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By W. G. Pfann

Formation and slow solidification of a molten zone in a homogeneous ingot produces a discontinuity in solute concentration at the boundary of the zone and a gradient of concentration within the zone. By using two solutes, one a donor, the other an acceptor, step or graded pn barriers can be produced. EXPLOITATION of the difference in concentration between a molten solution and the solid which freezes from it is continued in this paper. The normal segregation which results from this difference has been used for purification'.'2 and for pn barrier-formation in semiconductors."4 By means of traveling molten zones this difference has been used to effect multistage separations and to eliminate segregation in ingots." a It will be shown here that by means of stationary molten zones this difference can be used to produce discontinuities in concentration. In particular it will be shown that a pn or npn barrier can be made by the simple act of melting and refreezing a portion of a block of semiconductor, provided that the block contains proper concentrations of a donor and an acceptor in solid solution. Assumptions The operation of melting and slow refreezing will be called remelting and the interface between liquid and solid at the start of refreezing will be called the remelt bormdary. While remelting may be done for various shapes of starting crystal and remelt zone it will be assumed that the remelt zone is cylindrical, its length being measured in the direction of freezing, which is normal to the remelt boundary, and

Jan 1, 1955

By William D. Klopp

A study was conducted of the tensile and creep properties of are-melted and electron-beam-melted tungsten over the temperature range 2250° to 4140°F. The tensile and creep strengths vary with pain size, the finer -grained materials being stronger. The activation energy for transient creep at 2250° to 3500°F is 95,000 cal per (g)(mole), while that for steady creep at 3300o to 4000°F is 141,000 cal per (g)(mole). At 2875o to 3300°F the apparent activation energy for steady creep decreased with decreasing temperature. It is suggested that cross slip is rate controlling during transient creep, while dislocation climb is rate controlling during steady creep above 3300° F. ALTHOUGH considerable data have been generated on the high-temperature tensile properties of powder-metallurgy tungsten,1 relative few data have been published on the properties of are-melted2 or electron-beam melted tungsten.3 Furthermore, the available data show a fairly large amount of scatter at high temperatures. The effects of structure and purity have not yet been defined, aside from observations that high-purity electron-beam (EB)-melted tungsten3 and single-crystal-plasma-flame tungsten4 are weaker than less pure powder-metallurgy tungsten. The creep behavior of powder-metallurgy tungsten has been studied, but no data are available on the creep behavior of tungsten consolidated by melting. Green5 found that powder-metallurgy tungsten exhibited linear creep at 4082° to 5072°F with an activation energy of 160,000 cal per (g) (mole). The similarity of this activation energy to that of 153,000 cal per (g)(mole) for volume self-diffusion6 suggests that recovery of strain hardening by dislocation climb is rate controlling in this temperature range. In studies at 4800°F, Sutherland and Klopp7 concluded that grain size significantly affected the creep behavior of commercial powder-metallurgy materials. This grain-size effect was suggested by sherby,8 although the opposite relation had been reported for aluminum.' The purpose of the present study was twofold: first, to provide baseline data on high-temperature tensile and creep properties of are-melted and EB-melted tungsten for concurrent tungsten alloy studies, and, second, to determine the phenomenolo-gical effects of structural and purity variables on the high - temperature mechanical properties of tungsten. EXPERIMENTAL The are-melted tungsten materials were prepared by melting sintered 1-1/4-in.-diam electrodes into 2-1/2-in.-diam by 5-in.-long ingots. The are-melting facility has been described previously.' Eight ingots were melted, machined, and hot-extruded at

Jan 1, 1965

According to published reports, controlled production from the Masjid-i-Sulaiman and Haft Kel fields for the year 1934 was as follows: No new development of major importance in either of these areas is reported. Owing to the increasing- demand in North Iran, particularly for gasoline and kerosene, it has been arranged that this market shall in future draw its supplies from a newly constructed refinery at Kennan-shah. This refinery is connected by a 3-in. pipe line, 145 miles long, with the Naft-i-Shah field, on the borders of Iraq and Iran. The crude oil will pass through a topping plant at the field and the products will be stored and pumped to the Kermanshah refinery in such proportions and quantities as are demanded by actual market requirements.

