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By W. C. Ellis, M. E. Fine

In the iron-cobalt system there are several property-composition relationships of theoretical importance. The alloys are ferromagnetic exhibiting a maximum saturation at approximately 33 at. pct cobalt;' the interatomic distances at room temperature show large positive deviations from Vegard's rule;2 and in the region of 50 at. pct cobalt there is an order-disorder transition2 which has a large effect on other properties. The thermal expansion data for the system have been reported only from room temperature to 100 C and for this temperature range a minimum in coefticient in the rcgion of 50 at. pct cobalt is well established.3,4,6 This paper reports the expansion properties in the iron-cobalt system for temperatures from 30 to 850°C, shows the correlation with other physical properties, and proposes an explanation for the tcrrelation. Experimental Procedure The thermal expansivities of iron-cobalt alloys in 10 pct steps in composition were determined from 30 to about 850°C. The major series of alloys was prepared by induction melting from Armco iron and cobalt rondelles with the addition of about 0.6 wt pct manganese. This series was supplemented by three compositions: high purity electrolytic iron, high purity electrolytic cobalt, and an alloy of 52.1 wt pct cobalt prepared from the high purity electrolytic iron and electrolytic cobalt by vacuum-melting. The compositions of both series of alloys are given in Table I. The specimens for measurement were prepared from the cast ingots by first hot-swaging followed in most instances by cold-swaging to 0.180-in. diam rod. Some alloys in the region of 5c pct cobalt, because of inherent cold-brittleness, were swaged hot to the final diameter. The expansion specimens were approximately 2 1/2 in. long with flat-ground and polished ends. Before measurement they were annealed in hydrogen for I hr at 900°C. In the apparatus used for the thermal expansivity measurements (Fig I), a modification of that designed by Hidnert and Sweeney,? the differential expausion between the specimen and vitreous silica was measured by a dial micrometer reading to 0.001 mm. An invar fixture supported the dial gauge, and this fixture, in turn, was connected to a split invar collar which supported the outer silica tube. A pyrex "T" was placed over the outcr silica tube and was sealed to the invar collar by Apiezon Q sealing compound. A movable hollow vitreous silica rod which rested on top of the test specimen slid through a flexible rubber diaphragm cemented to the top of the pyrex "T." The specimen rested On a vitreous silica pedestal fused into the outer tube and was kept centered by a silica sleeve. A nonoxidizing atmosphere was obtained by admitting hydrogen into the

Jan 1, 1949

By Richard L. Ash, J. H. Kelley, W. H. McCracken, William Bellano, J. D. Morgan, D. R. Mitchell

Rubber-tired shuttle cars are finding increased use as gathering units in under-ground mines. As the name implies, they shuttle back and forth between the loading machine and the section transportation system. In many mines their use has in-creased the efficiency of the loading and gathering operations, resulting in increased output per man employed and decreased costs per ton. A car motivated by cable reel is shown in Fig I, and a car with battery power in Fig 2. A brief chronological account of the development of rubber-tired haulage equip-ment follows: 1936. J. H. Fletcher1 installed his first unit for the Blue Bird Coal Co., Carrier Mills, Illinois. This unit consisted of a battery-powered industrial tractor and a trailer with drop-bottom doors. The coal was transported for short distances to a hopper for discharge into mine cars. 1938. A battery-powered self-contained shut-tle car with rubber-tired wheels was built by the Joy Manufacturing Co. Two units were installed for the Katherine Coal Co., Lumberport, West Virginia. The shuttle car consisted of a materials-receiving body with a conveyor in the bottom to move the coal forward during loading and to unload the coal at the discharge station. 1938. A cable-powered shuttle car with four-wheel drive and four-wheel steer was developed by A. L. Lee. Special wheels mounted on king pins on the car permitted all four wheels to be steered. The cable reel was driven by a torque motor to give the cable proper tension during the payout and uniform tension when rewound on the reel. 1940. A Joy shuttle car equipped with trolley poles and a cable reel was developed to operate on a double trolley-wire system.2 This unit was installed at a potash mine of the International Minerals and Chemicals Corporation at Carlsbad, New Mexico. It is believed to have been the first installation of its kind, and the first of shuttle cars for handling material other than coal. 1941. Truck haulage3 was introduced in the Tri State District by the Eagle-Picher Mining and Smelting Company. Storage battery electric trucks were rebuilt to handle a bottom-dump trailer. 1941—Joy shuttle cars were installed in a mine of the United States Gypsum Company, Oakfield, New York. 1943—Joy shuttle cars were installed in several bauxite mines.4,5 1945—Joy shuttle cars installed in a lime-stone mine. Because of the nature of bituminous coal beds trackless equipment has had its greatest application in bituminous coal mining. Usually a shuttle car is filled by a continuous type of loader. In Fig 3 is shown a continuous machine loading coal into a battery-powered shuttle car, 48-in. coal. However, intermittent types of load-

