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By Nicholas J. Grant, Robert H. Kane, Bill C. Geissen

The platinum-rich section of the Ta-Pt Phase diagram was worked out by metallographic and X-ray techniques, using twenty alloys; solidus temperatures and solid solubilities were determined. Besides s phase, four intermediate stoi-chiometric phases were found: TaPt (high-temperatzrre phase, structure unknown); TaPt2 (ortho-rhomhic, new type); a TaPt3 (orthorhombic, TiCu3 type); ß TaPf3 (monoclinic, new type). TaPt2 and TaPt3 melt congmentlv; two eutectic, two beritec-tic, and one eutectoid reactions occur in the inuestigated range. THE Ta-Pt system was examined as part of a study of the phase diagrams and the intermediate phases of niobium and tantalum with Group VIII transition metals. Since the main scope of this work was the relation among the ordered close-packed phases occurring in the concentration range from AB to AB3 (A = Nb, Ta; B = Pd, Pt, and so forth), only the platinum-rich side of the diagram was worked out thoroughly, while spot checks were made to determine the approximate outline of the tantalum-rich portion. The system has not been thoroughly treated in the literature; Greenfield and Beck1 in their survey of refractory metal intermediate phases found a o phase between -18 and 37 at. pct Pt which coexists at higher platinum concentrations with an unidentified intermediate phase. Browning2 worked on the range between 55 and 100 at. pct Pt; however, this preliminary study did not cover temperatures >1500°C, and included no conclusive crystallo-graphic data. More recently, a study of the platinum-rich section of the Ta-Pt system has been carried out by Ray,3 while Hartley, Parsons, and steedly4 have studied the tantalum-rich section. The crystal structures of TaPt2, a TaPt3, ß TaPt3, and others have been reported in abstracted form by Giessen and Grant.5 EXPERIMENTAL METHODS The methods of alloy preparation and processing and of the metallographic and X-ray procedures follow very closely those described by Ritter, Giessen, and rant' for other refractory metal-noble metal systems, and will not be repeated except where significant differences exist. Starting materials were: tantalum rod of 99.95 pct purity, supplied by National Research Corp., and platinum sponge 99.9 pct, supplied by The International Nickel Co. An oxygen analysis of the tantalum rod showed 30 ppm O; other impurities were C, N, Si, Nb, Mo, none of these exceeding 50 ppm, and 100 ppm W. Homogeneous buttons were readily prepared as follows: ingots of 5 g were weighed and arc-melted under welding grade argon, followed by two remelts after inverting the button. Between 40 and 100 at. pct Pt, weight losses during the first melt rarely exceeded 0.1 pct, in which case the alloy was rejected, and were nil on subsequent remelts. Since the vapor pressure of platinum far exceeds that of tantalum, the small evaporation loss could be assumed to consist of platinum; therefore, a corresponding amount of platinum was added on the second melt. Since it was confirmed on pure tantalum samples that no weight gain through oxygen or impurity pickup occurs (see Ref. 71, it could be assumed that the uncertainty in the composition did not exceed ±0.1 at. pct for the range 40 to 100 at. pct Pt. For the four control melts from 0 to 40 at. pct Pt, an error of ±0.5 at. pct was calculated. Microscopic examination of the whole buttons after melting showed no inhomogeneity other than local coring. This was eliminated by a homogenization heat treatment. A total of twenty-three alloys were prepared. Heat treatments were carried out in a tantalum

Jan 1, 1965

An extensive ferromanganese nodule field south of New Zealand (Fig. 1) has existed beneath the flow of the Deep Western Boundary Current (DWBC) and Antarctic Circumpolar Current (ACC) for at least the past 15 m.y. West of 174°E, between 59°S and 48°S, the field is 300-500 km wide, but east of 174°E, where current flow impinges on the eastern slope of the Campbell Plateau, the field narrows from ca. 200 km at 55°S to ca. 120 km at 51°S. This narrowing coincides with deflection of current flow eastwards, and consequent reduction in bottom-current velocity and eddy kinetic energy. Based on sea floor photographs, dredge samples and 3.5 kHz profile data, six distinct seafloor substrates are delineated, forming broadly parallel zones eastwards from the lowermost Campbell Slope (Fig. 2).

