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By J. W. Matousek, C. S. Samis

me thermodynamic properties of the components in the Cu-Ni-S ternary system have been studied by measuring the H2S/H2 ratios in a gas phase equilibrated with the molten matte at 1200°C. The measured sulficr potentials have been extended by extrapolation and combtned with previously published data to permit the thermodynamic analysis of the ternary system by integration techniques. A simplifying assumption of ideality in the Ni3Sz-CuS pseudobinary has been made. SINCE copper and nickel metals were first extracted from their sulfide ores the interaction between the two metals and sulfur has been of industrial and theoretical interest. The processes developed to extract copper and nickel have been reported periodically in the literature. The theoretical studies on the Cu-Ni-S ternary have been restricted to problems directly associated with production. It is only within recent years that the binary systems related to this ternary have been analyzed thermodynamically and mathematical devices developed to permit a corresponding treatment of the ternary system. In 1910 an investigation of the role of nickel in Cu-Ni-Fe-S mattes was reported,' It was concluded from this work that nickel is not an element replacing iron in mattes, but rather that Cu-Ni alloys behave as a single metal during the matte converting operation. The laws which govern the converting of Cu-Ni alloys were found to be the same laws as those that govern the conventional converting of copper alone. Although considerable information has been reported on the theoretical aspects of copper smelting, little data is available on the smelting of Cu-Ni ores except for phase diagrams and process descriptions. The available information on the Cu-Fe-S ternary and related systems prior to 1953 has been summarized by Ruddle.2 Since Ruddle's publication, Rosenqvist and coworkers have completed phase equilibria investigations of the Fe-S, Co-S, and Ni-S binaries,3 the Fe-S-O ternary,4 and the Cu-Fe-S-O quaternary.4 An extensive investigation of the Cu-Fe-S ternary system at smelting temperatures has been made by Krivsky and Schuhmann.5 At these temperatures, the Cu-Fe-S system is very similar to the Cu-Ni-S ternary with respect to the location of the single-phase and two-phase liquid matte regions. In the former, however, the area of immiscibility is more extensive and the region of liquid matte-liquid metal-solid metal equilibria is much more extensive. Krivsky and Schumann observed that the Cu2S-FeS pseudobinary system is, for practical purposes, an ideal mixture. This investigation has served as a valuable guide in analyzing the Cu-Ni-S ternary. EXPERIMENTAL The experimental procedure employed in this investigation was the conventional technique of equilibration of the sulfide matte with gaseous mixtures of hydrogen and hydrogen sulfide at 1200°C. The thermodynamic properties of sulfur (S2) in the gas phase are determined by means of Eq. [I] and the activity of sulfur in the matte by Eq. [2].

Jan 1, 1963

By John Tabor, R. B. Spivey

INTRODUCTION Shafts have been sunk in a number of ways. By and large, however, most have been sunk by drill, blast, muck, and slip form methods. Most methods have provided satisfactory means of completion. The major drawbacks to the conventional methods have been: (1) un- reliable cost estimates for completion; (2) unreliable rates of progress; and (3) unreliable integrity of the shaft lining. Recently , engineering design studies have been done that strongly indicate, that, by the use of existing proven technology, shafts can be completed faster, more cheaply, safer, and with a better lining that has been possible using conventional methods. THE METHOD The alternate method utilizes the techniques of big hole drilling, shield excavator tunnelling, and pre-cast concrete lining. Big Hole Drilling - Drilled shafts have been completed for several years and I don't intend to address the technology, only to point out the use of drilled shafts is an alternate method of shaft sinking. Ideally, in a multi-shaft mine, the first shaft would be blind bored and cased to a size, whereby it can handle the muck from the excavation of subsequent shafts. It should be an 8' to 10' completed diameter. The cost of such a shaft in sedimentary formations compares favorably with a conventional shaft. However, it is much faster. For additional shafts, the role of the drill is to complete small holes (3' to 4') to the desired horizon. This is merely a hole to allow slashing to a larger size with easier muck removal from an underground drift connected to the previously drilled shaft. The smaller holes, just as the larger one, should be cased and grouted into place so as to assure a smooth open hole. SHIELD EXCAVATOR Theory and History - Tunnelling Shields have been in common use for several decades, and steam powered shields were used in England over 100 years ago. The purpose of the shield is to provide temporary ground support, protect personnel during operation, and to house the excavating equipment. The excavating equipment approach has varied from attempts at a continuous or semi-continuous boring machine, the tunnel mole, to a mechanical or hydraulic digging arm. The shield may or may not be an active digging element. If not active, the boring tool or cutter is advanced ahead of the shield to full diameter and then the shield is moved up to the cutter, or the shield advances along with the cutter. If active, the cutter opens a conical free face, and the shield is either jacked forward so that the leading edge spalls rock to the free face or poling plates are thrust forward to spall to the free face. In either approach, various methods of muck removal are incorporated into the shield. These may be gathering arms, conveyors, or with moles, hydraulic transport, where the cuttings are sufficiently fine. The efficiency of the muck removal system is critical as excavation is rapid and the system could easily become muck-bound. Application to Shaft Sinking- Technically, the use of a shield to sink vertical shafts is no different than a tunnelling operation. It is, in fact, less complex in most respects. Muck and water removal are difficult in a blind-driven shaft. However, where an underground opening is available for handling muck and water, rapid shield sinking is possible, using a pilot hole for muck and water to fall through and be removed from below. The Tabor Mining Shield is basically a caisson approach, whereby, the ground is only seen at the face or bottom. It does have hydraulic jacks to force it downward and has segmented cutting edges which allows forepoling ahead of the shield. The hydraulic excavator allows mining in the center of the face around the previously drilled hole with the forepoling segments slabbing to this free face. First, the drilled hole is cased and grouted. The machine is a tube with an excavator. The work cycle is as follows: (1) The casing and grout are cut at predetermined intervals by shaped charges. (2) The casing is crushed by the excavator, removed from the hole and hoisted to the surface. (3) The excavator then rips out a face around the hole by digging the formation in tension and breaking the pieces

