Search Documents

By E. W. McAllister

The decision of the California Oil Umpire's' office to accept well potentials established from subsurface-pressure data has brought to the attention of many operators for the first time the application of such data to production problems. This paper reviews the development and application of subsurface-pressure data in kettleman Hills, where these data have been used in production work since 1932. The Kettleman Hills field was discovered in October 1928. The first subsurface-pressure surveys were made early in 1932 and by the middle of 1933 the two major operators in the field were making routine pressure runs in thcir respective wells. Shortly therefter, these companies agreed to exchange pressure data, and later a third company entered into the agreement. This made it possible to analyze the effects upon the reservoir pressure of past production practices, which are discussed in the latter part of the report. It is interesting that many of the applications of subsurface-pressure data discussed herein were anticipated by C. L. Clark and C. B. Kimberlint in a report2 dated March 1933. PRODUCTIVITY Index In the majority of individual well problems involving the use of subsurface- pressure data, it is necessary to use a factor known as the "productivity index," which indicates the ability of a well to produce and represents the number of barrels of oil or oil and water per day that will enter the well bore for each pound per square inch pressure drop made opposite the producing interval of the well. The productivity index may be determined from pressure and production data for one or more rates of flow and a static or shut-in pressure survey or from production and pressure data for two or more rates of flow. Productivity indices may be based upon either net oil or total fluid production. All indices referred to in this report are based upon net oil. The value of the productivity index for Kettlcman Hills wells decreases generally with time and it is not uncommon for the index to vary in value at any specific time fur various rates of flow. The decline of the productivity index may be attributed partly to the loss of reservoir pressure. As the reservoir pressure decreases, gas originally in solution is released from the oil, congesting the flow of fluids through the formation and increasing the viscosity and surface tension of the oil; thus the flow resistance of the fluids through the formation is increased, resulting in a decrease of the productivity index. For convenient reference, the productivity indices for each well are plotted against time and a curve is drawn through the points showing the trend of the average index value. Plotted on Fig. I are index values for five

Jan 1, 1941

By Glenn C. Waterman, Scott Hazen

3.1-1. Purpose. Development drilling is directed towards determining the grade, volume, and three-dimensional outline of a mineralized zone previously located by exploration. It is distinguished from exploration or reconnaissance drilling, which has as an objective the discovery of new mineralized areas. A development drilling and bulk sampling program should furnish the following information: 1. Geology of the mineralized zone. 2. Quantitative data on grade and tons of material within pertinent cutoff limits. 3. Physical size and shape of the deposit. Areas underlain by waste as well as ore. 4. Mineralogica1 and metallurgical characteristics of the ore. 5. I1hysical characteristics of the ore. 6. Hulk samples for metallurgical testing and grade check. 7. Data on other factors that could affect mining operations, such as ground water, ground conditions, etc. 3.1-2. Procedures. There are numerous methods and techniques that are used to acquire pertinent data on a potential ore deposit. The total cost and accuracy of a development program depends on significant evaluation of the geology, proper selection of a drilling method, thorough analysis of sample data, and accurate evaluation of all the information. Knowledge of the geology of the mineral deposit is important in planning the development testing program and in evaluating the data obtained from such a program. Exploration work usually provides some significant information on the general size, shape, thickness, grade, and geology of the deposit. In some cases no further data are needed; in many cases geological complexities suggested by the preliminary work require further geological study. Of considerable importance may be the following:

Jan 1, 1968

By W. O. Philbrook, A. H. Jolly, T. R. Henry

In the course of a study of slag-control methods, the authors devised a laboratory technique by which the basicity of basic open-hearth furnace slags could be estimated with sufficient accuracy to make the method of possible interest for slag-control purposes. The process consisted of measuring the slight basicity of an extract of the powdered slag in water. The procedure was presented, for the benefit of those who might care to experiment with it, in a paper submitted in the 1944 McKune Award Contest and presented at the 1944 Open-hearth Conference in Pittsburgh.' Interest expressed at that meeting and favorable operating experience with the method at a neighboring plant, as described in a related paper by Michael Tenenbaum and C. C. Brown,; have encouraged us to present further details of the development, technique, and pitfalls of the method and some additional data not included in our first report. At the outset it must be admitted that this method of estimation of slag basicity is not perfect; neither is any other that we know of. It is subject to somewhat greater errors than would be desirable, but compares favorably with other techniques in this respect. It has the disadvantage that it places the final determination of slag basicity in the hands of a laboratory technician rather than the furnace oper- ator, although it can be operated in a control laboratory adjacent to the furnaces. It requires close attention to certain details, as do most laboratory procedures. On the favorable side, it eliminates the "human factor" of personal judgment inherent in estimation from pancakes or powdered samples. The procedure can be placed in practice much more rapidly and presents much less of a training problem than visual methods of estimation. The principle and operation of the method are so simple that it would not be surprising if others have experimented with the same or similar procedures, but we are not aware of any prior publication of details. Our own shop has been operating fairly successfully for several years with the aid of slag pancakes. Since the past year or two have not been favorable for the expenditure of time and energy where the need was not urgent, we have not introduced the new practice for actual slag control and hence do not have extensive operating data. In presenting this method, therefore, we make no claims as to theoretical justification nor guarantees as to practical applicability, but offer it as an empirical observation. We believe it has sufficient merit to warrant its investigation by those just instituting a slag-control program or in search of supplementary methods for judging slag basicity Principle of the Method It is well known that when burnt lime or slaked lime is shaken up with water it

