sulfur is not. Pyritic sulfur particles must be freed from the surrounding coal if they are to be removed in the washing process. Generally, this requires crushing; the required degree of crushing being a function of the size and distribution of the pyrite. The present investigation was conducted to 1) determine the size, distribution and association of sulfur-bearing minerals in the coal and 2) to determine whether the organic sulfur has a preferential association with specific coal entities. If a preferential association exists, this knowledge may be useful in devising a method for organic sulfur reduction. In addition, a study was made to determine the extent to which pyritic and organic forms of sulfur in the coal contribute to the sulfur in the coke produced from this coal. A representative 3%-ton sample of raw coal was crushed to —3 in. and screened into square-mesh sizes: 3 to 1 in., 1 to % in., % to % in., % in. to 10 mesh, 10 to 35 mesh, and 35 to 200 mesh. Separate samples of the -200-mesh fraction and the all-size composite were also collected. Each size fraction (except the -200 mesh) was fractionated on heavy liquids of the following specific gravities: 1.280, 1.300, 1.325, 1.350, 1.375, 1.400, 1.425, 1.450,

Jan 1, 1963

By J. G. Morlet, A. R. Troiano, H. H. Johnson

Delayed failure in steel occurs by controlled initiation and growth of a crack. The incubation period for crack initiation was measured. Crack initiation and Propagation are controlled by interaction between hydrogen concentration and triaxial stress state. The incubation period results from stress-induced diffusion of hydrogen to the point of crack initiation. THIS paper describes the phenomenon of delayed failureland the kinetics of crack initiation and propagation in uniformly hydrogenated steel. The investigation was prompted by service failures of high-strength-steel components at applied stresses far below the conventional yield strength, after the part had successfully sustained equal or higher stresses for an extended time interval. This delayed failure phenomenon, often termed static fatigue, is associated1 with the presence of hydrogen. This is not surprising, since the voluminous literature on flaking and hydrogen embrit-tlement demonstrates that hydrogen may exert a potent embrittling influence. Flakes, or internal hairline cracks in alloy-steel forgings, are related to the presence of hydrogen retained from the steel-making process.2 On the other hand, embrit-tlement arises from hydrogen introduced during surface treatments such as acid pickling or electroplating. Laboratory delayed failures have been induced in notch specimens of 4340 steel by electrolytic charging of the specimen surface with hydrogen and subsequent static stressing.l,4,5 Static fatigue was observed over a wide range of applied stresses; however, failure did not occur below a critical applied stress, designated as the static fatigue limit. In this instance the extended time for failure resulted primarily from the macroscopic redistribution of hydrogen. A crack formed to the depth of the hydrogen enriched "case* upon loading; further crack growth naturally awaited diffusion of hydrogen into the material ahead of the crack. The crack-propagation rate was therefore controlled primarily by the macroscopic diffusion of hydrogen into the specimen interior during the test. Thus, in this instance static fatigue was associated with the movement of a gross concentration gradient. However, delayed failure in steel of uniform hydrogen concentration is of greater fundamental interest, as well as commercial significance. With a homogeneous hydrogen distribution a different mode of delayed failure might be expected. As the heterogeneous hydrogen distribution became more uniform,4,5 the initial crack-propagation rate decreased markedly and, in fact, what appeared to be an incubation period for crack initiation was observed. However, the macroscopic concentration gradient precluded an accurate evaluation of these observations. Although the existence of an incubation period, up to this time, has not been established definitely, it is not amiss to point out some implications of the concept. It is of interest in all hydrogen-induced embrittlement phenomena, such as flaking, delayed failure, and hydrogen embrittlement. Much discussion has centered around the possibility of an

Jan 1, 1959

By AIME AIME

THANKS to the excellent preliminary work of: the Division officers and the local committee the fall meeting of the Coal Division at Bluefields was a brilliant success. West Virginia was at its best with its beautiful mountains, good roads and foliage, just beginning to turn to autumn colors. The meeting began on time and was kept on time throughout. The papers were-well prepared and, being on subjects of live interest, provoked 'lively discussion. One factor in bringing this about was that they were not too numerous and that with one exception copies had been submitted in advance a sufficient time to permit of them having been studied. The Friday morning session was presided over by John .J. Lincoln of the Crozier Coal and Coke Co. and president of the Pocahontas Coal Operators Association. After brief announcements by Thomas H. Clagett, chairman of the local committee, regarding plans for the two day's involved, and by Howard N. Eavenson, chairman of the Division, the reading of papers was taken up.

