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By A. Taylor

NICKEL-RICH alloys hardened with small additions of titanium and aluminum and centered around that region of face-centered-cubic primary solid solution, 7, where the atomic ratio of nickel chromium is about 3 to 1, form an important series of high temperature materials because of their resistance to oxidation, high temperature strength, and excellent creep characteristics at elevated temperatures.' With suitable heat treatment it is possible to precipitate a phase based on the ordered face-centered-cubic y' phase, Ni,,Al, from the randomly ordered nickel-rich y face-centered-cubic primary solid solution. By adjusting the ratio of aluminum to titanium the precipitation of needles of the hexagonal intermetallic compound n-NiaTi may be effected. In the present investigation, the main interest was focused on the interrelationships among the y, y, and 7 phases. By using alloying elements of high purity and a suitable melting technique the presence of carbides and nitrides could be completely obviated, although it is realized that their presence is probably essential in alloys of commercial importance. As a preliminary to the study of the Ni-Cr-Ti-A1 system, it was necessary to establish the constitutional diagrams of the three contiguous systems, Ni-Cr-Al, Ni-Ti-Al, and Ni-Cr-Ti, which together form the nickel-rich corner of the Ni-Cr-Ti-A1 quaternary tetrahedron. Accounts of this work on the ternary systems have already been published by the author. A preliminary disclosure of the author's investigations on the ternary and quaternary systems has been made by Hignett2 and has since stimulated work in the same field by Nordheim and Grant." Before the detailed account of the quaternary system, a brief description will be given of the three contiguous ternary diagrams on which it is ultimately founded. Ni-Cr-Ti System The 750° and 1000°C temperature isothermals of the Ni-Cr-Ti system as determined by Taylor and Floyd' are shown in Figs. 1 and 2. The single phase fields of greatest interest are occupied by the extensive face-centered-cubic y primary solid solution of chromium and titanium in nickel, and the close-packed-hexagonal Daltonide, n-Ni3Ti, from which tie-lines radiate across the very wide y+n two-phase field. The chromium-rich primary solid solution is represented by a. As may be seen from the diagrams, the extent of the y phase field increases quite appreciably as the temperature is raised from 750" to 1000°C. Isoparametric lines for the y phase are shown in Fig. 1. They are interesting in that they are almost parallel to the y/r+n boundary. Thus the determination of the tie-line directions is entirely dependent on the exact stoichiometry of composition of the 7 phase. Ni-Ti-Al System The 750°C isothermal section of the nickel-rich portion of the system Ni-Ti-A1 as determined by Taylor and Floyd is shown in Fig. 3.' The regions of greatest interest are the extensive y primary solid solution of titanium and aluminum in nickel, the y' Berthollide phase field based on the ordered face-centered-cubic structure of Ni3Al, and the hexagonal Daltonide phase, n-Ni3Ti. As may be seen from Figs. 3 and 4 the area occupied by the y phase is substantially the same at 750°C as at 1000°C, but the area of the 7 phase is considerably greater at the higher temperature. The tie-lines radiate from n-Ni3Ti across the y+v and y' + v two-phase fields and are evenly distributed

Jan 1, 1957

By C. S. Blair

THE development of a geological department as an integral part of the Tennessee Coal, Iron & Railroad Co. in the Birmingham district, Ala-bama, in 1908 was an innovation probably unique for any mining organ-ization south of the Ohio River, At that time a few of the oil companies were starting geological departments, but there were no geologists among the larger industrial organizations. The Tennessee Coal, Iron & Railroad Co. had been a substantial producer of pig iron for many years and was just getting a start in the manufacture of steel products from the pig iron. It had large acreages of coal land, ore land and timber land, but there was lacking the detailed knowledge of the possible mineral resources of much (of its territory, and this territory at that time covered an area in the neighborhood of 500,000 acres in the states of Alabama, Tennessee and Georgia. Those responsible for the operations of the mining properties, not being familiar with the kind of help that might be expected from geologic work, were decidedly skeptical of the utility of the work. Therefore it was necessary to develop cooperation with the mining organization, by not infringing upon its functions, and by developing the work in such a manner as to demonstrate to the mining department its practical utility. In the first instance the man responsible for a mining operation is apt to assume the attitude that the geologist is a blockader, because the geologist, unfortunately, is obliged to point out difficulties in the carrying out of mining operations in the way originally planned. He points out the dangers of mining out one coal or metal seam at the ultimate expense of an upper or lower seam, and he develops structural features, such as faults and folds, which call to the attention of the mining operator diffi-culties in his future operations, which the latter had not anticipated. Eventually, however, -the value of having these data became ap to the Tennessee Coal, Iron and Railroad Co., and the operators not only welcomed the geological work and cooperated fully in its acquisition of essential information, but reached the point of asking assistance and advice; and complete coordination of the two activities has existed for many years.

