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CONTENTS [The Nature of Strain Markings in Alpha Brass (paper by J. E. BURKE and C. S. BARRETT, Met. Tech. Feb. 1948. TP 2327). ......................... 2 X Ray Studies of Twinning and Untwinning in Magnesium Alloys (paper by J. B. HESS and R. L. DIETRICH, Met. Tech. Feb. 1948. TP 2328) ................ 11 ]

Jan 1, 1948

By Alvin M. Cummings

During the period from June to September 1943, the Bureau of Mines undertook an investigation of the Vial mine, near the village of Linden I in Iowa County, Wis., as part of a general investigation of the zinc-producing area of southwestern Wisconsin. The work was done when every effort was being made to increase the production of metals important to the Nation's war program. The Bureau completed 29 churn-drill holes aggregating 3,348 feet of bore in areas immediately north and west of the old Vial mine, which had been worked intermittently between 1908 and 1939. The drilling supplemented a previous drilling program by the lessees and disclosed the presence of two significant zones of mineralization.

Jan 1, 1948

By W. J. Millard

It is well known that the excellent mica from Brazil played a most vital role in World War 11. Increased production from Brazil was necessary and with the assent of the Brazilian Government engineers from the united' States went down to study the problem. One of the means advocated by a former Chief of Metals and Minerals Branch of the Board of Economic Warfare, Mr. Carl Schwegler, was the use of stripping machinery. On a pegmatite dike favorable for stripping, a small number of miners could produce as much mica per day as an underground mine employing many times the same number of men. Type of Deposit Favorable The most favorable type of pegmatite for stripping is that which lies more or less parallel to the surface of the slope of a hill. Pegmatite dikes that cut directly into a hill are not suitable for stripping except where a long outcrop extends along the slope. Then just the outcrop could be stripped to some advantage. The ratio at no time should be greater than 1: 3; i.e., I cu. m. of pegmatite to 3 cu. m. of waste, unless the pegmatite was extremely rich throughout the dike. The Economics of Stripping When a mineral is badly needed to win a war, not too much heed can be paid to costs. The object is to produce rapidly and in quantity. Again, a deposit containing a superior grade of mica may be worked although the pegmatite is more costly to work than another deposit containing a considerably greater amount of poor-grade mica. For example: one deposit may have a little mica that will run 11 per cent good stain and better when trimmed and contain 2 kg, of crude mica per cubic meter of pegmatite, while another may run 0.2 per cent of good stain and better, and have a mica content of 10 kg. per cubic meter. A number of stripping projects were found in the vicinity of Raul Soares, Minas Gerais. These were sampled in the various old workings by cutting out a cubic meter of pegmatite and saving the usable mica. Because of the proximity to the surface of most stripping projects, about 34 per cent only of the mica is good for trimming. The other 66 per cent is too badly weathered to be of use. In Table I are given data collected on five properties near Raul Soares. In Table I, "value per cubic meter of pegmatite" means the value of the qualified mica obtainable. This does not include the half trimmed and splittings. By "value per cubic meter of matter" is meant the value of the qualified mica per cubic meter of material excavated; i.e. both waste and pegmatite. This is the figure that should be used to determine whether the operation will be profitable or not, as the cost of stripping is about $0.14 per cubic meter. With new machinery not subject to breakdowns, the cost per cubic meter should be around $0.10 to $0.15. To this should be