Jan 1, 1949

By Gloyd Wiles

THE eastern section of the Park City dis-trict is drained to an elevation of 6300 ft. by means of the Ontario drain tunnel owned and maintained by Park Utah Consolidated Mines Co. This elevation represents a depth of 920 ft. in Park City Consolidated Mines Company's property. When development below the goo-ft. level was commenced, sinking was at-tempted through an incline winze in the fissure. The intensely fractured quartzite forming the walls of the vein could not be held after water was encountered, so it was decided that further sinking should be done in a vertical winze far enough in the foot-wall to be in the more solid quartzite out-side the fractured zone containing the vein. Progress was good until a depth of 200 ft. below water table was reached, where the flow of water became serious. Sinking pro-gressed 300 ft., to the 1200-ft. level, and there approximately 750 gal. of water per minute was being pumped. The last 20 ft, of sinking cost more than $200 per foot direct charges, whereas the over-all cost for the 300 ft. was $100 per foot. This high last cost was caused by excessive lost time in pumping. Crosscutting to the vein disclosed the fact that the 1050-ft. level had been devel-oped only after drainage had been effected from the 1200-ft. level, but that the 1200-ft. level could not be opened because of the

Jan 1, 1940

[Abrahamson, Guy C., (5'43) Lt., 0-880861, Hammer Field Fresno Calif. Adamec, LAW, (8'42) c/o Braden Copper Co., Rancagua, Chile. Adams, Karl S., (9'43) Univ. of New Hampshire, Durham, N. H. Agnilar, Roperto Q., (9'43) Texas College of Mines, El Paso Texas. Ainge, B'onald Farra, (943) Otago University, Dunedin, New Zealand. Akaran, Cevdet A,, (9'44) Pennsylvania State College, State College, Pa. Albala, Americo, (5'44) School of Mineral Industries, The Pennsylvania State College, State College, PA.]

Jan 1, 1946

By L. Grånäs, B. Aronsson

In a recent study of Fe-Si alloys Walter and Koch' observed pairs of dislocations in Fe-6.25 pct Si crystals rolled to reductions of 10 pct and subsequently annealed for 1 hr at 650°C. They interpreted these pairs as being due to stacking faults. From the spacing of the dislocations a stacking-fault energy of 15 to 19 ergs per sq cm was calculated. These authors did not pay any attention to the order -disorder phenomena in Fe-Si alloys. It is the purpose of this communication to report some preliminary results from a study of ordering and its influence on the plastic deformation of Fe-6.0 pct Si alloys. Our results suggest that the dislocation pairs observed by Walter and Koch may have been due to long-range order rather than to stacking faults. The alloys used in this investigation were prepared from pure electrolytic iron and silicon and were melted in argon. After hot working, strip tensile specimens were machined to a gage length of 70 mm and a thickness of 1 mm. The specimens were heat- treated for 2 hr at 1000°C, furnace-cooled, and subsequently deformed 2 pct in tension at 400°C. The strain rate was 5.6 X lo-' sec-'. Thin foils of the deformed specimens were prepared by means of the Bollman technique and examined in a Siemens Elmis-kop I at an accelerating voltage of 100 kv. In the Fe-Si system long-range ordering occurs at a silicon content of about 6 pct Si. In our alloys evidence for ordering is given by selected-area diffraction patterns which show extra reflections consistent with a super lattice of the D0 9 type. In addition, the dislocation arrangement in deformed specimens indicates that long-range order is present. A typical mi-crophotograph of a thin foil is shown in Fig. l. The dislocations are usually straight and parallel and laying in distinct directions. In the commonly observed (111) zone the projections of the dislocations are close to (113) and (oil). Observation of the image position indicates that the two dislocations in a pair have the same Burgers vector and hence constitute a superlattice dislocation.~ According to theory3 a superlattice dislocation in the DO3 structure consists of four ordinary dislocations connected by two types of antiphase boundaries as shown in Fig. 2. In our specimens only two dislocations in each pair have been observed, Fig. 1. The problem is to decide whether the area between these is a NN or a NNN antiphase boundary. Dark-field images using the (200) reflexion (the indices are defined with respect to the DO3-type superlattice) can