Jan 1, 1982

By Herman Garlichs

THE development of the extensive Southeast Missouri deposits greatly preceded that of the Iowa and Wisconsin deposits. It began about 1720 at Mine La Motte and other localities, and has continued uninterruptedly to the present time. It is estimated that this district produced about 184,000 tons of lead in the year 1916, having a value of approximately, $25,000,000. This is a new high record. The Southwest Missouri orebodies were hardly known before 1845 and were not extensively developed until 1870: Lead production in South-west Missouri for 1916 was 38,788 tons. Oklahoma produced 14,399 ,tons and Kansas 2,345 tons, making a total for Southwest Missouri, and the adjoining districts in Oklahoma and Kansas, of 55,532 tons of lead. The first metallurgists in Missouri were the Indian and the hunter. The early settler learned to procure his bullets either by melting ore in his camp fire or by throwing pieces of galena on an old stump and depending on the usual roasting and reaction method to obtain the metal. A little more refined method was to arrange two flat stones in the form of a V. Wood and galena were placed in the furnace and ignited. As heat developed, some of the metal was extracted and molded into bullets. The next step in utilizing the lead ores was the log hearth. All lead in Missouri was smelted in this crude affair before 1820. It was built on sloping ground, and consisted of a hearth of stone, surrounded on the front and two sides by a stone wall. The wall was 7 ft. high in front. The top and rear end were left open and in front an arch or opening was made, forming the eye of the furnace. In front of this a pit was dug in the ground, to receive the molten metal. Large logs were rolled in at the back and made to rest upon ledges formed inside, to raise them from the hearth and to give a draft. These logs filled the entire width of the furnace. Small split logs were then set up around the two sides and the front, and the ore was piled on until the furnace was full. Finally, the mass was covered with logs and fuel until the ore was completely surrounded. A gentle heat was started, which was raised very gradually. After 12 hr. the heat was increased and continued for 12 hr. more, 24 hr. being required, for each smelting. The furnace treated about 5,000 lb. per 24 hr. and the ore yielded about 50 per cent. of its metallic lead. A considerable quantity of the ore was not desulphurized, and fell between

Jan 7, 1917

By Hsun Hu

Direct observations were made on the recrystalli-zntion of a cold-rolled (110) [001] crystal of 3 pct Si-Fe by annealing thin sections of the specimen inside the electron microscope during examination. Subgrain formation, presumably by polygonization, occurs rapidly, followed by a slow increase in the subgrain size before recrystallization can be detected. No migration of the subgrain boundaries was observed, and the formation of recrystallization nuclei by coalescence of subgrains is suggested. Re-c~ystallization characteristics of various texture components were individually examined. AS reported in a previous paper: orientations of the various deformation texture components in a 70 pct cold-rolled (110) [0011 crystal of 3 pct Si-Fe were completely confirmed by transmission electron diffraction patterns taken at various areas of the specimen. These orientations can be summarized as follows: Over most areas of the deformed crystal (111) planes are parallel to the surface. These areas correspond to either $- the two main defor: mation textures, the (111) [112] or the (111) [112].2-5 Within a few areas, (100) planes parallel to the surface are frequently found. These areas represent the minor deformation texture component, the (100) [Oll] type, which is derived from rotated Neumann bands.2'4'5 Only in very rare instances are (110) planes found parallel to the surface. This orientation corresponds to the very weak texture component of the (110) [001] type. In a pole figure, this (110) [0011 orientation is located within the orientation spread between the two main (111) [1121 type textures. Usually the electron diffraction patterns taken at various areas of the specimen are very simple ones, i.e., for any one area the pattern consists only of sharply defined spots of a single orientation. In some areas, however, more or less continuous rings of weaker intensities are superimposed on a simple pattern, indicating the presence of a wide range of additional orientations. In order to understand in detail the recrystallization of this cold-worked crystal, the recrywtallization characteristics of each of the various deformation texture components must be examined. This can best be done by using a high-temperature stage and transmission electron microscopy techniques, whereby small areas corresponding to the various texture components can be selected individually for examination. The present paper presents results obtained by this method. MATERIAL AND METHOD Specimens were prepared from a previously studied1 single crystal of a commercial grade 3 pct Si-Fe alloy with an initial orientation (110) [00l], cold-rolled 70 pct from 0.014 in. to 0.004 in. in thickness. Thin foils were prepared by electrolytic polishing as described in the earlier paper.' Specimens of this thin foil were then mounted in a high temperature stage similar to the one designed by Whelan.6 A direct current was used as the heating source. The temperature of the heating grid was calibrated by means of a Pt vs Pt-10 pct Rh thermocouple in a vacuum chamber (0.01 to 0.02µHg). Maximum variation within the temperature range of 250o to 000°C was about ± 25°C. After the desired areas were selected for examination, the specimen was heated to a low temperature for some time while the selected area was constantly watched. When no further change in structure and orientation could be observed, a higher temperature was applied. Photomicrographs and diffraction patterns were taken from time to time during examination, with the heater temporarily turned off* while the photographs were

