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By J. PARKCEH CHANNING

Discussion of the paper of Lewis T. Wright, presented at the New Haven meeting, February, 1909 (Trans., xl., 492 to 495). J. PARKE CHANNING, New York, N. Y. (communication to the Secretary*):-Mr. Wright, in his introductory paragraph, says: " It is commonly believed by metallurgists that in copper-smelting, the copper in the slags, which is irreducible by continued smelting, is retained in the form of 'prills' of matte." On a recent visit to Greenwood, B. C., I was discussing with J. E. McAllister, the general manager of the British Columbia Copper Co., and the former metallurgist of the Tennessee Copper Co., this very question; and he held strongly to the idea that the copper in slag was in two forms-one portion contained in occluded matte and the other as an oxide. He believed that this same rule would apply to the silver in the slag. From my experience I am strongly inclined to agree with Mr. McAllister; and it appears quite reasonable that in furnaces treating oxidized copper-ores, and to a less degree in furnaces treating sulphide ores, there is always bound to be present a certain amount of copper oxide and silver oxide, which will behave like any base and get into the slag as such. We all know that in Arizona, in the early days, when it was customary to produce black copper, because the ores were nearly all oxidized or carbonates, and sulphur was scarce, the slags as a rule carried 2.5 per cent. of copper and seldom ran less than 1.5 per cent. of copper. The economic point was determined by the extra coke necessary to produce the reducing action. In connection with the extra saving in copper, Arthur L. Walker, now professor of metallurgy at Columbia University, and formerly general manager of the Old Dominion Copper Mining Co., at Globe, told me that with coke at $60 per ton it never paid him to make slags containing less than 2.5 per cent. of copper. *Received Oct. 8, 1909.

Feb 1, 1911

By Dorsey A. Lyon, Robert M. Keeney

Page I. Introduction..........234 I1. Chemistry oF Copper Smelting...235 1. Native Copper Ores.......235 2. Oxide and Carbonate Ores....235 3. Sulphide Ores.......236 (a) Elimination of sulphur ...236 III. Reverberatory Copper Smelting...237 IV. Substitution of the Electric Furnace for the Reverberatory Furnace..................... 238 V. Smelting of Fine Michigan Native-copper Concentrates in the Electric Furnace.................. 239 1. Conclusions.................240 VI. Proposed Process for Electric melting of Fine Native-copper Concentrates.................. 241 VII. OrdINary Copper Blast Furnace 'Smelting......... 245 1. Prerequisites.................. 245 2. Losses .................... 247 VIII. Experimental Data on the Electric Smelting of Raw Sulphide. Oxidized and Carbonate Ores............. 248 1. Vattier's experiment................ 248 2. Wolkoff's experiments................ 249 3. Experiments of Stephan.............. 250 (a) Conclusions of Stephan ............ 251 IX. Smelting of Sulphide Ores in the Electric Furnace...... 251 1. Ores Treated.................. 252 2. Conclusions................... 253 X. ELectric Furnace vs. the Reverberatory Furnace and Blast Furnace as a Melting Agent.............. 253 XI. Smelting of Sulphide Ores without Roasting or Partial Pyritic Smelting............254 XII. The Electric Blast Furnace vs. the Copper Blast Furnace for Partial Pyritic Smelting.......259 1. Metallurgical conditions......259 2. Mechanical considerations......259 3. Comparison of costs.......260 (a) General considerations .....260 (b) Cost of plant........260 (c) Amount of electrical energy...260 (d) Cost of smelting.......261 (e) Cost of electrodes........262 XIII. The Use of the Electric Furnace in the Smelting of PyRITIc ORES : 262 XIV. Summary...................... 264

Jan 1, 1914

By R. G. Aspden

Crystals with an (001) [loo] initial orientation of an iron-base alloy containing 3 pct Si were cold rolled with and without the use of constraints. A major difference in the rolling and annealing textures was observed between crystals rolled with and without constraints. These data show that the contribution of constraints at grain boundaries in a poly crystalline sheet should be considered in applying textural data on single crystals to grains in an aggregate. SILICON-iron alloys with a cube texture have been recently developed and their magnetic characteristics reported.1-4 Of interest in the development of this texture were the textural changes of single crystals accompanying rolling and annealing and the influence of constraints at grain boundaries in an aggregate on the behavior of individual grains. The present study was primarily concerned with the effect of constraints during rolling on the textures of 3 pct Si-Fe crystals initially (001)[100]. Barrett and Levenson5 were among the first to observe an influence of constraints at grain boundaries on the textural changes of individual grains during deformation. They tested Taylor's6 theory of plastic deformation of face-centered-cubic metals in which deformation textures were predicted. About one-third of the grains in poly crystalline aluminum did not rotate as predicted. Grains of the same initial orientation were observed to rotate in different directions under the influence of applied stress and anisotropic flow of neighboring grains. Recently, the various inhomogeneities of flow of crystals in an aggregate have been studied7'8 and reviewed.9-11 Barrett and Levenson" rolled (001) [loo] iron single crystals inserted in close-fitting holes in copper to limit lateral flow and to simulate rolling of grains in an aggregate. Deformation bands were formed after a 90 pct reduction in thickness, and the cold-rolling texture contained two components described by rotating the (001)[100] about 35 deg in both directions around the normal of the rolling plane. No annealing textures were reported. Chen and Maddin13 rolled molybdenum single crystals initially (001) [loo]. The crystals were mounted between two hardened silicon-iron plates and 96 pct reduced in thickness by rolling at a low rate of reduction, about 0.0001 in. per pass. The deformation texture had the mean orientation of (001) [loo], and the azimuthal spread included orientations described by rotating (001) [loo] about 35 deg in both directions about the pole of the rolling plane. The presence of deformation bands were not reported by Chen and Maddin or detected in subsequent work of Ujiiye and Maddin.14 The ideal orientation of the annealing texture was (001) [loo]. Recently, Walter and Hibbard 15 reported on the textures of 3 pct Si-Fe alloy crystals initially near (001) [loo]. Each crystal was in an aggregate cut from a columnar ingot. After 66 pct reduction by rolling, the texture consisted of two symmetrical components which had the orientations described by rotating (001) [loo] about 30 deg in both directions about the pole of the rolling plane. Annealing texture was near (001) [loo]. In the above work, the textures of body-centered-cubic crystals were studied after rolling under the influence of constraints. The deformation textures varied from (001) [loo] to near the (001) [110] type and appeared sensitive to the manner in which the crystals were rolled. No textural data were available on the effect of rolling (001) [loo] crystals with and without constraints. The purpose of the present work was to evaluate the influence of constraints during rolling on the textures of 3 pct Si-Fe crystals initially (001) [loo]. Rolling and annealing textures were studied for a) crystals rolled with no constraints at different rates of reduction, and b) crystals rolled with constraints imposed by neighboring grains and by plates between which a crystal was "sandwiched". PROCEDURES AND EXPERIMENTAL TECHNIQUES Data are presented on four crystals which are representative of several crystals studied. The orientation of each crystal prior to rolling was (001) [loo] as determined by the Laue X-ray back-reflection method," i.e., each crystal had an (001) within 3 deg of the rolling plane and [100] within 3 deg of the rolling direction. These crystals were obtained from two iron-base alloys containing 3 pct Si by weight which were prepared by vacuum melting electrolytic iron and a commercial grade of silicon. Crystals 1, 2, and S-1 were cut from a large single crystal grown from the melt of one alloy by the Bridgman technique17 in an apparatus described by