Jan 1, 1933

By J. L. Lamont, W. Crafts

Studies of the effect of composition of steel on hardenability by Grossmann,' and as-quenched hardness by Field2 and by the authors, have made it possible to predict the results of quenching when the composition and cooling rate are known In order to extend the application of the hardenability calculation to the quenched and tempered condition in which steel is finally used, a survey has been made of the hardness changes caused by tempering. From this study a method has been developed for the prediction of hardness and tensile strength of steel in the tempered condition. The method of calculation was developed from Rockwell C hardness tests on quenched and tempered Jominy specimens and is expressed in Rockwell C hardness units. It is of an additive type in which the degree of softening from as-quenched hardness is estimated. Tem-pred properties may be calculated within about plus or minus 5 Rockwell C hardness or plus or minus 15,000 lb. per sq, in. tensile strength. This degree of accuracy is less than that necessary for the control of heat-treatment, but is adequate for the comparative evaluation and selection of alloy steels. The formula indicates that tempered hardness is primarily dependent on as-quenched hardness and, as in the hardening reaction, the progress of tempering is controlled mainly by the carbon content. Alloys tend to retard softening by a secondary hardening mechanism in which the alloy appears to migrate from ferrite to the carbide phase. The factors for the effects of alloys in tempering vary so much from the effects of alloys in hardening that different steels of equivalent hardenability may, after like heat-treatment, differ considerably in tensile strength. Procedure In developing the method for calculating tempered Rockwell C hardness, simple and complex steels were tested over a range from 0.08 to 0'65 per cent carbon and up to 2.68 per cent manganese, 2.18 silicon, 3.50 chromium, 4.94 nickel, and 1.06 molybdenum. The steels were made as small induction-furnace heats at the Union Carbide and Carbon Research Laboratories, Inc' The accuracy of the method was finally checked on a variety of carbon, low-alloy and high-alloy types of heat-treating steels produced under commercial conditions. The method was based on a study of tempered Jominy bars, and its validity was tested by correlation with the center hardness and tensile strength of oil-quenched and tempered bars from 1/2 to 4 in. in diameter. The method of calculation was derived from the differences in Rockwell C hardness between as-quenched and quenched and tempered Jominy specimens. Prenormalized Jominy hardenability speci-menS quenched from a normal temperature for austenitizatios were used to obtain a range of as-quenched Rockwell C

Jan 1, 1948

By W. Crafts, J. L. Lamont

Studies of the effect of composition of steel on hardenability by Grossmann,' and as-quenched hardness by Field2 and by the authors, have made it possible to predict the results of quenching when the composition and cooling rate are known In order to extend the application of the hardenability calculation to the quenched and tempered condition in which steel is finally used, a survey has been made of the hardness changes caused by tempering. From this study a method has been developed for the prediction of hardness and tensile strength of steel in the tempered condition. The method of calculation was developed from Rockwell C hardness tests on quenched and tempered Jominy specimens and is expressed in Rockwell C hardness units. It is of an additive type in which the degree of softening from as-quenched hardness is estimated. Tem-pred properties may be calculated within about plus or minus 5 Rockwell C hardness or plus or minus 15,000 lb. per sq, in. tensile strength. This degree of accuracy is less than that necessary for the control of heat-treatment, but is adequate for the comparative evaluation and selection of alloy steels. The formula indicates that tempered hardness is primarily dependent on as-quenched hardness and, as in the hardening reaction, the progress of tempering is controlled mainly by the carbon content. Alloys tend to retard softening by a secondary hardening mechanism in which the alloy appears to migrate from ferrite to the carbide phase. The factors for the effects of alloys in tempering vary so much from the effects of alloys in hardening that different steels of equivalent hardenability may, after like heat-treatment, differ considerably in tensile strength. Procedure In developing the method for calculating tempered Rockwell C hardness, simple and complex steels were tested over a range from 0.08 to 0'65 per cent carbon and up to 2.68 per cent manganese, 2.18 silicon, 3.50 chromium, 4.94 nickel, and 1.06 molybdenum. The steels were made as small induction-furnace heats at the Union Carbide and Carbon Research Laboratories, Inc' The accuracy of the method was finally checked on a variety of carbon, low-alloy and high-alloy types of heat-treating steels produced under commercial conditions. The method was based on a study of tempered Jominy bars, and its validity was tested by correlation with the center hardness and tensile strength of oil-quenched and tempered bars from 1/2 to 4 in. in diameter. The method of calculation was derived from the differences in Rockwell C hardness between as-quenched and quenched and tempered Jominy specimens. Prenormalized Jominy hardenability speci-menS quenched from a normal temperature for austenitizatios were used to obtain a range of as-quenched Rockwell C

Jan 1, 1948

By R. W. Smith, K. H. Lee

Bond type simulated closed-circuit grindability tests and batch grindability tests were run on six pure minerals plus a limestone and a taconite ore. The simulated closed-circuit tests were made at five different meshes of grind. The minerals tested varied widely in hardness and breakage pattern. For all simulated closed-circuit tests both Bond and Holmes work indices were calculated and compared. The Bond grindability testlm6 has been widely and quite successfully used for predictions of such items as ball mill power requirements and as an aid in plant comminution equipment selection. The test is interesting in that although run as a series of batch tests it is over-all a simulated closed-circuit test with a circulating load of 250%. The simulated closed-circuit feature is probably the primary reason for the notable success of the test. Perhaps the major criticism that can be brought against the test is that it is time consuming, since it normally takes a trained technician about an eight-hour shift to perform it.4 For many purposes this is an acceptable time requirement. However, if a grindability test is to be used for mill control purposes it should be performed in as short a time interval as possible, preferably under an hour. Sometimes in industry strictly batch tumbling mill tests are run and an attempt is made to correlate such batch tests with plant mill operation. Then an attempt is made to adjust operating tumbling mills based on data from the tests. The tests, however, frequently do not correlate well with operating data. The purpose of the present work was to explore some of the similarities and differences between a simulated closed-circui.t grindability test and a batch type test. Six different minerals plus a limestone and a taconite ore were tested in the experimentation. These materials varied widely in hardness, breakage characteristics, etc. Standard dry Bond grindability tests were mn on all materials at 48, 65, 100, 150 and 200 meshes of grind. Feed material was prepared by stage cmshing to -6 mesh. Initial feed volume was 700 cc. Care was taken to maintain the standard constant weight (about 20125 gm) and number of balls (283) during all the tests. Rotational speed was 70 rpm. Table I lists the materials studied along with pertinent data regarding the materials and Bond grindabilities obtained. In addition, dry batch tests were performed on all materials. These tests were made in the same mill as was used for the Bond tests. An identical ball charge and an identical feed volume were used. After rotating the mill for a predetermined number of revolutions at the same rotational speed as for the Bond test, the material was removed and separated from the ball charge. A screen analysis was performed on this product. Knowing the original size distribution of the feed and the size distribution of the product, it was possible to calculate the grindability or grams of a particular mesh produced per revolution of the mill. After the screen analysis of the product, the various fractions were recombined and placed back in the mill for further grinding. After grinding for additional revolutions of the mill the procedure was repeated. Thus, it was possible to obtain grindabilities for each material at a number of meshes of grind and for various revolutions of the mill. Grind-abilities obtained from batch tests are shown in Table 11. COMPARISON OF BOND AND BATCH GRINDABILITIES Figs. 1A through 1D show relationships between batch and Bond grindabilities. Figs. lA, lB, 1C and 1D show respectively graphs indicating the relationships for batch tests run for 100, 300, 500 and 700 revolutions of the mill. For a particular material on each graph, points are shown for 200, 150, 100, 65, and 48 meshes of grind except for several missing coarse meshes of grind for galena which did not fall conveniently on the figures. For each material and graph the point nearest the origin is for 200 mesh, the next for 150 mesh, the next for 100 mesh, etc. Considering Fig. lA, it is interesting to note that all relationships shown are essentially straight lines. There are, however, considerable differences