Jan 1, 1962

By G. P. de Forest, G. O. Hiers

The object of the experiments reported in this paper is to determine the eutectic composition of copper and tin and to determine the location of the part of the liquidus line immediately above the eutectic composition in copper. Three methods were used in the investigation: (1) the method of dissolving copper in a bath of molten tin at constant temperature, with the idea that when the bath was saturated with copper its copper content, by analysis, together with the temperature at which the saturation took place would locate a point on the liquidus line; (2) the method of the differential cooling curves, and (3) the method of microscopic examination. This paper is principally concerned with the results of the last two methods mentioned. The results of the first method, that of dissolving copper in molten tin, did not agree with the results of Miller,' by the same method, nor with the results of other investigators by other methods, nor yet with the results of the cooling-curve experiments and microscopic examination reported here. Solution of Copper in Molten Tin To determine the amount of copper that would be dissolved by molten tin at a given constant temperature the apparatus shown in Fig. 1 was used. It consists of a cast-iron pot a, 6 in. deep and 8 in. dia., well cleaned inside and dressed with graphite. This pot, which holds 15 lb. of tin, g, rests on a hot plate b with resistance winding c. The thermocouple d leads to a Wilson-Maeulen temperature regulator for maintaining the temperahre at the heating coils at a fixed temperature to within ±5°C. There is a shallow bath of solder e on the top of the hot plate to improve the thermal contact between the plate and the pot. An insulating jacket f of asbestos prevents the irregular radiation of heat from the apparatus. The copper stirrer h slowly rotates in, and is dissolved by the bath of molten tin g. This bath is protected from excessive oxidation by a layer of charcoal j. The stirrer h is fixed to the lower end of the iron tube k, which is so mounted that it can be driven by a motor with reducing gear. The temperature of the bath is indicated