Jan 1, 1960

By N. Jaffee, J. E. Dorn

An activation energy of 27,400 5 1000 cal per mole was obtained for the creep of poly crystaLline aluminum over the temperature range of 273° to 350°K, at strains varying from 0.003 to 0.230. Stresses varied fronz 2250 to 6000 psi, strain rates from 0.1145 x 10-5 to 29.5 x 10~s min.-1. A strain temperature-compensated time correlatzon was oblained indicating that the structure was depedent solely upon the strain at a given stress; X-ray diffraction analyses and tensile tests on precrept specimens verified this. PREVIOUS investigations1 have revealed that the apparent activation energies for creep of high-purity polycrystalline aluminum are insensitive to the applied stress, strain, and strain rate. On the other hand, the apparent activation energy is dependent on the temperature1,2 as shown by the data reproduced in Fig. 1. Each plateau of this curve is associated with a unique rate-controlling dislocation mechanism for thermally activated creep. It is now very well-established that over the high-temperature range, where the activation energy is about 35,500 cal per mole, the creep rate is controlled by the dislocation climb mechanism. 3 Over the intermediate range of temperatures, from about 260° to 370°K, another unique activation energy, namely 28,000 cal per mole, is obtained. Current evidence strongly supports the contention that cross slip is the rate controlling mechanism for creep in this case: 1) Crystal recovery can occur by a number of mechanisms prominent among which are climb and cross slip. Astrom,4 using the Borelius micro-calorimeter, reported rapid crystal recovery of cold-worked polycrystalline A1 at 340° to 370°K giving an activation energy of 28,000 cal per mole and additional recovery at 455° to 478°K giving an activation energy of 3 6,000 cal per mole. Whereas the latter recovery process was clearly associated with the dislocation climb mechanism, the former, 28,000 cal per mole process could have arisen as a result of the cross-slip mechanism. 2) Single crystals of high-purity A1 favorably oriented for octahedral glide exhibit an activation energy for creep of about 3400 cal per mole over the range from 0° to 400°K and that of about 28,O:O cal per mole over the range from 600° to 775°K . Whereas the lower activation energy process is accompanied by the development of sharp slip-band traces, the 28,000 cal per mole process was uniquely characterized by extensive cross slip. 3) Theoretical calculations by Seeger and schoeck6 suggest an apparent activation energy for cross slip in A1 of about 24,000 cal per mole and another theoretical estimate by Friedel 7 gives the observed 28,000 cal per mole as the theoretical activation energy. It is therefore consistent with this evidence to, at least tentatively, ascribe the 28,000 cal per mole creep process in A1 to the operation of cross slip as the rate controlling mechanism. Although creep has been rather well-explored in the dislocation climb range, very little, excepting perhaps the activation energy, is known about creep in the cross slip range. This investigation was undertaken, therefore, to reconfirm the activation energy for creep in polycrystalline A1 in the intermediate temperature range and to study the strain-time relationship for creep at a constant stress in the cross slip region. EXPERIMENTAL PROCEDURE AND TECHNIQUE The three types of high-purity A1 alloy specimens described in Table I were used in this investigation. Specimens were milled from 0.100 in. thick cold-rolled sheets with their tensile axes in the rolling direction following which they were given a recrys-tallization and grain growth heat treatment at temperatures well above those which were subsequently employed in the creep tests.