Jan 1, 1969

By G. W. Preckshot, J. W. Gorman

MOLTEN metals offer an excellent medium for the study of molecular diffusion and interphase transfer. In the absence of intermetallic compound formation, solutions of molten metals are solutions of elements. Therefore, complicating factors such as molecular structure, dissociation, association, polarization, and so on, are generally absent. Further, metals remain in the liquid state over a wide range of temperature. Thus, the effect of temperature on transport processes can be studied in more detail. Although some data are available for the diffusion of metals in the molten state at high temperatures,"""".'!' the most useful and accurate data are those for the diffusion of various elements in mercury at room temperature. The effective radius of diffusion for the various elements in mercury may be calculated from the Stokes-Einstein relation.'-These calculated radii are compared with the atomic and ionic radii calculated from X-ray data on crystals." The effective radii of diffusion of the alkali and alkali-earth metals correspond to those of the neutral atom; those of cadmium, silver, lead, zinc, thallium, bismuth, and tin correspond to ionic radii; and those of copper and gold indicate larger radii due to association with mercury or more likely, faulty diffusion data. Independent evidence"' supports the notion that alkali metals are uncharged in amalgams. The concept of metals moving free of their conductance electrons in the liquid state has been discussed by Eyring, et al,'" in connection with studies on the viscosity of pure molten metals. Considering the approximate nature of the tabulated atomic and ionic radii and the variance in the diffusion measurements, the application of the Stokes-Einstein equation to the diffusion of metals in mercury gives gratifying results. It is noted that a number of these metals have article radii eaual to or less than those of the solvent mercury.".'*,"' Soba1,'" sing the inaccurate Roberts-Austin20 data, indicated that the effective radius of diffusion might well be a function of temperature. Additionally in studies of mass transfer between phases, many conflicting opinions have been expressed concerning the assumption of equilibrium at the interface." ,","," The results of the present work on the diffusion of solid copper in molten lead throws light on the tempera- ture variation of the diffusion radius as well as the extent of the interfacial resistance. The theoretical background and experimental procedure are presented here in detail. Results on other systems will be presented in subsequent papers. Nomenclature The following nomenclature is used in the paper: A—cross-sectional area of diffusion column, sq cm C, C,—concentration of solution and saturated solution respectively, g per cc D,D'—-diffusivity, sq cm per sec e—Naperian base h—ratio, k/D k—mass-transfer coefficient, cm per sec L—diffusion column length, cm n—series index q—flux, g per sq cm-sec r—radius, cm or A R—-gas constant t—time, sec T—temperature, K V—volume, cc w—weight of copper-lead alloy on copper cylinder WT—-total copper transferred to lead diffusion, column, g Z—distance from top of diffusion column, cm 2'—distance from copper-lead interface, cm (2' = L-2) u,,—roots of diffusion equation Mathematical Apparatus The model shown in Fig. 1 applies to the diffusion of solid copper into molten lead. Here Fick's first law may be written where all terms are defined in the nomenclature. It is assumed that the volume of the lead column is constant. This is justified since experimental densities compared favorably with calculated densities based on perfect interstitial mixing. Thus there is no counter diffusion of lead. It is clear that 2 in Equation [I] should be measured relative to the interface which moves relative to the diffusion cell but not to the liquid phase. Hence, the frame of reference is the plane of the copper-lead interface. The diffusivity calculated at constant volume, D", is that defined by Hartly and Crank" and is designated herein as D. No Interfacial Resistance—The two situations of the presence and absence of interfacial resistance will be treated. The simplest is the latter which re-