Jan 1, 1935

By T. F. Witherbee

The subject of an article in the Engineering and Mining Journal for October 13th, 1877, namely, Blast Furnace Treatment of Silicious Iron Ores,, is of great interest to myself, and doubtless to many others who hare been troubled with a seemingly unaccountable percentage of silicon in their pig iron. It is to be regretted that the author of the article did not give us complete details of his charges, and thus prevent a possibly fatal error in the very beginning of the discussion of the points raised in his communication. As it now stands, I think we will have to fall back upon the statement that, " every conceivable mixture of fluxes and ores (A and B) has been used." Let us first, in order to simplify the calculations, do with she

Jan 1, 1879

By H. H. Johnson, Eric W. Johnson

A newly developed ac technique was used to measure the electrical-resistivity changes associated with both cyclic stressing and subsequent annealing of high-purity and OFHC copper. The early stage of push-pull fatigue was explored at 293° and 4.2°K, with stresses chosen for an expected fatigue life of 2 to 8 X 105 cycles. For high-purity copper, fatigue at 293°K was accompanied by an initial increase in resistivity. After about 5000 stress cycles, no further increase was observed. Annealing studies showed that the resisticity increment could be accounted for solely in terms of an increased dislocation density, to about 2 x 1010 disl pev sq cm. Fov 42°K fatigue, however, the resistivity increased continually, with no evidence of saturation, as approximately (No. of stress cycles)1/4. This is intevpreted to result from both dislocation multiplication and vacancy production; estimated vacancy concentrations are 10-4 to 10-9. The continuous resistivity increase was inhibited by interspersed annealing treatments at 293°K, which also raised the level of the stress-strain curve. These effects are attributed to annealing of vacancies into dislocation loops. In contrast, OFHC copper fatigued at 4.2°K showed saturation of electrical vesistivity after only a few hundred cycles, and it appeared that vacancies annealed out completely between 77° and 293°K. MaNY observations suggest that both dislocation multiplication and point-defect production are important aspects of fatigue deformation. An increase in dislocation density with number of stress cycles has been demonstrated or implied by several experimental techniques, including optical metallography with lithium fluoride,' electron microscopy with copper,2 thermal conductivity of Cu-Zn alloys,3 and peak-stress measurements during cycling at constant plastic strain amplitude.4,5 At low amplitudes and/or stresses saturation of fatigue hardening and dislocation density occurs early in the expected fatigue life,4-6 frequently by a few hundred cycles. The evidence for point-defect production is somewhat less extensive, but it is nonetheless persuasive. Rosenberg7 reported briefly that, for equal stresses, copper fatigued at liquid-hydrogen temperature displays an increase in electrical resistance some ten times that which accompanies unidirectional stressing. Further, the extra resistance saturates after a few hundred stress cycles, and does not anneal significantly until a temperature of about 170°K is attained.' Point-defect production during fatigue is also suggested by 1) the softening during cyclic stressing of initially strain-hardened polycrystal-line copper9, 2) the strong temperature dependence of the yield strength of fatigue-hardened mono- and polycrystalline copper,579"0 and 3) the influence of intermediate annealing upon slip-band formation in copper fatigued at 90°K.8 It is evident that characterization of the fatigued state requires a knowledge of both point-defect and dislocation densities. For this purpose electrical-resistivity measurements are useful, but they are difficult to perform because of geometric limitations. Push-pull fatigue is necessary to obtain a macroscopic ally uniform deformation over the specimen cross section; however, the specimen must then have a low slenderness ratio to avoid plastic buckling, while the standard dc resistivity technique requires a very slender specimen for