Jan 1, 1962

By Harold M. Feder, Irving Johnson

Tkermodynamic functions for dilute solutions of uranium in liquid cadmium were obtained from galvanic cell measurements. Deviations from Henvy's law were observed at concentvations down to 2 x 10-3 atom fraction. Unusually large negative values of the pavtial molal relative enthalpy and excess entvopy of uranium, suggestive of the occu,vvenctwe of partial ovdering of cadmium around uranium in these solutions, were found. The negative relative partial molal enthalpy of uranium at the liquidus accounts for the observed retrogvade solubility of metallic uranium between 473" and about 630°C. The standard free enevgy of formation of UCd,, from a-uranium and liquid cadmium is given by F,° THE constitutional diagram of the U-Cd system,' Fig. 1, is unusual in several respects. The near equality of the Goldschmidt atomic radii2 (Cd, 1.52A U, 1.56A) might be expected to give rise to extensive mutual solubility of the elements except for the restrictions imposed by differences in valence and electronegativity.3 The latter factors must be strongly influential because mutual solubilities in the terminal solid phases are probably negligible, and the solubility of uranium in liquid cadmium in the readily accessible range is only about 1 at. pct. Interestingly, the solubility of uranium in liquid cadmium actually decveases slightly with increasing temperature from 473" to about 630°C. Such a retrograde solubility in a liquid phase is a rare phenomenon in metallic systems; no other authenticated examples were found among the nearly 700 binary phase diagrams compiled by Hansen and Anderko.2 These observations suggested the existence of some unusual interaction between uranium and cadmium. Greater insight into the nature of this interaction was sought by measurement of the thermodynamic properties of the system. EXPERIMENTAL PROCEDURE The emf of cells denoted by the scheme was measured. Each cell consisted of two uranium and two alloy electrodes immersed in a molten lithium chloride-potassium chloride eutectic mixture containing uranium trichloride. The physical arrangement is shown in Fig. 2. The potential between any pair of electrodes could be measured with a Leeds and Northrup Type K-3 potentiometer. The measurements were made at points marked in Fig. 1. It should be noted that points on the phase boundaries represent the concentrations of uranium in the saturated liquid phases of heterogeneous alloys. Uranium Electrodes. Uranium, which is not a particularly soft metal, is capable of retaining residual stresses, and its potential may depend upon its physical condition. Electrodes were prepared from 3-mm diam rods of hot-rolled high-purity4

Jan 1, 1962

By W. C. Hellyer, R. A. Campbell

In single-stage autogenous grinding, the buildup of a critical size fraction in the media can be corrected by removing this material through pebble ports, crushing it below the critical size range, and recycling the crusher product back to the mill. The rate at which the critical sire fraction is crushed affects the size distribution of the grinding media and this in turn affects the sire distribution of the grate discharge. This provides a means for controlling the grind produced by a single-stage autogenous grinding unit. The pilot-plant investigations on a very hard copper ore were carried out at the facilities of the Institute of Mineral Research, Michigan Technological University. In an autogenous grinding circuit in which feed at approximately 9 in. top size is reduced to a size suitable for subsequent processing, the build-up of a "critical size" fraction in the media causes problems. A "critical size" fraction has been defined as, "media too small to effect reduction by impact grinding of ore coarser than a quarter of an inch and too large to be broken by the largest size of media in the charge."' The buildup of a critical size fraction reduces the capacity of the mill, increases the grinding power requirements per ton of finished product, and generally produces a finer grind than is desired. It is general practice to overcome this problem by the addition of large-diameter steel balls to the grinding charge. This has certain disadvantages, such as 1) an increase in mill liner wear, 2) wear on the steel balls, and 3) some loss in flexibility in grinding circuit opera-tions resulting from difficulties in removing the steel balls by means other than by grinding out. A number of investigators' have suggested the use of a small crusher as an alternative to the use of large steel balls for control of a critical size fraction. With this technique the critical size fraction is removed from the mill continuously through suitable-sized pebble ports, crushed below the critical size range, and returned to the mill. An external crusher would cause no increase in mill liner wear, the wear on the crusher would be less expensive than the wear on the steel balls, and the flexibility of the grinding circuit operations would be enhanced since the crusher can be cut into and out of the circuit at will. This paper describes an autogenous grinding pilot plant investigation on a very hard copper ore, which led to the selection of an autogenous grinding flowsheet incorporating a crusher in the circuit as the preferred method for grinding this ore. Further development of the technique demonstrated that the crusher could be employed to control the product produced by an autogenous grinding unit. Description of Pilot Plant These investigations were carried out at the pilot-plant facilities of the Institute of Mineral Research, Michigan Technological University, Houghton, Mich. The autogenous grinding unit was a 6-ft-diam by 2-ft-long Hardinge Cascade mill with %-in. slotted steel grates. Pebble ports cut into the grates allowed passage of pebbles having a top size of approximately 21/2 in. The grate and pebble-port discharge material passed over a 3/16-in, trommel screen with the trommel under -size being pumped to the classification device. DSM screens were employed for classification in the early investigations, but were replaced later by a small Dorr rake classifier. The trommel oversize material returned to the mill via scissor conveyors. When the crusher was incorporated into the circuit, the trommel oversize material passed to a double-deck vibrating screen fitted with 2-in. and 3/4-in. square mesh screen cloths. The $ 2-in. and the —2 +3/4-in. size fractions were combined and crushed to a nominal 1 in. in a 4 x 6-in. jaw crusher. The crusher discharge and the —3/4-in. screen undersize returned to the mill. In the final investigation, a flap gate fitted to the top deck of the screen directed the — 21/2 +2-in. size fraction to the crusher or back to the grinding mill. The flap gate operated manually on a 15-min cycle, with this material being crushed for so many minutes out of each cycle. This was found to be a more reliable method of controlling the amount of — 21/2 +2-in. material crushed than attempting to make a split of a small weight of material on a continuous basis. Operational Techniques: Each sample was sized into three fractions and the Cascade mill feed was reconstituted from these size fractions in the proportions existing in the original sample. Sufficient feed for 15 min operation was weighed out and fed by hand over a 15-min period. The gross mill power draft was recorded every 15 min, corrected for tare power and drive efficiency, and reported as net kilowatt-hours per ton. A recording kilo-watt-hour meter provided a continuous visual record of the power drawn by the mill. Pulp densities of the trommel undersize and the classifier overflow (or DSM undersize) were taken every 15 min. The classifier overflow was sampled automatically. Timed samples of the classifier sands were taken every 15 or 30 min, weighed, and returned to the circuit. After correcting for moisture content, the weights were converted into percent circulating load. Timed samples of the +2 in., the —2 + 3/4-in., —3/4-in. screen fractions, and the crusher discharge, were