Jan 1, 1959

By R. Schuhmann

Available thermodynamic data applicable to copper smelting systems are collected and tabulated, and the important gaps are pointed out. A few examples are given of estimations which can be made from the available data. An experimental research program is proposed to supply the thermodynamic data that appear most essential to better quantitative understanding of the chemistry of copper smelting. The proposed program is designed also to shed specific light on the practical problems of slag losses and magnetite behavior. OPPER smelting, from flotation concentrates to V-/ blister copper, is conspicuous among the large scale chemical processes which are conducted with only relatively incomplete knowledge of the physical chemistry involved. For example, no common metallurgy text explains adequately why copper enters the matte phase while iron enters only to the extent that sulphur is left over after satisfying the copper. Explaining this important phenomenon as a manifestation of the greater affinity of sulphur for copper than for iron is unsound because the affinities of copper and iron for sulphur are about the same at smelting temperatures. As will be shown, other affinities are really decisive in the relatively clear-cut separation of copper in the matte. In contrast, thermodynamic studies have contributed much to our understanding of copper refining, zinc oxide reduction, magnesium production, steelmaking, etc. For each of these processes, the important reactions are clearly recognized, and fair to good quantitative values of the free-energy changes and equilibrium constants are available. Such data have proved to be of constantly increasing practical value in process development and improvement. Work in other fields has furnished much thermodynamic data applicable to copper smelting systems. In fact, several promising starts have been made in applying such data to specific smelting problems. Kelleyl made an appraisal of the possibilities of recovering elementary sulphur from low grade matte. His calculations and compilations of thermodynamic data have represented the starting point for much of the work in this field, including the present survey. Huang and Hayward2 nd Aksoy3 used thermodynamic methods in the study of copper losses in reverberatory slags. Peretti4 used thermodynamic data in explaining the chemistry of converting. The recent publications of Darken and Gurry"." and of Darken,' dealing with the iron-oxygen and iron-silicon-oxygen systems, respectively, present equilibrium data that are applicable to copper smelting systems. Also additional data were reported recently on the affinity of sulphur for copper, manganese, and iron8 and on the sulphur pressures of iron-sulphur melts." The survey presented in this paper was made as the basis for planning an experimental program on the thermodynamics of copper smelting. Few researches on the chemistry of copper smelting have been reported in recent years, so that a reappraisal and coordination of old and new data are essential if further work is to take the directions of maximum value. The experimental program is now in progress, and the plan of attack is outlined at the end of this paper. Progressive Oxidation and Desulphurization of Copper-bearing Liquid Phases: The chemical activities of sulphur and oxygen are two of the most important thermodynamic yardsticks to be applied to copper smelting processes. Virtually the entire smelting and refining sequence involves a series of systems characterized by decreasing sulphur activity and increasing oxygen activity. In this section, therefore, an attempt is made to define and explain these activities in terms of equilibrium partial pressures of SO2, s2, and O,. Also, estimates of these quantities are presented, the estimates being based largely on calculations presented in a later section of this paper. The sequence of steps from raw flotation concentrate to fully oxidized copper ready for poling involves progressive and controlled oxidation. Iron is oxidized and enters the slag. Sulphur is oxidized and leaves in the gas. Table I summarizes several important features of this oxidation and desulphurization sequence, starting at the beginning of the matte blow in the converter. The top line gives in order the principal smelting and refining stages up to fully oxidized copper, plus the additional step, not used commercially, of oxidizing all the way to Cu,O. In the second line are shown the principal copper-bearing liquid phases which characterize the process. Through most of the sequence the copper

Jan 1, 1951

By G. W. Tracy, R. D. Carter

Numerical calculations were made to determine the behavior of reservoirs with high-pressure drawdown and wide well spacing where the initial productivity is low and the wells are completed by hydraulic fracturing. The two-phase flow equations were solved for the flow into a single well. This well was assumed to be producing from a reservoir with hydraulically created horizontal fractures (four different systems with fractures were studied). For comparison purposes, additional two-phase flow calculations were made assuming a reservoir with uniform rock properties. The two-phase flow results were also compared with the conventional calculation methods, which do not include the effect of saturation gradients resulting from a simultaneous flow of oil and gas which are normal to this type reservoir. It was found that the conventional methods predict (1) a high and too optimistic value of ultimate recovery, (2) a high producing rate and a high reservoir pressure at a given oil recovery and (3) a low trend of gas-oil ratio with oil recovery. Included in the two-phase flow calculations were provisions to control the oil production rate by an allowable rate and, also, by a gas-oil ratio penalty rule. For the systems with hydraulic fractures, the producing rate was controlled by the gas-oil ratio penalty rule for most of the life. This is in contrast to the system with uniform rock properties which went "on decline" almost immediately. An unexpected characteristic of the systems which included fractures was the early rise in producing gas-oil ratio from 730 cu ft/bbl to approximately 1,200 cu ft/bbl, followed by a "leveling off" before the normally expected gas-oil ratio rise began. Additional features which are a result of hydraulic fracturing are (I) greater ultimate recovery, (2) higher average producing rates and (3) a lower average reservoir pressure at a given oil recovery. INTRODUCTION Some oil fields discovered during the past few years are producing from certain volumetric ally controlled reservoirs (often referred to as solution or internal gas- drive reservoirs) which are characterized by high-pressure drawdown at the wells. Since the available pressure drawdown at a well is limited by the static reservoir pressure and the producing rate is controlled by the available drawdown, wells completed in this type of reservoir usually produce at a rate less than the allowable from the time of completion. Because of this, this type of reservoir is referred to as a low productivity reservoir. Economic considerations require the use of wide well spacing and well stimulation by hydraulic fracturing to make commercial wells in this type of reservoir. Performance predictions for volumetrically controlled reservoirs have been made using a combination of two standard equations. 1. The "Schilthuis" or "Muskat" type material balance equation is used to relate the average reservoir pressure and the cumulative oil recovery. 2. The results from the material balance equation and the productivity factor as described by Pirson' are used to relate the cumulative recovery with producing rate and time. The material balance equation assumed uniform pressure and liquid saturation conditions throughout a reservoir. The steady-state radial flow formula allows for a pressure gradient toward a well but assumes uniform liquid saturation. These calculation methods are adequate for application to reservoirs wherein the drawdown at the well to realize satisfactory producing rates is small compared to the total pressure. In low productivity, volumetrically controlled reservoirs, the pressure drawdown at the well is large cornpared to the total pressure. Although a precise number cannot be given for the magnitude of a large pressure drawdown, values in excess of 1,000 psi would definitely be included. For practical considerations, this usually occurs when the formation flow capacity is less than about 100 md-ft. However, this limit of formation flow capacity will vary with the well producing rate. The low pressure in the neighborhood of the well which results from a high drawdown causes evolution of large volumes of gas. This causes the gas saturation to be higher near the well than at a greater distance—-hence, a non-uniform gas saturation. Also, the relationship between the relative permeability to oil (K/K) and gas saturation is nonlinear but decreases approximately in an exponential way with increases in gas saturation. Because of this, the following chain reaction is established.

