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By C. P. Fiske

The pottery of the New Jersey Zinc Co. (of Pa.) is equipped to make three classes of refractories; namely, spelter vessels, spelter condensers, and high-grade fire-brick. The most important of these are the spelter vessels, on account of the extreme severity of the service which they must perform, necessitating great care in the selection of materials and in the whole process of manufacture. Next to the spelter vessels in importance come the condensers, as large numbers are used in the manufacture of spelter, and improvements in quality result in material decreases in working cost. While the fire-brick which form the laboratory, or high-temperature portion of the spelter furnaces, are placed at the foot of the list, they are nevertheless of much importance, since high quality will decrease the frequency of the always expensive shutdowns for repairs. To promote simplicity, the manufacture of these three products will be described separately in this paper, except in so far as materials and the crushing and handling system common to all are conccrned. Materials Bond Clay.—The bond clay used entirely at the Palmerton pottery is a St. Louis clay furnished by the Grand View Fire Clay Co. It is stored in a system of covered bins, and no attempt is made to weather it in the accepted sense of the term " weathering." Some years ago our clay was weathered before pugging, but we did not find the benefit sufficient to offset the additional handling and the loss of clay incurred. Furthermore, we believe that the rotting of the pugged mix (described later) accomplishes the same results to better advantage. That there is some action taking place in the clay while stored in the covered bins, however, is evidenced by an increase in volume which at times has caused damage to the retaining walls of the bins. We have at present a stock of bond clay weathering in the open, and when this is put into service an opportunity will be afforded for making comparisons with the clay as usually handled.

Jan 1, 1918

The following is an incomplete list of members and guests who called at Institute headquarters during the period July 10, 1919, to Aug. 10, 1919. Howard C. Arnold, Washington, Pa. J. S. Lane, New York City, N. Y. L. M. Banks, Denver, Colo. H. W. Lawrie, Washington, D. C. George S. Blair, Pittsburgh, Pa. J. L. McAllen, Latouche, Alaska. Richard H. Catlett, Newark, N. J. David F. McFarland, Urbana, Ill. George L. Clark, Council Bluffs, Iowa. James F. Magee, Lancaster, Pa. Capt. Frederick S. Cook, Portland, Ore. A. G. McGregor, Warren, Ariz. John J. Croston, Atlanta, Ga. Emory M. Marshall, Santa Fe, N. M. Thomas Davis, Seattle, Wash. Benjamin L. Miller, Bethlehem, Pa. J. C. Dick, Salt Lake Cite, Utah. 2d Lieut. C. 0. Noble, U. S. Army. Chris. G. Dobson, Vancouver, B. C. Irving Perrine, Hutchinson, Kans. H. S. Emlaw, Grand Rapids, Mich. Fred S. Porter, Seattle, Washington. F. A. Fahrenwald, Cleveland, Ohio. Capt. H. R. Robbins, Peking, China Lieut. James L. Head, St. Louis, Mo. Joseph P. Ruth, Jr., Denver, Colo. A. A. Holland, Napanu, Canada. William A. Simkins, San Francisco, Calif. W. E. Hopper, Washington, D. C. T .V. Smirnoff, Siberia. Harry Hotchkin, Chicago, Ill. Harrison Souder, Cornwall, Pa. L. L. Hutchison, Tulsa., Okla. F. J. Sur, Denver, Colo. Paul S. King, Wilmington, Del. Harry Wolf, Denver, Colo. George K. Anderson, Jr., lies accepted a position as manager for Wight & Co., producers of air-cooled blast-furnace slag, at Longdale, Va. J. M. Armstrong, with the Caribbean Petroleum Co., has been transferred from Philadelphia to Maracaibo, Venezuela. Fred Sherman Baker, who has been with the Oriental Consolidated Mining Co., at Unsankinko, Chosen, Korea, is now at Columbia City, Ind. Ray J. Barber has removed from Roxbury, Mass., to New York City, and is now consulting engineer for the Roxbury Cold Mines, Inc.