Jan 1, 1945

By P. K. Koh

Chi phase (body-centered cubic, a = 8.89A) was found in as-cast 23 pct Cr-10 pct Mo-Fe alloy as well as in heat-treated 316, 316L, 317, and modified 446 stainless steels. Chi phase resembles sigma phase in its metastability, chemical composition, and brittleness. Microstructure and some of physical properties of chi phase were developed. MOLYBDENUM has been established as a desirable alloying element for the enhancement of corrosion resistance in Cr-Ni steels and creep and rupture strength in high temperature alloys. On the debit side, however, the molybdenum addition has been known to stimulate u formation. When the molybdenum-bearing steels were subjected to prolonged heating at elevated temperatures, losses of corrosion resistance and/or ductility have been observed. Andrews1,2 first found the existence of a new phase among the anodic FeC13 extracts from type 316 stainless steel solution treated and then reheated from 800" to 1100°C. This phase was designated x and was identified by diffraction patterns as of body-centered cubic structure with a value equal to 8.89? and similar to that of a manganese. He found that the x phase in this steel always coexisted with s phase, but failed to distinguish one from the other by metallographic technique. No definite conclusion was drawn regarding the equilibrium temperature ranges of these two phases. The appearance of Andrew's work recalled an un- explained structure observed several years ago at the Allegheny Ludlum Research Laboratories during the investigation of an experimental alloy containing 25 pct Cr, 10 pct Mo, and balance Fe. X-ray diffraction patterns on the as-cast material revealed the presence of a body-centered cubic phase having a lattice parameter of 8.89?. The structure was sufficiently similar to that of y chromium that for a time it was thought that the structure might be caused by the presence of some undissolved electrolytic chromium particles. The alloy as-cast was so brittle that it shattered after the slightest hot reduction by hammer cogging. This paper gives the results of some metallographic and X-ray diffraction studies on x phase as found in several experimental alloys as well as a few molybdenum-bearing commercial stainless steels. The analyses of materials studied were as shown in Table I. Experimental Procedure Metallography: Solid specimens were polished mechanically and etched in boiling 50 pct HC1 solution in order to best delineate x and s phases in the alloys studied. X-Ray Diffraction: The difficulty of satisfactorily developing diffraction lines for s phase in solid specimens has been previously reported by Bindari, Koh, and Zmeskal.V or this reason all specimens used in this investigation were subjected to anodic

Jan 1, 1954

By Erwin H. E. Gaertner

The brown-coal opencast mines in Germany's Rhineland have to cope with several problems. Predominant are densely populated areas with highly productive farmland, many railroads, highways, and rivers, as well as towns and farm settlements which have to be removed before or during the lifetime of a mine. Also the ground- water table has to be lowered to keep the deep opencast mines dry and stable. These problems urged a development towards deep opencast mines with powerful machines; the large material movement and the high coal production are the basis for the economic success in working the Rhenish brown-coal deposits. Up to now, bucket wheel excavators with a daily capacity of 100,000 m3 (130,000 cu yd) are in operation. Machines of twice this capacity are under construction. Corresponding to the excavators development, the stackers' capacity has to be increased too. The new stackers will reach a daily capacity of 240,000 m3 (310,000 cu yd). The haulage in the mine is performed by conveyor belt systems, which are arranged in such a manner that the belt lines meet at a belt switching point. Here the incoming lines may be connected to the outgoing ones so as to direct the material flow to the desired receiving stations. At the moment there are conveyor belts in operation with a width of 2.2 m (7.2 ft); the new belts will have a layout for 37,000 t/h (40,770 stph), the width of which amounts to 3.0 m (9.8 ft). All belts and driving stations on benches are shiftable.