By B. Silliman

THESE ores occur in the Laurentian rocks in the town of Clifton, St. Lawrence County, New York. The Clifton Mining Company have opened these magnetites upon their estate of 23,000 acres, on the waters of the Grasse River, an affluent of the St. Lawrence, which flows for about nine miles across the estate, exposing a section of the formation, and affording ample water-power at numerous points. In approaching this district from De Kalb Junction, upon the Watertown and Ogdensburg Railway, the observer passes from the zone of the specular ores, which abound to a great extent in Jefferson County, across a belt of non-metal-liferous limestones, continuing for some miles intermingled with metamorphic rocks, and gradually giving place to heavy beds of granitic and porphyritic gneiss, the weathered surfaces of which present a gray and rough appearance, from the washing out of the feldspar; leaving the quartz very prominent, and completely disguising the real character of the rock, which, when freshly broken, has a lively reddish appearance from the predominance of flesh-colored and red feldspar. Beds of fine-grained quartzite appear also intercalated in the granitic gneiss. These are the dominating rocks of the magnetite region, but the magnetite is always found, so far as the speaker's observation extended, associated with beds of calcite, carrying hornblende, black and green, biotite, and brown garnet interspersed with yellow and magnetic pyrites. The strike of these rocks is about northeast, and their dip about 40° to 45° southeast.

Jan 1, 1873

By Sidney Cornell

THE technical- advantages of adding oxygen to air and producer gas, or using it as a reactive agent, producing 400 B.t.u. gas instead of. the present 150 B.t.u., with higher flame temperatures and a reducing flame possible, with a smaller volume of gases passing through a furnace, will result in quicker melting, sharp working, and the production of a higher quality steel. Probably electric-furnace quality steel can be produced in an open-hearth furnace using oxygen. A modern, standard, 100-ton open-hearth furnace, with all equipment, costs from $350,000 to $500,000; there are about 1000 open-hearth furnaces in this country. When considering the use of oxygen in the open-hearth furnace, it is necessary to study: (1) The use of oxygen in the present furnaces, such that modification of existing furnaces and equipment cost the minimum, although the full advantages of oxygen are not thus obtainable. (2) The use of oxygen in a furnace designed for its use, producing quality and quantity steel at the lowest possible cost. This entails the junking of approximately one-half the value in existing equipment-writing off $250,000,000. The basis of such an estimate is the present cost of manufacturing open-hearth steel. Objection may be taken to some of the unit prices here given, but while some companies obtain their ores for much less, the prices quoted are -market price. They are used to show the maximum economy obtained by oxygen at that point.

Jan 11, 1924

By Warren Smith

IT has been 5 years since the writer left the Philippine Islands and while in that country his chief work did not lie in this field, though he has visited all but one of the localities mentioned in this article. The principal field studies relating to oil were made by his colleague, Mr. Wallace E. Pratt. The writer's investigations dealt with the general stratigraphy and paleontology of the Philippines. With the exception of the investigations made by Mr. Pratt and the writer very little information on this subject is available. A number of geologists in the employ of large oil companies have visited the Islands from time to time but, following the general rule, the public has scarcely ever been permitted to learn anything of the results. The writer is indebted to the Bureau of Insular Affairs, Washington, for late information regarding recent legislation in the Islands relating to petroleum. All the known oil seeps, petroleum residues, such as ozocerite, and natural-gas emanations are associated with Tertiary sediments. The chief seeps and most promising prospects are located as follows: 1. Bondoc Peninsula (lower end), Tayabas Province, Southeast Luzon. 2. The west coast of the island of Cebú from Alegria north to, and perhaps beyond, Toledo. 3. Central Mindanao not far from Lake Lanao. 4. The ozocerite veins on the Island of Leyte in the northwestern part in the vicinity of the town of Villaba. 5. Natural gas from some deep wells in Tertiary shale formations on the eastern flank of the Cordillera and extending out under the plain on the island of Panay.