Jan 9, 1919

Development: Iron Ore In Peru Utah Construction Co., of San Francisco and a Peruvian Government corporation have signed an agreement for development of a large iron orebody in Peru. Most tests show 60 pct iron or better. Most of the ore is magnetite. Although ore will be shipped mainly to Eastern U. S. blast furnaces, shipments to the West Coast may develop as the orebody opens up. The construction firm indicated three American steel companies have shown interest in the project. The orebody is located 300 miles south of Lima and about 15 miles from the coast at the same longitude as New York City. Allen D. Christensen, executive vice-president of Utah Construction noted that the distance is shorter between Peru and the East Coast. The ore deposit measures about four to five miles in one direction and 10 miles in the other. As many as 15 outcroppings within these limits have been reported. Several of the outcroppings are thousands of feet long by several hundred feet wide, Christensen said. Shipments between 100,000 and 200,000 tons of ore per month are projected by the company. Cost of opening the mine to production may go over $5 million, Christensen said. It would include building of docks for ocean going vessels, housing, and heavy duty mining and processing equipment. The contract with the Peruvian government gives Utah Construction six to eight months for additional tests and a 20 year concession for mining the orebody. Peru will collect royalties on all ore. The company will also supply about 300,000 tons per year for an electric furnace to be installed by the Peruvian Government at Chimbote. No steel is produced in Peru now. Utah Construction formed the Argonaut Mining Co. last year, and was shipping iron ore from British Columbia to Japan at a rate of more than 50,000 tons a month, before winter weather decreased operations.

Jan 1, 1952

By C. P. Fiske

THE pottery of the New Jersey Zinc Co. (of Pa.) is equipped to make three classes of refractories; namely, spelter vessels, spelter condensers; and high-grade fire-brick. The most important of these are the spelter vessels, on account of the extreme severity of the service which they must perform, necessitating great care in the selection of materials and in the whole process of manufacture. Next to the spelter vessels in importance? come the condensers, as large numbers are used in the manufacture of spelter, and improvements m quality result in material decreases in working cost. , While the fire-brick which form the laboratory, or high-temperature portion of the spelter furnaces, are placed at the foot of the list, they-are nevertheless of much importance, since high quality will decrease the frequency of the always expensive shutdowns for repairs. To promote simplicity, the manufacture of these three products will be described separately in this paper, except in so far as materials and the crushing and handling system common to all are concerned. Materials Bond Clay.-The bond clay used entirely at the Palmerton pottery is a St. Louis clay furnished by the Grand View Fire Clay Co. It is stored in a system of covered bins, and no attempt is made to weather -it in the, accepted sense of the term " weathering." Some years ago our clay was weathered before pugging, but we did not find the benefit sufficient to offset the additional handling and the loss. of clay incurred. Furthermore, we believe that the rotting of the pugged mix (described later), accomplishes the same results to better advantage. That there is some action taking place in the clay while stored in the covered bins, however, is, evidenced by an increase in volume which at times has caused damage to the retaining walls of the bins. We have at present a stock of bond clay weathering in the open, and when this is put into service an opportunity will be afforded for making comparisons-with the clay as usually handled.

Jan 10, 1917

By Ralph Nichols

THE mines are near the town of Gilmore, in the Texas mining district. This district was organized in 1880. The present producing mines are near the terminus of the Gilmore & Pittsburg railroad. This railroad connects with the main line of the Oregon Short Line at Armstead, Mont. The distance from Gilmore to Armstead is 70 miles. The town of Gilmore is near the head of the Lemhi river at an elevation of 7,000 ft., while the mines are from 400 to 1,300 ft. higher. The range of mountains upon which the mines are situated forms the water shed for the Little Lost river on the west and the, Lemhi river on the east. The winters are severe and the snow fall is heavy, but the mines ship right through the year. There is an abundance of timber suitable for all mining purposes in the immediate vicinity. Water is piped into the camp from a lake about 2.5 miles distant. This lake is about 1,400 ft. above the town of Gilmore. The water is exceptionally good for boiler and domestic purposes. The mining companies now producing are. the Pittsburg Idaho Co., Ltd. ; the Latest Out Mining & Smelting Co:, Ltd. ; and the Gilmore Mining Co., Ltd. Besides the above properties, there are a number .of prospects in different stages of development. The producing mines of the district are fissures in limestone, with a general northeast-southwest course .and a dip to the west of from 40° to 70°. These fissures are at about right angles to the bedding of the limestones. The veins are regular and the ore shoots strong, both laterally and in depth, so far as developed. The ore shoots have a rake or pitch to the south. The character of the formation is shown in Fig. 1, a section of the Latest Out mine.