Jan 1, 1965

By Robert T. Ault

The nature of fracture in unnotched tensile and notched tensile sheet and round specimens and V -notched and precracked Charpy-type sheet specimens of both wrought stress -relieved and re-crystallized molybdenum was investigated over the temperature range of -78° to 300°C. The sharp rise in fracture stress, for unnotched tensile samples, as the temperature is increased above the brittleness transition temperature (Tb), is found to be a result of an increase in flow stress due to plastic constraint and an increased strain rate which results from the onset of necking at Tb . Quasi-brittle fracture in notched tensile samples is found to occur when the tensile stress at the elastic-plastic interface, beneath the root of the notch, reaches a maximum critical value, which is independent of test temperature over the range from —78° to 25°C, but dependent on microstructure. In the temperature range between 150° and 300°C, unnotched tensile and notched tensile samples alike are found to fracture by a ductile fibrous tearing process which is discontinuous in nature, as a result of the competition between the processes of tearing, through continued plastic flow, and local work hardening. Results from the V-notched and fatigue-cracked Charpy-type impact tests demonstrate that crack initiation is the governing factor in the low-temperature (-78° to 25°C) fracture process for molybdenum. In recent years, there has been considerable interest, and a commensurate number of investigations, concerning the duc tile-to-brittle transition in refractory metals. There has not developed, however, a satisfactory explanation for the uniaxial tensile fracture-stress transition which accompanies the well-known ductility transition. This matter therefore warranted investigation. In a similar manner, the increased importance of notched tensile strength values in evaluating materials and in design criteria requires a better understanding of the factors which control the fracture behavior of notched tensile samples. In a previous investigation1 of the nature of initial yielding and fracture in notched sheet molybdenum at room temperature, it was suggested that plastic constraint was the controlling factor in governing the fracture behavior of notched samples. The final portion of this investigation was concerned with the fracture toughness of molybdenum. Of particular interest was the comparison of effective surface energies for fracture in V-notched samples with fatigue-cracked, Charpy-type samples in order to ascertain the relative importance of the initiation and propagation phases of the fracture process. MATERIALS AND TEST PROCEDURE Sheet tensile, notched tensile, and V-notched Charpy specimens were prepared from 50-mil, stress-relieved molybdenum sheet.* Half of these specimens were vacuum-annealed for 1 hr at 1200°C and 7.5 x l0-5 Torr and furnace-cooled to produce a relatively uniform grain size of 0.30 mm diameter. Round notched and unnotched tensile specimens were prepared from warm-rolled and swaged 5/8-in.-diam bar.* These specimens were vacuum-annealed for 1 hr at 1350°C and 5 x lom5 Torr and furnace-cooled to produce a grain diameter of 0.11 mm. Material analyses in ppm were: C N O H Sheet 290 10 4 1 Bar 50 30 30 2 The unnotched tensile sheet specimens had a 1/4-in. gage width and a 1-in. gage length. The unnotched tensile round specimens had a 9/32-in. gage diameter and a 1-1/4-in. gage length. The sheet and round notch tensile specimens are shown in Fig. 1. All of the notched tensile specimens tested were tandem-notched in order to study the location and mode of fracture initiation. The tandem-notched specimens were carefully machined so that the largest variation in notch depth on a single specimen was 0.001 in. Thus, when fracture occurs, the extent of plastic deformation that exists in the un-fractured notch section is that which exists just prior to fracture. All of the tandem-notched sheet specimens were electrolytic ally polished and chemically etched prior to testing. The V-notched Charpy-type specimens machined from the 50-mil sheet material had the standard dimensions of 2.125 in. long, 0.394 in. deep, and a 0.010-in. root radius. The unnotched tensile and notched tensile tests were conducted over the temperature range of -78° to 300°C, in a 10,000-lb Instron Universal Testing Machine, at a constant crosshead speed of 0.020 in. per min. The V-notched Charpy-type

Jan 1, 1964

By John T. Eash

DURING the course of the war the supply of tin in this country has steadily decreased and a continued effort has been made since the beginning of the emergency to use alloys that are either tin free or contain appreciably lower quantities of it. Even with the use of emergency materials, the tin supply has become particularly acute, and it is estimated that after recovery of the far eastern mines probably one to two years will be required before our normal annual supply is attained. One of the major tin-consuming fields is the bronze industry. Recommendations were made several years ago by the War Production Board toward the use of bronzes of lower tin content and manganese bronze and yellow brass for the customarily used tin-bronze alloys. The substitution of 88-5-5-2 copper-nickel-tin-zinc for composition G received considerable attention before the war, and consequently this alloy and modifications of it have been adopted for some current structural and pressure castings requiring high mechanical properties. Another of the commonly used bronzes is the 10 per cent tin, 10 per cent lead copper alloy for heavy-duty bearings. It has been known for some time that about 2.5 per cent nickel could be substituted for an equal amount of tin with retention of equivalent casting and tensile properties; however, it was desirable that a still greater conservation of tin be made. Consequently the present paper deals with the development of an alloy of low tin content that could be used for bearings. SELECTION OF COMPOSITION While the final criterion of the suitability of an alloy for a bearing must 0f course be determined by practical applications where all the ramifications of service conditions are encountered, certain short cuts in the initial selection of compositions are necessary to avoid an extensive and time-consuming investigation. The present program has been to match certain metallurgical properties of the 80-10-10 bronze that have been shown by previous investigations to be related to good bearing qualities. Such properties as microstructure, hardness, compressive yield strength, ductility, and castability have been considered, and certain simple wear tests have been made. Of the readily assessed characteristics desired in a good bearing alloy, one can be certain that a uniform distribution of lead is required in a leaded bronze. It has generally been found that segregation of lead is one of the chief difficulties with this type of material. Tin has a powerful effect on lead distribution, and with 10 per cent present produces a random distribution of chunky particles, such as are shown in Fig. I. In addition, the tin bronze contains the hard alpha + delta eutectoid constituent distributed through-