Jan 2, 1920

By Otto H. Henry, Edward L. Badwick

Up to the present time, according to Hansenl and Haughton,² the constitution of the bismuth-indium system has not yet been published. The generally accepted value for the melting point of indium, as listed in most scientific journals and textbooks, is 155°C. Recently, however, H. M. Davis3 has determined a new melting point for indium of 157.3²C O.l°C. Since bismuth has been used in preponderance as an alloying element in the majority of low-melting alloys, and indium has been used with considerable success in lowering the melting points of low-melting alloys, 4,6 an interest has been developed in the application of these two metals, making a complete knowledge of the constitution of its alloys desirable. In this investigation, the melting point of indium reported by H. M. Davis has been found to be substantially correct. A value of 157. ²°C ± O.5°C was determined. The two eutectic compositions have been placed at 34 and 50 pct bismuth with their respective eutectic temperatures as 72.4°C ± 0.5°C and 90°C ± 0.5°C. The bismuth-indium phase diagram has been determined by thermal and microscopic analysis and is shown in Fig I. Thermal Analysis Granulated bismuth and indium shot, having a guaranteed purity of 99.99 and 99.95 pct, respectively, were melted in a carbon crucible by means of a laboratory heater. To minimize oxidation of the alloys, a heavy mineral oil (Nujol) having a high- flash point was used. A platinum-platinum rhodium thermocouple of fine wire encased in a thin-walled Pyrex glass tube, about 1/8 in. in diameter, was connected to a type K potentiometer for temperature measurements. The thermocouple was calibrated against the melting points of tin, bismuth, and the silver-copper eutectic and the boiling point of water. Since the temperature readings could be reproduced to 0.²°Cl and the accuracy of the temperature standards in this range is slightly better than 0.5°C, it scems reasonable to assume that the temperature readings are reliable to approximately 0.5°C. The cooling rate could be controlled by varying the amount of insulation surrounding the crucible. Three or four layers of asbestos paper, 1/16 in. thick, sufficed to give a cooling rate of from I° to ²°C per minute. Undercooling was slightly noticeable. This condition was minimized by vigorous stirring, since the advantage of using slower rates was too small to justify the greater expenditure

Jan 1, 1947

By Otto H. Henry, Edward L. Badwick

Up to the present time, according to Hansenl and Haughton,² the constitution of the bismuth-indium system has not yet been published. The generally accepted value for the melting point of indium, as listed in most scientific journals and textbooks, is 155°C. Recently, however, H. M. Davis3 has determined a new melting point for indium of 157.3²C O.l°C. Since bismuth has been used in preponderance as an alloying element in the majority of low-melting alloys, and indium has been used with considerable success in lowering the melting points of low-melting alloys, 4,6 an interest has been developed in the application of these two metals, making a complete knowledge of the constitution of its alloys desirable. In this investigation, the melting point of indium reported by H. M. Davis has been found to be substantially correct. A value of 157. ²°C ± O.5°C was determined. The two eutectic compositions have been placed at 34 and 50 pct bismuth with their respective eutectic temperatures as 72.4°C ± 0.5°C and 90°C ± 0.5°C. The bismuth-indium phase diagram has been determined by thermal and microscopic analysis and is shown in Fig I. Thermal Analysis Granulated bismuth and indium shot, having a guaranteed purity of 99.99 and 99.95 pct, respectively, were melted in a carbon crucible by means of a laboratory heater. To minimize oxidation of the alloys, a heavy mineral oil (Nujol) having a high- flash point was used. A platinum-platinum rhodium thermocouple of fine wire encased in a thin-walled Pyrex glass tube, about 1/8 in. in diameter, was connected to a type K potentiometer for temperature measurements. The thermocouple was calibrated against the melting points of tin, bismuth, and the silver-copper eutectic and the boiling point of water. Since the temperature readings could be reproduced to 0.²°Cl and the accuracy of the temperature standards in this range is slightly better than 0.5°C, it scems reasonable to assume that the temperature readings are reliable to approximately 0.5°C. The cooling rate could be controlled by varying the amount of insulation surrounding the crucible. Three or four layers of asbestos paper, 1/16 in. thick, sufficed to give a cooling rate of from I° to ²°C per minute. Undercooling was slightly noticeable. This condition was minimized by vigorous stirring, since the advantage of using slower rates was too small to justify the greater expenditure

Jan 1, 1947

By James Herbert Bartlett

The MANUFACTURE of Iron in the PROVINCE of Quebec. The St. Maurice Forges.—The deposits of iron-ore in the St. Maurice district, in the rear of Three Rivers, were probably known to the Indians and .Jesuits, but it is not till 1667 that any discovery is reported. The affairs of the colony were then in the control of the " West India Company," who amongst other monopolies had a right to all mines and minerals." Specimens of the iron-ore were examined in Quebec. in 1668 by the Sieur la Potardière, who reported unfavorably on them. In 1671, Louis de Buade, Comte de Frontenac, was appointed governor of La Nouvelle France ; and in the original manuscripts relating to his administration some allusion is occasionally found to the St. Maurice iron mines.? In a letter to the Imperial Government, dated 2d November, 1672, he says: "The iron mine of which I have already spoken is of great consequence. I have visited it myself, in order that I may be enabled to give a more accurate account of its nature. I am gratified to learn that another mine has been discovered in Champlain, which is much richer than the Cap de la Madelaine mine, and the ore is in greater