Jan 1, 1976

By William Fink

THE investigation of the phase relations of high-purity aluminum-base alloys is a part of the fundamental research program of the laboratories of the Aluminum Company of America. The results of a number of binary and ternary systems have previously been published in the transactions of this society. The aluminum end of the binary aluminum-titanium system has been studied. This system has received relatively little attention. To the authors' knowledge, there are two previous papers on this subject. E. van Erckelens1 determined the constitutional diagram reproduced in Fig. 1. .According, to this diagram, the liquids represents the separation of TiAl4 (Ti = 30.7 per cent.). He concluded that this is the composition of the compound, because in his work the eutectic horizontal terminated at 30.7 per cent. titanium and the metallographic examination of a sample of this composition revealed only one phase. According to the constitutional diagram developed by W. Manchot and A. Leber,2 reproduced in Fig. 2 from data given in their article, the liquidus represents the separation of the compound TiAl3 (Ti = 37.2 per cent. ). The composition of the compound was derived in a manner similar to that utilized by van Erckelens. Moreover, Manchot and Leber confirmed the composition of the compound by analysis of the constituent left as a residue after dissolving the alloys in weak sodium hydroxide solution. Although these are the only publications describing the constitution of the system, other investigators have studied the composition of the titanium compound. L. Wöhler3 says that there is a compound TiAl3 which was later confirmed by W. Manchot and P. Richter4 and by

Jan 1, 1931

By Edwin T. Myskowski, Alfred H. Hesse

THE development by the U. S. Bureau of Mines of a process for purifying manganese by electrolysis has stimulated interest in manganese alloys. While recent publications have shown some alloys of manganese and copper to possess rather unusual properties, none have included information on tensile properties, which have an important bearing on the usefulness of the alloys. Manganese-copper alloys can be heat-treated, since the single phases that exist at high temperatures may be retained by quenching. According to Persson,1 when quenched alloys containing from about 30 to nearly 100 per cent manganese are reheated, alpha manganese is precipitated (Fig. I). Calculations based on Persson's X-ray data indicate that this decomposition is accompanied by a change in volume of as much as 5 per cent. The reason that this large volume change has not been noted probably is the slow rate at which equilibrium is reached by these alloys. In an investigation on hardening by cold-work and heat-treatment, Dean and Anderson' found that quenched alloys increase in hardness with manganese content and that high-manganese alloys harden very slowly with cold-work, while high-copper alloys harden much more rapidly. Heat-treatment has little effect on manganese-copper alloys containing less than 40 per cent Mn, and affects those alloys containing more than 40 per cent in accordance with their manganese content and heat-treating temperature. Dean and Anderson4 also found that cold-work increased electrical resistance of alloys containing from 50 to 80 per cent Mn, and that heat-treatment of alloys containing from 40 to 96 per cent Mn altered the electrical resistance considerably. Dean, Anderson and Potter5 have shown that the high vibration damping capacity that characterizes manganese-copper alloys containing 40 to 90 per cent Mn is affected by cold-work and heat-treatments. Highest vibration damping capacity at low stresses was obtained by heating for 2 hr. at 450°C. (842°F.) after quenching from solution temperatures. Preparation of Alloys.-The alloys for these tests were melted in a 17-lb. basic-lined high-frequency induction furnace using electrolytic manganese, grade A ingot copper and a 70-30 Cu-Mn master alloy (69.80 per cent Cu, 0.12 Fe, 0.04 Si, 0.006 S, 0.07 C, balance Mn). The manganese was heated slowly to expel excess dissolved gases and melted before the addition of copper. No fluxes or covers were used. The melts were cast in tapered iron molds and the ingots were reduced 90 per cent when forged to ½ -in. rounds. These rods were given 2-hr. solution treatments at temperatures that depended upon the composition (Table I). Tensile Tests.-The tensile specimens, 0.313-in. diameter (1.25-in. gauge length), which were enclosed in evacuated Pyrex

Jan 1, 1942

By Rex B. McLellan, W. A. Oates

FROM a consideration of the application of the Gibbs-Duhem equation to multicomponent systems arken' derived the following equation applying to dilute solutions of a solute, 2, dissolved in a binary solution, 1-3:

Jan 1, 1969

By M. H. Caron

BASIC U. S. Patent 1,487,145 on ammonia leaching of nickel ores was issued to the author on March 18, 1924. Equivalent patents in other countries were obtained later. The Dutch Syndicate Brikcarbo became interested in the process in 1935; and a couple of years later the late Dr. H. Foster Bain sent samples of Surigao laterite iron ore to Delft and witnessed tests made there on this material for the Philippines Commonwealth. As a result, the author was asked in April 1940 by Dr. Bain to go to the Philippines to erect and take charge of a pilot plant for the process. The World War prevented carrying out this proposal. The Dorr Co. also became interested in the process at about the same period. The investigations by the Freeport Sulphur Co. on the process as applied to the Cuban nickeliferous laterites, which resulted in the Nicaro enterprise, and the results of this operation have been well described elsewhere.1,²,³,4,5 This Nicaro plant was, in operation for almost two years, and during this period produced about 10 pct of the world nickel production from laterites containing 1.35 pct nickel. This plant and its operation were war measures and, in view of this, activities were suspended in April 1947. The results obtained fully demonstrated the technical feasibility of the process and its economical aspects on a commercial scale. In this respect, it should be understood that it is probable that improvements may be made by further development, and that there are possibilities for advantageous application of the process to garnierite and similar ores with higher values in nickel than the laterite iron ores at Nicaro. While the articles cited above have given a certain amount of information, no general article containing all the important process data has been pub- lished. Since the process is of more than local interest, a fuller knowledge of the fundamental and practical factors of the method may be welcome to those interested in this new field of metals technology. The author, accordingly, takes pleasure in submitting this article to the AIME as it represents in a condensed form the results of many years of research. The brief outlines of the fundamental and other factors and the explanation of observed phenomena are presented with as little discussion of details as possible, consistent with clarity. It is a special satisfaction to the author to make some contribution of his own in return for the benefits of the many valuable publications issued by the Institute during his thirty years of membership. Ores Adapted for Ammonia Leaching: All nickel and cobalt ores which originate from weathering of peridotites or similar basic rocks having sufficient values are suitable for treatment by the ammonia leaching process after a preliminary reduction under proper conditions. The formation of these deposits was probably as follows: In the course of time the basic rocks were attacked by atmospheric agencies; MgO and SiO2 were gradually leached out, and secondary nickel minerals formed, such as garnierite, a hydrated silicate of nickel and magnesia. These secondary products are, however, not stable. They decompose in a further stage of weathering and ultimately only a relatively small residue of insoluble oxides remains, known as laterite iron ore with a small nickel content and very little cobalt. Under these mantles of laterite, richer nickel values may be found, usually indicated by the occurrence of garnierite. The more the ore is disintegrated by nature, the higher the iron content and the better the nickel extractions that may be expected therefrom. Table I shows extractions that may be expected from different types of ores, assuming treatment of -200 mesh products and that all precautions have been taken for obtaining maximum extractions. As for the distribution of the various nickel minerals and compounds that may be present, great variations may occur from locality to locality as well as vertically in a deposit. From such "run of

Jan 1, 1951

By Robert L. Akright

LOCATION AND HISTORY The Candelaria silver mine is located in an arid, isolated desert setting in Sections 3 and 4, T.3N., R.35E., and Section 33, T.4N., R.35E., Mineral County, Nevada (Fig. 1). Silver was discovered in 1863 by Mexican explorers, and high grade silver veins were mined from the early 1870s through the 1880s. Essentially all mining had ended by 1890. The veins were mined by underground methods typical of the period. As is the case in many mining districts, Candelaria was studied and explored by a number of companies. All these efforts were terminated short of mine development because of low silver prices and difficult metallurgy. On March 1, 1976, Occidental Minerals (Oxymin) entered into a Limited Partnership agreement, known as Candelaria Partners, as General Partner with Congdon and Carey Ltd. No. 4. Prior to the agreement, Congdon and Carey had acquired the candelaria properties as part of a bullish silver program and did considerable exploration which indicated the possibility of a large reserve of shallow silver mineralization. GEOLOGY AND ORE DEPOSITS Silver and gold mineralization in the Candelaria district occurs in high-grade fissure veins with silver values ranging from 0.6 to 7.8 kg/t (20 to 250 oz per st). These veins, combined with the low grade disseminated mineralization in the intervening material between the veins, form Manto- like deposits 15 to 46 m (50 to 150 ft) thick. Three such deposits are found in the Candelaria district: the Lucky Hill, Mount Diablo, and Northern Belle deposits (Figs. 2 and 3). The average ore grade of the Lucky Hill and Mount Diablo [ ]