Jan 1, 1945

By John T. Eash

During the course of the war the supply of tin in this country has steadily decreased and a continued effort has been made since the beginning of the emergency to use alloys that are either tin free or contain appreciably lower quantities of it. Even with the use of emergency materials, the tin supply has become particularly acute, and it is estimated that after recovery of the far eastern mines probably one to two years will be required before our normal annual supply is attained. One of the major tin-consuming fields is the bronze industry. Recommendations were made several years ago by the War Production Board toward the use of bronzes of lower tin content and manganese bronze and yellow brass for the customarily used tin-bronze alloys. The substitution of 88-5-5-2 copper-nickel-tin-zinc for composition G received considerable attention before the war, and consequently this alloy and modifications of it have been adopted for some current structural and pressure castings requiring high mechanical properties. Another of the commonly used bronzes is the 10 per cent tin 10 per cent lead copper alloy for heavy-duty bearings. It has been known for some time that about 2.5 per cent nickel could be substituted for an equal amount of tin with retention of equivalent casting and tensile properties; however, it was desirable that a still greater conservation of tin be made. Consequently the present paper deals with the development of an alloy of low tin content that could be used for bearings. Selection or Composition While the final criterion of the suitability of an alloy for a bearing must of course be determined by practical applications where all the ramifications of service conditions are encountered, certain short cuts in the initial selection of compositions are necessary to avoid an extensive and time-consuming investigation. The present program has been to match certain metallurgical properties of the 80-10-10 bronze that have been shown by previous investigations to be related to good bearing qualities. Such properties as microstruc-ture, hardness, compressive yield strength, ductility, and castability have been considered, and certain simple wear tests have been made. Of the readily assessed characteristics desired in a good bearing alloy, one can be certain that a uniform distribution of lead is required in a leaded bronze. It has generally been found that segregation of lead is one of the chief difficulties with this type of material. Tin has a powerful effect on lead distribution, and with 10 per cent present produces a random distribution of chunky particles, such as are shown in Fig' 1. I' In the tin bronze contains the hard alpha + delta

Jan 1, 1946

By John T. Eash

During the course of the war the supply of tin in this country has steadily decreased and a continued effort has been made since the beginning of the emergency to use alloys that are either tin free or contain appreciably lower quantities of it. Even with the use of emergency materials, the tin supply has become particularly acute, and it is estimated that after recovery of the far eastern mines probably one to two years will be required before our normal annual supply is attained. One of the major tin-consuming fields is the bronze industry. Recommendations were made several years ago by the War Production Board toward the use of bronzes of lower tin content and manganese bronze and yellow brass for the customarily used tin-bronze alloys. The substitution of 88-5-5-2 copper-nickel-tin-zinc for composition G received considerable attention before the war, and consequently this alloy and modifications of it have been adopted for some current structural and pressure castings requiring high mechanical properties. Another of the commonly used bronzes is the 10 per cent tin 10 per cent lead copper alloy for heavy-duty bearings. It has been known for some time that about 2.5 per cent nickel could be substituted for an equal amount of tin with retention of equivalent casting and tensile properties; however, it was desirable that a still greater conservation of tin be made. Consequently the present paper deals with the development of an alloy of low tin content that could be used for bearings. Selection or Composition While the final criterion of the suitability of an alloy for a bearing must of course be determined by practical applications where all the ramifications of service conditions are encountered, certain short cuts in the initial selection of compositions are necessary to avoid an extensive and time-consuming investigation. The present program has been to match certain metallurgical properties of the 80-10-10 bronze that have been shown by previous investigations to be related to good bearing qualities. Such properties as microstruc-ture, hardness, compressive yield strength, ductility, and castability have been considered, and certain simple wear tests have been made. Of the readily assessed characteristics desired in a good bearing alloy, one can be certain that a uniform distribution of lead is required in a leaded bronze. It has generally been found that segregation of lead is one of the chief difficulties with this type of material. Tin has a powerful effect on lead distribution, and with 10 per cent present produces a random distribution of chunky particles, such as are shown in Fig' 1. I' In the tin bronze contains the hard alpha + delta