Jan 1, 1948

By George A. Morrison

A deposit Of limestone Was known to exist at a depth of 2000 ft under the property of the Columbia Chemical Division of the Pittsburgh Plate Glass Co. at Barber-ton, Ohio. A 662-acre site was selected for the projected mine, some two miles northwest of the plant and near the main line of the Erie Railway. An electric railway 8800 ft in length connects the mine and the plant. The deposit is overlain by 18 ft of un-consolidated (sand and gravel), 100 ft of water-bearing sandstone, and finally 2080 ft of shale into which thin layers of sandstone are embedded. The limestone bed itself is 345 ft thick and overlies 200 ft of gypsum. A 300-ton-per-hour capacity mine was decided upon. Two shafts were deemed advisable, one for production and the second for service; and these were sunk 550 ft apart, one to a depth of 2323 ft and the other to 2258 ft. Each is 16 X , ft in the clear and timbered with 6-in. steel H-sets on 8-ft centers, using 3/8 X 4 X 4-in. angles as studdles. Twelve inch I-beams were used at 160 ft intervals as bearers set in deep hitches and concreted to a vertical depth of 8 ft. Wooden sheathing was first employed but sloughing of the shale created a pressure on this lining and finally concrete was deemed necessary. Water was encountered only within the first 315 ft, flowing at 60 gpm and by grouting this was reduced to 15 gpm. Temporary wooden sinking headframes were installed with 300 cu ft capacity storage pockets and with doors automatically controlled by air cylinders by hoistmen. Each shaft was operated on four 6-hr shifts. The usual V-cut was drilled at the center of the shaft and the round consisted of 34 holes- A 40 pet gelatin dynamite was employed with some 60 pet in the cut holes. Drilling crews consisted of four to six men. Shale drilled easily but became more bed in depth and sandstone wore away the gauge Of the bits rapidly. A room-and-pillar mining system was adopted whereby rooms turned off on 80-ft centers, each room 40-ft wide and 50-ft high. Rib pillars between rooms were cut into 100 ft lengths by frequent holes. Ventilation is achieved through the two shafts, the down-draft through the service shaft, and the return through the pr0duction shaft. The hoist has a a Of 300 tons per hour and is equipped with 10-ton balanced skips requiring. a 2-min cycle Or 30 trips Per hour- Hoisting is designed to Operate by automatic loading and no hoistman is required for normal Operation. The 48 X 60 in. jaw primary crusher is underground. It is fed through chute and chain feeders and is flood-oil lubricated and cooled. An inclined belt conveyor returns oversized material to the crusher. Tramming is by Diesel-powered trucks.

Jan 1, 1948

Jan 1, 1948

By P. L. Menaul, P. P. Spafford

This paper discusses the most dangerous form of corrosion encountered in condensate-well oil production, the discovery of the agent causing this corrosion and the remedial chemical treatment proved effective by field use. The injection of ammonium hydroxide, even as little as one quart per day, has proved effective in preventing gas-condensate well corrosion. The injection of "Bone oil" is applied to wells producing brines containing calcium and magnesium salts. Introduction The severe corrosion of equipment in high pressure gas condensate wells has been a problem for several years and numerous papers have been presented on the subject. The term "gas condensate" or "distillate well" is applied to a distinct variety of hydrocarbon producing well, the characteristics of these wells being high bottom-hole (formation) pressures and temperatures and high gas production. These wells usually produce about 40 gal of water per million cubic feet of gas and the gas contains from 0.1 pct to 2 pct carbon dioxide. In the operation of this type of well, the operators soon noted severe internal corrosion of the pipe carrying the production. In some instances the dl tubing has been found to be corroded to perforation in eight months. The high well head pressures of 2500 to 8000 psi makes any weakening of the well equipment by corrosion extremely hazardous. Bursting of the well equipment results in very expensive repair operations and may result in the loss of valuable production, loss of the costly well, or even loss of life. In November 1944, a gas-condensate well in a Gulf Coast field producing from 9500 ft showed a tubing leak. The well was "killed" by filling it with mud through the tubing, and then the tubing was pulled. Corrosion was found to have caused several perforations in the tubing, Fig I. While the mud was being circulated to kill the well, a high pressure jet from the corroded hole in the tubing perforated the casing and damaged the well to the extent that the cost of repairs amounted to $75,000. In November 1944, another gas-con-densate well in a Gulf Coast field producing from 11,400 ft showed tubing failure. This well had been producing an average of 2,500,000 cu ft of gas daily for two years. In the workover of this well, it was found that a hole one inch in diameter had corroded through the tubing and that corrosion had reduced the wall of the tubing to less than half the standard thickness over large areas of the interior of the tubing. It was found that corrosion had also caused two leaks in the casing. After four months spent in an attempted workover failed to make the well safe for further production, the well was plugged and abandoned at a total loss of $300,000,00.