Jan 1, 1968

By Robert E. Brand, Strathmore R. B. Cooke

For a study of the bonding that occurs in magnetite pellets during oxidation, cubes of magnetite, hematite, and quartz were prepared, each with one surface polished and nearly optically plane. These faces were fired together under pressure and at various temperatures to determine whether bonding would occur. Sections of surfaces that bonded were examined microscopically. An attempt was made to measure the strength effected by this treatment. EARLIER work1,2 has shown that the strength developed by fired magnetite pellets containing silica or silicates is due to three separate effects. Two of these are associated with the formation of slag in the interior of each pellet at temperatures above 1000°C, where the atmosphere is substantially though temporarily neutral. The third is independent of the silica content and is responsible for the development of a hard shell around the pellet during the initial stages of firing, a condition which permits the pellet to assume the increasing load of the overlying burden and prevents extrusion of the slagged and semi-liquid interior. Pellets made from hematite concentrates do not show these effects unless specific additives are incorporated, and it is these differences in behavior which in large part account for the difficulty in producing satisfactory pellets from such concentrates. Elsewhere it has been stated that "the most important factor responsible for the development of strength in fired magnetite pellets containing no additive is the oxidation of magnetite to hematite and the subsequent recrystallization and grain growth of this constituent."' The object of the investigation reported in the present paper was to study in detail the mechanism described above and to evaluate its contribution to the strength of the fired pellet compared with other possible mechanisms, excluding slag bonding. This was effected by placing together polished faces of two cubes prepared from specific minerals and then heating them under constant pressure at different temperatures both in air and in nitrogen. The materials used were magnetite, hematite, and quartz. The magnetite was coarsely crystalline material from western Montana, the hematite was Itabira ore from Brazil, and the quartz was of optical grade but of unknown origin. A high-grade Lebanon (Pennsylvania) magnetite concentrate was used to prepare briquets for the investigation of bonding strength. Analyses of the iron oxide materials are given in Table I. The specimens for furnace treatment were prepared with a carborundum cut-off wheel, which cut the minerals into cubes approximating % in. along each edge. Care was taken to make the sides as parallel as possible. The edges of the faces to be polished were beveled, and the cubes were mounted in brass rings with sealing wax, this being essential to prevent the specimens from rocking during polishing. To prevent excessive friction, the wax was recessed so that only the brass ring and the face of each specimen would be in contact with the polishing lap. Each specimen was then ground with No. 600 carborundum to produce a level surface before polishing. The Graton-Vanderwilt process was not used for polishing because of the long time required by this method, but several other methods were tried. Cloth laps using levigated alumina or chromium oxide did not produce the desired polish. Satisfactory results were obtained by means of a pitch lap: which was used for all the mineral cubes, with levigated alumina as the polishing medium. The lap was operated in a machine similar to those used in the preparation of single optical surfaces. It was run at low speeds and moderate pressures, and polishing was stopped occasionally to prevent overheating of the lap. A good polish with flat surfaces could be obtained in about 5 min. After polishing the sealing wax was removed either in hot water or in a solvent such as turpentine.

Jan 1, 1955

New York Meets first Wednesday after first Tuesday of each month. DAVID H.-BROWNE, Chairman, JOHN H. JANEWAY, Vice-Chairman. F. E. PIERCE, Secretary, 35 Nassau St., New York, N. Y. P. A. MOSMAN, Treasurer. LEWIS W. FRANCIS. Boston Meets first Monday of each winter month. HENRY L. SMYTH, Chairman. . ALFRED C. LANE, Vice-Chairman. HENRY A. WENTWORTH, Secretary-Treasurer, 60 India St., Boston, Mass. ROBERT H. RICHARDS, ALBERT SAUVEUR. Columbia Holds four sessions during year. Annual meeting in September or October. FRANK A. ROSS, Chairman. RUSH J. WHITE, Vice-Chairman. LYNDON K. ARMSTRONG, Secretary-Treasurer, P O. Drawer 2154, Spokane, Wash. FREDERIC KEFFER, FRANCIS A. THOMSON.

Jan 11, 1915

By A. Kenneth Schellinger, O. Cutler Shepard

Overgrinding, Or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncon rolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about go pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, over-grinding is detrimental to the flotation process. Overground slime coats larger partitles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The rem- edy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. Development oF Apparatus and Experimental Method Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7½ in. in diameter and 8 in- deep. A cast-iron disk Was mounted on the end of a vertical shaft SO that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffee-mill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, more charges of Ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned On so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1949