Jan 11, 1913

By Darwin Soler

This chapter will outline the operating procedures at Freeport Kaolin Co.5 kaolin mining operations in Georgia, with emphasis on explaining new techniques recently introduced to reclaim mined-out areas for reforestation. Kaolin is found in the central Georgia area (See Fig. 1) in lenticular, sedimentary deposits of high purity. The beds are Cretaceous in age and vary from 4 to 60 ft in thickness, under a cover of sedimentary overburden ranging from 5 to 150 ft. The kaolin beds contain the mineral kaolinite (a hydrous aluminum silicate) and have a particle-size distribution ranging from about 20µ to the submicron level and averaging about 60% below 2µ. The only contaminants in the deposits are coarse mica, quartz particles, and minor amounts of iron and titanium oxides. These kaolin deposits vary widely in quality from deposit to deposit and also vary within individual deposits. Some of the controlling quality specifications which are determined during exploration are viscosity in water suspension, optical brightness index as measured by reflectance, amenability to bleaching, and particle-size distribution. For this reason, only a small portion of the total kaolin reserves in the Georgia belt are suitable for wet processing for the present markets being served. Kaolin clays are selectively mined by different mechanical methods and are beneficiated by both wet and dry processing to produce a number of products used as fillers and coatings in the paper industry; fillers or additives in the paint, plastic, and rubber industries; and as basic raw materials for the ceramic industry. The State of Georgia is the largest kaolin producer in the United States, with 3,206,454 tons produced in 1966, according to U.S. Bureau of Mines statistics. Freeport Kaolin Co’s processing plant is located in Gordon, Ga., about 20 miles east of Macon, Ga.

Jan 1, 1969

By R. F. Evans, J. M. Avery

MUCH has been written of the glamour of magnesium from sea water, the Aladdin-like creation of a huge magnesium plant in the Nevada desert using cheap hydroelectric power from Boulder Dam; the marvels of the Hansgirg process for the electrothermal reduction of magnesium oxide by carbon, and the ferrosilicon process for producing magnesium from dolomite. This is the story of the production of -magnesium by the electrolytic process, using steam-electric power supplied by a public utility, in a plant at Painesville, Ohio, 30 miles east of Cleveland on the shore 'of Lake Erie. The story is not without interest, since it involves several unique developments that provide the answer to a question frequently asked: How is it possible to produce magnesium metal economically at such a location as Painesville, Ohio? Basically, the answer lies in making use of facilities already in existence for the production of soda ash, which makes it possible to manufacture magnesium chloride, the principal raw material, cheaply enough to offset the high cost of electric power. To this is coupled the successful conversion of by-product chlorine into a product required for the war program. The source of magnesium for this operation is the enormous surface deposit of high-grade dolomite in the region south of Toledo, within reasonable freight haul of Painesville. This same geological stratum underlies the Painesville plant, but unfortunately at a depth of about 3000 ft. apparently too great a depth to permit underground mining to compete with the delivered cost of rock from surface quarries in other areas. The cost of dolomite obviously is an important factor, since the rock contains only some 12 per cent magnesium (20 per cent MgO) so that one dollar on the cost of stone means more than $8 in terms of its magnesium content, and more than that in terms of recoverable magnesium metal. PRODUCTION OF MAGNESIA FROM DOLOMITE In the summer of 1941, the Diamond Alkali Co. was well along with construction of a plant intended for the commercial production of refractory and other grades of magnesia from dolomite. The process to be used was designed to make maximum economic advantage of existing plant and operations by tying the production of magnesia in with the manufacture of soda ash in such a way as to make use of the lime (CaO) value of calcined dolomite. The process finally adopted after a lengthy research and development program has much in common with the old Clerc-Nihoul process, but includes certain innovations found to be necessary or

Jan 1, 1945

By R. F. Evans, J. M. Avert

Much has been written of the glamour of magnesium from sea water, the Aladdin-like creation of a huge magnesium plant in the Nevada desert using cheap hydroelectric power from Boulder Dam; the marvels of the Hansgirg process for the electrothermal reduction of magnesium oxide by carbon, and the ferrosilicon process for producing magnesium from dolomite. This is the story of the production of magnesium by the electrolytic process, using steam-electric power supplied by a public utility, in a plant at Painesville, Ohio, 30 miles east of Cleveland on the shore of Lake Erie. The story is not without interest, since it involves several unique developments that provide the answer to a question frequently asked: How is it possible to produce magnesium metal economically at such a location as Painesville, Ohio? Basically, the answer lies in making use of facilities already in existence for the production of soda ash, which makes it possible to manufacture magnesium chloride, the principal raw material, cheaply enough to offset the high cost of electric power. To this is coupled the successful conversion of by-product chlorine into a product required for the war program. The source of magnesium for this opera- tion is the enormous surface deposit of high-grade dolomite in the region south of Toledo, within reasonable freight haul of Painesville. This same geological stratum underlies the Painesville plant, but unfortunately at a depth of about 3000 ft.— apparently too great a depth to permit underground mining to compete with the delivered cost of rock from surface quarries in other areas. The cost of dolomite obviously is an important factor, since the rock contains only some 12 per cent magnesium (20 per cent MgO) so that one dollar on the cost of stone means more than $8 in terms of its magnesium content, and more than that in terms of recoverable magnesium metal. Production of Magnesia from Dolomite In the summer of 1941, the Diamond Alkali Co. was well along with construction of a plant intended for the commercial production of refractory and other grades of magnesia from dolomite. The process to be used was designed to make maximum economic advantage of existing plant and operations by tying the production of magnesia in with the manufacture of soda ash in such a way as to make use of the lime [CaO)., value of calcined dolomite. The process finally adopted after a lengthy research and development program has much in common with the old Clerc-Nihoul process, but includes certain innovations found to be necessary or advantageous. Essentially, the process is based upon the following series of reactions:

Jan 1, 1949

By Philip N. Adler, Harold Margolin

Univariant equilibria between the Ibcc, IIhcp, and UIfee phases of thallium have been determined and indicate a triple point at 110°C and 37 kbars. Calculated II —I transformation curves are compared with the experimental. Transformation volumes and/or enthalpies are calculated and enthalpies are compared with available data. ThE triple point of solid thallium has been reported by Bridgman as 153°C and 38 kbar.' More recently, Jayaraman et at.2 have reported the triple point as near 115°C and 39 kbars. Meyerhoff and Smith,3 assuming the elevated-pressure modification of thallium is fcc, as Kaufman suggested,4 and that it is continuous with the room-temperature fcc phase in the Tl-In system, have used pressure data up to 5 kbars to predict the triple point at 162°C and 35.8 kbars. Piermarini and weir5 have indeed found the elevated-pressure phase to be fcc. The effect of variations of pressure and temperature in the IIhCp — Ibcc transformation in thallium have also been reported by werner6 and more recently by Ponyatovskii.7 The work reported here is part of an investigation of the effects of temperature and pressure on the Tl-In system. I) EXPERIMENTAL PROCEDURE The onset and progress of transformation were determined by changes in resistance. A commercially available tetrahedral-type pressure apparatus was used. Specimens of 0.060 in. diam, representing 0.63 pct of the cell volume, were encased in AgCl and mounted in pyrophyllite, as shown schematically in Fig. 1. It can be seen that the specimen is in series with the graphite furnace. Resistance changes of 0.1 pct could be detected with a Kelvin bridge. Resistance of the sample was of the order of 1 milliohm. Other pertinent details in regard to calibration and cell-assembly design are given elsewhere.' Temperature calibration of the cell was carried out in two ways as a function of input wattage. Through a hole drilled in the graphite furnace, a chromel-alumel thermocouple was inserted into the AgCl to a point 0.010 in. from the sample, with the

Jan 1, 1964

By Ralph Hultgren

FEW phase diagrams of alloys composed of two transition metals have been adequately studied, probably because of the high melting points involved. Transition metals are the elements that have inner shells incompletely filled with electrons. They comprise groups IIIa to Villa of the periodic table, including more than half of all metals. We have chosen to study two systems whose comparatively low melting points allowed the alloys to be melted in a molybdenum resistance furnace. The nickel-palladium system proved to consist of a simple series of solid solutions, but iron-palladium manifested several remarkable features. The main characteristics of the latter diagram were decisively established, but lack of time prevented precise determination of phase boundaries. The gamma to alpha transformation in this system is unusual and warrants further study. Grigorjew1 determined the liquidus and solidus of the iron-palladium system by thermal methods. He found a minimum melting point, but no eutectic. He also found that palladium lowered the gamma to alpha and raised the gamma to delta transformation temperatures of iron, with a complete series of solid solutions existing between gamma iron and palladium. His measurements also indicated a compound or super-lattice near the composition FePd3. Jellinghaus2 found that annealing at 500° C. greatly increased the magnetic coercive force of an alloy of composition FePd. X-ray dif-fraction photographs showed that a tetragonal ordered phase had formed during the anneal.

Jan 1, 1939

By H. J. LONGMORE

MORE improvements have probably been made in the diamond-drill field in the past decade than were accomplished in the entire prior period since diamond drilling was discovered in 1864 by a French engineer named Leschot. True, machines were improved, and new accessories were adopted from time to time up to 1929, but no radical changes were made. The basic principle instituted during the early years of the industry were still in use in 1929. At that time, practically the only diamonds used were "black diamonds" or "carbon." These came from Brazil through a rather round- about route, and by the time they got to the drilling trade their cost was excessive-$175 per carat and even higher at the peak. Perhaps because of this. or perhaps because of increased competition, the industry sought ways to reduce expenses even in the face of constantly increased labor costs due to higher wages and reduction in hours worked. The solution seemed to rest in two things-use of a cheaper cutting medium to replace the carbon diamond and the development of new machines to drill faster and thus produce more hole per shift of operation.