Jan 1, 1948

During the course of an investigation into the fundamental chemistry of the cyanide process, it became evident that the usual method of determining protective alkalinity was not completely satisfactory. T.he routine method for the determination of protective alkali by first removing free cyanide by the addition of silver nitrate, followed by titration of alkali with dilute acid using. phenolphthalein as indicator, can be shown to give correct results only when the alkali is a strong base. For other sUbstances, the alkali so determined is not the protective alkali. In view of the complications associated with the presence of zinc in cyanide solutions the discussion will be restricted in the first place to solutions free from zinc. Until the chemistry of simple cyanide solutions is placed on a quantitative basis it will be difficult to elucidate the behaviour of complex mill solutions.By the use of physico-chemical theory, it is possible to treat certain aspects of 'the chemistry of cyanide solutions in a manner which not only illustrates the important variables involved, but which also shows the need for care in interpretIng results obtained by conventional methods.In the following, comparatively simple systems are discussed; the principles established should be of use in interpreting the behaviour of more complex solutions. The methods described are not intended for routine work, but may be regarded as supplementary to the usual analysis carried out for the control of plant solutions, or as alternatives to be used when the latter do not appear to be satisfactory.

Jan 1, 1948

By S. M. Molloy

The postwar demand for insulating material such as mineral wool, in-creased in proportion to the rapid construction of new buildings, placing an enormous burden on the existing mineral-wool industry both in known sources of raw materials and in available plant facilities. In view of the increased demand on the industry, the Bureau of Mines in its Mississippi Valley Experiment Station" Rolla, Mo., undertook an investigation to aid the industry by establishing resources of raw materials through actual laboratory wool-blowing tests and by developing certain methods for evaluating the physical proper-ties of a mineral wool. As material was received from the public, laboratory-evaluation technique had to be developed, the results of which would reflect in the mineral-wool-industry standards. The methods of testing used by the mineral-wool industry, not only would have required a pilot-plant installation capable of producing a large tonnage of mineral wool, but each wool-blowing campaign would have been very expensive and necessitated considerable personnel and equipment. The present laboratory wool-blowing equipment can produce as high as eight wool-blowing campaigns a day. The nature, simplicity, and size of the physical testing equipment permit testing as small a sample of wool as 25 grams and give information as reasonably accurate as in a 40-pound bag of wool. The purpose of this report is to describe the laboratory method of production and the evaluation of the physical properties of small quantities of mineral wool. Considerable work has already been done, and the interest of the public and industry is measured in the number of samples being submitted for testing.

Jan 1, 1948

By C. H. Mathewson, Robert Maddin, W. R. Hibbard

A SATISFACTORY mathematical relationship between shear, elongation and change of orientation in the axial straining of a single crystalline rod or wire may be based on a simple model in which blocks, representing glide lamellae, glide over one another and rotate so as to maintain the original fixed position of the axis as the wire elongates.1 Obviously such a model gives no information concerning strain-hardening or the atomic adjustments at the sliding surfaces and their influence on the internal structure of the blocks. Moreover such an operation could be conducted with rigid blocks only by using self-aligning grips or a flexible attachment which would permit the adjacent blocks to freely assume their changing orientations. There are many unsolved problems (see Andrade2) concerning the manner in which nature departs from the simple model in permitting the essentials of slip, rotation and elongation to occur along with strain-hardening and the appearance of glide lamellae of various dimensions and special characteristics, not fully appreciated at the present time. This paper describes part of an extensive investigation on the development and characteristics of glide ellipses in single crystalline alpha brass rods and is especially concerned with the observation of cross-slip in crystals oriented wholly within the range of slip on a single system according to the generalizations first presented by Taylor and Elam in 1923.3 EXPERIMENTAL PROCEDURE Half inch rods of alpha brass (nominally 70 pct Cu, 30 pct Zn) * were converted into single crystals by melting and gradual solidification in a graphite mold according to the Bridgman method.' In order to remove any coring, the crystals were homogenized at 800°C for 16 hr and tested by etching with a 50 pct nitric acid solution. The back-reflection Laue technique as described by Greninger5 was used for determining the orientations of the crystals. Standard stereographic operations were employed to locate the specimen axis in a triangle of the Taylor and Elam3 projection so that the slip plane and direction could be predicted. The lineage structure reported by Buerger6 was present in the crystals used in this investigation. However, only crystals were used in which the internal differences in orientation did not exceed two degrees. Very little is known concerning the effect of the presence of a lineage structure on the detailed mechanism of plastic deformation.