Jan 1, 2003

By T D. Sullivan

Risk lies at the heart of all mining and nowhere more so than in pit slope design. Mining is principally about two areas of risk; safety and economics. This paper addresses the main elements of pit slope design and presents a view of the state of the art in terms of safety, design, prediction of performance, success and pit slope management. There has been a recent trend for serious injury and fatalities due to stability in open pit mines. The causes and contributing factors are analysed in relation to geotechnical knowledge, pit slope design and management. Currently there is no standard classification system for mine slope designs and it is considered doubtful if a universal system is possible. Designs are usually formulated in accordance with many different methodologies, various engineering principals and a number of different philosophical approaches, with results often presented in terms of Factors of Safety or Probabilities of Failure. However wide experience has demonstrated that even with apparently detailed engineering investigation and analysis at Feasibility Level, at the Operating Stage mine slope angles are regularly subject to large changes. The general experience is that these changes are ± 5° and 16° overall, which is considerably outside the accepted order of accuracy. For large deep mines these are huge changes and represent significant risks to the operations. The current state of practice in pit slope design is examined for two stages, feasibility and operating.

Jan 1, 2006

By R. R. Vandervoort, W. L. Barmore

Tensile creep experiments were conducted on high-purity, poly cvystalline rhenium from 1500" to 2300°C at stresses from 1500 to I0,OOO psi in a vacuum of 10-a torr. The apparent activation energy for creep was 60 kcal per mole, and the steady-state creep rate varied directly with stress to the 3.4 power. Dislocation substructure that developed during creep was studied by transmission electron microscopy. Possible rate-controlling deformation mechanisms are discussed. The creep behavior of most metals at elevated temperature can be represented by the following equation:&apos;&apos;&apos; t = Cf(s)(^)(s/E)nD [1] where i = steady-state creep rate, C = constant, f(s) = a function involving microstructure, s = applied stress, E = the average elastic modulus at test temperature, n = constant, D = diffusion coefficient According to this well-established relationship, metals with higher elastic moduli and lower diffusion coefficients should have greater creep resistance at the same stress and temperature and equivalent mi-crostructures. While no diffusion data are available, the diffusivity of rhenium should be less than that for most other refractory metals because of its high melting point and hcp crystal structure. The Sherby-Simnad relation for calculating atomic mobility in metallic systems3 predicts that the diffusion coefficient for rhenium is less than that experimentally determined for tungsten4 in the temperature region 1500. to 2200°C. At these temperatures the elastic modulus for tungsten5 is only slightly larger than the extrapolated modulus for rhenium.6 Thus, rhenium is a good possibility for a a high-temperature structural material, but few data on the creep of rhenium have been reported. This investigation was undertaken to study the high-tempera-ture deformation behavior of rhenium in detail. EXPERIMENTAL TECHNIQUES The material used in this study was consolidated from high-purity powder. After cold pressing the powder to a plate a in. thick, the billet was sintered in hydrogen at 2250°C for 24 hr. The plate was reduced to 0.100 in. by cold cross rolling with intermediate anneals at 1650°C for 20 min between passes. The plate was further reduced to 0.060 in. by unidirectional cold rolling with similar heat treatments between passes, and then finally stress-relieved in hydrogen at 1650°C for 30 min. Specimens tested at 1900°C and below were pretest-annealed at 1900°C for 2 50 hr. Specimens tested above 1900°C were pretest-annealed at 2400°C for 5 hr. The impurity content in the "as-received" plate was quite low, table I. Essentially no change in impurity levels was detected in specimens after creep testing. All creep tests and annealing treatments were conducted in a vacuum of 10-8 torr in a test furnace heated by a tungsten mesh element. The load was applied to the specimens through a bellows, and stresses were maintained to ±1 pct of the selected value by periodic corrections for changes in specimen cross-sectional area during creep and for changes in the bellows spring force due to load column extension. One-inch-diameter tungsten force rods were used in the hot zone of the furnace. Deformation at temperature was measured by optically tracking gage marks on the specimen. Temperature was measured by a calibrated optical pyrometer and was determined to ±5"C. Grain sizes were determined by the linear intercept method and specimens were examined in the "as-polished" condition, using polarized light. Specimens annealed at 1900°C had a grain size of 52 ± 5µ , and those annealed at 2400°C had a grain size of 148 * 11 µ. Pieces were cut from the gage section of creep-tested specimens and planed to a thickness of about 0.010 in. by spark discharge machining. Thin foils for viewing by transmission electron microscopy were obtained by electropolishing in a solution of 6:3:1 ethyl alcohol, perchloric acid, and butoxy ethanol, respectively, using the window technique. Bath temperature was —4OoC, and the cell potential was 35 v. The foils were examined in Siemens Elmiskop I, operating at 100 kv. RESULTS AND DISCUSSION In order to analyze the results from creep experiments, Eq. [I] is rewritten in the following form: <=Kf(s)ne-/RT [2] where K = constant, ?// = apparent activation energy for creep,