Jan 1, 1931

By J. L. Huiatt, D. G. Foot

The Bureau of Mines, U.S. Department of the Interior, investigated methods of recovering potash values from process and waste brines. Laboratory pan evaporation of four chloride brines produced crude salts containing predominantly sylvite, halite and carnallite. Six sulfate- chloride brines produced crude salts containing primarily schoenite, kainite, leonite, sylvite, carnallite, and halite. Potash grades ranged from 7.2% to 22.2% K20, and recoveries by crystallization ranged from 85% to 99%. Sylvite flotation from chloride evaporites with amine collector, recovered 90% to 97% of the potash in a concentrate containing 54.3% to 60.3% K20. Fatty acid flotation of the high sulfate evaporite recovered 78% of the sulfate minerals in a 28%-K20 concentrate. Flotation of the remaining chloride minerals with amine collector recovered 80% of the potash in a 59.7%-K20 concentrate. Solar evaporation of 37 854 L (10,000 gal.) of brine recovered 99% of the potash in a crude evaporite containing 24.5% schoenite and 20% sylvite. Flotation in a 45-kglh (100-lblh) continuous flotation unit recovered over 95% of the potash in schoenite and sylvite concentrates, containing 28% and 62.3% K20, respectively. An economic evaluation is included.

Jan 1, 1985

By L. Thomassen, N. Spooner, J. M. Thomas

NI-CR and some Ni-Cr-Fc alloys, when used as electrical resistance heating elements in reducing atmospheres, at times suffer rapid breakdowns due to so-called "green rot." These reducing atmospheres are most frequently used in bright annealing and heat-treating furnaces which are kept for long periods of time at 1500' to 1800°F. The green rot is a preferential oxidation of the chromium in the alloy to such an extent that the remaining metal frequently becomes ferromagnetic. The Curie point (magnetic transition point) at room temperature for a pure Ni-Cr alloy is known to be at about 7 pct Cr.¹ This represents a severe loss in chromium for one popular grade of resistance alloy, which is nominally 80 pct Ni-20 pct Cr. The formation of oxides along the grain boundaries makes the ribbon brittle and gives the fracture a green, earthy appearance—hence the name. The authors have studied this phenomenon, simulating the industrial conditions by heating resistance ribbons in various reducing atmospheres, such as moist hydrogen or the atmosphere specified in the ASTM test.² This atmosphere contains 16 pct H2, 10 pct CO, 4 pct C02, 1 pct CH1, 1.5 to 2.5 pct H2O, balance N2. Little success was had in producing severe corrosion with the commercial 80 pct Ni-20 pct Cr alloys. It had been noticed, however, that 90 pct Ni-10 pct Cr alloy wires deteriorated very rapidly when placed in the bottom of narrow thermocouple protection tubes. The bottoms of the tubes with the wires were heated in air. This took place both in porcelain tubes and in Inconel* tubes either with the cold end open or loosely stoppered with asbestos. On examination the wires showed microstructures which would be classified as green-rot corrosion. Subsequent investigations with various alloy wires confirmed the observation that corrosion occurred in some wires, including 80 pct Ni-20 pct Cr alloys, much more rapidly in such tubes than had occurred previously in test atmospheres mentioned above. When 90 pct Ni-10 pct Cr alloys are oxidized with an abundant air supply, the oxide coat consists of both nickel and chromium oxides. In this case thermodynamic equilibrium between the oxides is not established. However, when oxidized 90 pct Ni-10 pct Cr alloy specimens were heated to 1500" to 1800°F inside small diameter tubes, the oxide coat transformed to a bright metallic outer layer of nickel or nickel-rich alloy, beneath which appeared a green oxide, followed by intergranularly attacked metal. The obvious explanation is that in the narrow tubes, thermodynamic equilibrium is being established according to the equation NiO + Cr (in alloy) ? Ni + Cr2O3 This reaction is favored due to the fact that the oxygen pressure over nickel oxide is more than 10"' times the oxygen pressure over the green chromium oxide. The initial oxygen in the tubes is depleted very rapidly by oxidation of the wire and of the tube, if it is metallic. A confirmation of the occurrence of the reaction, reduction of NiO by the chromium in the alloy, was obtained by putting oxidized nickel foil into an evacuated quartz tube along with a piece of bright 90 pct Ni-10 pct Cr alloy ribbon, and heating the tube at 1820°F for a number of hours. The nickel oxide on the foil was completely reduced to pure nickel, leaving a bright foil. The 90 pct Ni-10 pct Cr alloy became strongly magnetic and showed the typical green-rot structure. Weights of the samples before and after testing showed the weight loss of the foil to approximately equal the weight gain of the ribbon. Evidently the bottom of long, narrow protection tubes, stoppered or not, may under certain conditions act just as the sealed quartz tubes in respect to being a confined space in which oxygen is not freely replenished. This condition of oxygen depletion can be prevented by introducing a small amount of air into the bottom of the tube. Many conditions becloud the green-rot phenomenon, such as carburization, sulphidization, and the presence of other corrosive agents. However, these experiments have shown that the basic reaction is a case of internal oxidation which can occur simply by adjusting the oxygen pressure to a low enough level so that the atmosphere will leave the nickel intact and oxidize the chromium. The action of the other agents in commercial atmospheres can then be taken up as individual cases. From these and numerous other tests completed to date or still in progress, it is believed that a more critical evaluation of electrical resistance alloys for their resistance to green-rot corrosion can be made. A much better understanding of the mechanism for this type of corrosion should also result. A more complete report on these experiments will be published later. In the May 1953 issue: TP 35213 Discussion: Solubility of Carbon and Oxygen in Molybdenum by W. E. Few and G. K. Manning, discussion by N. A. Gokcen: p. 747, first column, the fifth, sixth, and seventh lines should be the first, second, and third lines, followed by the present first, second, third, fourth, eighth, ninth, and tenth lines.