Jan 1, 1886

By Gloyd M. Wilkes

The eastern section of the Park City district is drained to an elevation of 6300 ft, by means of the Ontario drain tunnel owned and maintained by Park Utah Consolidated Mines Co. This elevation represents a depth of 920 ft, in Park City Consolidated Mines Company's property. When development below the 900-ft. level was commenced, sinking was attempted through an incline winze in the fissure. The intensely fractured quartzite forming the walls of the vein could not be held after water was encountered, so it was decided that further sinking should be done in a vertical winze far enough in the foot-wall to be in the more solid quartzite outside the fractured zone containing the vein. Progress was good until a depth of 200 ft. below water table was reached, where the flow of water became serious. Sinking progressed 300 ft., to the 1200-ft. level, and there approximately 750 gal. of water per minute was being pumped. The last 20 ft. of sinking cost more than $200 per foot direct charges, whereas the over-all cost for the 300 ft. was $100 per foot. This high last cost was caused by excessive lost time in pumping. Crosscutting to the vein disclosed the fact that the 1050-ft. level had been developed only after drainage had been effected from the 1200-ft. level, but that the 1200-ft. level could not be opened because of the great amount of water encountered when the fractured zone was intersected. The ground in the fractured zone is so finely crushed that it precludes any possibility of economic support when it is wet. Furthermore, mining costs were found to be prohibitive after caving of vein walls had taken place as a result of an attempt to drive drifts before complete drainage had been effected. More than 3000 gal. per minute was pumped from the 1200-ft. level and the water table remained 75 ft. above this plane even after five crosscuts were driven to the fractured zone at 50-ft. intervals along the strike of the fissure from a footwall drainage drift. It was evident, therefore, that complete drainage must be effected before the fractured zone could be opened in any future development. After extensive investigations in this connection, a well-drilling campaign was decided upon to effect drainage of the fissure in advance of sinking and level development, so that all mining operations would be carried on in dry ground. Drilling Procedure A crosscut was driven through the hanging wall of the fissure on the 1050-ft. level, which was the deepest dry level at the time, to a predetermined location from which a vertical hole should reach the fractured zone at about 400 ft. below this level. A drill station was made and a raise was driven over the location for the hole, to act in lieu of a derrick or drill tower. This raise

Jan 1, 1943

By Gloyd M. Wilkes

The eastern section of the Park City district is drained to an elevation of 6300 ft, by means of the Ontario drain tunnel owned and maintained by Park Utah Consolidated Mines Co. This elevation represents a depth of 920 ft, in Park City Consolidated Mines Company's property. When development below the 900-ft. level was commenced, sinking was attempted through an incline winze in the fissure. The intensely fractured quartzite forming the walls of the vein could not be held after water was encountered, so it was decided that further sinking should be done in a vertical winze far enough in the foot-wall to be in the more solid quartzite outside the fractured zone containing the vein. Progress was good until a depth of 200 ft. below water table was reached, where the flow of water became serious. Sinking progressed 300 ft., to the 1200-ft. level, and there approximately 750 gal. of water per minute was being pumped. The last 20 ft. of sinking cost more than $200 per foot direct charges, whereas the over-all cost for the 300 ft. was $100 per foot. This high last cost was caused by excessive lost time in pumping. Crosscutting to the vein disclosed the fact that the 1050-ft. level had been developed only after drainage had been effected from the 1200-ft. level, but that the 1200-ft. level could not be opened because of the great amount of water encountered when the fractured zone was intersected. The ground in the fractured zone is so finely crushed that it precludes any possibility of economic support when it is wet. Furthermore, mining costs were found to be prohibitive after caving of vein walls had taken place as a result of an attempt to drive drifts before complete drainage had been effected. More than 3000 gal. per minute was pumped from the 1200-ft. level and the water table remained 75 ft. above this plane even after five crosscuts were driven to the fractured zone at 50-ft. intervals along the strike of the fissure from a footwall drainage drift. It was evident, therefore, that complete drainage must be effected before the fractured zone could be opened in any future development. After extensive investigations in this connection, a well-drilling campaign was decided upon to effect drainage of the fissure in advance of sinking and level development, so that all mining operations would be carried on in dry ground. Drilling Procedure A crosscut was driven through the hanging wall of the fissure on the 1050-ft. level, which was the deepest dry level at the time, to a predetermined location from which a vertical hole should reach the fractured zone at about 400 ft. below this level. A drill station was made and a raise was driven over the location for the hole, to act in lieu of a derrick or drill tower. This raise

Jan 1, 1943

By Joseph Daniels

The Fritz Engineering Laboratory was built under the direction of John Fritz, and presented by him to the University. A view of the building, looking east, is shown in Fig. 1. The building was started in 1909, and completed in 1910, although all of the equipment was not placed until later. Mr. Fritz gave his personal attention to the details of construction and equipment, and it was his custom to drive over every day to the University from his home in Bethlehem and spend some hours watching the work, offering suggestions, making changes, and planning new work. The result is a building and an equipment which embody his practical ideas. The laboratory structure is of the steel mill-building type, of light-colored brick, 91 by 114 ft., of which a section is shown in Fig. 2 and the floor-plan in Fig. 3. The steel frame carries the roof and traveling crane-way. Ample light has been provided by numerous windows in the side and end walls, in the clerestory, and by a skylight 84 ft. long and 9 ft. wide in the north roof. The main aisle of the building is 49 ft. 2 in. between centers of crane-columns, and has a clear height of 40 ft. The remainder of the width is taken up by two side aisles, 18 ft. high. The laboratory consists of four sections: (1) A general testing-section containing the testing-machinery, a small machine shop, and an office; (2) a cement-testing room; (3) a room for making and storing concrete test-specimens; (4) a hydraulic section.