Jan 1, 1970

IRON AND STEEL COMMITTEE ALBERT SAUVEUR, Chairman A. A. STEVENSON, Vice-Chairman HERBERT M. BOYLSTON, Secretary, Abbot Bldg., Cambridge,- Mass. Guilliaem Aertsen, Robert W. Hunt, E. Gybbon Spilsbury, John Birkinbine, J. E. Johnson, Jr., J. S. Unger, William H. Blauvelt, Charles Kirchhoff, Felix A. Vogel, Charles A. Buck, Richard Moldenke, Leonard Waldo, F. D. Carney, Charles F. Rand, William R. Walker, James Gayley, Joseph W. Richards, William R. Webster, Henry D. Hibbard, . C. F. W. Rys, . Frederick W. Wood. Henry M. Howe, Messrs. Guilliaem Aertsen, C. A. Buck, and F. D. Carney have accepted membership on the Committee.

Jan 3, 1914

By R. G. Davies, R. H. Richman

Several techniques, X-ray diffraction, transmission electron microscopy, conventional, metallography, and mechanical property measurements are combined to investigate decomposition in the suPersaturated solid solution Fe-27 at. pct Be. The following sequence of structures for aging temperatures below 400°C is established: massive short-range order, pseudo-FeBe (L2o) order, Fe3Be (DU3) order, sinusoidal modulation of composition, general or continuous precipitation, and discontinuous or cellular precipitation. In addition to this sequence itself, several features of the individual stages are noteworthy; for example, when coherent modulations occur, their development is parallel to (100) everywhere in the specimen apparently without incubation, and both the wavelength and the composition amplitude of the modulations increase with aging time. Two types of discontinuous reaction are observed, depending upon the aging temperature. At low temperatures (<550°C) the boundary-nucleated variety of discontinuous precipitate predominates; but at higher temperatures (>600°C) the crystallographic variety occurs instead. Hardness and flow-stress measurements clearly reflect the quoted structure sequence. Some qualitative aspects of the structure changes and of the relationship between mechanical properties and structure are discussed. BERYLLIUM in solid solution markedly strengthens iron.&apos;" A recent study demonstrated that not only were supersaturated single crystals of Fe-27 at. pct Be very strong, but they deformed primarily by mechanical twinning.3 Such behavior, coupled with the decrease in solid solubility of beryllium in a Fe from about 33 at. pct at 1165°C to about 10 at. pct at 500°C4 as shown in Fig. 1, leads naturally to questions about property changes with aging in this system. In 1929 roll,1 and later Seljesater and Rogers,5 reported that useful increases in strength could be obtained by aging quenched Fe-Be alloys, but it was not until quite recently that any of the details of precipitation in these alloys were studied. Richman6 observed that both an Fe3Be superlattice (with a DO3 crystal structure the same as Fe3A1, illustrated in Fig. 2) and a modulated structure preceded the appearance of a discrete second phase in the 20 at. pct Be composition. Here, a modulated structure refers to a periodic composition fluctuation within one crystal, while its lattice remains coherent. Subsequent work on a Fe-Be alloys by Heubner7 and von Heimendahl and Heubner8 confirmed the existence of the modulated structure, although only an FeBe (L20) ordered structure was detected (also shown in Fig. 2). Both Rich-man and Heubner corroborated the large increases in hardness during aging reported earlier, but other mechanical properties were not examined. Since modulated systems are encountered relatively infrequently, the influence of coherent modulation upon mechanical properties is not well-known. With the accumulation of information about age-hardening systems, it has become clear that many precipitation reactions cannot always be described by the concept of heterogeneous nucleation at some lattice discontinuity followed by diffusional growth. Rather, complex sequences involving prior "preparation" of the matrix are quite common; that is, a transition structure occurs in the parent phase, instead of in some "nucleus" form. Delineation of such subtle sequences appears to be a worthwhile step toward the understanding of solid-state precipitation. Since not many structural progressions of comparable nature have been reported, the objectives of the present survey are to identify and elucidate the various