Jan 1, 1954

By H. A. Lesser, H. L. Stone, G. H. Bruce

ESSO PRODUCTION RESEARCH CO., HOUSTON, TEX. A mathematical model that represents the conduction heating of a rock formation of limited permeability is formulated. Heat is introduced by the injection of a hot condensing gas into horizontal fractures through the formation. The fractures are assumed to be equally spaced; the flow of fluid is linear. The model consists of three coupled partial differential equations and two functional relationships. Temperature histories for both fractures and formation are obtained by numerical procedures. Application of the model in predicting in situ heating of oil shale is illustrated with four sample cases. Influences of shale thermal diffusivity, fluid pressure in the fracture and fracture spacing upon heating rate are demonstrated. In all cases considered, several years were required to heat all the formation to 700F with 1,000F steam. Less than 32 per cent of the injected heat was utilized for heating the shale. To facilitate extensive calculations of this type, a set of seven dimension-less quantities are derived. With this dimensionless system, the effect of controllable parameters on economics can be explored. INTRODUCTION Vast deposits of hydrocarbons are contained in oil shale formations in the western United States and in other countries. These hydrocarbons, known as kerogen, are difficult to recover because they are immobile solids and because the rock formations in which they are found have very limited permeability. The kerogen can be converted through pyrolysis to recoverable liquids and gases if the formation temperature is elevated to the range 600 to 800F. To accomplish this heating, the formation might be artificially fractured and a hot fluid injected into the fractures. If the injection and production wells are vertical and the fractures horizontal, interwell communication will be established and the hot fluid will flow from an injection well to a production well. In this process, some of the heat in the fluid will be given up and conducted into the formation. Products of pyrolysis will flow from the heated rock matrix into the fractures and be produced, commingled with the injected fluid. The practicality of such a process is strongly influenced by the rate at which heat can be transferred from the hot fluids to the formation. An analysis of this heat conduction problem is the subject of this paper. Linear flow of a condensing gas through a set of horizontal, equally spaced fractures is considered. Temperature histories for both fractures and formation are obtained from numerical procedures. The analysis also applies to formations containing streaks of natural permeability instead of fractures. Heat flow problems similar to the one treated here have been studied by other workers. Thomas1 has solved a radial problem in which he considers the injection of a noncondensing gas into a single, horizontal fracture and approximates the temperature profile in the fracture by a step function. The solutions to other problems, which are less similar to the present problem than the one treated by Thomas, have been reviewed by Spillette.2 In addition to these mathematical studies, a field test of an in situ shale oil recovery process using hot natural gas as the injected fluid has been reported.3 The first part of this paper is concerned with the mathematical development of the problem. The differential equations, boundary conditions and functional relationships which describe the physical model are formulated and the equations are cast into finite difference form. The method of solution of the difference equations is discussed (Appendix B). The second part of this paper presents results computed for a typical oil shale formation for which steam at 1,000F is the injected fluid. Fracture spacing, horizontal and vertical thermal diffusivities