Jan 1, 1946

By M. C. Flemings, T. Z. Kattamis

Dendrite morphology and grain size were studied in bulk samples of iron and nickel base alloys under-cooled up to 300°C. In the alloys studied, dendrite morphology gradually changes with increasing undercooling from the usual dendritic structure to a structure composed of cylindrical dendrite arms. At a critical undercooling of approximately 170°C in these alloys there is an abrupt transition to a fine-grained structure of spherical morphology. Final dendrite-arnz spacing of melts nucleated at less than the critical undercooling decreases with increasing undercooling and with decreasing distance from a chilled surface. Grain size of melts nucleated at greater than the critical undercooling decreases similarly with these factors. It is concluded that structure coarsening, with reduction of surface area as driving force, is the principal mechanism determining final dendrite-arm spacing in melts nucleated at small undercoolings and grain site in melts nucleated at large undercoolings. IN solidification of highly undercooled melts, nuclea-tion and growth of the solid is accompanied by a rapid release of latent heat of fusion. The heat release is sufficiently rapid in bulk specimens that it takes place essentially adiabatically with respect to the surroundings; hence the specimen heats up ("recalesces") to some maximum recalescence temperature. For bulk liquid metals and alloys, a simple thermal balance shows the undercoolings obtainable in bulk samples are not sufficient to lower the maximum recalescence temperature below the equilibrium solidus temperature. Hence, some liquid always remains at the end of recalescence; this liquid must solidify by rejection of heat to the surroundings. Metals, including iron and nickel base alloys, normally freeze with only a few degrees of undercooling. In general, nucleation begins on impurity particles present in the metal, at a "heterogeneous nucleation temperature" that is quite close to the equilibrium liquidus temperature. Very small droplets of metal can be undercooled considerably before nucleation occurs,1-4 presumably because subdivision of the liquid reduces the probability of solid impurity in a given droplet. Following early work on metal droplets,' a series of studies has shown that it is possible, without great difficulty, to undercool bulk samples of many metals and alloys. Undercoolings obtainable are approximately two tenths of the melting point for nickel, iron, copper, and other metals and alloys.5-9 Specimen sizes have generally been in the range of 50 to 500 g. Studies on undercooling to date have been concerned primarily with techniques of undercooling and with nucleation and growth kinetics. Only limited consideration has been given to dendrite morphology in under- cooled alloys. One such study, in bulk Cu-Ni alloy,8 showed that at small undercoolings a typical dendritic structure was obtained. At undercoolings the order of 175°C the structure of a Cu-40 pct Ni alloy became "ray or starlike", and a dense network of subbound-aries appeared. At still larger degrees of undercooling, of the order of 250°C, the structure became more random and included an ensemble of globular forms. Comparable results were obtained by walker10 on Cu-Ni alloys, and by Kamenetskaya et al.11 in research on Cu-Ni, Sb-Bi, Fe-Ni, and Fe-Cr alloys. In this work, study was made of the processing parameters influencing dendrite morphology and grain size in undercooled alloy melts. Experiments were specifically designed to show the effects of coarsening ("ripening") during solidification on final casting morphology. EXPERIMENTAL PROCEDURE Apparatus used was patterned after that of Walker,10 Fig. l(a). Melting was by induction under reduced pressure of helium. Two different types of crucible materials were used; there was no difficulty in obtaining substantial undercooling in both types of materials. Alumina crucibles were used for the bulk of the runs made, with a glass coating separating the metal from the crucible walls. Other runs were made using fused silica crucibles with or without slag coating. A graphite susceptor was employed around the crucible to aid in melting and to reduce melt stirring. The major portion of the experimental work was on

Jan 1, 1967

By J. P. Zannaras

Referring to the article by R. J. Charles and P. L. de Bruyn, let us assume that W = weight of glass bar; P = weight of hammer; e = total deformation; E = unit of deformation; K = potential stress energy; E = modulus of elasticity; L = length of the bar; 7, = coefficient of inertia; h = height, ft; V = velocity, fps; and a = cross section area. The portion of kinetic energy which is effective in producing stress energy in a fixed bar struck horizontally is given by the formula [ ]

Nov 1, 1956

By Orville R. Lyons

THE continuous vacuum filter, whether it be a vertical or horizontal disk, inside or outside drum, when properly applied is a very effective device for dewatering fine solids. Because of the number of variables that affect its performance, however, it is not possible to obtain consistently maximum moisture reduction in an operating plant. This paper presents a method for determining the average cake moisture and the variations in cake moisture to be anticipated when dewatering fine solids by continuous vacuum filters under normal operating conditions.