Jan 1, 1940

By E. F. La Vigne

ON Aug. 10, 1884, the Retsof Mining Co. began the sinking of an 18 by 12-ft. shaft at Retsof, N. Y. Rock salt was reached in 1885, in September, at 1008 ft. below the surface. The first salt was shipped in the fall of 1885. A timber breaker was built in 1886. In July, 1895, there was a consolidation of the Retsof Mining Co., the Greigsville Salt & Mining Co., the Livonia Salt Co., and the Lehigh Salt & Mining Co. These consolidated salt companies are now known as the Retsof Mining Co., a subsidiary of the International Salt Co. Eventually all of the mines were closed except the mine of the Retsof Mining Co., at Retsof, Livingston County, New York. On July 6, 1921, about ½ mile south of the original shaft and plant of the Retsof Mining Co., a new 9 by 28-ft. concrete-lined shaft was started. The shaft is 1063 ft. deep from the top of the shaft collar to the bottom of the salt bed. At the same time a reinforced-concrete breaker and other buildings were erected. This new plant began opera-tions on Dec. 15, 1923. It was designed for a capacity of 3000 tons in 8 hours.

Jan 1, 1936

By A. B. Wilder

We study the past Because it is a guide to the present and a promise for the future. The struggle for a better world is strengthened By the hopes, ambitions, and deeds Of those who were before us. As we look backward Our attention is directed forward. ONE hundred years ago, the manufacture of steel from molten pig iron without the use of fuel first began. In the earlier part of the nineteenth century the crucible process was the principal steelmaking method. It was a costly process, producing only a few pounds of steel at a time; so, most of the ferrous metal output was in the form of wrought or pig iron. With the perfection of the converter technique it became possible to produce low cost steel by the ton, thereby completely changing the structure, not only of the ferrous metal industry, but also of industrial production in general. This was the turning point from the Iron Age to the Age of Steel. William Kelly (Fig. 2), Sir Henry Bessemer (Fig. 1), and Joseph G. Martien experimented with the process before 1856. William Kelly began in Eddyville, Ky., in 1847 and has been given credit for first discovering the principle of the pneumatic converter process. Several years later in England Henry Bessemer independently conceived a steelmaking process similar to Kelly's. Bessemer made public his invention at the annual meeting of the British Association for the Advancement of Science held at Cheltenham, England, August 1856. As a result of his paper, "On the Manufacture of Malleable Iron and Steel Without Fuel," and his progressive leadership in the years to follow, the process became permanently identified with the name, Bessemer. Credit for the commercial adaptation of the process must be given two additional persons. Robert Mushet (Fig. 3) in 1856 in England recognized the necessity of deoxidation and recarburization of the converter product and evolved the technique for adding high manganese iron following the blow. In Sweden, in July 1858, G_ F. Göranson [ ]

Jan 1, 1961

By Paul A. Bailly

2.1-1. The Exploration Function in the Mineral Industry. With regard to a new mine, exploration can be technologically defined as all the activities and evaluations necessary before an intelligent decision can be made establishing size, initial flowsheet, and annual output of the new extractive operation based on a certain mineral deposit. The activities that follow such a decision are called first development, then operation (90). Thus, in the U.S.A., exploration may continue in the technological sense after a mineral property has entered the "development stage" as defined for tax purposes. Regulations under Sections 1.615.1 and 1.616.1 of the U.S. Internal Revenue Code state that the "development stage" has been reached when "deposits of ore or other minerals are shown to exist in sufficient quantity and quality to reasonably justify commercial exploitation by the taxpayer" (95). There are, of course, many instances where additional later exploration discloses additional reserves which justify a plant much larger, more efficient and more profitable than the minimum feasible commercial operation. After a surface mine is put into operation, additional search generally will increase the reserves and thus warrant expansion of the rate of output and/or extend the life of the operation. The methods and techniques in such search are similar to those used in predevelopment exploration. This activity has produced a major part of new reserves for all mineral commodities but is not considered as exploration in the context of this chapter. The objective of exploration is, briefly stated, to find and acquire a maximum number of new economic mineral deposits with a minimum cost and in minimum time. The search must be so organized as to maximize the overall probability of discovering commercial mineral concentrations within a minimum time with the funds available. As shown by Celasun (27) these statements are neither adequate nor complete from a pragmatic point of view. Each actual exploration venture must define its objectives in terms of a minimum rate of return on investment for any deposit found