Jan 1, 1930

MEMBERS, JUNIOR MEMBERS, ASSOCIATES AND JUNIOR ASSOCIATES LIST OF MEMBERS ALPHABETICALLY ARRANGED EXPLANATION OF SYMBOLS ** Junior Member. t Associate. || Junior Associate. Members are not marked. Capital letters signify Honorary membership. Student Associate and Rocky Mountain members on pages 165-167. Abadilla. Quirico A Catanawan. Tayabas, P. I. Abbey. Robert Graham. District Mgr The W. W. Sly Mfn. Co., 60 Church St New York. N. Y. Abbott. A. N Mines Supt Mazapil Copper Co Ltd . .Conception del Oro. Zac., Mexico. Abbott. Ckrence E Mgr Mines & Quarries, Tenn. Coal. Iron & R. R. Co 1242 Brown Marx Bldg Birminrham. Ala. Abbott. Louie Benjamin. Chief Enar Consolidation Coal Co Elkhorn Div .Jenkins, KY. Abeel. Gsorge H Jr., Cons. Min. Enar .Box 427. Santa Monica. Cal. Abel. Albert E Asst. to Pres Valley Mould & Iron Corp Hubbard. Ohio. Abel. Raymond L. Asst. Prof., Dept. of Pet. Refin Univ. of Pittsburgh Pittsburgh. Pa. Abernathy. George Elmer. Prof. of Geol. & Min Kansas State Teachers College. Pittsburg. Kans. IlAbernethy. D. Carl, Jr. Min. Engr Hudson Coal Co .Scranton. Pa. Abey, Hiroshi. Min. & Met. Engr., Min. & Met. Dept., The Higher Technical School, Osaka. Japan. tAbarn, Robert H Instructor in Met Harvard University Cambridge, Mass tAbozeid,Mahmood. Pet. Production E r . . Marland Oil Co. Ponca City, Okla. Abraham. Arthur W Geol., Richfield Oil Co Box 788. Bakersfield, Cal. Abraham, W. E. V Member, Staff of Burmah Oil Co Ltd. Nyaunghla. Upper Burmz, India. Abrahams. Alexander I Gen'l Mgr., Racon Electrical Co Slough, Bucks England. tAbrahammn. C. W Min. Engr The Mond Nickel Co Ltd . Coniston, Ont Canada. IlAbmn. Guy W., Asst. Metallographist, Youngstown Sheet & Tube Co., Research Dept., Youngstown. Ohio. Abstein. Henry T., Engr. & Mgr Profile-Tamarack Mines Co . .Yellow Pine, Idaho. tAcker. E. O'C Met I28 East Washington Lane. Germantown, Philadelphia. Pa. Ackerman. D. E Met. Research Lab International Nickel Co Inc Bayonne. N. J. tAbckerman. Ernest R Pres., Lawrence Portland Cement Co . .Plainfield. Union Co N. J. ||Ackerman. George F Jr. Min. Engr El Potosi Min. Co. .Apt. 13. Chihuahua, Chih Mexico. Acres, T. H Geol. Dept Valuation & Pet. Engr., Richfield Oil Co. 655 S. Flower St Los Angeles. Cal. Adachi. Jinzoo Address Wanted. Adair. Arthur C Valuation Enar., Internal Revenue. Treasury Dept., Washington, D. C. Adam. John. Min. Ennr., The Zinc Corp. Ltd., Box 4 C. Broken Hill, N. S. W Australia. Adami. Arthur E V. P Montana State School of Mines.Butte, Mont. Adami. Charles J Min. Ensr St. Joseph Lead Co 4540 Lindel Blvd., St. Louis, Mo. Adams. C. W.729 South Normandie St LaGranada Apt Los Angeles, Cal. Adams. Cmyler Geerwwd. Minn. Adams. F. J Gen'l Anent. Gulf Production & Pipa Line Cos .Fort Worth. Tex. ADAMS. FRANK DAWSON. Logan Prof. of Geology and Dean, Faculty of Applied Science, McGill Univ Montreal. Que Canada. Adama, Gale L Research Enar., General Pet. Corp. of California, Higgins Bldg Los Angeles. Cal. Adams. George I, Prof. of Geol. & Mineralogy Univ. of Alabama, Tuscaloosa, Ala. Adams. H. H Cons. Geol Room 416, W. T. Waagoner Bldg Fort Worth. Tex. Adams, Henry. Cons. Min. Ennr 1050 Edgewood Ave., Plainfield, N. J. Adams, Henry H.285 Madison Ave., New York. N. Y. Adams, Henry Farnum. Engr., Charge. Experimental Work. Inspiration Concentrator, Inspiration Cons. Copper Co Box 354. Inspiration. Ariz. Adams. Huntington. Min. Enar I41 Broadway, New York. N. Y. Mama, Leland D., V. P., Leslie-California Salt Co 166 Montgomery St., San Francisco, Cal. Adams, Louis W., Illinois Steel Co .Gary. Ind.

Jan 1, 1929

By R. V. Norris

It was stated in a former paper1 that an intensive study of anthracite costs was being made by the engineers of the United States Fuel Administration. The results of this study are now available and are offered as a supplement to the former paper. Anthracite costs as reported for the 6-months period, December, 1917, to May, 1918, inclusive, as compiled by the Federal Trade Commission, were used as a base, and charts similar to those described in the former paper were made for the standard white ash anthracite and for red ash and Lykens valley coals. Adjustments The adjustments of cost from a reported to a price-fixing basis, as described for the bituminous methods, were applied but showed only minor adjustments as necessary. The great spread in anthracite prices on the varyingsizes, which for the 6-mo. period under review ranged in average from $5.244 for nut to $2.074 for barley coal, makes the question of the percentage of sizes produced at the different collieries a vital one. The realization with the same prices for each size must be within very wide limits, when it is considered that the percentage of prepared coal reported from different collieries varied from over 80 per cent. to below 30 per cent. for fresh-mined coal. Hence, as the spread in prices for the various sizes must be predicated on some percentage, it is essential to find some method of adjustment to allow for this variation. The logical method of adjustment is to calculate actual costs to costs as of the standard percentage of sizes, so that the margin between the adjusted costs and the average realization shall be the actual margin for each colliery between its actual costs and actual realization due to its particular percentage of sizes. As a base for realization the actual percentage of sizes for fresh-mined coal for the 6-mo. period was adopted. This percentage is given in Table 1.

Jan 1, 1920

By R. V. Norris

IT was stated in a former paper1 that an intensive study of anthracite costs was being made by the engineers of the United States Fuel Administration. The results of this-study are now available and are offered as a supplement to the former paper. Anthracite costs as reported, for the 6-months period, December, 1917, to May, 1918, inclusive, as compiled by the Federal Trade Commission, were used as a base, and charts similar to those described in the-former paper were made for the standard white ash anthracite and for red ash and Lykens valley coals. ADJUSTMENTS The adjustments of cost from a reported to a price-fixing basis, as described for the bituminous methods, were applied but showed only minor adjustments as necessary. The great spread in anthracite prices on the varying sizes, which for the 6-mo. period under review ranged in average from $5.244 for nut to $2.074 for barley coal, makes the question of the percentage of sizes produced at the different collieries a vital one. The realization with the same prices for each size must be within very wide limits, when it is considered that the percentage of prepared coal reported from different collieries varied from over 80 per cent. to below 30 per. cent. for fresh-mined coal. Hence, as the spread in prices for the various sizes must be predicated on some percentage, it is essential to find some method of adjustment to allow for this variation. The logical method of adjustment is to calculate actual costs to costs as of the standard percentage of sizes, so that the margin between the adjusted costs and the average realization shall be the actual margin for each colliery between its actual costs and actual realization clue to its particular percentage of sizes. As a base for realization the actual percentage of sizes for fresh-mined coal for the 6-mo. period was adopted. This percentage is given in Table 1.