Jan 1, 1948

Jan 1, 1948

By Walter E. Duncan

The ores treated at the concentrating plant of the Mahoning Mining Co. at Rosiclare, Ill., come largely from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately inter-grown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four coyentrates: (I) lead, (2) zinc, (3) acid-grade fluorspar and (4) metallurgical-grade Auorspar.2-3 Although fluorspar had been floated in Canada in 19214 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Alines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1948

By Walter R. Hibbard, Robert I. O’Herron, Fred D. Rosi, Howard T. Clark

INTRODUCTION AND PREVIOUS WORK IN the search to find a copper-base alloy with high strength properties, it was considered that the addition of a small amount of an age-hardening element to a binary system with useful strength properties might result in a series of alloys of theoretical and practical importance. [ ] The copper-nickel system was selected for its basic properties and chromium was added as the age-hardening element in view of the small percentage involved and its decreasing solubility with decreasing temperature. A study of the Cu-Cr binary alloys was undertaken as a preliminary investigation. However, in view of the response of these alloys to heat treatment, the scope of this phase of the investigation was widened. [ ]

Jan 1, 1948

By Curtis L. Graversen, J. H. Hamilton, John C. Nixon, John R. Lewis

RECENTLY there has been much speculation concerning the advantages of using oxygen enriched air or pure oxygen in pyrometallurgical processes. The advantage of using oxygen in the iron blast furnace and elsewhere in ferrous production metallurgy has been given careful consideration and, as a result, several experiments have been tried and a number of papers have recently appeared on this subject.1.2 The use of oxygen in nonferrous smelting likewise poses interesting problems, therefore any experimental data obtained in this area should likewise be of general interest.3 Ashcroft4 in 1933 claimed to have obtained improved results when copper concentrates were roasted in pure oxygen. However, his' paper gives neither data nor experimental details. A few years ago a research program was started in this laboratory which had for its objective the determination of the mechanism of oxidation of metallic sulfides under varying conditions. To date we have studied the oxidation of galena, 5 sphalerite, and chalcocite. It is a well known fact that copper sulfide minerals are oxidized slowly by weathering agents, so that the sulfides of copper which find their way into a tailing pond are gradually changed into watersoluble and acid-soluble compounds. Commercially, the oxidation of copper sulfide minerals is accomplished by roasting them at elevated temperatures in air using a suitable (usually multiple-hearth) roaster. Most copper smelters are equipped with such roasters. In this phase of the study we were especially interested in determining the effect of temperature on: (I) the air oxidation of chalcocite and (2) the oxidation of chalcocite using commercial oxygen. The amount of sulphur removed as sulphur dioxide, the percentages of water-soluble copper, and of acid-soluble copper were determined for the air oxidation tests. In the tests using commercial oxygen, the percentages of sulphur dioxide, water-soluble copper, acid-soluble copper, water-soluble sulphur and acid-soluble sulphur were determined. For these experiments the temperature was varied from 250 to 700°°C. EXPERIMENTAL Preparation of Chalcocite (Cu2S) Natural chalcocite is not pure. Minerals such as bornite, covellite, and chalcopyrite may be present. It was decided, therefore, to prepare some chalcocite in our own laboratory. An intimate mixture of sulphur and finely divided metallic copper in the ratio of 4 g of copper to I g of sulphur was prepared. To this was added a slight excess of