Jan 1, 1968

By C. M. Ball

The very considerable progress made during the past three years in the crushing and concentration of ores, lends special interest at the present time to any means of a more efficient character than such as are already generally known and in use, for the rapid and regular reduction of large quantities of the harder and heavier grades of mineral-bearing rock. Such means are of exceptional importance in connection with the preparation of iron-ore for separation by the magnetic process, since in this operation the capability of dealing with large volumes of material and of securing by ready means of adjustment, a granulation of the material under treatment to any required mesh or fineness, is of the greatest possible value. With these, however, must be coupled the further considerations of the amount of power expended for the work done, and the wear and tear, represented in the cost for renewal of parts of the crushing machinery, per ton of material reduced to a given degree of fineness. In general terms it may be stated that the operation of a crushing train, capable of reducing an ore-rock of average density and hardness, from the condition of mine lumps to pass 1/16-inch screenopenings at the rate of 20 tons per hour, should not call for an expenditure of more than 125 horse-power, and that the wear and tear of the crushing machinery, all told, should not exceed 2 cents per ton. In the design of a crushing-train for an ore-dressing plant, there are, moreover, considerations of great importance which relate to the simplicity or complexity of construction and arrangement, the wear and tear of few or many elements, the easy ability or difficulty of maintaining the working adjustment of the various elements of the train, and the number of hands required for the operation of the system. With these points in view, it is thought that the Cummings ore-

Jan 1, 1893

By R. F. Mehl, C. E. Birchenall

R. E. Hoffman and D. Turnbull—The authors have presented evidence which they have interpreted as indicating that the rate of self diffusion is not intrinsically more rapid at grain boundaries than within the grain. Grain-size effects which apparently exist are attributed rather to impurities concentrated at the grain boundaries. In view of our own experiments and the existing evidence, we believe that the support for this hypothesis is not convincing. We have in progress an investigation in which the rate of self diffusion of silver is being measured over an extended temperature range in both single-crystal and polycrystalline specimens. The results of the single-crystal experiments and some preliminary data on fine-grained polycrystalline specimens have already been reported:' and it is anticipated that a complete report will be published in the near future. The self-diffusion coefficient of large-grained polycrystalline silver (1 grain per sq mm) has previously been measured by Johnson" between 730" and 940°C. The diffusion coefficients which we have measured in single crystals (210 plane normal to diffusion direction) agree within experimental error with values calculated from an extrapolation of Johnson's curve down to temperatures as low as 500 °C. However, it has been demonstrated that the overall self-diffusion rate in fine-grained polycrystalline specimens (initial grain size of 0.003 cm) becomes measurably larger than the overall rate in a single crystal at a temperature of 600°C, and the discrepancy between the two rates becomes greater as the temperature is further decreased. In fact, it has been possible to obtain satisfactory penetration curves for polycrystalline specimens using the sectioning technique at temperatures as low as 400°C. At this temperature, the penetration is 50 to 100 times greater in the polycrystalline specimens than in a single crystal. Fisher" has developed an analysis whereby the ratio of the rate of the unit diffusion process at the grain boundary to the corresponding rate within the grain can be calculated from the penetration curves and an assumption as to the width of a grain boundary. This analysis applied to our data indicates that the unit process at the grain boundary is faster by a factor of 10' at 475°C when the grain boundary width is taken to be 5. The silver used in most of these experiments was obtained from the Handy and Harmon Co. and listed as 99.97 pct pure. Preliminary experiments on 99.999 pct silver from the Jarrel-Asch Co. indicate a grain-size effect of the same order of magnitude as in the less pure silver. Nominally, these purities are as good, at least, as that of the carbonyl iron used by the authors, but of course if an impurity effect does exist its magnitude might be very dependent upon the nature of both major and minor constituents. The authors have cited the work of other investigators who have found no grain-size effects. Neither Steigman, Shockley and Nix" nor Maier and Nelson8 were able to correlate self-diffusion coefficients of copper with grain size. However, all their measurements were performed at or above 750°C; and on the basis of our work with silver, no grain-size effect would be expected at temperatures above about 0.7 of the absolute melting temperature unless the grain size were exceedingly small. Likewise, in the investigation of the self diffusion of lead by Seith and Keil,14 the lowest temperature at which the diffusion coefficient was measured in polycrystalline specimens was 207°C, which is still sufficiently high so that the lack of a grain-size effect is not surprising. Finally, in those experiments on iron from which they concluded that there was no grain size effect, Drs. Birchenall and Mehl seem to have no information as to the actual grain sizes immediately prior to and following the diffusion anneal. Without this information, we believe that their own experiments offer little support for their hypothesis. F. S. Buffington, I. D. Bakalar, and M. Cohen—The results given in this paper agree in order of magnitude with those tentatively reported by us.27 However, significant differences exist in the two sets of data, and it may be well to make an explicit comparison. The diffusion studies at M.I.T. were conducted on somewhat higher purity iron (99.98 pct Fe) than the grades used by the authors, but this is undoubtedly not the answer. Fig. 4 shows the diffusion results of both laboratories for the gamma phase, omitting the authors' data on the commercial steels, while fig. 5 presents a similar comparison for the alpha phase. The divergence is much more marked in the latter case than in the former. In connection with the M.I.T. determinations, all of the runs in the gamma range and those above 800 °C in the alpha range were conducted with specimens having a relatively thick (0.002 cm) electrodeposit of radioactive iron. This practice minimizes any possible error due to extraneous diffusion that may occur during the heating to and cooling from the operating temperature. An exact solution of Fick's law for these boundary conditions was used in calculating the diffusion coefficients. At a later time, three runs were made below 800°C, using very thin electrodeposits similar to those of the authors, and the points fell considerably below the values expected from the extrapolation of the results based on the specimens with the thick deposits (compare dash-dot line in fig. 5). However, in the runs with the thin deposits, deviations of 100 pct were found between the individual specimens, whereas the maximum deviation with the thick deposits was less than 25 pct. Accordingly, it is not known at the moment whether the M.I.T. points below 800 °C should be given as much weight as those above 800°C. If this were done, the frequency factor would be of the order of 400 cm2 per sec, which is quite high. In other metals, the frequency factor for self diffusion lies between about 0.1 and 10 cm2 per sec. As the points below