Jan 2, 1919

By Donald Cummings

MEASURES appropriate for the control of the silicosis hazard in mining cannot be formulated precisely, but sufficient knowledge1-19 has accumu-lated during the past quarter century to permit the suggestion of useful, though partly empirical, rules for its alleviation. The industrial hygienist today recognizes that the peril incurred by the inhalation of harmful dust is a function of two variable factors-the degree of harmful exposure and the specific susceptibility of the exposed individual to pulmonary injury. Con-trol of the danger arising out of prolonged contact with pernicious air-borne particulate matter is consequently best exercised by influencing both of these recognized variables. Obviously the problem has two broad aspects-one, medical; the other within the province of the engineer.

Jan 1, 1938

By Frederick Laist

IN the course of the past 75 years the treatment of copper ores has undergone the most profound changes. To a lesser degree, this is true of all the nonferrous metals, but the rapid increase in the demand for copper due to the phenomenal growth of the electrical industry, and the discovery of vast new ore deposits in Michigan, Montana and Arizona, provided an exceptional incentive for improvements in smelting practice and for large-scale operations. Because these must perforce be carried on in regions where labor was scarce and wages high, the need for mechanization and larger smelting units soon became apparent. At a later date the porphyry coppers were discovered, with ores so low in grade that enormous tonnages had to be treated daily if they were to be made to pay at all. In some instances these ores were associated with, or overlain by, oxidized ores, which could not be concentrated mechanically and, of course, could not be smelted directly. This led to the development of the copper-leaching process. While the principles of the various smelting and leaching processes are for the most part of European origin, and have long been known, modern practice is essentially American and was evolved on this side of the Atlantic. The ores of copper may be divided into three general classes; viz, native copper, sulphide and oxidized ores. By far the greater number of commercially important ore deposits come under the classification of sulphides. Commercial deposits of native copper are limited to the Lake Superior

Jan 1, 1947

By Herman Garlichs

The development of the extensive Southeast Missouri deposits greatly preceded that of the Iowa and Wisconsin deposits. It began about 1720 at Mine La Motte and other localities, and has continued uninterruptedly to the present time. It is estimated that this district produced about 184,000 tons of lead in the year 1916, having a value of approximately $25,000,000. This is a new high record. The Southwest Missouri orebodies were hardly known before 1845 and were not extensively developed until 1870. Lead production in Southwest Missouri for 1916 was 38,788 tons. Oklahoma produced 14,399 tons and Kansas 2,345 tons, making a total for Southwest Missouri, and the adjoining districts in Oklahoma and Kansas, of 55,532 tons of lead. The first metallurgists in Missouri were the Indian and the hunter. The early settler learned to procure his bullets either by melting ore in his camp fire or by throwing pieces of galena on an old stump and depending on the usual roasting and reaction method to obtain the metal. A little more refined method was to arrange two flat stones in the form of a, V. Wood and galena were placed in the furnace and ignited.. As heat developed, some of the metal was extracted and molded into bullets. The next step in utilizing the lead ores was the log hearth. All lead in Missouri was smelted in this crude affair before 1820. It was built on sloping ground, and consisted of a hearth of stone, surrounded on the front and two sides by a stone wall. The wall was 7 it. high in front. The top and rear end were left open and in front an arch or opening was made, forming the eye of the furnace. In front of this a pit was dug in the ground, to receive the molten metal. Large logs were rolled in at the back and made to rest upon ledges formed inside, to raise them from the hearth and to give a draft. These logs filled the entire width of the furnace. Small split logs were then set up around the two sides and the front, and the ore was piled on until the furnace was full. Finally, the mass was covered with logs and fuel until the ore was completely surrounded. A gentle heat was started, which was raised very gradually. After 12 hr. the heat was increased and continued for 12 hr. more, 24 hr. being required for each smelting. The furnace treated about 5,000 lb. per 24 hr. and the ore yielded about 50 per cent. of its metallic lead. A considerable quantity of the ore was not desulphurized, and fell between

Jan 1, 1918

By C. E. Beecher

The application of gas or air under pressure to obtain more oil from a sand which has been practically exhausted by ordinary production methods has been practiced to a limited extent for many years. Until recently such methods were confined to the Eastern fields, principally those of Ohio and West Virginia, where in many cases the production obtained after applying pressure for a few years was nearly equal to the total production prior to that time which had been obtained by the usual production methods. As early as 1917, J. 0. LewisL called attention to this method of increasing the recovery of oil and cited the phenomenal results obtained by H. E. Smith, I. L. Dunn and 0. C. Dunn, who pioneered the application of air or gas pressure recovery methods. With this information available it would appear that other oil-producing sections of the country have been extremely slow to realize the advantages of repressuring. However, after a few successful projects in the Mid-Continent area, the oil producers appreciated t' e value of this method for obtaining further and profitable production from their practically exhausted leases and are now making extensive use of it. In general, it has been the practice to apply pressure to the sands when production has declined to a small yield per well. The purpose of this paper is to discuss some of the advantages which may be expected from returning gas under pressure to the oil sands during the early stages of development, before the rock pressure has been dissipated and the gas exhausted, and to stimulate discussion of this subject on which comparatively little actual field data are available. The Function of Gas Gas associated with or dissolved in the oil is the source of energy causing movement of the oil to the well. When this energy is dissipated production ceases. The exception to this general statement is found in fields where hydrostatic water pressure may supply energy after the gas