Jan 1, 1912

By J. L. Ratliff, R. I. Jaffee, H. R. Ogden, D. J. Maykuth

An experimental program was carried out to improve the low-temperature ductjlity of tungsten through the combined use of dispersed oxides for grain-size control and Groups VII and VIII metal additions for matrix softening. Separate binary -alloy studies showed that increasing amounts of finely dispersed thoria or zirconia and critical amounts of rhenium, osmium, and iridium were effective in significantly reducing the ductile-to -brittle bend transition temperature of tungsten as well as in increasing its re crystallization temper - LIKE most of the high-melting point bcc metals, tungsten undergoes a marked transition from ductile to brittle behavior with decreasing temperatures. The material factors which influence the transition temperature include grain size, metal purity, and the presence of dispersed phases. ature and decreasing its recrystallized grain size. The beneficial effects shown by each of these additions on transition temperature and grain size in binary combinations appeared to be additive in tevnary combinations. Thus, combined additions of 2 to 8 vol pct thoria with 5 pct Re, 0.87 pct Os, or 0.30 pct Ir were effective in lowering the transition temperature of tungsten pom 2000 to 75"C, increasing its re crystallization temperature from 1600 to 1800°C, and decreasing its recrystallized gain size from 1175 to 6800 grains per sq mm. The mechanism by which these factors are operable in tungsten is not yet clearly understood. Nevertheless, it has become increasingly apparent that the interrelationship of grain size and matrix composition governs to a large extent the degree of low-temperature ductility which can be achieved. Though it is not possible to accurately assess the relative importance of these variables, the available evidence suggests that, at low temperatures, grain-size effects predominate over purity effects. This was indicated in recent studies by Allen, Maykuth, and Jaffee,' involving single crystals containing 20 to 40 ppm 0 and 1 to 30 ppm C. Thus, while each

Jan 1, 1964

By T. C. Russell

The Conda phosphate mine, eight miles north of Soda Springs in Caribou County, Idaho, was opened up by the Anaconda Copper Mining CO. in 1920. Except for brief periods, during the 20's and early 30's, the mine has maintained a continuous production, and 2,100,000 short tons of phosphate rock have been shipped to the company's phosphate fertilizer manufacturing plant at Anaconda, Mont., or to independent processing plants on the Pacific Coast; a small tonnage of the rock has been shipped to the Hawaiian Islands. The Phosphoria formation on the Conda property consists of two, commercial-grade phosphate beds separated by 147 ft of black to brownish colored phosphatic calcareous shales interbedded with softer mudstones, clays, and occasional narrow seams of oolitic phosphate. The main lower phosphate bed in the series averaging 7.1 ft in thickness, is called the "foot-wall bed," which lies coriformably on the Wells limestone formation, the 1atter often containing irregular bands of chert. The upper or " hanging-wall bed" of phosphate rock averaging 7.2 ft in thickness, is overlaid by thick quartzite beds, locally called the " Rex chert" member in the upper portion of the Phosphoria formation. (Fig 1.) There are two anticlines on the property with axis striking approximately parallel to the mountain range beneath which extensive development on phosphate rock has been conducted in the past, and is now in progress. The phosphate beds have been eroded to a large extent from the tops of these anticlines, leaving at least three well-pronounced outcrops of the phosphate beds exposed within the boundaries of the Anaconda Copper Mining Cornpany's properties. At present, mining operations in the Conda No. 3 mine are confined to the two phosphate beds along the east flank of the easternmost anticline, where the dips on the beds vary from 50° to 60° east, thereby affording favorable conditions for the sublevel mining system now in use. Mining operations formerly were confined to the footwall or lower oolitic phosphate bed because of its higher phosphate and vanadium contents (32 pct P²O5 and 0.29 pct V2 O 5); this bed averages 7.1 ft in thickness. The oolitic portions of the upper bed were of satisfactory grade, but the shale and clay partings (Fig 2) diluted the mined rock and could not he separated economically by selective mining or by sorting. Experiments in beneficiating this rock were carried on for a number of years. Recently, a simple washing process was adopted, which gave a satisfactory grade of concentrates and a high recovery of P² O5. In 1947 it was decided to expand Anaconda's production of treble superphosphate and phosphoric acid by approximately 90 pct. This rock will be obtained from the footwall and hanging-wall beds, and the rock from the hanging-wall bed will be benefici-ated at -Anaconda.l In order to take advantage of the previous development on the 500 secondary haulage level, short crosscuts will be driven at approximately 950 ft intervals from the existing lateral drifts through the hanging-wall shale beds. From the inner end of each crosscut, short lateral drifts will he driven in both directions, so that four working raises can he put up on the dip of the phosphate bed at proper intervals along each of the lateral drifts. The mining operation on the footwall bed at Conda

Jan 1, 1950

By Joel H. Watkins

The terms kaolin, china clay, ball clay, and paper clay are more or less loosely and interchangeably applied to a large class of white-burning clays. These clays are made up chiefly of hydrous amorphous (colloidal) aluminum silicates with variable amounts of free silica and other impurities in small quantities. Occasionally also the term kaolinite is wrongly applied to a white-burning clay, for kaolinite is a mineral of definite chemical composition and definite crystalline form which occurs but sparingly in nature. It is not the purpose of this paper, however, to define or classify clays, or to enter into a discussion of the chemistry of clays. It is rather a general description of the occurrence and methods of mining of a kindred group of clays, and a discussion of their future economic aspect. In the Southern Appalachian States, the mining of white-burning clays has in recent years become an important industry. Although these clays are of several distinct types, which occur in as many localities under different geologic conditions, they are for the most part primarily of the same origin. They are essentially the same in elementary constituents, and are largely consumed in the game industrial arts. From the standpoint of a mining engineer, therefore, and for convenience in comparison and correlation, it should be interesting to discuss them under one heading. Kinds of Clay The most important types of clay which are now being actively mined in the Southern Appalachian States and with which this paper is particularly concerned are the North Carolina, Virginia, and Georgia kaolins or china clays (residual); the South Carolina and Georgia kaolins or paper clays (sedimentary); and the Florida and Tennessee plastic kaolins or ball clays (sedimentary). Other varieties which are interesting but which are as yet of little importance are the pocket deposits of white clay in Augusta County, Virginia, the bauxitic kaolins of Tennessee,

Jan 1, 1915