Jan 1, 1967

resident Schloss called the meeting to order at 9:30 a.m. Bob Dean, manager of the Hotel Colorado, gave a brief explanation of hotel policies governing the swimming pool, and also extended an invitation to the Institute to return in 1963, and stated that the dates June 30 through July 3 were available for next year. He also stated that June 28 through July 1 were available for 1964. He felt that holding the convention during the 4th was not good policy, and that dates prior to or following the 4th would make for better attendance. John A. Reeves, chairman of the Resolutions Committee presented the following resolutions: RESOLVED: That the members of the Rocky Mountain Coal Mining Institute in convention assembled, hereby ratify and confirm all official actions of the officers, the Executive Committee and other committees of the Institute between June 28, 1961, and the present time.

Jan 1, 1962

By J. D. Vincent, Pierre R. Hines

INTRODUCTION OF FROTH FLOTATION INTO THE UNITED STATES "The introduction and development of the flotation process have proved to be of such momentous importance to the mining industry of the United States that they deserve to be considered historically ." T. A. Rickard l James M. Hyde installed the first froth flotation process in the U. S., about Aug. 1, 1911, at the Basin Reduction Co. plant, Basin, Mont., which was then under lease to the Butte and Superior Copper Co.2 Two months later, the Minerals Separation Ltd.. filed suit in the U. S. District Court of Montana against Mr. Hyde char& him with infringement of their Patent 835, 120/1906. The U. S. District Court decided in favor of Minerals Separation Ltd. on July 28, 1913. This decision was reversed by the Ninth Court of Appeals in San Francisco on May 4, 1914. Then the U. S. Supreme Court, on Dec. 11, 1916, upheld the patent on three counts only but restricted it to &apos;the use of less than one percent of oil.” The Miami Copper Co. was sued by Minerals Separation on Oct. 10, 1914, for infringement of Patents 835,120 and 962,678, and 1,099,699. The Miami Copper Co. was operating under John M. Callow&apos;s froth flotation process Patent 1,104,755/~uly 23, 1914. Miami lost this suit in the District Court of Wilmington, Del., and appealed to the Court of Appeals, Third Circuit, Philadelphia. The Third Court of Appeals decided against Miami, May 24, 1917. The case was referred to a Master and hearings to determine profit and damages were not settled until 1923. Nevada Consolidated Copper Co. was sued Sept. 9, 1919, and Magma Copper Co. was sued Jan. 10, 1920. Thus, the first ten years of froth flotation in the U. S. were beset by patent litigation. This litigation suppressed the free circulation of new discoveries or useful information that might be cited in a patent suit against the originator; it accounts for many of the blank spaces in the historical record. T. A. Rickard3 has reviewed the history of flotation and has given the chain of discoveries, patents, and processes that resulted in froth flotation. He has re- ported also on the patent litigation beginning in 1911 and ending in November 1923, when Minerals Separation&apos;s Patent 835,120 expired. He comments on