By E. I. Larsen

Powder metallurgy has been described as being "as old as the pyramids and yet as new as the latest bomber." While this may be true literally, it has been only in the last Io or 1 5 years that widespread use has been made of this method of fabricating various articles One of the first items fabricated by powder metallurgical methods on a commercial scale was electrical contacts, which were fabricated from metal powders as early as 1920. The reason why these contacts have been made for so many years by powder metallurgy is because the compositions of most of these materials arc such that they could not be made by the conventional methods of melting, casting, rolling, drawing, etc. Classes of Contacts In general, electrical contacts fabricated from metal powders fall into two classes, both from the point of view of the method of fabrication and the components that make up the chemical composition. One class is composed of silver or copper and a refractory metal such as tungsten, molybdenum, or compounds thereof such as carbides. The second class contains a predominant amount of silver and a semi-refractory material such as cadmium oxide, nickel, cobalt, or graphite. Electrical contacts of the first class have been developed primarily as facing materials for heavy-duty circuit-interrupting devices. This class of material unites the high melting point, high boiling point, great resistance to electrical erosion, and the refractory characteristics of molybdenum, tungsten, or their carbides with the high thermal and electrical conductivities and low contact resistance properties of silver or copper. The second class of material is used for lighter duty or specialized applications such as heavy-duty, direct-current aircraft relays. Varieties Electrical contacts fabricated from metal powders are made in a wide variety of compositions, shapes, and sizes. In general, the sizes of electrical contacts fabricated from powders vary from contacts weighing approximately 0.001 oz. up to those weighing 3 to 4 Ib., the contacting face being as small as 0.001 sq. in. up to approximately 6 sq. in. in area. The shapes consist of: sound or rectangular buttons, some of which have radii or bevels on the contacting faces; rings, cups, rectangular bars, the contact face of which may have a radius across the width and/or at one end; and form-pressed pieces of a wide variety such as angles, both acute and obtuse, irregular shapes comprising a radius blended into a flat, which in turn may have steps. The size or shape of the contact is dictated principally by the mechanical design of the manufacturer, who in turn may have problems to be met such as: (I) size, weight, or space limitations; (2) electrical conditions such as blowout coils arc chutes, baffies, gas passages.

Jan 1, 1945

By Edwin Ludlow

THERE has been so much misinformation sent out through the newspapers, and I find so few people who are really acquainted with the true facts in regard to the coal strike, that I feel it would be advan-tageous to present the facts as I have been able to gather them. One of the great points that the miners are now mak-ing is the fact that the operators of the central competi-tive field, consisting of the states of Illinois, Indiana, Ohio and Pennsylvania, have refused to meet them in a joint conference such as has been customary in previous years. There are two reasons why the operators de-clined to hold such a meeting. In the first place, the last time they held a meeting of this kind they were all indicted under the Sherman Act for making an agree-ment in restraint of trade. In the second place, the miners at their convention did not authorize their dele-gates to negotiate a new contract, but instructed them that nothing less than a scale based on the present wage, with the addition that there should be a six-hour day, a five-clay week, time j and a half for over time, and double time for Sundays, would be accepted. These additions would very much increase the cost of coal over the present wage scale. The operators could not consider renewing a contract on the present scale, as it would mean ruin for them, and as the miners were not authorized to negotiate a new scale on a fair basis that would enable the operators to mine coal in competi¬tion, with the open-shop mines the operators did not feel that there was any use in meeting the delegates. Thirty per cent. of the output of bituminous coal comes from the open-shop mines of West Virginia, Alabama and Pennsylvania. During the strike, in 1919, this 30 per cent. was increased to 50 per cent. and it has affected the operations of the union mines to such an extent this year that they have been able to run only 15 to 30 per cent. of their time, while the open shop mines have run from 60 to 80 per cent. as the wages in these mines had been reduced to the 1917 scale.

Jan 5, 1922

By E. K. Olson

Eagle Mountain mine of Kaiser Steel Corp. began production in 1947 when the direct shipping ore contained 51-56% Fe. The high grade, wide ore face and low stripping ratio ores were rapidly depleted and Eagle Mountain required processing plants for upgrading and more efficient mining equipment. During 1970-71 total annual pit tonnage averaged 56 million tons of material ranking the Eagle Mountain Mine as the fourth largest in the country. Ore production averaged 12.5 million tons requiring a stripping ratio of 3.5: 1. During this period mining activities were scheduled for a three-shift, seven-day operation. The present mining rate is about 30 million tpy on a two-shift, five-day schedule.