Jan 1, 1951

By M. E. Wadsworth, R. C. Peterson, W. M. Fassell

The temperature and pressure dependence of the reaction of tantalum in oxygen were investigated from 500° to 1000°C at pressures from 10 mm Hg to 600 psi total oxygen pressure. Tantalum was found to oxidize linearly under the above conditions. Three distinct regions of temperature dependence were found with different energies of activation. From 500° to 600°C the rate of oxidation of tantalum was found to be essentially independent of the oxygen pressure at the pressure investigated. The oxidation rate increases rapidly with an increase in pressure from 600' to 800°C. The dependence of the oxidation rate on the bulk concentration may be expressed by V = k'?, where k' is the specific rate constant and ?=k1Co2/(l + K1Co2), where k, is the equilibrium constant for the adsorption of oxygen on tantalum. ONLY a limited amount of data has been presented in the literature on the oxidation behavior of tantalum metal. Some values of its oxidation rate were reported in the Corroston Handbook' and by McAdams and Geil.' The first extensive investigation of tantalum was the work of Gulbransen and Andrews." They determined the reaction rates of tantalum in O2 and N2. Their results indicate that this metal follows the parabolic law in the temperature range of 250" to 450°C in O2 at pressures of 7.6 cm Hg with an activation energy of 27,400 cal per mol. In nitrogen, the metal follows the parabolic law from 500" to 800°C at 7.6 cm pressure. All experimental determinations were of 2 hr or less duration and deviations from the parabolic rate conceivably could occur at longer times. A spectrophotometric study of the oxidation of tantalum by Waber, Sturdy, Wise, and Tipton' indicated good agreement between this method and the microbalance technique. The experiments were conducted in air from 220" to 350°C. The metal was found to follow the logarithmic law up to 320°C and the parabolic law at higher temperatures. The activation energies in the logarithmic region were reported as 12.57 kcal per mol and 12.0 kcal per mol, respectively, for the microgravimetric and spectrophotometric methods. In the parabolic region the activation energy was found to be 27.2 kcal per mol, in good agreement with the value of Gulbransen and Andrews." The present paper is concerned with the behavior of tantalum from 500" to 1000°C in oxygen at pressures ranging from 1/2 to 40.8 atm pressure. Earlier preliminary data by McKewan5,6 indicated the following facts concerning the behavior of tantalum: 1—At temperatures from 500" to 600°C, the metal followed the linear law. 2—At 575°C the rate of oxidation was found to increase with O2 pressure. 3—The Arrhenius plot in this temperature region was nonlinear. The nonlinearity of the Arrhenius plot was indicative of a more complex behavior than originally suspected for this metal.