Jan 1, 1929

By H. P. Reid

Portland cement is manufactured under either of two general processes, the wet or the dry. The raw materials in general consist of limestone, shells, marl, cement rock, clay, shale, blast-furnace slag, etc., so proportioned that when calcined the desired composition will be obtained. In the wet process the combined raw materials are ground to extreme fineness in the presence of water, to a resulting slurry of from 30 to 50 per cent moisture content. The moisture of the slurry may be reduced by thickening or filtering processes for feed to the cement kiln or the slurry may be fed to the kiln with its full percentage of water. The dry process differs from the wet in that the surface moisture of the raw materials is removed prior to grinding. For the drying of these materials heat is required and in most cases is produced by burning of raw fuel either on grates or stokers before the dryers or in suspension in the form of pulverized coal, oil or gaseous fuel. The power required for making the cement is sometimes purchased from outside sources; sometimes it is generated from the heat in the waste gases leaving the cement kilns, or generated at the plants direct from raw fuel. We have, then, as the possible requirements of fuel in the cement industry: (1) fuel for power generation, (2) fuel for drying raw materials, including coal before pulverizing, and (3) fuel for burning the raw materials to cement clinker. There is no essential difference in quality required in fuel for burning the materials by the wet and,dry processes. Fuel fob Power Generation As there is nothing peculiar to steam generation in a cement mill power plant as compared with other power plants this use classification of such fuel will be found in other reports of the Technical Committee. Fuel for Drying Raw Materials and Coal Any fuels capable of producing moderate temperature gases may be used for drying raw materials. There is no problem of material contamination from fuels at this part of the cement manufacture to be guarded against.

Jan 1, 1932

By B. M. Larsen, W. E. Shenk

This paper describes a method of reversal control of the open-hearth furnace that obtains in practice those effects considered below as essential to a completely automatic control, without appreciable interference with the natural rise and fall in temperature of the regenerative portions of the furnace system. Growing out of some studies of radiation pyrometry in open-hearth regenerators, it was developed during 1934-1938, and has been successfully applied to a number of commercial furnaces. The goal of open-hearth furnace operation is the highest rate of production of steel of proper quality consistent with good refractory life and fuel economy. This requires rapid heat transfer from furnace to charge, which in turn is dependent upon heating the furnace refractories to the highest safe operating temperature. The greater the heat input to the melting chamber, the sooner this limiting temperature will be reached, so that control methods in general that are designed to protect against overheating of refractories should help to maintain maximum fuel input and production rate. For the typical open-hearth system, with fixed regenerator capacity and with rate of fuel input usually restricted in its range of variation, our basic assumption is the idea that the only other important variable —that is, the time interval between successive reversals—should also be held constant or within a small range. The present reversal control was designed therefore so that, under normal operating conditions, no arbitrary limitation is imposed other than to keep the reversal interval within the optimum time range. However, since a near approach to completely automatic operation involves protection of refractories through the balancing of furnace end temperatures as well as emergency protection against temperature peaks above some upper limit, it was also felt essential to employ, as an integral part of the control unit, a temperature-measuring instrument to register the true temperature of the refractory surfaces at some critical region in the hot ends of the regenerative zones. As the basic elements of the control, we thus have one "maximum" and one lLminimum" time relay and a potentiometer recorder-controller, combined with switches and relays to operate as follows: I. Reversal normally occurs within an optimum range limited by the maximum and minimum settings to about 4 to 6 min. variation, but the actual interval usually is less variable and closer to the minimum setting. 2. The margin between minimum and maximum time is utilized, together with a "floating temperature limit" in the potenti-ometric control instrument, to correct any tendency toward temperature unbalance between ends of the furnace. 3. A second "maximum temperature