Jan 1, 1967

By J. C. Marquardt L., J. A. Bratt, J. P. Hunt

The Quebrada Blanca copper/molybdenum prospect lies on a dissected plateau in the high Andes of Northern Chile. The Superior Oil-Falconbridge Group began exploration of the Property in 1977 under a joint venture foreign investment agreement with the Chilean government. Quebrada Blanca is a relatively high grade secondarily enriched porphry copper deposit. Primary mineralization occurred 38 million years ago and accompanied quartz monzonite intrusive. Early disseminated Primary mineralization was followed by hydrothermal mineralization, mineralized breccia, and post-mineral faulting. After supergene enrichment Quebrada Blanca was buried under late Tertiary gravels and then partially re-exposed in Quaternary time. During the past five years, a successful exploration program has developed an open-pit mining reserve, provided samples for bulk metallurgical testing, and located more than adequate supplies of water for future operations. The main orebody is amenable to conventional metallurgical treatment and contains more than 400 Mt (441 million st) of recoverable open-pit ore at about a 1:1 overall stripping ratio. The average millhead grade during the first six years of operations will exceed 1.5% Cu, plus recoverable molybdenum values. Several hundred million tons of additional reserves are available in the surrounding districts

Jan 1, 1984

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemisorption.

Jan 5, 1950

By Allan T. Gwathmey, J. Bruce Wagner, Kenneth R. Lawless

Between 250°and 550°C in oxygen pressures of 10 to 760 mm Hg, the relative oxide thicknesses formed per unit time on the (100), (111), (110), and (320), decreased in this order. The predominant oxide orientation was (001)Fe3O4 and [110] Fe304 A small amount of a-Fe2O3 which exhibited a fiber pattern was observed in all oxides except those formed at 250°c in low oxygen pressures. IN order to understand the properties and behavior of thin films, particularly their role in oxidation reactions, there is needed more experimental information on their structure, composition, and kinetics of formation on clean surfaces of known geometry. The present study was undertaken to obtain such information for thin oxide films formed on single crystals of a body centered cubic metal, iron. Crystals in the form of spheres were used in order to expose all possible crystal planes to the reactant gases simultaneously. Using this procedure the relative rates of oxidation of all planes could be estimated, and those exhibiting, maximum or minimum oxidation rates could be selected for a more intensive study of composition and expitaxy. Since the three types of information were obtained from the same experiment using a single crystal substrate, any differences between rate, composition, and expitaxy on the several planes may be attributed to the influence of the substrate structure. EXPERIMENTAL The starting material for this study was Armco iron rods, 6 in. long and 3/8 or 1/2 in. in diam. The analysis was reported to be 99.8 pct Fe, with major impurities listed as carbon 0.018 pct, manganese 0.027 pct, phosphorous 0.005 pct, sulfur 0.029 pct, silicon 0.005 pct, and copper 0.11 pct. These rods were cleaned with ether and heated in an atmosphere of moist hydrogen at 950°C for three days. The purpose of this treatment was to decarburize the iron and to provide a consistent, uniform grain size for the growth of single crystals. Crystals in the form of rods, up to 1/2 in. in diam and 2 in. long, were prepared by a method similar to that reported by Leidheiser and Buck.2 Spherical crystals, 3/8 and 1/2 in. in diam and having a handle about 1/4 in. in diam by 112 in. long, were machined from the single-crystal rods. After machining, each crystal was etched in aqueous nitric acid to remove the cold-worked metal at the surface. Before each experiment, the crystal was polished mechanically with metallographic papers through number 4/0, and electropolished in a perchloric acid-acetic anhydride mixture by the method of Jacquet and Rocquet.3 The crystal was washed in running distilled water for at least 5 min. The excess water was blotted with a soft paper tissue, and the crystal was placed in an all-glass reaction vessel (See Fig. 1). The crystal was heated at 550°C in hydrogen for at least 8 hr prior to the oxidation in order to reduce any surface oxides and to help relieve any stresses in the metal. For the structure and composition studies, flat faces were cut parallel to (loo), (111), and (110) planes on the iron crystal spheres. The orientation of these faces was checked by a back-reflection X-ray diffraction technique. The faces were within 2 deg of the desired orientation. The flat faces were treated as described above except that prior to electropolishing they were given an additional polish on metallographic felt saturated with twenty-minute levigated alumina. After the crystal was heated in hydrogen for at least 8 hr, the temperature was adjusted to the desired value, and the hydrogen was evacuated. Commercial oxygen which had been purified by successively passing through tubes containing magnesium perchlorate, Ascarite and magnesium perchlorate was admitted to the desired pressure as measured by a mercury manometer. A resistance-wound furnace surrounded the oxidation chamber, and a temperature controller maintained the temperature within ±3°C. The furnace was provided with a small window for observing the crystal. The changes in thickness of the oxide films were followed by observing the interference colors. The oxidation was allowed to proceed until the desired oxide thickness was attained on a given crystal plane. The experiment was terminated by evacuating the oxidation