Jan 12, 1972

By C. H. Mathewson

BERYLLIUM, magnesium, zinc and cadmium, together with mercury, constitute a coherent sub-group of the periodic system and these metals, excepting mercury; have been studied in sufficient detail by the new X-ray diffraction methods to prove their common adherence to the hexagonal close-packed type of crystal structure with a progressive increase in axial ratio as we pass through the group from beryllium to cadmium. The first four columns of Table 1 summarize the data on crystal structure of these metals reported in the International Critical Tables.1 The last two columns are based on results obtained in the present investigation. TABLE 1.-Lattice Constants and Twinning Law of the Be-Cd Group of Hexagonal Metals Dimensions of Unit Axial Cell, A Ratio Description of Twinning Element a° c° (c°/a°) Composition Axis Plane Beryllium 2.28(3) 3.60(7) 1.58 {1012} Probably 1 {1012} Magnesium 3.22 5.23 1.621 ,1 {1012} Zinc 2.65(7) 4.94(8 1.86(2) 11012 } 1 {1012) Cadmium 2.98 5.63 1.88(5) {1012} Probably 1 {1012} Twinning has often been observed in zinc and cadmium. The general aspects of this twinning are abundantly illustrated by photomicrographs, diagrams, etc., in the papers by Romig, and Mathewson and Phillips reported to this Institute in 1927.2. A determination of the composition

Jan 1, 1928

By A. G. Preban, R. J. De Angelis, J. B. Cohen, P. Gordon, D. C. Brillhart

Changes in stored energy, resistivity, density, X-ray line broadening, and dislocation arrangements (from transmission electron microscopy) have been mensured on copper specimens shock-loaded at 75 to 435 kbars. The stored energy in shock-loaded copper and changes in density and resistivity are larger than those after conventional deformation. A high density of twins makes an appreciable contribution to all the measured quantities. Dislocation densities rise to values of 1011 lines per sq cm at the highest pressure. The resistivity, density, and X-ray In a study of the effects of shock loading on copper, changes in stored energy, resistivity, density, hardness, lattice parameter, line broadening, and dislocation structure have been examined after subsequent annealing of specimens shocked to various pressures. This paper presents certain aspects of the data, taken by two groups, one at the Illinois Institute of Technology, and the other at Northwestern University, and compares the changes in these properties, and what they indicate about the effects of shock deformation. measurements yield dislocation densities in substantial agreement at all pressures, after correcting for the effects due to the twins. At the higher pressures, electron microscopy yields dislocation densities that are lower by a factor of about 5 to 10. Dislocation densities from stored -energy measurements are high; this finding can be rationalized by the consideration that the deformation twin boundaries present are largely incoherent. There is no evidence for homogeneous dislocation nucleation during shock loading. Shock loading any metal or alloy, such as copper, causes a large increase in yield strength, with dimensional changes of the order of only a few percent."' Other properties have not been extensively studied, but this result clearly indicates the importance of examining the substructure. Previous studies of the substructure of conventionally deformed specimens have been made using electron microscopy,3-6 X-ray diffraction,7,8 and etch pitting.9,10 The dislocation density can reach values of about 10" per sq cm; and the dislocations are arranged in cells approximately 0.15 in size. The substructure varies widely, however, from point to point in the specimen. The substructure of explosively loaded copper consists of cells of about the same dimensions6,11 but their size is more uniform.6 Profuse twinning occurs12,13 and the twins are often too small to be detected optically.6,13 The boundaries outlining these twins appear to be incoherent.14 There is now ample evidence 13-15 that twinning is a strengthening mechanism in fcc metals but this is only part of the strengthening effect in shock-loaded copper.* From the twinning systems which operate in shock-loaded single crystals,13 and from changes in texture on shock loading,17 it has been shown that both the

Jan 1, 1968

US 4,201,749-In the Bayer process production of alumina wherein aluminum hydroxide is precipitated from a sodium aluminate solution for use as seed, sodium oxalate and caustic components are removed from the precipitate by washing it first with a washing liquor which is saturated with sodium oxalate but has little content of caustic, and secondly with hot water. Substantially no sodium oxalate is recycled, yet loss of sodium aluminate is minimal. K. Yamada, T Harato, and H Kato, assigned to Sumitomo Aluminum Smelting Co Ltd May 6, 1980 8 pp. (423-127). Filed. 5-8-79 (37,145) Priority. Japan, 5-8-78 (53/54,803). Same¬British 2,020,642, dated Nov. 21, 1979.