Jan 1, 1948

By T. A. Klinefelter, Glenn C. Truesdell, Richard W. Smith, R. G. Meara, O&apos

The investigation of domestic talcs was undertaken by the Bureau of Mines, Department of the Interior, in cooperation with the University of Alabama, at the request of the U. S. Army, on Dec. 1, 1941: "Considerable concern has been expressed by the Supply Arms and Service over maintaining continued availability of supplies of the necessary raw materials. These ceramic products are used by the Signal Corps in all army radio equipment, and to a greater degree in the new ultra-high-frequency equipment. The Air Corps has become a user of large amounts of radio-grade ceramics for' aircraft spark-plug covers required for high-altitude operations. In addition, the use of this material is essential to the proper functioning of the radio air-navigation aids of the Civil Aeronautics Board." History The use of steatite and talc in ceramic bodies goes back a great many years. If pure, talcs shrink little under heat-treatment and are very resistant to heat shock. The massive form, commonly called "steatite" by the trade, cuts and works easily, and this was naturally the first method used in obtaining articles of talc. When gas was in general use for illuminating purposes, a. large business was carried on in the manufacture of tips for gas burners.' With the development of uses for electricity, low-tension insulators were made in a similar manner; and, in fact, the forming of shapes from massive talc continues. However, as the business increased, cuttings and waste became more and more of a problem. In the twenties, one of the larger producing firms had a cooperative agreement with the Bureau of Mines at its Columbus station to work out a body using the finely powdered material. Somewhat earlier, manufacturers of spark plugs used the finely ground talc as part of the body and to some extent in wall-tile bodies. Somewhat later it came into use for kiln furniture and similar products and still more recently it has been used in some dinnerwares. With the development of the radio and the high-frequency fields generated, it was found that a dense steatite porcelain was far better than even high-tension electrical porcelain, which is a clay-flint-feldspathic type of body. Until the war increased the demand, the several pioneer steatite-porcelain manufacturers were able to furnish adequate supplies. When war began in Europe and Asia, three important foreign sources of raw material were cut off; i.e., the French, Italian, and Man-churian talcs. Domestic producers of high-grade porcelain are few, and the trade

Jan 1, 1948

By W. L. Fink, H. C. Stumpf, L. A. Willey

VERY few studies of the aluminum-sodium system have been reported. Heycock and Neville1 were unable to detect any solubility of sodium in liquid aluminum. Mathewson2 prepared an equilibrium diagram consisting of two horizontal lines at the melting points of the two metals. He was unable to detect solubility at either end of the system. Scheuer3 first developed a method for the chemical determination of sodium in aluminum and then studied the solubility of sodium in liquid aluminum. His method involved heating aluminum with excess sodium under hydrogen until the two liquids were at equilibrium, chilling the melt, and analyzing the aluminum-rich layer. He reported the solubility to be 0.l0 pct sodium at 700°C, 0.115 pct at 750°C, and 0.128 pct at 800°C. This paper on the aluminum-sodium system is the twenty-first in a series from the Aluminum Research Laboratories describing equilibrium relations in aluminum alloy systems. MATERIALS The alloys for this investigation were prepared from electrolytically refined aluminum (99.99 pct) and sodium of reagent quality conforming to specifications of the American Chemical Society.4 Chemical analysis of the aluminum showed the following impurities: 0.002 pct Si, 0.002 pct Fe, 0.002 pct Cu, <0.001 pct Mn, 0.002 pct Mg and 0.001 pct Na. The sodium content of all alloys was determined by either chemical or spectrographic standard procedures5 No. IIA and No. SII respectively. DETERMINATION OF LIQUIDUS The boundary of the miscibility gap was determined by a modification of the method of Scheuer3 and of that used by Henry and Cordiano.6 As finally developed the method was as follows: A fused alumina crucible 25 mm id with 75 mm depth and 1.5 mm wall thickness was placed on the bottom of a container made from a 4-in. length of 1 ½ -in. iron pipe and pipe caps. The space around the sides of the crucible was filled with 40 mesh crystalline alundum. A charge of 60 g of aluminum and 10 to 15 g of sodium was placed in the crucible, the sodium first being trimmed and rinsed with iso-pentane to remove kerosene. The container was tightly capped, filled with argon (99.6 pct pure), and heated to the desired temperature (between 670 and 800°C). A positive pressure of argon (about 8 mm of mercury) was maintained through a small pipe to the upper cap. After soaking for from 2 to 5 hr, the container was quenched by directing a stream of water onto the base, so that solidification of the aluminum-rich layer was rapid and essentially unidirectional. Analyses were made on drillings taken from the bottom of the ingots after machining off the surface layers. Pre-