Jan 1, 1962

By W. H. Moore, E. T. Powell

IN the early days of mining in the Anthracite field, only the thicker and better seams of coal were mined, because of the limited mining and coal-cleaning facilities, therefore many of the thinner and less productive seams were permitted to remain unmined. With improvements in mining methods, the introduction of modem preparation plants, and the increased market demand for small-size coal, the thinner and more laminated seams have now become assets. In many instances these seams overlie the mined-out beds of the thicker and better seams, from which all recoverable coal was removed by the conventional breast-and-pillar method of mining. Some collieries in certain sections of the Anthracite field are now relying on these thinner and laminated seams. For the purpose of illustration, a large area in the Hickory Swamp-Ridge Basin of the western middle coal field has been selected, where as early as 1898 the No. 4, or Little Buck Mountain, seam was mined and pillars removed. The No. 5, or Buck Mountain, seam lying on a comparatively light pitch and with intervening strata of but 55 ft. of rock was not mined until 1935. Farther east, in the same basin the Nos. 8 and 9 seams, which are two splits of the Mammoth seam, have been mined and robbed and the workings have been abandoned since 1896. The overlying No. 9½ seam, or the upper split of the Mam- moth, and the No. 9¾ or 4-ft. seam remained virgin, with only an interval of 50 ft. of rock separating them, and with 40 to so ft. of rock strata separating the No. 9½ seam from the completely mined area of the Nos. 8 and 9 seams. These seams lie on a pitch of 45° on the south side of the basin and on 75° on the north side of the basin. Mining has been carried on in the 9½ and 9¾ seams for 3 years on both dips of the basin, for a distance of 3500 ft. In some sections the overlying strata of the thicker seams previously mined have settled without any apparent breaks in the upper strata, as indications have shown in the top split of the Mammoth; while in other areas the roof and bottom of this top split of the Mammoth have settled unevenly, thereby causing the bottom and roof to break into blocks. In attempting to mine over a large area of the No. 5 or Buck Mountain seam, a similar settlement has been encountered; the roof of the seam has been broken and a considerable amount of timbering has been necessary in order to make the mining of the seam possible. The conventional breast-and-pillar method, using shaker conveyors, has been used in the No. 5 seam, and a slant-breast method in the mining of Nos. 9½ and 9¾ seams. Mining OF Light-pitch Seams The No. 4 seam was mined and robbed by the conventional breast-and-pillar system. In this area the seam pitched from 5° to 25° on the north dip, averaged 8 ft. in thickness and had a slate bottom and a