Jan 1, 1982

By Heribert Wiedemeier, Ali Khan

The tetnperature-conposition phase diagram of the syste,n MnS-CdS has been investigated over a range fro)n 600° to 1000° C. The small degree of solid solution on the MnS rich slde cornpared to the large CdS rich single-plzase region could be esplained by the difference in strzrctures of the cotnponents and the relative sizes of the metal ions. The change in lattice parameter as a function of cot)zposition is linear atzd jollous Vegavd's rule within experittzental limits of error. Mns and CdS have gained considerable attention in recent years because of their interesting solid state properties. In order to investigate such properties and thermodynamic quantities of mixed compositions of transition metal chalcogenides, it is necessary to establish their phase diagrams. Two earlier publications are concerned with the MnS-CdS system. Schnaase' found complete miscibility when MnS and CdS were co-precipitated in aqueous solutions at room temperature. However, there is evidence that the mixed crystals were not stable over the entire range of composition. roger's' investigation of the ternary system ZnS-CdS-MnS at 900" C indicates a solution of 48 mole pct MnS in CdS. During the course of this work which covers a temperature range from 600" to 100O° C, emphasis has been placed on the use of high-purity materials and well-defined modifications of starting products. EXPERIMENTAL Starting Materials. The stable green modification of MnS with NaCl structure was prepared by direct fusion of the elements and by chemical transport reaction. Stoichiometric amounts of manganese of 99.99+ pct purity and of sulfur of 99.999 pct purity from Gallard-Schlesinger Corp. were sealed in an evacuated quartz tube at a pressure of 10"5 mm Hg and annealed for 48 to 60 hr at 800°C. The quartz tubes used throughout this work were outgassed under high vacuum at a temperature of 90OU to 1000° C for 12 to 18 hr. For the synthesis of MnS by chemical transport3 stoichiometric ratios of elemental manganese and sulfur and small amounts of iodine were sealed under vacuum in a quartz tube, which was placed in the temperature gradient, 900/800-C, of a two zone furnace. Single crystals of MnS in form of platelets and octahedra were deposited at the cooler end of the tube. The manganese content was determined by EDTA titratin. A typical result of several analyses was 63.05 pct Mn which agreed well with the computed value, 63.14 pct. Commercially available CdS of 99.9+ pct purity, a mixture of the cubic and hexagonal forms, was converted into the hexagonal high-temperature modification by annealing CdS under vacuum in quartz tubes at 700° C for about 36 hr. As a second method, cadmium sulfide was transported in a temperature gradient, 70O° /650" C, with in vacuo sublimed iodine from the hot to the cooler zone of the tube. Large sized crystals in form of hexagonal columns and some platelets were obtained. An average result of the cadmium analysis by EDTA titration4 was 77.93 pct in good agreement with the computed value, 77.81 pct. Only the NaCl modification of MnS and the wurtzite modification of CdS were used for the solid solutions. Preparation of Solid Solutions. Finely powdered MnS and CdS samples were thoroughly mixed in molar ratios and compressed into pellets of about 6 mm diam and 2 mm in height at a pressure of approximately 13,500 atm.

Jan 1, 1969

By J. C. Bierlein, R. A. Dodd

Endurance strengths at 10' cycles, fatigue-hardening rates, and endurance strength/0.2 pct proof stress ratios have been determined jbr a range of Cu(Ni), Cu(Si), and S.R.0. Cu(Au) solid solutions. Some douht is cast on simple cross slip models of fatigue hardening in view of opposite composition dependencies of hardening displayed by Cu(Si) and Cu(Au). The temperature dependencies of hardening also are the opposite of predictions made on the basis of a cross slip model. A correlation apparently exists between the fatigue strength/0.2 Pc~ proof s1.re.s~ ratios and rates of fatigue hardening. An increase in PS/PS due to alloying or temperature change is accompanied by a decrease in hardening rate, and vice versa. In recent years there has been much interest in the determination of the dislocation structure of fatigued metals and alloys. The accumulated evidence, e.g., Refs. 1 to 5, suggests that the structure may be dependent on the stacking fault energy, 7, of the fatigued metal in the same way that fatigue hardening rates have also been shown to apparently depend on u .6' 7 The above research followed the various earlier theories relating crack nucleation to ease of cross slip, e.g., Refs. 8 to 131 so "ggesting that y may indeed be the parameter of principal significance in all facets of the fatigue process. However, comparatively little attention has been paid to correlating engineering fatigue data with observations of the above type. Certainly, it would be useful if potential engineering fatigue performance could be assessed from a knowledge of easily determined alloy properties, and the present research originated from this standpoint. TO investigate the widest Possible range of 7 would require the use of two solvent bases, e.g., aluminum and copper, but a reasonable coverage can be provided by copper-based solutions alone. Therefore, it was decided to work with Cu(Ni) and Cu(Si) solid solutions, the approximate stackillg fault energies of which are given in Table I. The y range extends from low, Cu(7.5 at- Pet Si), to moderately high in Cu(2-5 at. ~ct ~i) and two of the Cu(Ni) alloys. The values listed in Table I should be regarded as qualitative only, being derived as follows. Dillamore and smallman14 quote a value of 85 i 30 erg.cm-2 for the stacking fault energy of pure copper, based on an earlier value due to Howie and swannl5 corrected by Brown's formula.'' Since the ratios (alloy)h(pure copper) have been reported17,18 for Cu(Ni) alloys containing up to 30 pct Ni, approximate ) values for these alloys can be computed, and a rough estimate of 7 for Cu(50 at. pct Ni) obtained by extrapolation. The relatively low y value obtained in this way for Cu(50 at. pct Ni) coincides with the observation of Nakajima19 that the stacking fault probability is a maximum at this composition. Likewise, the y values for the three ~u(~i) alloys are estimates based on values given by Swann and Nuttingo corrected on the basis Of Brown's estimate that the true 7 values are probably 2.3 times greater than those originally computed. In addition to the above alloys it was decided to investigate short-range ordered (s.R.O.) CU(AU) alloys containing up to 25 at. pct Au. This last alloy has been shown to have a well-defined planar arrangement of dislocations when deformed,21,22 probably due to cross slip being restricted by the S.R.O. Engineering fatigue data was to be represented by the determination of endurance strengths, and these were to be correlated with fatigue hardening and mechanical property data. EXPERIMENTAL I, order to study the desired properties and property changes, the following alloys were prepared: Cu(2.5 at. pct Si), Cu(5.0 at. pct Si), Cu(7.5 at. pct Si), Cu(5.0 at. pct Ni), Cu(25.0 at. pct Nil, Cu(50.0 at. pct Ni), Cu(5.0 at. pct AU), Cu(15.0 at. pct AU), Cu(25.0 at. pct AU). The copper and gold were zone-refined, while the silicon was semiconductor grade. The nickel was of 99.95 p,t. purity.* All alloys were induction-melted in a *Kindly supplied by the International Nickel co. helium atmosphere, appropriate precautions being taken to avoid segregation in the Cu-Au series. Rod stock was obtained by rolling and swag,ng. A few Bridgman single crystals were grown for fatigue-hardening studies, but most material was machined into polycrystalline fatigue speciments of the design shown in Fig. 1, and into polycrystalline tensile and fatigue-hardening specimens of simple cylindrical design. All specimens except Cu(Au) were annealed; the latter were quenched from above T, to produce short-range order. A few of the quenched alloys were examined for long-range order by step-scanning over