Jan 1, 1971

By Y. Dardel

Since the first determination of Dumas1 in 1880, many authors have tried to measure the solubility of hydrogen in solid aluminum, or at least the amount of dissolved gas in it. However, the interpretation of the results which have been obtained is still difficult. Sieverts? then Baukloh and Oesterlen,a during their gassing experiments could not determine any solubility of hydrogen in solid aluminum. On the other hand, Portevin, Chaudron and Moreq4 leaving the determination of the hydrogen solubility beyond the scope of their experiments, measured only the amount of gas that they could extract by high voltage positive ion bombardment and found up to 151 cm3 of gas per I00 g of metal. Winterhager,6 by repeated vacuum extraction during severe cold rolling, extracted up to 161 cm3 per 100 g. In both series of experiments, the extracted gas contained chiefly hydrogen with a small amount of nitrogen, carbon monoxide, carbon dioxide and methan. Later, Schmidt and von Schweinitze as well as Chaudron and Moreau7 showed the importance of the adsorbed layer of hydrogen, the extracted quantities being proportional to the surface and not to the weight of the sample, when its cleaning was not perfect. In earlier experiments, Steinhausers extracted from more massive samples nothing but a little amount (< 0. 5 cm3 per I00 g) of hydrogen, carbon monoxide and methane. Thus, according to these experiments the quantity of dissolved gas is probably very small and the effect of the surface could falsify the results of the extraction. However, owing to the care taken by Chaudron and his coworkers, it was clear that all the hydrogen extracted could not come from the superficial layer, or that the links which bound hydrogen to the surface were so strong that the usual method of cleaning could not destroy them. After the main parts of the experiments reported here had already been carried out, Eborall and Ransley9 published the results which they obtained with an analysor having an accuracy of 0.03 cm3 per I00 g, while that of the usual apparatus was only about 0.2 cm3 per I00 g. During all their experiments the adsorbed layer would have been a difficulty, for the value of the adsorbed hydrogen was approximately equal to that of the dissolved amount. Table i—Amount of Adsorbed and Dissolved Hydrogen (After Eborall and Ransley) Extraction Hz Adsorbed „ h2 dissolved Temperature Calculated Extraction OCcma 500 0.057 0.050 600 0.062 o.096 Since always keeping the same adsorbed layer is very difficult, an error in the correction for the adsorption can be made when using the apparatus of. Eborall and Ransley, as well as the earlier apparatus.

Jan 1, 1949

By W. J. Joslin

The frontal-advance equation can determine how the fluid withdrawal rate and subsurface operating pressure influence oil recovery from pressure-maintained reservoirs having characteristics favorable for efficient vertical segregation. For effective vertical segregation in formations having low dip angles, the requirements are: thick, massive sands, low viscosity oil and high vertical permeability with few discontinuous shale members. This approach has been successfully applied to the LL-370, which is a massive Eocene sandstone reservoir containing crude with an average gravity of 26" APT, a dip of only 3" and vertical permeability of ½ darcy. The average gas cap saturation was calculated for various rates of vertical gas-oil contact movement corresponding to a wide range of reservoir producing rates. This method also evaluated the effect of different pressure maintenance levels on oil recovery. Applicability of this procedure for calculating the present oil saturation in the LL-370 gas cap has been substantiated from field performance, where the oil saturation was computed by comparing oil recovery with the gas cap volume vacated. INTRODUCTION During the natural depletion history of the LL-370 reservoir, located in Lake Maracaibo, Venezuela, excellent vertical segregation resulted in the formation of a thin gas zone over a substantial portion of the oil-productive area. Subsequently, gas injection forced this thin gas-saturated zone slowly downward, reducing the oil saturation in the gas cap to a low value. The oil recovery has been raised considerably above earlier predictions where vertical segregation effects were under-estimated. In order to determine the maximum efficient producing rate and optimum operating pressure for this reservoir, a vertical gas frontal-advance analysis has been made through application of the fractional flow equation. LL-370 RESERVOIR CHARACTERISTICS The reservoir consists of a truncated monocline of massive. highly-permeable Eocene sandstone, called the B-b formation, covering an area of more than 13,500 acres. The original oil in place was 2.17 billion STB. Internal non-sealing faults subdivide the structure into four blocks called the B-6-x.6, 10, 11 and 13, which are effectively bounded by major faulting on the flanks, an unconformity updip to the west, and an immobile aquifer to the southeast. The producing formation occurs at an average depth of 5,250 ft subsea and has a closure of more than 1,500 ft. A typical producing section has a net oil sand thickness of 170 ft. Scattered throughout the section are a few randomly-positioned shale lenses which increase in thickness and areal extent downdip. The reservoir rock and fluid properties vary with depth. The permeability and produced oil gravity are 1600 md and 28" API at the top, compared to 300 md and 18° API at the water-oil contact. The productivity of updip wells is, therefore, much higher than the reservoir average. The reservoir oil was originally saturated at the crest, grading to a slight degree of undersaturation at the water-oil contact. There was no original gas cap; and material balance and the field performance indicate no water influx. Except for the northern B-6-x.6 block, the fluid and rock properties have already been published.&apos; Average LL-370 reservoir properties are shown in Table 1. RESERVOIR PERFORMANCE UNDER NATURAL DEPLETION Fig. 1 shows production performance from 1939 to date. Early development was gradual, with 30 wells pro-