Jan 1, 1985

By C. A. Stetefeldt

In May and June, 1885, I was engaged in examining the goldores of Las Minas, in the State of Vera Cruz, Mexico, with a view of finding a cheap and efficient method for extracting the gold. The following minerals could be plainly seen in the ore : magnetite, pyrites of iron, pyrites of' copper, quartz, and garnet. Neither galena nor zinc-hlende mere visible, but the ore contained a trace of lead, and in some samples of ore from the Muertos mine, 1.19 per cent. of zinc mas found. Antimony and arsenic were not present. Native gold could scarcely be seen in specimens of' the ore, and it only appeared—mainly in extremely fine particles—after pulverizing and washing. Magnetite was in every instance the most prominent mineral, its quantity varying between 43 and 67 per cent. Determinations of copper and sulphur showed that from 3.5 to 7 per cent. pyrites of copper, and from 3 to 22 per cent. pyrites of iron were present. The gold-value of the ore varied between 0.3 oz. and 0.8 oz. per ton. I do not speak of specimens, but of average samples of large quantities of ore put through the battery. In ore that had been crushed through a No. 40 screen, the gold appeared, in part, as free gold, the rest being thoroughly mixed with the magnetite and pyrites, both minerals containing about the same percentage of gold. If this free gold was brought in contact with pure quicksilver it (lid not amalgamate readily, even if the quicksilver was charged with sodium amalgam.

Jan 1, 1886

By C.F. Pascoe, H.J. Roast

The object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4.Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests.' 8. Brinell tests, at temperatures from 80' to 400' F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: Lead. ' Antimony, Tin, CoppeR ARSEnic, NumBeR PeR CENT. PER CENT. PeR Cent. PeR CeNt. PeR Cent. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.04 59.87 2.65 none I 3 none- 0.13 ------------1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 . 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sul-furic acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thio-cyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference. With the exception of No. 5, the alloys were taken as received from the manufacturer. About 40 lb, were melted in a graphite crucible and recast into small ingots to insure uniformity of composition; samples

Jan 1, 1923

By C. F. Pascoe, H. J. Roast

The object of this investigation was to compare the arsenical antimony-lead alloy with some of the regular bearing-metal alloys. With this end in view, the following tests were made: 1. Chemical analyses to show the composition. 2. Thermal analyses. 3. Macroscopic examination of fractures. 4.Microscopic examination. 5. Tensile tests. 6. Crushing tests. 7. Scleroscope tests.' 8. Brinell tests, at temperatures from 80' to 400' F. Tests 3 to 8 were conducted on both sand-cast and chill-cast specimens. For the purposes of this paper the alloys have been numbered from 1 to 6; their composition is as follows: Lead. ' Antimony, Tin, CoppeR ARSEnic, NumBeR PeR CENT. PER CENT. PeR Cent. PeR CeNt. PeR Cent. 1 82.64 12.95 4.40 0.01 none 2 27.54 9.04 59.87 2.65 none I 3 none- 0.13 ------------1.47 4 82.95 16.85 none 0.20 none 5 83.78 15.35 none 0.05 0.82 6 0.15 7.87 . 84.13 7.85 none The analysis, in the case of the antimony, was made by solution in sul-furic acid and titrated with permanganate. For tin, the ferric-chloride method was used; for copper, the alkaline separations, and the thio-cyanate methods; for arsenic, volatilization of arsenic and titration with iodine. The lead, with the exception of alloy 6, was taken by difference. With the exception of No. 5, the alloys were taken as received from the manufacturer. About 40 lb, were melted in a graphite crucible and recast into small ingots to insure uniformity of composition; samples

Jan 1, 1923

By H. B. Goodrich

Anthony I?. Lucas died suddenly at his home in Washington, D. C., on Sept. 2, 1921. Captain Lucas, as he was known to us, was born in Dalmatia, Austria, in 1855, of Montenegrin ancestry. He was graduated as an engineer at the Polytechnic of Grate and served in the Austrian Navy as second lieutenant. In 1879, he obtained leave of absence and visited an uncle in the United States. After an extension of this leave of absence, in order to undertake an engineering engagement in the lumber district of Michigan, where he resided, he decided to become an American citizen. He was naturalized in May, 1885. His name was Luchich, but as his uncle had adopted the name of Lucas, which was more easily pronounced by Americans, from his entrance to this country, he used this Anglo-Saxon form. Without knowing this fact, upon first meeting him a person was sometimes surprised to note the rather Germanic pronunciation of the Captain. Although he subsequently revisited Austria with Mrs. Lucas, he made his permanent home at Washington, D. C. His son served with distinction in the A. E. F. during the World War. His activities in this country as a mining engineer were at first in Colorado and later in salt mining at Petit Anse and Belle Isle, La. During his salt investigations, his attention was directed toward the possibility of oil in the Gulf Coast region and in January, 1901, his well, the "Lucas Gusher," on Spindle Top, Tex., started a new era in the oil business and his reputation as discoverer made him famous throughout the world. Captain Lucas became a member of the A. I. M. E. in 1895. During 1914, 1915, 1918, and 1919, he was chairman of the Petroleum and Gas Committee of the Institute and was at all times prominent in Institute affairs. As to the personality of Captain Lucas, the lasting impression is of a courteous hospitable gentleman, genial, affable, obliging, and helpful with his advice br assistance to any colleague. He was sincere, honest, firm against all obstacles, backing his judgment with his own hard work along any course which he had determined to be correct. His value in the engineering world lies mainly in the petroleum industry. In the oil business all wild-catters are pioneers that deserve credit and gratitude upon their success. There are, however, names that particularly stand out in our history. Drake conquered such obstacles as ridicule, lack of finances, and started the oil business. In 1901, Captain Lucas had the conviction that Spindle Top, a dome rising about 12 ft. above the coastal prairie south of Beaumont, Tex., contained commercial oil. He was scoffed at by practical oil men of the East. Noted geologists