Jan 1, 1969

By John Wulff, J. H. Hollomon

THE opinion has been widely held that the corrosion of alpha brass occurs by the selective solution of zinc. As late as 1939, Fink1 and Evans2 suggested that in the initial stage of the corrosion the selective solution of zinc produces copper-rich areas, which subsequently act as cathodes that facilitate the deposition of copper. Previously Abrams3 had shown that most of the copper found on corroded brass was redeposited and not residual, although he did not consider the nature of the initiation of the corrosion process. Enough data exist to indicate that any metal surface is sufliciently heterogeneous to provide the anodes and cathodes necessary for the initiation of corrosion .4,5 Thus, for alpha brass we need postulate not an initial selective solution of zinc, but only the existence of local differences in electrochemical potential. The heterogeneity responsible for potential differences may arise from oxide and other films as well as from alloy concentration gradients. Even though selective corrosion may be possible in some alloy systems, previous data on alpha brass as well as the experimental work reported in this paper do not require this supposition. Bengough and May,6 in a discussion of Abrams' paper, first suggested the addition of small amounts of arsenic (0.03 per cent) to alpha brass to prevent the redeposition of copper (dezincification). This suggestion was later substantiated by Nixon.7 In their studies of the inhibiting influence of arsenic, Evans and Fink were led to the selective-solution hypothesis in order to explain the initial corrosion of such alloys. "I hey believed that an arsenic film covered anode areas and prevented the formation of copper-rich spots capable of acting as cathodes for the deposition of copper. They (lid not consider, however, the mechanism by which the arsenical film was formed. Bengough and May,8 in another publication, suggested that the commonly found gray or black arsenical film may act to: (1) carry oxygen to the brass surface, (2) increase the discharge potential of copper on brass, (3) precipitate cuprous chloride from colloidal suspension in chloride solutions. Masing9 assumed that the film found upon arsenical brass was metallic arsenic. His measurements of hydrogen overvoltage of bulk arsenic led hint to believe that the inhibition of dezincification occurs because the overvoltage of a brass surface is lowered by the arsenic film. He reasoned that this film would permit the more rapid reduction of hydrogen at the cathode, which would prevent the redeposition of copper. In a recent paper,10 we have pointed out that the film formed (luring the corrosion of arsenical brass and copper in stagnant solutions is essentially arsenious oxide (AS208)