Jan 1, 1965

By A Andersen

DURING an investigation of the copper-rich portion of the copper-silicon-iron system as part of an extensive research program on P.M.G. alloys, which was begun in 1937 in the research laboratory of the Phelps Dodge Corporation, it became apparent that the binary copper-silicon system below 7 per cent Si needed revision, therefore an investiga-tion was undertaken. Both X-ray methods and microscopic technique were employed. Despite the many investigations that have the clarification of this sys-tem as their object, no two successive reports on the system have agreed in placement of the alpha-phase solubility limit, and uncertainty about the secondary phase first to appear has been voiced several times.3,4,5 It was therefore not surprising, when T. Isawa&apos;s6 and M. Okamoto&apos;s7 works were reported at the time we were finishing up our experimental work, to find essential disagreement between their results and ours. But when C. S. Smith&apos;s latest work on the copper-silicon system8 was published just as we were finishing our laboratory report, it was indeed gratifying to find that the latter&apos;s diagram and ours checked in all essen-tial features, except the temperature of the peritectic line above 830° C. Many attempts were made to retain the beta phase during the present investigation, but none was successful. Isawa,6 however, reports that he has been able to obtain X-ray photograms of the beta phase. The structure, he states, is body-centered cubic with a lattice constant of 2.848 A.. at 7.2 per cent Si. Okamoto7 and Smith8 both use the designation Kappa for the hexag-onal phase, and to avoid confusion this designation will be adhered to in the present paper.

Jan 1, 1939

By William L. Fink, L. A. Willey

Since so many different curves have been published for the solid solubility of zinc in aluminum, it seems desirable to definitely establish the correct curve by two or more independent methods. The curve published in part I1t was based entirely upon X-ray measurements because the other methods employed at that time were not satisfactory. The present paper (the seventeenth of this series on high-purity aluminum alloys) describes the determination of the solid solubility of zinc in aluminum by the measurement of electrical resistivity at temperature and by microscopic examination. Since the publication of our part I in 1932, an investigation of the solubility relations in the aluminum-zinc system has been reported by Schmid and Wassermann2, who obtained their solubility results from lattice-parameter measurements on heat-treated and quenched specimens. The diagram that they published is shown in Fig. 1. In this investigation they also found by X-ray measurements that the parameter of the gamma phase changed continuously with zinc content and at such a rate that it would be identical with that of the beta phase at 60 atomic per cent zinc. It was concluded therefore that the beta and gamma phases, from the point of view of structural relationship, are one single phase. They also suggested a gap of immiscibility between 270" and 350" C. The complete boundary in this region could not be determined satisfactorily by the method employed and therefore was not shown in their diagram. Aluminum-zinc Alloys Materials Used.—Electrolytically refined aluminum3 was used in preparing the alloys for the work reported in this paper. Chemical analysis showed the following impurities: 0.004 to 0.005 per cent silicon, 0.006 to 0.008 per cent iron and 0.005 to 0.006 per cent copper. Also,

Jan 1, 1936

By AIME

AS with Phil Kraft, referred to on this page last month, travel has always held a great fascination for Raymond Frank Baker and for that reason he determined to become a geologist. He had heard that geologists travel and that was what he wanted to do. Now as chief geologist for The Texas Co. his early wish has been fulfilled-he spends almost one third of his time journeying&apos; from one oil field to another. Mr. Baker was born in New Haven, Conn., on Jan. 11, 1892, and went through the public schools in New Britain. A stalwart New Englander, he attended Yale and in 1912 received a degree in civil engineering. For a year following this he worked as a civil engineer for the New York Central Railroad, but his early urge to travel widely could not be satisfied here. So in 1913 he re-entered Yale and until 1916 studied mining and economic geology. He had a little taste of field work, which later became so enjoyable to him, in the summer of 1915 when he worked with the U. S. Geological Survey as a geologic aid. In 1916 he made his first contact with The Texas Co. when he entered their Tulsa office as a geologist. This early career was interrupted by service in the

Jan 1, 1947

By B. H. Bergstrom

The Gaudin-Meloy1 size distribution equation requires the evaluation of two constants, xo and r. The common log-log representation of a screen analysis as cumulative weight percentage passing versus size, readily yields the required constants as illustrated for curves A and B in Fig. 1. The 100% passing (y = 1.0) size is simply x,. If the straight portion of the curve be extrapolated to the 100% passing (y=1.0) abscissa, the intercept is the Schuhmann size modulus, k. The ratio of xo to the size modulus, k, is r as shown: The Gaudin-Meloy equation requires that the slope, a, of the size distribution approach unity in the finer sizes, as is the case for curves A and B. Often distributions are better characterized by slopes other than unity. The schuhmann2 equation is able to represent these distributions if the upper portions of the curves are ignored. Empirically, the Gaudin-Meloy equation can be modified to yield slopes other than unity by writing it as Here, a is the familiar Schuhmann slope, and xo is the 100% passing size. The value of r is still given by Eq. 2. Eq. 4 with its easily determined constants is more versatile than either the Schuhmann or the Gaudin-Meloy equations, since a wide variety of size distributions can be closely described. It is easier to fit than the Rosin-Rammler3 equation, and has the advantage that the top size, x,, is finite. REFERENCES A. M. Gaudin and T. P. Meloy: Model and a Comminution Distribution Equation for Single Fracture, AIME Transactions, 1962, Vol. 223, pp. 40 to 43. R. Schumann, Jr.: principles of comminution, I., size Distribution and Surface Calculations, AIME Technical Publication No. 1187, 1940. P. Rosin and E. Rammler: The Laws coverning the Fineness of Powdered Coal, J. Inst. Fuel, 1933. Vol. 7, p. 29.