Jan 1, 1955

By W. L. Austin

AT the Steptoe metallurgical plant, where ore of the Nevada Consolidated Copper Co. is beneficiated, a small copper-leaching annex has been in operation treating flue dust from roasting-furnace dust chambers, together with a siliceous oxidized ore. This leaching plant was designed to obviate smelting low-grade siliceous flue dust, which interfered with the operation of the reverberatories by forming a blanket on the charge The major portion of the copper content of the flue dust accumulating in the upper chambers of the. roaster flue, and at the stack, is water soluble, and also contains much free sulphuric acid and soluble sulphate It was found that sufficient acid could be obtained by mixing this material with water to provide a lixiviant for leaching the siliceous copper o mentioned. This was the sole source of the lixiviant used in the operatic to be described. An analysis of the flue dust showed it to be composed as follows: Per Cent. Copper 3.00 SiO2 23.41 Fe : 4.08 CaO + MgO : 2:72 Al203 6.07 Free H2SO4 19.94 S03 combined as sulphate 25.10 H2O combined With the sulphates 10.67 0 2.50 Total 97.49 Copper soluble in dilute H2SO4 2,56

Jan 7, 1914

By W. L. Austin

At the Steptoe metallurgical plant, where ore of the Nevada Consolidated Copper Co. is beneficiated, a small copper-leaching annex has been in operation treating flue dust from roasting-furnace dust chambers, together with a siliceous oxidized ore. This leaching plant was designed to obviate smelting low-grade siliceous flue dust, which interfered with the operation of the reverberatories by forming a blanket on the charge. The major portion of the copper content of the flue dust accumulating in the upper chambers of the roaster flue, and at the stack, is water soluble, and also contains much free sulphuric acid and soluble sulphates. It was found that sufficient acid could be obtained by mixing this material with water to provide a lixiviant for leaching the siliceous copper ore mentioned. This was the sole source of the lixiviant used in the operation to be described. An analysis of the flue dust showed it to be composed as follows : Per Cent. Copper.............................................. 3.00 siO2................................................. 23.41 Fe................................................... 4.08 CaO + MgO......................................... 2.72 Al203................................................ 0.07 Free H2S04........................................... 19.94 SO3 combined as sulphate.............................. 25.10 H2O combined with the sulphates....................... 10.67 0................................................... 2.50 Total................................... 97.49 Copper soluble in dilute H2SO4.......................... 2.56 The oxidized ore, locally known as carbonate, has the following composition:

Jan 1, 1915

By H. E. Montgomery

Engineering trends for some years have been toward higher temperatures, both in process work and in steam generation. Progress in design has been dependent upon the development and evaluation of suitable alloys. Elementary knowledge of the limitations of copper-base alloys indicate their unsuitability for the higher ranges of operating temperatures; say, for steam to 950°F. Emphasis upon the more severe conditions perhaps is a reason why more complete evaluations have not been made for copper-base alloys over their range of suitability. However, the few studies made show clearly the need for more accurate evaluations. Some copper-base alloys long looked upon favorably because of superior room-temperature properties are not as desirable as some others perhaps not so "strong" and "hard," as measured by usual acceptance tests. Load-carrying capacity and structural stability arc two important factors. Structural stability as influenced by temperature, by presence of corrosive media, or by the effects of stress itself, frequently is a limiting factor. An early reference to creep properties of the cast copper-base alloys is that of Brown1 (Admiralty gun metal, Cu 87.5, Sn 10.1, Zn 2.3 and Pb 0.08). His tests were similar to present-day stress-rupture tests without measurement of creep elongrttion. Loading was in steps with usually 24 hr. between steps, and continued until failure. He advocated a "useful strength value" of 61/2 tons (I4,560 Ib.) per sq. in. at 572°F. (300°C.). A specimen broken in 1/2 hr. at 572°F. had a strength of 21 tons (26,880 Ib.) per sq. in. The strength values as determined by 1/2-hr. tests and those by the longer stress-rupturc start to differentiate at 392°F. (200°C). This is the temperature at which time at load began to play an important part. Bolton, Price and Hebemann2 give creep curves on 88-10-2 (A.S.T.M. B-60, CU 88, Sn 10, Zn 2) and steam or valve bronze (A.S.T.M. B-6I, Cu 87.50, Sn 6.00, Zn 3.75, Pb 1.70) at 500°F. Direct comparison of the two alloys under 7500 lb. per sq. in. stress showed the steam bronze (A.S.T.M. B-61) superior in creep strength to the 88-10-2 (A.S.T.M. B-60). The steam bronze shows no creep after 700 hr. (1540-hr. test). The 88-10-2 shows a rate of about 7.0 per cent in 100,000 hr. over the last 700 hr. of the test. Creep curves on the steam bronze are also given at the same temperature (500°F.) with loads of 6500 lb. per sq. in. and 8000 lb. per sq. in. These indicate "creep strengths" (creep rate within I per cent in 100,000 hr.) in excess of 8000 lb. per sq, in. for this alloy. It was found that higher percentages of tin are detrimental to creep resistance. This is caused by the presence of or formation during test of considerable alpha-delta eutectoid in the higher tin alloys. Bolton also states that the high

Jan 1, 1945