Jan 1, 1962

By A. Stanley, J. C. Mead

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus % in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9 1/2 in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9 1/2 in. bed sinters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite recovery). These tests have proved successful in increasing ilmenite recovery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was plus from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By W. Hrynewich

"The Central Canada Potash facility, a division of Noranda Metal Industries Limited, is located 3 kilometres south of the junction of Highways 16 and 2, about 75 kilometres south-east of Saskatoon, Saskatchewan. The site occupies 26~ hectares of land and includes well-equipped machine and electrical _shops, a 325,000-tonne muriate storage and reclaim warehouse, a stores warehouse, processing plant (mill), two mine shafts (one for production, the other for servicing and personnel transport), as well as a two track product loading complex. Access to the site is by road and railway (CN and CP Rail). Suppliers of operating and maintenance use both means of transportation, whereas muriate shipments are entirely by railway.Currently five continuous mining machines - one a Goodman 526 borer and four Marietta 911-A units produce some 10,500 tonnes of ore per operating day. A smaller Voest-Alpine Af.1-100 boom type miner is used for floor cutting which was previously done by a production unit. In order to maintain safe, competent ground conditions only 35 percent of the total ore is extracted. At this rate mine reserves are deemed adequate for about 100 years. The plant has operated since August 1969 and has a current practical annual capacity of 1,270,000 tonnes of muriate. As of June 30, 1981, some 11.9 million tonnes of muriate had been produced from 31.4 million tonnes of ore. Of the total muriate, 6.1 million tonnes was granular, 4.8 million was coarse, and 1.0 million was standard. No standard product has been made since August 1978."

Jan 1, 1982

By R. W. Ludt, C. C. DeWitt

The use of froth flotation for the separation of minerals has become one of the most important of ore dressing processes. Its particular adaptability to the enrichment of low grade ores has made the process an important factor in the national economy. The methods have been extended to the recovery of a great number of minerals. Among the few minerals which have resisted efforts toward industrial flotation is chrysocolla, a hydrated partly colloidal copper silicate. Chrysocolla, being a product of natural oxidation, has been found to occur in small quantities with many ores which are recovered by flotation methods. In present practice, these small quantities of copper silicate pass off with the tailings and are lost. The advantages to be gained by a satisfactory process for the recovery of chrysocolla is apparent. Any application of principles which points a way toward the satisfactory industrial flotation process for copper silicate would be of advantage. This paper presents an attack on this problem. Two methods for the recovery of chrysocolla have been developed by the United States Bureau of Mines.1,2 They have been successful on a laboratory scale but have been seriously restricted in industrial application by critical requirements in the procedure. In one of the Bureau of Mines methods,&apos; the ore is activated with sodium or hydrogen sulphide in an aqueous solution at a pH of 4. Amy1 xanthate is then used as a collector with pine oil as a frother in the flotation process. An excess of sulphide acts as a depressant and the state of optimum conditions is difficult to control industrially. In the second Bureau of Mines method,2 soap is used as the collector at a pH of between 8 and 9. The diffi- culties with this process are that soap is not a specific collector, that heavy metal or alkaline earth ions cause the formation of insoluble soaps, and that a more acid solution causes the formation of a free acid which does not act as a collector for chrysocolla. The problem of recovering chrysocolla by flotation involves the selection of a suitable collector. The collector molecule must be composed of an active polar group that has an attraction for chrysocolla, and of a hydrocarbon chain. Certain dyes have been shown to have an attraction for certain minerals. Suida3 found that hydrated silicates are colored by basic dyes. Dittler4 showed that chrysocolla, among other colloidal minerals of acid reaction, preferentially takes up such basic dyes as fuchsin B, methylene blue, and methyl green. Endell5 gave information to show that the colloidal material in clay may be determined by its selective adsorption of fuchsin. A simple experiment, likewise, illustrates the difference in the adsorptive power of chrysocolla and of silica for the basic triphenyl methane dyes. When a mixture of chrysocolla and silica is immersed in a very dilute dye solution, less than 5 ppm, the chryso-colla is rapidly dyed and the silica is dyed more slowly. The difference is substantial but one of degree. Dean2 showed that the dyes, crystal violet and toluidine blue, are taken up by quartz in an adsorption type process. The difference in the adsorptive power, however, offers the means by which a new collector may act. To form such a collector, a hydrocarbon chain must be attached to the dye molecule. This involves a process of organic synthesis. Butyl, hexyl, and octyl hydrocarbon chains were selected for substitution in the malachite green molecule. For the purpose of identification, the alkyl-substituted dyes formed are called: butyl-malachite green; hexyl- malachite green; and octyl-malachite green. An outline of the procedure for their synthesis is given in the appendix. It is generally recognized in the preparation of this type of dye that the chemical structure of some of the dye molecules varies. However, a uniform formula is attributed to the dye. Such a procedure has been followed in specifying the structure of these alkyl-substi-tuted malachite green dyes. The structure is given on the basis of their properties as an homologous series of dyes, on their method of preparation, and on the purity of intermediates used. Structure of substituted alkyl malachite green is: C6H4 N(CH3)2 p-R C6H4 CH C6H4 N(CH)2 Procedure The flotation cell is a Bureau of Mines 100-g, batch unit provided with an air inlet at the bottom above which is a variable speed agitator. The agi-