Jan 1, 1967

By E. F. Koch, J. L. Walter

oriented crystals of iron and iron with 3, 5, and 6.25 pct Si were rolled to reductions of 10 and 70 to 97 pct at room temperature. Similarly oriented crystals were deformed in tension. Dislocation substructures of the deformed crystals were observed by transmission electron microscopy to determine the effect of silicon on the formation of substructures. Pole figures were obtained to relate orientation changes to substructure. When rolled 10 pct, the iron crystals and the 3 pct Si-Fe crystals formed cells, 1 and 0.2 u in diameter, respecliuely. Cells were absent in the higher-silicon crystals. Extended dislocations and possible stacking faults were observed in the 6.25 pct Si-Fe crystal rolled 10 pct and annealed at 650°C. The stacking-fault energy was estimated to be 20 ergs per sq cm. Rolling to 70 pct resulted in the formation of sub-bands (0.9 µ wide) ill the iron crystals and transition bands (containing 0.2-µ-wide subbands) in the 3 pct Si crystals. No subbands formed in the 5 pct Si-Fe crystal until it was ankzealed. SliP occurred on (112) planes ill tension. The slip traces on the 3 pct Si crystal were wary while those on the 5 pct Si crystal wvere straight. The strain-hardening coefficient for the 5 pct Si crystal was nearly zero. Cells did not form, at least at elongations up to 10 pet. The results suggest that cross slip of iron is restricted by additions of silicon beyond about 3 pct possibly by formation of immobile extended dislocations. IN a previous paper&apos; the authors described the substructures developed in (100)[001]-oriented crystals of 3 pct Si-Fe which were rolled to reductions of 10 to 90 pct at room temperature. At low reductions (10 to 20 pct) cells, approximately 0.2 to 0.3 ja in diameter, were formed. The cell walls consisted mainly of edge dislocations. With increasing reduction (up to 50 pct) the cells were seen to elongate in the rolling direction. In certain regions of the crystal there were significant reorientations which were characterized as rotations about an axis normal to the (100) or rolling plane. These regions were called "transition bands". The regions in which there were no reorientations were called (&apos;deformation bands". At reductions of 60 to 70 pct the elongated cells in the transition bands became sub-bands separated by low-angle tilt boundaries with angles of disorientation of about 2 deg. The elongated cell structure in the deformation band was replaced by a general distribution of dislocations. It was noted that the width of the subbands in the transition bands remained 0.2 to 0.3 µ; i .e., the width of the subbands was the same as the initial cell diameter for reductions up to at least 70 pct. From this, and from considerations of the mechanism of formation of the transition bands,&apos; it was concluded that the subbands evolved directly from the initial cells. In order to check this conclusion, it was decided to examine the relationship between initial cell diameter and width of subbands produced by large rolling reductions. Cell size is known to be dependent upon the temperature of deformation.2,3 However, preliminary experiments with 3 pct Si-Fe crystals indicated that the change in cell size with increasing temperature of deform,ation was not sufficient for the present purpose. On the other hand, cell diameters generally reported for iron deformed at room temperature2&apos;3 range from 1 to 2 p, a factor of 3 to 10 larger than the cells in 3 pct Si-Fe rolled to 10 pct reduction,&apos; indicating the possibility of a marked dependence of substructure (at least in terms of cell size) on the amount of silicon in iron. Thus, the investigation was enlarged to include the study of the effects of varying silicon content on substructure in lightly rolled as well as in heavily rolled crystals of iron and iron with 3, 5, and 6.25 pct Si. The crystals used in this study all had the same orientation, (100)[001], with respect to rolling plane and rolling direction. These were rolled to reductions of from 10 to 97 pct and the substructures determined by electron transmission microscopy in both the rolled state and after annealing. In addition, stress-strain curves were obtained from (100)[001]-oriented crystals of iron and 3 and 5 pct Si-Fe to determine the effect of silicon on tensile properties. The dislocation substructure of the tensile specimens was also determined for Samples pulled to 2 and 10 pct elongation at room temperature for comparison with the substructures produced by rolling. 1) EXPERIMENTAL PROCEDURE Crystals with 3, 5, and 6.25 pct Si were prepared by annealing 0.012-in.-thick sheets of high-purity Si-Fe in purified argon at 1200°C to effect growth