Jan 1, 1948

By Stuart H. Ingram

The term lightweight aggregate implies material which may be substituted for the usual rock, sand and gravel commonly used as the major part of concrete, but distinguished by being much lighter in weight. Background The art of making and using concrete is old and well developed. Research and experience have pretty well determined the qualities necessary in its constituents, and time has tested theories and applications. Elaborate specifications and tests have been devised and generally adopted governing all phases of concrete work, from selection of material, through engineering design and down to mixing and placing. Standard practices have become routine to executives and laborers, and operations have been simplified and costs lowered. Though many lightweight aggregates meet the basic specifications necessary for usefulness, their physical qualities are so different that they fiequently will not pass standard tests, and some of them cannot be mixed and placed with the usual formulas for water and slump. As a rule their use produces a concrete with a lower modulus of elasticity which requires design change, and being produced in much smaller quantities than ordinary aggregate, production costs are higher. This latter works in a vicious circle to restrict their use and maintain prices that are not competitive, except as against each other. Increased Demand In spite of such handicaps consumption has grown steadily for the past two decades and is apparently growing even faster today. It is evident that lightweight concrete must have better qualities, and be sufficiently better to Overcome the higher cost and the other difficulties and disadvantages pertinent to the use of a new and different material in an old industry. Hence the study of its main ingredient should be of interest to all engineers. Our group has a double interest in the subject since most of the materials which have been used as sources of lightweight aggregate are either mined, Or are by-products of mining operations. We are directly interested in their production as well as their use, and since available supplies of materials suitable are limited, the increasing use, and rising demand, make the discovery of new Or or the development of processes to adapt other natural materials to this use, a promising field; and one in which our profession should take an active part. Recent Development Descriptions of ancient work contain references to the use of spalls of tuffs Or pumice in concrete masonry, but it is only since the nineteen twenties that serious attempts have been made to produce and market lightweight aggregates. A period of rapid advance in theory and practice in cements, mortars and con-