Jan 1, 1950

By Hiram C. Draper

The Mining Division of the Bureau of Mines, in cooperation with the Obstacle and Demolition Branch, Engineer Research and Development Laboratories, Department of the Army, undertook some extensive tests to determine the effectiveness and limitations of using standard demolition-type shaped charges in underground Mining. The tests proved the shaped charge has no practical application in the functions of mining covered by this report. At an approximate cost of 0 per foot hole plus spall produced by the most efficient single shots with the M3 shaped charge, the cost is over 20 times that of producing a better hole with a rock drill. The concussion would prohibit its use in many mines, and the excess carbon monoxide gas liberated by the pontolite explosive would be an additional hazard when used underground.

Jan 1, 1948

By Dale A. Vaughan, Oliver M. Stewart, Charles M. Schwartz

Solid-solubility limits at 1500°, l000q and 500°C for carbon, nitrogen, and oxygen in high-purity tantalum were determined by X-ray lattice-parameter methods. For carbon, the solubility was found to be 0.17 at. pct at 1500°C, and less than 0.07 at. pct at 1000°C. A nitrogen solubility of 3.70 at. pct at 1500° C decreases linearly with temperature to 2.75 at. pct at 1000°C, and 1.8 at. pct at 500°C. In the case of oxygen, the solubility was found to be 3.65 at. pct at 1500°C, 2.95 at. pct at 1000°C, and 2.5 at. pct at 500°C. The phases Ta2C, the low-temperature modification of Ta205, and Ta,N of unknown composition but which has a superlattice structure based upon the original bcc tantalum lattice have been identified as the initial precipitates in the respective systems. Metallographic methods were employed to verify the X-ray analyses. The etching behavior of Ta is discussed in terms of lattice i?rzperfections and precipitate phases. The excellent fabricability, high melting point, and nuclear properties of tantalum are responsible for interest in this refractory metal. Data on the solid solubility of the interstitial elements (oxygen, nitrogen, and carbon) in tantalum and on the precipitate phases are somewhat limited. The significant contributions are discussed below. Because the purity of electron-beam melted tantalum (only recently available) is considerably higher than that used in previous studies, the present investigation was initiated. Gebhardt et all-3 have investigated the tantalum-oxygen and tantalum-nitrogen systems with particular reference to the changes in physical properties and to the rates of reaction between these gases and the metal. The solubility of oxygen in tantalum was reported2 to be 3.7 at. pet at 1500°, 2.3 at. pet at 100O°C, and 1.4 at. pct at 750°C. Schonberg4 reported that several oxide phases (Ta40, Ta,O, TaO and Ta,05) exist while X-ray studies by Gebhardtl showed only two oxides, Ta,05 and an unidentified phase which was associated with a platelet-type precipitate. La-gergren and Magneli,&apos; however, questioned the existence of compounds other than the two allotropic modifications of Ta,05 for the tantalum-oxygen system. In the case of nitrogen, the solubility was estab- lished by Gebhardt3 to be of the order of 7 at. pct at 1800°c. The solubility was reported to decrease rapidly with temperature, and, although no limits were established, a precipitate phase was observed by Gebhardt except when the high-nitrogen specimens were cooled very rapidly from the reaction temperature