Jan 1, 1965

By V. A. Josendal, J. W. Wilson, B. B. Sandiford

Two radioactive tracers have been tested as a means of determining core saturation in multiphase flow studies. Cesium chloride was tried as a water-phase tracer, but complications in its use in low permeability cores resulted from sorption of cesium by the core or water-wet pads. IodoI3&apos;ben-zene proved very satisfactory as an oil-phase tracer. The synthesis of iodobenzene from the sodium iodide as received from Oak Ridge is simple and direct. The tracer is insoluble in water and there was no evidence of sorption by any of the core materials used. use of the method to determine saturation profiles during capillary and dynamic desaturations and relative permeability measurements on oil-water and oil-gas systems is described. Comparisons of the dynamic and capillary methods of relative permeability determination were made using the tracer to check core saturation and saturation distribution. Other experiments are also described in which mobility of the oil phase at various saturations was measured by displacing labeled oil by flowing inactive oil. Similar experiments were made using water labeled with cesium 134. INTRODUCTION To be fundamentally sound. any laboratory method of determining relative permeability must meet, among others, the following requirements: (1) the core saturation between the pressure taps must be uniform, and (2) there must be uniform pressure difference between the phases in the region between the pressure taps. In practice it may eventually be shown that appreciable departure from either or both of these conditions may be permitted without introducing sensible error. but the magnitude of such error remains to be established. There is need then for methods of saturation determina- tion which will determine the saturation profile in a core as well as the overall average saturation given by the gravimetric or material balance methods usually employed. Methods which have received consideration include resistivity, X-ray absorption, gamma ray absorption, neutron diffraction, and radioactive tracers. The resistivity method has use in the determination of profile, but needs checking by independent means.&apos; The X-ray method has been used successfully, but requires elaborate equipment and calibration.23 4 At present gamma ray absorption is practical only for very large cores. The neutron methodo does not appear adaptable to measurement of saturation variations in short cores since it is a scattering method. Russell, Morgan and Muskat&apos; employed radiovanadium in a study of the mobility of interstitial water. Coomber and Tiratsook used radioiodine as an oil-phase tracer and measured profiles in unconsolidated sand packs. ADVANTAGES OF RADIOACTIVITY METHOD A thorough discussion of the theory of radioactive tracers and the technique of using them to study the movement of fluids in sands is given by Coomber and Tiratsoos and need not be detailed here. The method requires less equipment than the X-ray method; the counting problem is similar, but the generation and regulation of the radiation, which is a major problem in the X-ray method, is not required in the radioactivity method. In the X-ray method it is customary to add up to 20 per cent of absorber, usually an iodine compound, to the phase to be labeled. while in the radioactivity method a mere trace is sufficient. A possible disadvantage of the radioactivity method is that a disproportionate amount of the activity comes from the portion of the core nearest the counter, whereas the X-ray gives a true average saturation in the region traversed by the beam. 011 the other hand, this characteristic of the radioactivity method renders it peculiarly adaptable to study of certain types of radial variations in saturation. As with the X-ray method, a tracer may be added to either the oil or water-phase. An oil-phase tracer is more generally useful. however. since it permits work on oil and gas in the

Jan 1, 1952

By B. A. Wilcox, A. H. Clauer

DURING the course of an investigation on the high-temperature creep behavior of TD Nickel* (Ni + 2) vol pct ThO2), it was observed that the creep fractures were similar in appearance to low-temperature ductile fibrous fractures and creep ruptures by cavitation. Microfractographs of fibrous ruptures are typically characterized by a dimpled appearance, with second-phase particles frequently found near the bottom of the dimples.&apos;-4 These somewhat resemble creep ruptures by cavitation, in that fracture occurs by void formation, with the subsequent growth and final merging of the voids to form a complete fracture.&apos; It is the purpose of this communication to describe the creep fracture of TD nickel and suggest a mechanism for the rupturing process. The discussion is relevant to material crept to fracture in vacuo at a constant stress of 18,000 psi, over the temperature range 900" to ll00°C. The material used in this investigation was 1/2-in.-diam bar stock supplied by Du Pont, contain- ing 2.3 vol pct Tho2. The final fabrication treatment consisted of -95 pct reduction by swaging followed by a 1-hr anneal at 1000°C. Transmission and replica electron microscopy revealed that the material had a fibered structure with a transverse grain size of -1 µ and a longitudinal grain size of 10 to 15 µ. Selected-area diffraction indicated that the fiber axis was parallel to (00l), in agreement with the results of Inman et a1.6 The thoria particle size varied from 80 to 1000Å, with an average size of 400Å, and the mean particle spacing was about 1200Å. The remarkable structural stability7 was confirmed by annealing for 6 hr at 1300°C (about &apos; 0.9 the absolute melting point). Transmission electron microscopy revealed that the only structural change was a slight decrease in dislocation density. Furthermore, recrystallization or grain growth during creep has not been observed. All specimens were annealed in vacuo at 1300°C for 3 hr prior to creep testing. The following times to rupture were obtained for vacuum creep tests at a constant stress of 18,000 psi: 900°C, 854 hr; 930°C,

Jan 1, 1965

By Richard P. Rothwell

In January last I published in the Engineering and Mining Journal a table giving, with a considerable degree of accuracy, the production of anthracite coal for the year 1874. At that time it mas impossible to give any correct estimate of the production of bituminous coal, for the returns I had been able to secure from such distant and scattered districts were not always forwarded punctually, and the work of .obtaining anything like a true estimate of the coal mined in districts from which I was unable to obtain official returns was both laborious and tedious. As will be seen from the detailed reports furnished herewith, I have succeeded in obtaining tolerably accurate returns of 20,093,704 tons; and upon careful revision by numerous well-informed local authorities, my first estimate of the amount of coal not officially returned to me has grown very considerably; indeed, as I get more information on the subject I find that our bituminous coal production greatly exceeds the amount we have been accustomed to hear it placed at, aggregating as it does, 26,465,704 tons. Mr. Andrew Roy, State Inspector of Mines for Ohio, has reported the production of Ohio at 4,168,445 tons, while his first estimates were but 3,820,000 tons; and similarly, in other cases, the more full the returns direct from the mines the more incorrect do I find all previous reports. I am able now to furnish, not an accurate statement, indeed, but by far the most full and reliable one that has yet been published of our coal production. The amount of labor involved in obtaining this admittedly imperfect report has been simply enormous. In the following tabular statement it will be perceived that Pennsylvania produces nearly one-half of all the bituminous coal mined

By John R. Weeks, Dan McLachlan, John R. Lewis

DURING the operation of the iron blast furnace, aggregates of hard, infusible, yellow, cubic crystals frequently form in the interstices of the lining and on the hearth of the furnace. These were first examined by Wohler&apos; and were identified by him as a complex cyanonitride of titanium, having a formula Ti(CN)2 • 3Ti3N2, or more simply Ti10C2N8. However, recent investigators have suggested that this complex compound does not exist. Recent work by Hume-Rothery, Raynor, and Little2 on the carbide and nitride particles in titanium steels suggests that a complete series of solid solutions of the general type Ti<0+y>C0Ny exists between titanium carbide (Tic) and titanium nitride (TiN). Assuming this to be true, it seems likely that Wohler&apos;s cyanonitride of titanium is in reality a solid solution of 20 mol pct Tic in 80 mol pct TiN, .the nearly consistent carbide-to-nitride ratio probably resulting from temperature and pressure conditions in the blast furnace. The object of the present work was to investigate thoroughly this possibility of a solid solution by using the single-crystal methods of X-ray analysis. A spectrographic qualitative analysis of a sample of the yellow crystals furnished by the Geneva Steel Co. sh&apos;owed, in addition to the titanium, strong lines of vanadium and moderate to weak lines of iron, magnesium, silicon, aluminum, and calcium. The last five were probably slag impurities, but the vanadium was possibly included in the yellow crystals, owing to a great structural similarity of its carbide (VC) and its nitride (VN) to the titanium compounds, Table 11. A chemical analysis was made