Jan 1, 1921

By Carle Hayward

THIS work was first undertaken in the metallurgical laboratory of the Massachusetts Institute of Technology in 1907 by L. A. Dickinson, E. Phelps, and V. S. Rood, under the author's direction. They attempted to determine the diagram for the system Cu2S-NiS, but several inexplicable results were obtained. The following year, P. H. Mayer with Mr. Dickinson carried, on some experiments with nickel sulphide and found that the molten material slowly lost sulphur until it reached a composition corresponding to the compound Ni3S2, established by Bornemann.1 At this point the composition remained stable. The freezing points of the mixtures gradually fell off with the loss of sulphur. Since the above facts seemed to explain the erratic results obtained in the first experiments, Mr. Mayer with F. Jaeger attempted in 1910 to establish a correct equilibrium diagram for the Cu2S-Ni3S2 series. They prepared pure Cu2S and Ni3S2 and determined the cooling curves for various mixtures. The fusions were made in a platinum resistance furnace and the temperatures were measured with a platinum-iridium thermocouple connected with a Sullivan galvanometer. The curve was established for mixtures containing more than 40 per cent. Cu2S, but they were unable to determine certainly the direction of the curve beyond that point. Nothing further was attempted with this series until the present year, when the work was carried to a successful conclusion with the assistance of A. Butts, Jr., and E. H. Weil. The results of this work are given in the following pages.

Jan 1, 1914

By E. M. Elkin, J. H. Schloen

All known methods of treating and recovering the various components of copper refinery slimes are discussed. The slimes treatment processes presently used by five copper refineries are described and flowsheets are given. A bibliography is appended. THE recent increase in anode copper production at the Noranda smelter placed an additional load on the Montreal East plant of Canadian Copper Refiners Limited. The consequent increase in the amount of anode slimes led to a careful consideration of the various methods of slimes treatment. A critical review of these methods is presented here in the belief that it would be of interest to copper refiners. No such review has yet been published, although a number of individual flowsheets have been described. To make the information of greater value, a questionnaire was sent to all copper refineries. While the overall response to the questionnaire was not entirely satisfactory, the foreign refiners gave much detail in their answers and expressed willingness to cooperate further if needed. Although the data below do not cover all refineries, it is thought that examples of all flowsheets currently in use are included. A comparison of these flowsheets should prove of interest. The paper is divided into two parts. One part deals with a discussion of all known methods of treating and recovering the various components of the slimes. The second describes the processes now in use at various copper refineries for the treatment of raw slimes up to the production of Doré bullion. A brief description of the two main electrolytic methods of parting Doré bullion is appended. Since the first commercial application of electrolytic refining of copper by J. B. Elkington in 1865, the treatment of anode mud or slimes has undergone many changes. The object of treating slimes has always been the recovery and separation of precious metals. With the passage of time, increasing attention has been paid to improvement in recovery and quality of byproducts, mainly selenium and tellurium. The original methods of treatment were crude, although relatively direct. For example," the slimes were fused on a dolomitic clay bed in a reverbera-tory furnace; the slag was removed and the resulting metal cupelled to silver and gold. In another process, the dried slimes were oxidized and scorified on the lead bath of a cupelling furnace. Later, these pyrometallurgical methods were modified by sulphuric acid leaching, followed by fusion with niter in a reverb eratory furnace. A preliminary oxidizing roast improved the leaching step. Straight hydro-metallurgy has not found practical application. No matter what method is employed, the result is Dorb metal—a silver-gold alloy, with or without the metals of the platinum group. Characteristics of the Slimes Anode slimes are made up of those components of the anodes which are not soluble in the electrolyte. They contain varying quantities of copper, silver, gold, sulphur, selenium, tellurium, lead, arsenic, antimony, nickel, iron, silica, etc. The color of raw slimes is grayish black and the particle size is —200 mesh. However, Baraboshkin and Gaev (Ref. 1, p. 37), describe light gray colored slimes from the electrolysis of secondary copper which are composed predominantly of sulphates and of basic salts. In anode furnace practice primary and secondary metals are not usually segregated and such slimes do not commonly occur. Copper in the slimes originates to a large extent in the cuprous oxide of the anodes and is therefore dependent on the oxygen content of the latter: Cu2O + H2SO4 + CuSO4 + H2O + Cu [I] Opinions have been expressed that some of the copper is formed by decomposition of cuprous to cupric sulphate Cu2SO4 ? CuSO4 + Cu PI The product of either reaction is very finely divided. Depending on the slimes composition, much of the copper is combined with sulphur, selenium, tellurium, etc. The balance of the copper is in the free metallic state.