Jan 1, 1948

By Herluf P. Nielsen, Robert I. Jaffee

IN recent years much attention has been given to high-melting alloy systems, It has been of interest in this respect to investigate the alloys of platinum and tungsten, which were known from the literature to have very high melting points up to about 50 pct tungsten, but about which little had been published on the high-tungsten side of the system. Powder metallurgical methods were used throughout this investigation for the preparation of the alloys, utilizing electrical sintering techniques such as are used for tungsten and molybdenum. In the course of the work it was found desirable to determine fusion temperatures of the alloys, and it was possible from these and other data to draw a tentative constitutional diagram (solidus, mostly) applicable to the high-temperature part of the system. No attempt was made to determine solid solubility vs. temperature data, nor to investigate the lower temperature transformations in the solid state. Data on the microstructure, hardness, and a note on oxidation resistance of the sintered alloys were obtained. PAST WORK Müller1 determined liquidus temperatures of platinum-rich alloys to 50 pct tungsten, using a radiation method employing a photoelectric cell, and found that the melting points increased progressively from 1771°C for platinum to 2350°C for the 50 pct alloy. Microscopic evidence by Müller that complete solid solutions exist to 50 pct tungsten was not checked by Kremer2 who, according to Vines and Wise,3 found that "although the 8 pct tungsten alloy was a solid solution, the 26 pct tungsten alloy contained a eutectic or an intermediate phase." Hultgren and jaffee4 using X ray methods found continuous solid solutions to at least 33.6 pct tungsten (35 at. pct) after annealing in vacuum for 6 days at 1000°C. Vines and Wise3 report work communicated to them by E. C. Rhodes on wrought alloys to 7 pct tungsten, which showed that tungsten hardened platinum extremely rapidly, and was one of the most potent hardeners of the binary additions to platinum. Considerable quantities of 4 pct tungsten platinum wire have been produced for grids in radar tubes and for aircraft spark plugs. Some of the alloy for the latter application is made by sintering, the remainder by smelting and casting. MATERIALS The platinum powder used in this work was kindly supplied by the research and Development Laboratory of the Owens-Corning Fiberglas Corp., Newark, Ohio, It was prepared from scrap platinum. After dissolving in aqua regia and evaporating almost to dryness, the solution was worked up several times with concentrated HCI. the resultant platinic chloride was diluted considerably, and metallic zinc was added, which precipitated the platinum in a very finely divided form. After dissolving the

Jan 1, 1948

By P. H. Bohart

THe answer to the title question will be found by considering the ultimate influence of the petroleum engineers on industry and by considering the tools with which petroleum engineers must be equipped. Some 2000 to 3000 petroleum engineers are being absorbed by the industry yearly, indicating the magnitude of influence which they may exert on the industry. The petroleum industry is a highly technical one and it is inevitable that management be entrusted more and more to executives with technical training. The industry, therefore, expects the petroleum engineering schools to provide a broad basic training and a social and economic orientation which goes beyond merely supplying the minimum technical qualifications necessary to discharge responsibilities for a particular job. Because of the many branches of industry, emphasis must be placed on the Student&apos;s being trained in fundamentals, as well as grounded in other engineering branches, such as mechanical, civil and electrical engineering. The petroleum engineering graduate must know how to outline and organize the material for an engineering report and should be thoroughly grounded in English, both written and spoken. Even to the end that a five-year course may be required, they should receive instructions in auxiliary cultural and social science courses, industrial relations, and public speaking. Training in the method of approach to an engineering problem is something in which industry is deeply interested. A standard of ethics, a desire to arrive at the truth, and leadership training should be stressed. Introduction The basis of an an to any ques- tion concerning the petroleum engineering schools&apos; responsibility to the industry will be found by considering the ultimate influence of the petroleum engineers on the industry and by considering also the tools with which petroleum engineers must be equipped in order to achieve success and happiness in the role in which their training, individual talents and choice of a profession ultimately will cast them. The problem must always be considered as it relates to a large group and not as it relates to a particular individual possessing strong natural talents and destined, by virtue of them, to pursue a certain course. The question implies that petroleum engineering schools have a responsibility, and indeed they do, not only to the industry but also to the students and even to the public. Because the responsibility is concentrated in a relatively few schools whose graduates are absorbed largely by the one industry, the responsibility is more apparent and the effect of failure to discharge the responsibility fully is more serious than the effect of similar failure on the part of schools supplying a smaller portion of the industry&apos;s intake of engineering graduates. This fact is especially true with respect to an industry such as the petroleum industry whose ratio of technical to nontechnical employees is relatively high and is increasing.