Jan 1, 1952

By AIME AIME

THAT the copper industry is in serious straits is admitted. So are the lead and zinc industries, and both lead and zinc are tariff protected. Conditions in the Western lead, zinc and silver mining districts are desperate. These are illustrated by the present disastrously low prices of these metals compared with the levels of two and three years ago and by the comparative rates of metal production and the present activity of the western mining camps. The January production of copper was 32.2 per cent of the 1929 average rate, of steel was 32.3 per cent of the 1929 average, while January production of pig iron was only 27.6 per cent of the monthly average of 1929. The copper industry is only one of many industries that are suffering in this tragic depression. It has been exceedingly prosperous and will be prosperous again when business recovers. Its difficulties are not unique and it does not deserve the tariff equivalent of a dole. The burden of subsidizing through an effective tax or tariff the minority group of high-cost mines would be a heavy one on the consuming public of the United States. On the basis of 1929 consumption figures, an effective 5c. tariff would result in an increased copper bill to the country of over $109,000,000, of which the electric manufacturers would contribute $26,000,000, the telephone and telegraph companies $16,000,000, and the automobile manufacturers $14,000,000. And of this total bill, only $33,000,000 would go to those who are urging a tariff and the balance of $76,000,000 would be paid to those who do not admit either its necessity or usefulness under present conditions.

Jan 1, 1932

By J. D. Cooner

Since my working life of 32 yr has been spent in and about the anthracite mines of the Hudson Coal Co., and the previous 4 yr in a college school of mines, I can write best about the safety program of a coal mining company. However, most of the suggestions, statements and practices mentioned herein are applicable to the members of a safety organization and program of almost any industry. It is realized that industrial officials actually interested in real safety work are fully acquainted with the requirements of a practical safety program. Therefore, for such persons there may be nothing new in this paper. The Hudson Coal Co. of Scranton, Pa., employs 7000 men and produces 5,000,000 tons of anthracite annually at nine collieries scattered for a distance of 40 miles throughout the Lackawanna and Wyoming valleys of northeastern Pennsylvania. For the past 17 yr of the above mentioned 32, I have been engaged in safety work as a safety inspector, assistant mine superintendent, and at present the safety engineer. I have also had the privilege of visiting coal and metal mines in other parts of the United States and Canada. We Americans are the most careless people in the world. If any one doubts this statement why are 100,000 of our people killed accidentally annually? Or why during the almost four years we were engaged in World War II were more of our people killed accidentally at home than were killed in our army, navy, marine corps, coast guard, and other services during the same period? We are careless everywhere— in our homes, on the highway, at work and at play. The purpose of this paper is to outline the requisites of a safety department and its program to stop the accidents occurring to employees at work in mines or industrial plants. The first requisite for a safety program is that the management must be convinced of the need and value of a real safety program. By "management" is meant each department head from the president down. Otherwise a good safety department with a good safety program will not succeed because it will not have the proper personnel and equipment to work with, its rules and regulations will not be enforced, and the members of the safety department will be belittled in the eyes of most of the workmen and supervisors. Next, a safety engineer and his inspectors should have worked in the production end of the mines as workmen and also in an official capacity. They are then in a position to have practical and first hand knowledge of the accident hazards of the industry, how to prevent them, and in general the problems of both the workmen and officials. Some men are appointed as safety inspectors when they are physically unable to carry on as mine officials. Such men may know the safety end of the industry and even be safety-conscious, but if they are physically unable to be mine officials they are also physically unable to be safety inspectors, as a safety inspector in our mines must walk from 2 to 4 miles on each daily inspection and visit each working

Jan 1, 1949

By Leonard Waldo

The recent and rapid development of the physics of engineering materials at temperatures as low as that of liquid air and as high as that of the electric are, has drawn renewed attention to the absence of tables, properly printed and spaced, for the convenient and accurate translation of temperature from centigrade to Fahrenheit scales and vice versa. It is to be noted that the Fahrenheit scale is the almost universal usage in foundries and shops in the United States, though the laboratories of the same works employ the centigrade notation. The growing use of the electric furnace in Germany and France calls for much American citation all in centigrade high temperatures, and while the arithmetical calculation is simple, it becomes burdensome when often repeated. I have been surprised at the numerous reprints of Table I. from centigrade to Fahrenheit degrees, first privately printed for personal use, which have been made by works, blue-prints, text-books, and technical journals, and I hope that Table II., from Fahrenheit to centigrade degrees, will be equally useful. For mental calculation the easiest rule to convert centigrade into Fahrenheit is to add 32° to twice the centigrade reading less one-tenth of this product: Thus 1000&apos; C. = 2000" — 200" + 32" = 1832" F., and to convert Fahrenheit to centigrade, subtract 32" from the Fahrenheit reading, multiply by 10 and divide by 2 and 9 successively : Thus 1000° F. = 1000° — 32° = 968° X 10 = 9680° - 2 = 4840° - 9 = 537.77° C. Another and often very convenient method is by &apos;use of a graphic construction as follows :

Jan 1, 1913