Jan 1, 1951

By K. D. Jacob, W. L. Hill

THE bulk of natural phosphates is comprised of calcium phosphates, which are usually apatites;1 calcium aluminum phosphates such as pseudowavellite;2 and aluminum phosphates, which occur in extensive deposits, notably on Grand Connetable Island.3 Nearly all deposits of current commercial value are apatite, often with admixed material of other classes. In the trade coarsely crystalline material, mainly fluorapatite of igneous origin, moves as apatite, whereas the fine-grained material of sedimentary origin is marketed as phosphate rock. The latter classification includes the phosphate shales in western North America. Although several genetic varieties of phosphate rock are recognized,4 most known reserves are of marine origin. Calcium phosphate is concentrated from Florida land pebble and Florida hard rock deposits and from Tennessee brown rock deposits. Phosphate Rock as a Fluorine Carrier: More than two decades ago Reynolds and Jacob surveyed the fluorine content of commercial phosphate rock from various parts of the world.3,5 Later these samples were reanalyzed by a new and more reliable method of determining flourine.6,7 The revised results with interim additions of new samples, published a few years ago in a brief summary,' are shown in Figs. 1-3. Through the points representing the respective varieties of rock medial straight lines are drawn. In some instances the lines are well defined by the points; in others, where scattering renders the position of the best line uncertain, the greater weight is given points near the end of the covered range. Wherever distinct varieties are not recognized among regional deposits the custom is to identify the sample with the geographical location of the deposit.

Jan 10, 1954

By Sidney A. Swensrud

The writer has been interested for some time in trying to appraise our industry's prospects for gasoline consumption over the next half dozen years or so. Anyone who has even approached the problem knows the intimate relationship to it of the number of automobiles in use— by automobiles I mean both passenger cars and trucks. Probably at least 90 per cent of our gasoline is used in automotive vehicles. Some recent estimates secured from the Department of Commerce show the following breakdown: Per Cent Passenger cars............................................ 60.0 Trucks................................................... 27.4 Busses................................................... 2.9 Taxicabs................................................. 2.2 Subtotal............................................... 92.5 Aviation, motor boats and railroads......................... 0.5 Agricultural: farm tractors, etc.............................. 3.3 Balance (manufacturing, government use, construction, etc.)... 3.7 100.0 Up to the beginning of the depression there was, of course, a constant annual increase in the number of automobiles in use, for each year the industry added to the total inventory or ''population" more new units than were being scrapped, as may be seen in Fig. 1. Few of us in the oil industry needed to give much attention to the detailed figures, although the automobile industry seemed to interest itself in all sorts of calculations of replacement demand vs. new car demand, the average life of the car, the number scrapped, and the like. Probably many people in the oil industry, as well as outside it, were surprised at the way in which automobile registrations and gasoline consumption held up during the early years of the depression. Nor were they greatly perturbed when, as in the last two years, both registrations and consumption began to fall somewhat. That was natural enough; it