Jan 1, 1942

By W. H. Moore, E. T. Powell

IN the early days of mining in the Anthracite field, only the thicker and better seams of coal were mined, because of the limited mining and coal-cleaning facilities, therefore many of the thinner and less productive seams were permitted to remain unmined. With improvements in mining methods, the introduction of modem preparation plants, and the increased market demand for small-size coal, the thinner and more laminated seams have now become assets. In many instances these seams overlie the mined-out beds of the thicker and better seams, from which all recoverable coal was removed by the conventional breast-and-pillar method of mining. Some collieries in certain sections of the Anthracite field are now relying on these thinner and laminated seams. For the purpose of illustration, a large area in the Hickory Swamp-Ridge Basin of the western middle coal field has been selected, where as early as 1898 the No. 4, or Little Buck Mountain, seam was mined and pillars removed. The No. 5, or Buck Mountain, seam lying on a comparatively light pitch and with intervening strata of but 55 ft. of rock was not mined until 1935. Farther east, in the same basin the Nos. 8 and 9 seams, which are two splits of the Mammoth seam, have been mined and robbed and the workings have been abandoned since 1896. The overlying No. 9½ seam, or the upper split of the Mam- moth, and the No. 9¾ or 4-ft. seam remained virgin, with only an interval of 50 ft. of rock separating them, and with 40 to so ft. of rock strata separating the No. 9½ seam from the completely mined area of the Nos. 8 and 9 seams. These seams lie on a pitch of 45° on the south side of the basin and on 75° on the north side of the basin. Mining has been carried on in the 9½ and 9¾ seams for 3 years on both dips of the basin, for a distance of 3500 ft. In some sections the overlying strata of the thicker seams previously mined have settled without any apparent breaks in the upper strata, as indications have shown in the top split of the Mammoth; while in other areas the roof and bottom of this top split of the Mammoth have settled unevenly, thereby causing the bottom and roof to break into blocks. In attempting to mine over a large area of the No. 5 or Buck Mountain seam, a similar settlement has been encountered; the roof of the seam has been broken and a considerable amount of timbering has been necessary in order to make the mining of the seam possible. The conventional breast-and-pillar method, using shaker conveyors, has been used in the No. 5 seam, and a slant-breast method in the mining of Nos. 9½ and 9¾ seams. Mining OF Light-pitch Seams The No. 4 seam was mined and robbed by the conventional breast-and-pillar system. In this area the seam pitched from 5° to 25° on the north dip, averaged 8 ft. in thickness and had a slate bottom and a

Jan 1, 1942

By Orville R. Lyons, A. C. Richardson

F. X. Ferney—We are pleased that this paper was presented at this meeting and thank Mr. Richardson and Mr. Lyons for their effort and work in preparing it. We agree with the authors that it was unfortunate that more comprehensive data were not available when the paper was prepared, this being particularly true since we had no knowledge that the paper was being written and could not make available data which were in our process files. Further, it is our opinion that because the data are not sufficiently comprehensive the results which are presented are erratic and will lead to false conclusions. As we see it, there are three Specific reasons why the data which are presented cannot be considered accurate. 1. The preparation plant was not at equilibrium while the samples were being taken. This is indicated by the fact that the percent of solids in the circulating water varied from 3.9 pct to 10.7 pct during the period when the plant was sampled. 2. We do not believe that the samples of Bird feed which were taken were representative and could indicatc the true feed to the filter. Evidence of this is shown in the data where the dewatered product leaving the centrifuge contains more coarse material than the feed to it, that is, material which is +6-mesh in size. 3. We question if the data concerning screen analysis in the fine sizes is accurate and would conclude (from the data) that these tests were made by a dry-screen method. We have found that wet screening is always necessary to obtain a true picture of the quantity of —200-mesh fines which is present. The test work described in the Lyons-Richardson paper was conducted some time ago and the tests were carried on during a period when the preparation plant had not ironed out all of their start-up problems. Since that time most of these problems have been corrected and now this is a smooth running, well-operated prep-