Jan 10, 1951

By K. Leutwyler

The authors present an interesting review of various applications of rather well known theories. Their rearrangement of the original working equations is strictly computer oriented and serves well to that effect; however, the new appearance will be confusing to the slide rule-wielding members of the profession. In reviewing discrepancies existing between calculations made by Baker Oil Tools, Inc. field engineering staff members and computer output submitted based on the authors' program, the following error in concept was discovered in Eq. 4. The authors quote the length change due to helical buckling in the following form We appreciate Kurt Leutwyler's awareness of Archimedes' principle, and concur that the final term in the denominator of Eq. 4 results in a small error. Examples checked revealed less than I ft error, and corrective expressions for the four possibilities of displacement follow. The portion of the denominator within extra large brackets represents the original term W,, = weight of outside fluid displaced per unit length. The authors interpreted the expression "displaced" in general oilfield terminology as a fluid displacement, such as "mud displaced with water, etc." A review of the theory forming the basis for these helical buckling calculations discloses that this displacement is meant to be interpreted in the Archimedean sense, i.e., fluid displaced by the solid body. here the tubing and the seal nipple. Lubinski* illustrated the point by summing moments due to mechanical and hydraulic forces acting on a free body immersed in fluid. This moment summation, always in the Archi,medean sense, yields effects of the real forces, including W,, and the fictitious force on the bending mechanism. I would be interested to hear the authors' comments regarding this theoretical point, which nevertheless is not mute since its lack of recognition seems to result in erroneous calculations. We must not remain mute to Leutwyler's suggestion that these equations, although computer oriented, are confusing to the slide rule engineers. Such input information as density in pounds per gallon, pipe diameters in inches, pumping rate in barrels per minute and pipe length in feet is not confusing to most of us. Yet there may be some rugged individuals in field positions who carry tables of conversion constants and have committed to memory such items as coefficients of thermal expansion and Poisson's ratio; however, even this type of individual would welcome a simplified equation occasionally. M

By H. M. Zoerb

Based on the belief that operating details are a definite contributing factor to major economies, this paper traces the development of crushing cavity design in Symons cone crushers to attain maximum liner utilization. Wear rates are analyzed and compared in this presentation and drawings illustrate succeeding design changes. IN these times of rising labor and material costs, it has become more and more necessary that attention be paid to some operating details which, in their obscurity, may he the key to major economies. Liner wear in crushing cavities of secondary and tertiary crushers can become an appreciable cost item when the material to be crushed is hard and abrasive. This item of cost not only includes the value of the crushing members, but also more intangible costs such as labor and lost production due to more frequent replacement. The variables which are encountered in ores and minerals to be reduced; the design of plant and machine application; the sizes, shape, and fineness, characteristics of the crushed product; the moisture; hardness; friability; and abrasiveness of the material to be crushed are all influencing factors which must be taken into consideration in the selection of a crusher, and particularly in the design of crushing cavity and liners to be used in a crusher. Through a research program undertaken in cooperation with many operators of Symons cone crushers a new approach to crusher cavity design was made, resulting in the development of liners for specific operations which showed: 1—maximum utilization, as high as 70 to 80 pct of original weight of metal, and 2— maximum capacity of unit during the greater portion of its life. It has been found that liners so designed for a given operation will show added economies in power consumption, maintenance, and general wear and tear on the crushing unit. Initial work in the so-called tailoring of crushing cavitles was begun on the tertiary or fine crushing units where as a rule reduction ratios were low, varying from 3 to 6. Parallel or sizing zones in the lower portion of the crushing cavity were too long, resulting in a tendency to pack. It was found that very little additional crushing was done in the parallel zone after the initial impact in that zone and that a relatively small amount of' additional crushing was done by attrition, which required very careful feed control. A small amount of over-feeding would result in packing which not only consumed power but caused unnecessary liner wear as well. The illustrations which follow in this discussion will show only contours of crushing cavities, and for purposes of simplification the cavities will be considered only in their closed position. The first step, therefore. was to reduce the sizing zone to a minimum. This was done by removing the lower portion of the liner as shown in Fig. 1. The result of the change was a saving of 15 to 20 pct in liner cost, less power consumption, with no change in capacity. This change in design, while an improvement, did not go far enough. As wear took place, the change in the liner was not uniform throughout its entire length, resulting in a restriction of the feed opening and thereby loss of capacity. Furthermore, progressive wear of the liner had the effect of lengthening the parallel zone until finally the entire crushing cavity was all parallel zone, see Fig. 2. It is obvious from the reduced feed opening of the worn liner that the ability of the machine to receive material is lessened considerably. Furthermore, the long parallel zone with its worn, irregular profile did not operate at its highest efficiency. The first attempt to overcome this difficulty was carried out on a 5 1/2-ft crusher installed in a plant producing roofing granules. The material being crushed was a very hard graywacke and the crusher was closed-circuited with a screen having .232-in. slotted openings. A radical change in contour was developed, as illustrated in Fig. 3. Equal wear lines on both concave and mantle are designated 1, 2, 3, etc. The method of development of this contour is as follows: Since adjustment for wear is vertical, corresponding intersections of wear lines and vertical lines developed concave and mantle contours which maintained equal but lengthening wear surfaces in the parallel zone. The ideal contour, of course, is one in which the length of the parallel zone remains constant, but because of present foundry practice and heat treating characteristics this is impossible.