Jan 1, 1945

By B. M. Larsen, W. E. Shenk

THIS paper describes a method of reversal control of the open-hearth furnace that obtains in practice those effects considered below as essential to a completely automatic control, without appreciable interference with the natural rise and fall in temperature of the regenerative portions of the furnace system. Growing out of some studies of radiation pyrometry in open-hearth regenerators, it was developed during 1934-1938, and has been successfully applied to a number of commercial furnaces. The goal of open-hearth furnace operation is the highest rate of production of steel of proper quality consistent with good refractory life and fuel economy. This requires rapid heat transfer from furnace to charge, which in turn is dependent upon heating the furnace refractories to the highest safe operating temperature. The greater the heat input to the melting chamber, the sooner this limiting temperature will be reached, so that control methods in general that are designed to protect against overheating of refractories should help to maintain maximum fuel input and production rate. For the typical open-hearth system, with fixed regenerator capacity and with rate of fuel input usually restricted in its range of variation, our basic assumption is the idea that the only other important variable that is, the time interval between successive reversals-should also be held constant or within a small range. The present reversal control was designed therefore so that, under normal operating conditions, no arbitrary limitation is imposed other than to keep the reversal interval within the optimum time range. However, since a near approach to completely automatic operation involves protection of refractories through the balancing of furnace end temperatures as well as emergency protection against temperature peaks above some upper limit, it was also felt essential to employ, as an integral part of the control unit, a temperature-measuring instrument to register the true temperature of the refractory surfaces at some critical region in the hot ends of the regenerative zones. As the basic elements of the control, we thus have one "maximum" and one "minimum" time relay and a potentiometer recorder-controller, combined with switches and relays to operate as follows: x. Reversal normally occurs within an optimum range limited by the maximum and minimum settings to about 4 to 6 min. variation, but the actual interval usually is less variable and closer to the minimum setting. 2. The margin between minimum and maximum time is utilized, together with a "floating temperature limit" in the potentiometric control instrument, to correct any

Jan 1, 1945

By C. E. Beecher

THE application of gas or air under pressure to obtain more oil from a sand which has been practically exhausted by ordinary production methods has been practiced to a limited extent for many years. Until recently such methods were confined to the Eastern fields, principally those of Ohio and West Virginia, where in many cases the production obtained after applying pressure for a few years was nearly equal to the total production prior to that time which had been obtained by the usual production methods.' As early as 1917, J. 0. Lewis1 called attention to this method of increasing the recovery of oil and cited the phenomenal results obtained by H. E. Smith, I. L. Dunn and 0. C. Dunn, who pioneered the application of air or gas pressure recovery methods. With this information available it would appear that other oil-producing sections of the country have been extremely slow to realize the advantages of repressuring. However, after a few successful projects in the Mid-Continent area, the oil producers appreciated the value of this method for obtaining further and profitable production from their practically exhausted leases and are now making extensive use of it. In general, it has been the practice to apply pressure to the sands when production has declined to a small yield per well. The purpose of this paper is to discuss some of the advantages which may be expected from returning gas under pressure to the oil sands during the early stages of development, before the rock pressure has been dissipated and the gas exhausted, and to stimulate discussion of this subject on which comparatively little actual field data are available.

Jan 1, 1928

By Henry Fay

THE erosion of guns is a complex problem which can be solved only by a detailed study of all the factors involved. In the present paper it is proposed to submit the results of observations and experiments which have extended over several years. Certain facts have been established and preliminary reports have been published.' Conclusions have been reached since then which, it is believed, are worth presenting for the purpose of promoting discussion, and with the hope that ultimately the whole truth will be known. The particular phase of the problem which will be presented here is the hardening of the inner surface of the gun tube. It has been known for a long time that after firing a large caliber gun for some time, the surface of the metal becomes hard and brittle, cracks, and wears away.2 The character of this phenomenon may be learned by an examination of some results which are typical, obtained from a detailed study of a 12-in. gun. The gun was trepanned and rings representing various sections of the gun were cut out for examination. The first ring represented the metal at the muzzle end, and the other rings were removed at approximately 120, 240, 325 and 337 in., respectively, toward the breech end. The appearance of the surface of the metal in the different sections is shown in Fig. 1. The greatest amount of wear and maximum amount of hard surface layer was found in section marked E, which is nearest the powder chamber and at the beginning of the rifling; the maximum amount of cracking in D; the maximum amount of copper deposition in C. Section B showed heat cracks on both lands and grooves, and section A showed heat cracks on the driving edge of the lands only. There was much flow of metal in both sections E and D, and there was progressively less effect noticeable toward the muzzle end. The polished cross-section of these same pieces is shown in Fig. 2. The

Jan 12, 1916

By J. Chipman, T. P. Floridis

The equilibrium of gaseous H2O-H2, mixtures with liquid iron is used to establish the activity coefficient of dissolved oxygen as log f o = — 0.20 [%O]. Discrepancies in earlier work are explained or eliminated. Additions of cobalt, molybdenum, tungsten, and platinum increase the activity coefficient of oxygen whereas copper and gold decrease it. THE importance of dissolved oxygen as a refining agent in liquid steel has led to a number of studies of its thermodynamic properties. It is the purpose of this paper to examine the effects of a number of alloying elements on the activity of oxygen in the steel bath. As a preliminary to the study of the effects of alloying elements, it seemed desirable to examine certain minor discrepancies in the published data on the properties of oxygen in simple iron-oxygen solutions. The activity and free energy of oxygen in liquid iron were studied by chipman1 in 1933 by means of the reaction h2(g)+-o = H2O (g);K; = PH2O/PH2[% O] [1] The data were in fairly good agreement with calculations from results on the similar reaction CO (g) + O- = CO2 (g); K2 =P CO2/P CO [% O] [2] which had been obtained by Vacher and Hamilton2 as an incidental part of their study of the reaction of carbon with oxygen dissolved in the melt. It was found that the equilibrium ratio k1 was not a true constant but varied with the oxygen concentration in the manner shown in Fig. 1, which is copied from the original paper. This early observation has been refuted in subsequent work but is confirmed in the present paper, and an explanation of the discrepancy can now be presented. In Fig. 1, the two points on the extreme right were disregarded in drawing the