Jan 1, 1948

By B. M. Larsen, T. E. Brower

In two previous paperslJ dealing with the carbon-oxygcn reaction, and the simultaneous content of each, in liquid steel in the furnace, we have made use of the quantity A[O], defined as the excess oxygen beyond what would be in equilibrium with carbon then present if the pressure of carbon monoxide were I atm.; that is, if [C] X [O] is taken as 0.00222, A[O] = [O] - 0.0022[C] where [O] is the percentage of oxygen by weight as measured by the method described earlier.3 This quantity ?[O] proves to be substantially independent of carbon content over the range 0.08 to I,z per cent, as well as of the basicity of the slag, and is nearly always within the limits o.o15 to o.oz5 per cent when the bath is boiling normally and is in what we have termed a steady state, Apparently it must be at least about 0.014 per cent before boiling occurs; but this threshold value seems to depend upon the ease of bubble formation at the metal-hearth interface, hence upon the condition of the surface of the hearth. In any case, ~(01, on the average, changes considerably following any disturbance of the bath, as by addition of ore. The present paper deals with the average changes in A[O] level following addition of spiegel or crop ends, or after procedures to promote stirring, or by tapping of the steel from the furnace. The observations were made on a large number of commercial heats, and consequently are subject to the influence of various disturbing factors; the results have the compensating merit of showing the consistency obtainable under the limitations imposed by practical operation. Variability of Excess Oxygen in Normal Open-hearth Operation The preceding genera&apos; statements refer to the average value of A[0] particularly in heats that are boiling actively; on individual heats, which are less active, there may be a variable time lag between the disturbance and its effect, which complicates the interpretation of what is going on. For instance, if in the normal refining process boiling is sluggish, the slag usually is gaining Oxygen from the gas phase above it, Or from added Ore$ faster than it is losing oxygen to the metal; Yet when the bath starts to boil vigorously, the opposite condition may prevail for Some time. The consequence of this lag is a variation of A[o], the interpretation of which in a sing1e heat may not be immediately obvious. This condition is illustrated in Figs. I and 2. In Fig. 1, A[0] at the beginning of the test period was near the lower limit of its usual range (dashed lines) in the steady state as a consequence of the lime boil, which had ended about as indicated. The bath is now cold, carbon drop is slow, and the first small orc feed causes only a small rise in A[O]; as the temperature rises and as lime dissolves in the slag, and the bath becomes more active, the

Jan 1, 1948

By Edward N. Trump

Extensive beds of rock salt occur in New York, Michigan, Kansas, and Texas. Wells are drilled through the beds, cased, and equipped with a suspended center tube. By circulating water through such a well, the soluble saline is dissolved and a brine produced. Operation of a well results ordinarily in the formation of an inverted conical cavity in the saline, inasmuch as the dissolving action of the fresh water is more rapid at the top of the bed than at the bottom. Under continued operations, the side wall of the cavity becomes so flat that a blanket of mud is deposited thereon and if the bed is overlain with shale or a soft rock caving follows. Three methods of operating wells have been employed—the "Tully plan," "the Detroit plan," and "the Trump plan." Both the Tully plan and the Detroit plan involve the formation of the aforesaid conical cavity with constant caving of shale above and deposition of mud on the side wall; the Trump plan was developed to avoid both these circumstances and thus materially prolong the life of a well. Details of the three plans follow. The Tully Plan In the Tully plan, the water passes down the annular space between the casing and the center tube, the brine rising in the center tube (Figs. I and 2). An inverted conical cavity is formed, the side wall of which relatively soon becomes so flat that it is blanketed by mud and impurities from the rock salt; this blanket limits dissolving action to the perimeter of the base of the cone, situated at the top of the saline bed. Caving of the overlying strata follows, and the total production of brine from the well is comparatively small. The Detroit Plan In the Detroit plan, the water passes down the center tube, the brine rising in the annular space between the center tube and casing (Figs. I and 2). Initially, this flow forms a pear-shaped cavity in the bottom of the salt bed, but this cavity soon changes in shape to that of an inverted cone, as under the Tully plan, with the attendant disadvantages inherent in that plan. Application of the Detroit plan is limited to salt beds 300 ft. thick as a minimum. This depth of bed is necessary to prevent fresh water of lower specific gravity than the brine from rising in the

Jan 1, 1948