Jan 1, 1954

By R. A. MacGregor

The Stanrock mine of Stanrock Uranium Mines Ltd. is located on Quirke Lake in the Elliot Lake area of northern Ontario, about 25 miles north of Lake Huron. The property was developed and brought into production at a rate of 3000 tons of ore per day during the uranium boom of the mid-1950's. The mine workings are from 1200 to 2100 ft below sea level. Mining was done by the room-and-pillar method with 65 to 80% extraction. The plant is located at an elevation of 1400 ft above sea level. The orebodies are conglomeratic beds, 6 to 40 ft in thickness and dipping 5º to 40º. The ore is a quartz pebble conglomerate with a pyritized sericitic matrix. The ore minerals are brannerite, frequently intergrown with anatase or rutile, uraninite, and some monazite and minor coffinite and thucolite. Sulfides constitute 5 to 10% of the ore as pyrite with minor amounts of pyrrhotite, chalcopyrite and galena.

Jan 3, 1969

By Niels Hansen

Dispersion-strengthened sintered aluminum products are normally manufactured from surface-oxidized aluminum powder, SAP,' or from atomized aluminum powder.' The powder-blending technique consisting in mechanical mixing of the constituents has been used in the manufacture of numerous dispersion-strengthening alloys, but has never been successfully applied to aluminum alloys. This technique may, however, be applicable if it is possible to produce an unagglomerated mixture of the constituents in the form of fine powders, and the present study was initiated to test this assumption in the aluminum/aluminum oxide system.3 Table I lists the two grades of aluminum powder and the aluminum oxide powder used in the manufacture of the alloys. MD 105 A is an aluminum powder with an average grain size less than 5 IJ., which is isolated from the MD 105 powder by centrifugal separation in cyclohexanol. The aluminum powders and the aluminum oxide were blended in a high-speed blender and then subjected to cold compaction and hot compaction at 550" to 600°C. Subsequently the compacts were hot-extruded with an extrusion ratio of 15 to 1. The alloys were evaluated by tensile testing at room temperature and at 400°C, microscopic examination, and heat treatment at 500°C followed by tensile testing at room temperature. For the purpose of comparison, a commercial SAP quality supplied by Swiss Aluminium Ltd. was tested in the same manner. The tensile data at room temperature and at 400°C are plotted in Figs. 1, 2, and 3 as a function of the weight percent of oxide. The oxide content

Jan 1, 1964

By R. L. Vaughn, A. G. Moon

In 1972 a major renovation of the flotation and regrinding circuits was completed at Kennecott Copper Corporation's Chino Mines Division concentrator at Hurley, New Mexico. At that time a new lime distribution and control system was placed in operation. Lime is used to depress pyrite, and flotation at a high pH or alkalinity is required for optimum recovery of principal copper minerals, chalcopyrite and chalcocite with good rejection of the pyrite. Regrinding is also required to liberate chalcocite, much of which occurs as intimately associated chalcocite-pyrite middling grains. Accurate control of the lime addition is necessary to obtain both optimum metallurgy and minimum lime usage. At low pH levels, flotation cell control is difficult due to excess froth, but excessively high pH levels can go unnoticed because froth characteristics do not change from those experienced at the desired pH. With manual control of the lime addition, the desired pH can be exceeded for periods of up to two hours. Test Work Plant test work on a lime control system using conventional pH probes gave disappointing results with high electrode maintenance and poor control characteristics. For example, the rougher flotation cells are operated at a pH between 11.4 and 11.6 and the probes required cleaning at least once a day, and occasionally every shift. These high maintenance requirements were considered unreasonable for a plant-wide system using six or more control loops.

Jan 1, 1979

Insoluble slimes are present in nearly all sylvinite ore bodies and must be removed prior to potash flotation because they adsorb excessive amounts of potash flotation reagents and cause low potash recoveries and grades. The problems associated with insoluble slimes will become more pronounced as high grade-low insol potash ores are depleted and lower grade ores containing high insols are processed. The most common commercial practice for insoluble slimes removal is mechanical desliming (hydroclassification); however, most of these desliming circuits were designed to treat sylvinite ores containing less than 4 percent insoluble slimes. As Mr. Banks mentioned in his presentation, higher insoluble slimes concentrations (>4 pct) in the mill feed overtax the mechanical desliming circuit and result in excessive brine and potash losses. Increasing the capacity of the circuit would require redesigning of the entire potash circuit and additional costly equipment. Because of mineralization and liberation properties of the high insols sylvinite deposits, the expansion of the desliming circuit may not economically maintain high potash grades and recoveries. Selective flocculation-flotation is an excellent alternative for removing insoluble slimes from potash ores. Cominco's ore, containing 4.5 to 8 percent insoluble slimes, is particularly amenable to the slimes flotation process, probably because of the absence of expanding-type clays such as montmorillonite. The overall potash recovery of 86 percent and low slime flocculant and collector dosages of 0.03 to 0.04 pound reagent per ton of feed is good evidence of the success of the process. If appreciable montmorillonite were present, the metallurgy would not be as favorable and reagent consumptions would increase significantly.