Jan 1, 1959

By W. P. Blake

THIS furnace is designed especially for the oxidation of sulphur or arsenic in pyritic ores, but may be used for all oxidizing or desulphurizing operations, and for calcining, roasting and chloridizing ores, or for drying ores in large quantities preparatory to treatment by magnetic separators or otherwise. It is of the tabular, revolving type, with automatic feed and delivery, and is a modification of the Cornish calcining furnace, designed by Brunton, from which it differs, however, not only in construction, but in two or more radical features, for which novelty is claimed and patents have been applied for. The principal of these novelties are: 1. A series of concentric terraced floors, one above another. 2. The introduction of superheated air upon the charge without contact with the fuel. In the furnace which I have had in operation for more than a

Jan 1, 1893

By S. Storchheim

MORE and more attention is being paid to the bonding of metals in their solid states. For a better understanding of this technique for joining metals and how it is affected by changes in temperature, pressure, and time at temperature and pressure, a detailed report concerning nickel to aluminum bonding has been published.' In order to broaden the knowledge accrued, some additional work concerning solid state joining of aluminum to copper and aluminum to zirconium was performed. The investigation of the Al-Cu system was considerably more extensive than the investigation of the Al-Zr system. For the A1-Cu system, not only were tensile sudies made but intermetallic penetration rate investigations also were carried out. The effect of temperature on intermetallic penetration rate for the A1-Cu system was determined at 11 tsi pressure, held 2 min. Procedure Apparatus: The hot pressing technique was the means of solid state reaction used and required the equipment depicted in Fig. 1. The following procedure was involved: The two metals to be reacted were placed in an aquadag-lubricated 18-4-1 tool steel die, 16 in. high by 1.440 in. ID, between punches of 1.366 in. diam made of the same material. A thermocouple well was located in the die body 3½ in. down from the top of the die, while another well was located centrally in the bottom punch 8½ in. from the bottom of the die. This die assembly was located in three cylindrical ceramic heating furnaces placed in tandem. Each furnace was controlled individually by a Variac power transformer. In turn, the die and furnaces were placed in a water-cooled stainless steel pot which could be evacuated. A cover, which contained a centrally located Wilson seal with an 18-4-1 1 in. diam ram running through it, was bolted on the pot. After sealing, the pot was evacuated by a roughing pump to 200 microns pressure, after which a diffusion pump was used to bring the pressure down to 5 to 15 u. At this pressure, the furnaces were turned on. As soon as they started to heat, out-gassing of the entire unit raised the pressure to 30 to 400 p. By the time the specimens were at temperature ready to be pressed, approximately 4/2 hr, the vacuum pumps had re-established the 5 to 15 u pressure. Once the desired temperature was reached, the required pressure was applied for a predetermined length of time to the 1 in. ram, through to the top punch, and to the specimen. When the time for keeping the specimen under pressure had elapsed, the pressure was released, the energizing coil current turned off, and the assembly allowed to cool. After cooling, the die was removed from the pot and the specimen was ejected. Specimen Preparation: Two different types of specimens were made for this investigation. One was for subsequent tensile testing, while the other was for determination of intermetallic alloy zone penetration into the parent metals. Tensile Bars—Commercially rolled copper pieces in. thick or zirconium sheet pieces 1/32 in. thick and 1.366 in. diam were placed between commercial 2-S aluminum rod 1 in. thick and 1.366 in. diam. This sandwich in turn was slipped into a 2-S aluminum sleeve 1.438 in. OD and 1.370 in. ID. This sleeve lined the couple up and prevented the aquadag lubricant from getting in between either the A1-Cu or Al-Zr interfaces. Immediately prior to the specimen assembly, the copper or zirconium was abraded on the flat surfaces with 320 grit silicon carbide paper, producing clean smooth surfaces. The aluminum was chemically cleaned just before assembly by: l—degrease in acetone, 2—distilled water rinse, 3—immersion for 3 min in 5 pct NaOH at 70" to 80°C, 4—distilled water rinse, 5—immersion for 2 min in 50 pct HNO3 solution at room temperature, 6—distilled water rinse, and 7—drying in a blast of gas. After the A1-Cu sandwiches were hot pressed and ejected, the specimens were machined such that the aluminum sleeve was removed, and the remaining aluminum was then threaded; the bar so produced was tested later for tensile strength. In all the instances where Al-Cu couples were tested, the specimens broke during the test at the Cu-A1 interface and never within the aluminum or copper. The ultimate tensile strength values at times showed considerable scatter for a set of given reaction conditions. Because of this, as many as three to five specimens were made for a particular set of conditions. The trend of the average tensile strengths obtained was not as conclusive as was the trend of the maximum tensile strengths, the latter values being obtained under optimum reaction conditions. Therefore, the values of ultimate tensile strength given herein are maximums.

Jan 1, 1956