Jan 1, 1933

By Arthur H. Wilson

INTRODUCTION The stratified deposits of the coal measures are strongly laminated and exhibit a high degree of anisotropy. The stronger rocks in the sequence possess joints and shear planes, the coal itself has a closely spaced cleat pattern, and the softer rocks are capable of flowage. To calculate the general distribution of stress in such a medium by analytical means based on the theories of elasticity is difficult, and has not yet been satisfactorily achieved. Nevertheless, some estimate is required to aid the design of supporting pillars and permit the correct siting of main roadways. This is important when laying out districts for longwall extraction, particularly if the results of previous experience are not available in the area to be exploited. Although elastic stress analyses, including finite element analysis, may not be very appropriate, the basic equilibrium conditions will apply. These, combined with logical deductions based on observed fact, can be used to give appropriate solutions which experience has shown to be capable of giving acceptable results in the design of pillars associated with longwall workings. Constancy of cover load is one of the fundamental equilibrium conditions. When mining below ground, the cover is not removed, therefore the average vertical load after mining must equal the initial cover load. If mining produces a local decrease in load in any particular area, it must be balanced by an equivalent increase in load nearby. Bending moments in the strata above are capable of compensating for the turning moments which this redistribution of load may produce, and the existence of a "rear abutment", much discussed between 1950 and 1960, is not an essential equilibrium requirement. If the distribution of vertical stress is plotted three-dimensionally, then the load on any small element of area 6S will be o.6S, where o is the average stress. This also represents the volume of the "stress envelope" which lies above 6s. By making the individual areas very small and summing together, it is possible to show that the total load on any particular area can be found by calculating the volume of the stress distribution envelope which lies above it. A similar consideration applies to the stress distribution in a two-dimensional situation. If the stress distribution curve is drawn, then the area below any portion will represent the load per unit depth of section. For example, Figure 1 represents the possible stress distribution across a caved waste and associated ribside. If AW is the load deficiency [ ] These loads must balance, therefore area AW = AS, and if one can be deduced, the other will then be known. Attempts have been made to measure stress distributions in the ribside and in the waste. Stress meters in soft rocks such as the coal measures have not proved very effective. Those based on the measurement of strain have experienced difficulty in strata capable of flowage, high modulus stress meters suffer from cross sensitivity, and overcoming to obtain absolute stress is almost impossible. Load measuring instruments left in the waste have also proved largely ineffective, mainly as a result of the difficulties of protecting the leads in a caving environment. Certain logical deductions, however, can be made about the conditions in the waste. CONDITIONS IN THE WASTE Over the area of the waste the surface trough which develops is remarkably consistent, despite the geological variations which occur from coalfield to coalfield. In Britain, an accuracy of 210% is claimed for subsidence prediction based on width of extraction and depth of cover (National Coal Board, 1975). The subsidence in the USA is more variable (Adamek and Jeran, 1981), particularly with respect to the point of inflexion, but the same general pattern occurs. At the shallower depths the somewhat harder roof strata involved in the caving in the USA may suffer less compaction, giving reduced values of subsidence. Subsidence at intermediate horizons below the surface

Jan 1, 1982

By J. P. Nielsen

When two grains in a polycrystalline specimen meet at a point in the course of grain-boundary movements, and the new boundary created at the point is one of relatively low specific free energy, a nonequilibrium boundary condition occurs. The nonequilibrium is enhanced if the other grain boundaries involved at the point of meeting are relatively high in specific free energy. The nonequilibrium results in a correction by growth of the complex grain (two subgrains) to a large size, sufficient in many cases for it to become a self-propagating unit, i.e., a recrystallization grain. This mechanism, herein called "geometrical coalescence," is proposed as a logical origin for recrystallization nuclei, particularly for secondary recrystallization. RECENT publications1-' indicate that considerable progress has been made toward a complete understanding of the three microstructural processes in metals: recovery, recrystallization, and grain growth. However, a major problem persists, namely, the origin of recrystallization nuclei. It is generally accepted that once a recrystallization nucleus, a secondary recrystallization nucleus particularly, has reached a certain size, it will continue to grow at the expense of its neighbors, simply as a consequence of grain-boundary free-energy considerations. What has remained obscure, however, is the mechanism whereby such a nucleus comes into being and grows to the self-propagating size. Recrystallization Nucleation in a Two-Dimensional Grain-Boundary System In the left portion of Fig. 1 is shown schematically the meeting of grains 1 and 4 on disappearance of boundary 2-3, i.e., the boundary that existed between grains 2 and 3, as might happen in the course of normal grain-boundary migration. If, by chance —and the probability is not necessarily small— grains 1 and 4 are close together in their mutual orientations, or in some other special way mutually oriented so that the new common boundary has relatively a very low specific free energy (hereinafter referred to by s), then the grain boundary arrangement as depicted in the figure is highly unstable. This is particularly so if boundaries 1-3, 3-4, 4-2, and 2-1 are high in their s values. Assuming, for the sake of argument, that boundary 1-4 has a s equal to zero, and the s's of all the other boundaries present are equal to each other and constant, then the boundary arrangement on the right in Fig. 1 will be

Jan 1, 1955