Jan 1, 1979

By E. J. Duggan

It is a coincidence that the two plants in this country for treating copper ores by ammonia leaching came into existence at almost the same time, that of the Kennecott Copper Corporation at Kennecott, Alaska, in March, 1916, and that of the Calumet & Hecla Mining Co. in July of the same year. The success of both from their very beginning attests the merits of the process under widely different sets of conditions. As would be expected, in the time they have been operating these two pioneers in this field have greatly perfected their methods and improved their technic. The early difficulties inevitably encountered in undertakings of this nature have been largely surmounted and the process may now be considered to be well seasoned and matured. There have been no other commercial plants built in this country and but one elsewhere, the Bwana M'Kubwa in Northern Rhodesia, for the reason that this method of treatment may be employed on comparatively few ores. However, it has certain advantages, some of which are inherent to sand leaching processes generally and some to ammonia leaching in particular. Among these may be enumerated: (1) the avoidance of fine grinding; (2) the small amount of water required; (4) a high-grade product; (5) little likelihood of deleterious gangue constituents and fouling of solutions. It is true that great progress has been made in recent years in the flotation of the ores that come within the province of ammonia leaching. Nevertheless ammonia leaching has its possibilities and these may be overlooked in the present tendency toward the use of flotation for all ores. Probably the chief requisite to its further extension to suitable ores will be a cheap source of steam. Ammonia, the only reagent required, is unlimited in supply, readily transported and the price that is likely to obtain in the future should not make its cost at all onerous. Choice of Method for Kennecott Ores The Kennecott ores are typical limestone replacement deposits. The milling ores have a copper content of approximately 8 per cent in a limestone dolomite gangue. About half the copper occurs as sulfide, mostly chalcocite with some covellite, and the remainder in the form of the two carbonates, malachite and azurite. A comparatively coarse grind followed

Jan 1, 1934

By A. N. Holden, J. H. Hollomon

Inhomogeneous yielding during the early stages of plastic flow has been observed in many metals and has long been a subject of controversy. Low carbon steel, when strained at room temperature, exhibits a distinct yield point at which the metal ceases to behave elastically and begins to flow plastically without further increase in stress,† often with an actual decrease in stress from that existing at the yield point. One manifestation of this behavior is the familiar Lüders lines. The extent of the yield point elongation (the plastic strain which results without raising the stress above that at the initial yield point) may be several percent. Several explanations of this type of yielding in steels have appeared in the literature:1-3}; 1. The segregation of carbides or other constituents at the grain boundaries of ferrite forms cell-like blocks which break down at a higher stress than that at which the ferrite alone will flow. 2. The precipitation of various constituents within the ferrite itself in some way blocks initial flow until a higher stress is reached, after which a great deal of flow occurs without further increase in stress. 3. The restraining influence of neighboring, less favorably oriented grains in a polycrystal results in a storing up of energy, which on yielding is sufficient to cause the continued propagation of slip bands on the multiplicity of slip systems available in adjacent iron crystals without increase in stress. 4. Local segregation of carbon atoms increases the stress required for the initial propagation of dislocations. The stress required for further deformation is lower.3 Such an explanation would require the presence of upper and lower yield points in single crystals. Low and Gensamer2 have shown that heterogeneous yielding can be eliminated by removal of carbon and nitrogen from polycrystalline steels by wet-hydrogen treatment. They also demonstrated that the re-introduction of minute amounts of carbon or nitrogen caused a return of the yield point. Observations of the initial flow of single crystals containing small amounts of carbon and nitrogen- should provide a clue to the explanation of inhomogeneous yielding. Previous tests of iron crystals4-6 can not be considered conclusive because it is uncertain whether the carbon and nitrogen were satisfactorily removed by the decarburization technique employed. The purpose of the experiments described herein was to determine whether single iron crystals containing either carbon or nitrogen would exhibit in-homogeneous yielding. Experimental Procedures THE PREPARATION OF SINGLE CRYSTALS All crvstals were formed in the following way. Aluminum killed sheet material 80 mils thick was machined into specimens of the outer contour shown in Fig 1. These mildly tapered specimens were decarburized at 730°C for 16 hr in hydrogen that had been bubbled rapidly through water maintained at 70°. The resultant grain size after decarburizing was approximately ASTM 5. Several stress-strain curves were made of the decarburized niate-

Jan 1, 1950