Jan 1, 1951

By Orville R. Lyons, A. C. Richardson

F. X. Ferney—We are pleased that this paper was presented at this meeting and thank Mr. Richardson and Mr. Lyons for their effort and work in preparing it. We agree with the authors that it was unfortunate that more comprehensive data were not available when the paper was prepared, this being particularly true since we had no knowledge that the paper was being written and could not make available data which were in our process files. Further, it is our opinion that because the data are not sufficiently comprehensive the results which are presented are erratic and will lead to false conclusions. As we see it, there are three Specific reasons why the data which are presented cannot be considered accurate. 1. The preparation plant was not at equilibrium while the samples were being taken. This is indicated by the fact that the percent of solids in the circulating water varied from 3.9 pct to 10.7 pct during the period when the plant was sampled. 2. We do not believe that the samples of Bird feed which were taken were representative and could indicatc the true feed to the filter. Evidence of this is shown in the data where the dewatered product leaving the centrifuge contains more coarse material than the feed to it, that is, material which is +6-mesh in size. 3. We question if the data concerning screen analysis in the fine sizes is accurate and would conclude (from the data) that these tests were made by a dry-screen method. We have found that wet screening is always necessary to obtain a true picture of the quantity of —200-mesh fines which is present. The test work described in the Lyons-Richardson paper was conducted some time ago and the tests were carried on during a period when the preparation plant had not ironed out all of their start-up problems. Since that time most of these problems have been corrected and now this is a smooth running, well-operated prep-

Jan 1, 1951

By J. T. Eash, E. M. Wise

Platinum and palladium are the most generally useful, most ductile and least rare members of the platinum family. They have many important applications in the pure state but for other applications it is frequently desirable to harden them and otherwise modify their properties by alloying small amounts of the other platinum group metals with them. These alloys are frequently required to withstand very drastic forming operations and to possess certain mechanical properties to fulfill specific requirements. For these reasons a knowledge of the tensile properties and annealing characteristics of pure platinum and palladium as well as their more useful commercial alloys is important to the user and the fabricator. The present work was undertaken to fill this need. Rather brief (5 min.) annealing treatments were employed to render the data of maximum practical value and all tests were conducted upon 0.050-in. dia. wire, initially reduced 50 per cent in area by cold drawing, for the same reason. Conclusions 1. Pure (thermocouple grade) platinum reduced 50 per. cent by cold drawing exhibits the following tensile properties: ultimate tensile strength, 36,000 lb. per sq. in.; proportional limit, 27,000 lb. per sq. in.; elongation, 2.5 per cent in 2 in.; reduction in area, 95 per cent. When fully annealed by treatment at 1100°C. its properties are: ultimate tensile strength, 20,700 lb. per sq. in.; elongation, 30 per cent in 2 in.; reduction in area, 93 per cent. 2. The addition of alloying elements in moderate amounts markedly increases the strength and annealing temperature without much detriment to the ductility. The behavior and properties of variously hardened platinum alloys are considered in some detail. The behavior of a 20 per cent iridium

Jan 1, 1935

By J. T. Eash, E. M. Wise

Platinum and palladium are the most generally useful, most ductile and least rare members of the platinum family. They have many important applications in the pure state but for other applications it is frequently desirable to harden them and otherwise modify their properties by alloying small amounts of the other platinum group metals with them. These alloys are frequently required to withstand very drastic forming operations and to possess certain mechanical properties to fulfill specific requirements. For these reasons a knowledge of the tensile properties and annealing characteristics of pure platinum and palladium as well as their more useful commercial alloys is important to the user and the fabricator. The present work was undertaken to fill this need. Rather brief (5 min.) annealing treatments were employed to render the data of maximum practical value and all tests were conducted upon 0.050-in. dia. wire, initially reduced 50 per cent in area by cold drawing, for the same reason. Conclusions 1. Pure (thermocouple grade) platinum reduced 50 per. cent by cold drawing exhibits the following tensile properties: ultimate tensile strength, 36,000 lb. per sq. in.; proportional limit, 27,000 lb. per sq. in.; elongation, 2.5 per cent in 2 in.; reduction in area, 95 per cent. When fully annealed by treatment at 1100°C. its properties are: ultimate tensile strength, 20,700 lb. per sq. in.; elongation, 30 per cent in 2 in.; reduction in area, 93 per cent. 2. The addition of alloying elements in moderate amounts markedly increases the strength and annealing temperature without much detriment to the ductility. The behavior and properties of variously hardened platinum alloys are considered in some detail. The behavior of a 20 per cent iridium