Jan 1, 1942

By P. M. Blair, D. W. Peaceman

The shape and position of the gas-oil transition zone during downdip displacement of oil by gas has been calculated using flow equations which include the effects of gravity, relative permeability, capillary pressure and compressibility of the fluids. The calculations treat the problem in two space dimensions, and results are compared with data from a laboratory model tilted at 30° and 60° from the horizontal on displacements near and above the maximum rate at which gravity segregation prevents channeling of the gas along the top of the stratum. The good agreement between calculated and experimental results demonstrates the validity of the tecbnique as well as that of the flow equations. INTRODUCTION Knowledge of the fluid distribution and movement in an oil reservoir is important in producing operations and estimation of reserves. The history of the oil industry has included steady progress in improving the accuracy of calculations which provide the required knowledge. The earliest method of calculating reservoir performance consisted of material-balance equations based on the assumption that all properties were uniform throughout a reservoir.l For many reservoirs such a simple formulation is still the most useful. However, when large pressure and saturation gradients exist in a reservoir, the assumption of uniform values throughout may lead to significant error. To reduce these errors, Buckley and Leverett2 introduced a displacement eqation which considers pressure and saturation gradients. Methods available at that time permitted solutions to the Buckley-Leverett equation in one space dimension; these solutions have been very useful in solving many problems related to the production of oil. However, the one-dimensional methods are not adequate for systems in which saturations vary in directions other than the direction of flow. An example of such a system is the case of gas displacing oil down a dipping stratum in which the gas-oil contact becomes significantly tilted. Of course, the Buckley-Leverett displacement method cannot predict the tilt of the gas-oil contact. Recent improvements3,4 of the one-dimensional Buckley-Leverett method achieve some success in predicting the tilt of the gas-oil contact at sufficiently low flow rates. However, at rates high enough that the viscous pressye gradient nearly equals or exceeds the gravity gradient, even these improved one-dimen,sional methods incorrectly predict the shape and velocity of the contact. Further progress in estimating such fluid movements in a reservoir appears to require consideration of the problem in more than one space dimension. The recent two-dimensional method of Douglas, Peaceman and Rachford5 appears adaptable to calculate changes with time of the saturation distribution in a vertical cross-section of a reservoir. The movement of saturation contours should represent the moving fluid contacts and include the effects of crossflow due to gravity, as well as variations in the rock and fluid properties. The nonlinear nature of the equations used in the method has prevented proof of the validity of the solutions. Douglas, Peaceman and Rachford made some comparisons with experiment but did not include cases in which gravity was important nor cases involving displacement by the nonwetting phase.For these reasons, a test of the two-dimensional method for a case in which these factors are included would be very desirable. The test selected was a comparison of calculated results with those from a carefully controlled laboratory experiment on a model with measured physical properties. The model selected was one in which gas displaced oil down a tilted, rectangular sand pack. The model can be thought of as representing a vertical cross-section taken parallel to the dip of a reservoir. The displacement thus simulates gas displacing oil downdip that might result from gas-cap expansion or gas injection. This report outlines the calculation method used

By P. Dosmond

In ore mining technology, the loading and hauling operations are undoubtedly the area where the most spectacular progress has been made during the last twenty years. In the iron ore mines of Eastern France production of 62 million tons for 1962 and 19863 constitute a worthwhile experience worldwide. These deposits which are located in Lorraine are of sedimentary origin and they contain many mineralized horizons with workable seams varying in depth from 330 to 830 ft.

Jan 1, 1970

By C. M. Means

Considering the hazard involved in mining operations, statistics show that a very small percentage of accidents is chargeable to electricity. These accidents do represent quite a large percentage of those that are preventable and they are the direct result of the introduction, for purely economical reasons, of a dangerous element. The introduction of electricity in mines should decrease the hazard, and in no case should electricity be applied to the mechanical operation of equipment if by so doing the dangers incidental to mining are increased. The greater number of accidents are the result of persons coming in contact with exposed conductors at potentials varying from 258 to 508 volts direct current. The use of alternating current at higher potentials than these is quite common in large mines, but it is very rare that an accident happens from this source. This is explained by the fact that where these high voltages are warranted, proper precautions are taken, and the installations are directed by those who fully appreciate the dangers incidental to the work. Direct current is in much more general use than alternating current because of its adaptability to haulage locomotives and the low cost of installation in small operations. We are, however, coming to a more extended application of alternating current for the operation of mechanical devices used in connection with mining. Our underground direct-current wiring system is an evolution from the surface trolley and feeder systems. This is probably due to the fact that our earlier successful applications were electric locomotives, which required the use of a trolley wire. The trolley current has been almost banished from industrial establishments on the surface because of the hazard involved, yet we use it indiscriminately underground, where the dangers from accidental contact or fire are infinitely greater., This does not mean that the underground trolley systems should be eliminated, but it does imply the proper safeguarding of such equipment as is necessarily a part of the trolley systems. It is a fact that the men now employed in the installation and maintenance of mining equipment may be in many cases incompetent or not thoroughly familiar with the work, but they are the men who must do the work and, in order to do it properly, it is necessary that they be educated

Jan 1, 1915