Jan 1, 1935

By S. L. Hoyt

IN technological studies on steel considerable emphasis has been placed on the identification of the foreign inclusions, testimony of which is adequately given in the metallographic literature covering methods for their identification. The outstanding accomplishment in this field is the scheme devised by Campbell, Comstock and Wohrman for identi-fying various oxides, sulfides, silicates and other constituents by syste-matic etching tests and microscopic examination.1 The time involved in carrying out this procedure has led to attempts to secure quicker methods, or at least to shorten the time required to secure essential information. An example of this is the use of dark field illumination, but when this was tried out by the authors it was found that cavities and pits on the polished surface appeared bright on a dark field and could not be distinguished from small glassy inclusions. Still another means is the Leitz Ultropak, an instrument equipped with optical units for the illumination of the specimen by reflected light to secure an effect similar to dark field illumination. The structures are made visible in exceedingly sharp details by utilizing the inherent properties of the samples, such as refractive index, absorption coefficients and reflection powers. We have had no experience, with the Ultropak, but a recent paper by Benedicks and Löfquist3 describes methods for

Jan 1, 1934

By Richard P. Rothwell

The anthracite coals of Pennsylvania are all mined from large veins. A seam less than four feet. in thickness is generally considered as unworkable, those from which most of the coal now comes being from 8 'to 30 feet thick. The system of mining employed is, almost without exception, that known as " chamber and pillar work;" that is, the coal is taken out of chambers of from 18 to 30 feet in width, and pillars of from 12 to 20 feet are left between these to support the roof. In general, the entire thickness of the seam, even

By R. B. Hoy, Stanley J. LeFond, K. Reckling

World production of natural diamonds prob¬ably exceeds 50,000,000 carats a year. The Republic of Zaire is the leading producer, with an output which is primarily industrial rather than gem grade. The USSR ranks second, and the Republic of South Africa third. Diamonds are also mined in South-West Africa, Angola, Ghana, Sierra Leone, and other African coun¬tries. In the western hemisphere, Brazil and Venezuela are the principal producers (Fig. 1). A very small output of diamonds is mined today in India, which was the first source of commercial production. In the U.S., efforts at commercial diamond mining have been con¬fined to a small area near Murfreesboro, Ark. The first diamond was found in a kimberlite pipe there in 1906. Small-scale trial mining has not, however, proved economical. Since diamonds were first discovered over 2000 years ago, only about 130 tons have been mined. In order to obtain 1 g (5 carats) of diamonds, it is necessary to remove and process approximately 25 tons of rock. Recovering this small percentage involves a combination of highly developed techniques in mining, quarrying, and earth-moving, and extremely sophisticated recovery processes. End Uses Diamonds are used for two unrelated end uses; gem diamonds are jewels of great beauty while industrial diamonds are essential ma¬terials of modern industry. Although imitation stones are substituted for the gem diamond, none of these matches its properties sufficiently well to offer real competition. Synthetic indus¬trial diamonds are now of a quality that per¬mits wide substitution in the fine-grained sizes. Gem Diamonds A diamond is a natural prism. It has the ability to bounce light rays (reflection), to bend them (refraction), and to break them into all colors of the rainbow (dispersion). This ability to "modify" light gives the diamond three characteristics for providing beauty. First is brilliance, the reflection of light back to the eye. Second is fire, dispersion of light into the colors of the spectrum. Third is scintillation, the "twinkling" that occurs when the diamond is moved (Bauer, 1969). The skill of the diamond cutter is required to unlock the brilliant, fiery beauty of a gem dia¬mond (Fisher, 1966). Diamond-cutting skills have gradually developed over the centuries. The first cutters, in India, learned how to grind one rough diamond against another to remove their "skins" and to give them various shapes. This revealed some fire and beauty, but little compared to that of the diamonds that are cut today. About 1475, a Flemish cutter named Van Gerquem discovered that diamond dust on an iron wheel would polish facets on a diamond to provide more brilliance and fire. Cutters continued to experiment with facets, and shortly before 1700 a Venetian cutter named Peruzzi developed the 58-facet brilliant cut. With the development of the science of optics, cutters have learned more about the behavior of light within the diamond. Early in the 20th century, they learned how to angle Peruzzi's

Jan 1, 1975