of 1800c. He reported the initial precipitate phase to be a tetragonal distortion of the bcc tantalum lattice while Schonberg6 reported the phase lowest in nitrogen Jo be a cubic super-lattice with a cell size of 10.11 A. Two other nitride phases, Ta,N and TaN, were reported; these appear to be isomor-phous with the carbides of tantalum. The tantalum-carbon system was investigated by Ellinger7 and by Lesser and Braurer. Two compounds, Ta,C and TaC, were reported to exist, each with a range of composition. The solubility of carbon in tantalum was found to be practically nil at all temperatures. Thus, of the interstitial elements which are present in small amounts in high-purity tantalum, carbon might be expected to form precipitates. The present investigation was initiated to obtain additional data on the solid-solubility limits of these interstitials at 1500°, 1000°, and 500&apos; with particular emphasis on the distribution and the identification of the precipitate phases. EXPERIMENTAL WORK AND RESULTS In the present investigations of the solid solubility and of the precipitate phases in the systems tantalum-nitrogen, -oxygen, and -carbon, high-purity tantalum was reacted with high-purity gases, homogenized at 1800°c, and annealed at and quenched from 1500°, 1000,

Jan 1, 1962

By C. W. Spencer, B. F. Addis, G. H. Bishop, C. A. Steidel

STUDIES of the inter granular attack of metals by liquids have generally been confined to polycrys-talline specimens. This note reports the results of preliminary studies of the penetration of bismuth-rich Ni-Bi liquids into oriented bicrystal specimens of nickel. It is found that isothermal penetration follows a parabolic rate law and shows a definite anisatropy with respect to direction on the grain boundary surface. Nickel bicrystals containing < 100> tilt boundaries were grown using the Czochralski method. Relative misorientation of the parent crystals was controlled to ± 1deg. The boundaries are straight in the growth direction, i.e., the common <100>, and generally straight in the transverse direction, as shown in Fig. 1. The starting material, Inco HPM grade nickel was contained in a recrystallized alumina crucible and maintained under argon. Bicrystals typically contain 30 ppm C, less than 10 ppm S, O, N, and 50 ppm residual metallics, principally iron and silicon. Specimens, as schematically illustrated in Fig. l(c),were cut from bicrystal ingots using a thin water-cooled abrasive wheel. To insure removal of all residual strain, which accelerates the penetration, specimens were round through 600 grit silicon-carbide, chemically polished, and annealed for 4 hr at 1375°C under argon. The procedure described by Cheney et al.1 for exposing specimens to

Jan 1, 1962

The cleaning of copper containing fayalite slag using a CuFe alloy was studied under an oxygen pressure ranging from 10-7 to 5 X 10-11 atm at 1200º c. The reduction of copper from fayalite slags by Cu-Fe alloys was found to be very effective. For example, at a P 02 of 10-7 atm, the copper content in the slag was reduced from 4.1 to 0.402 wt. percent after 90 min. of reaction time using a Cu- 5.27 wt. percent Fe alloy. The copper reduction rates were found to be independent of P02 in the gas phase when using Cu-Fe alloys which exhibit an iron activity of greater than 0.4. A mechanism for slag c1eaning is proposed based on the concept that the Cu-Fe alloy lowers the oxygen potential of the slag according to Eq.1 1, fol1owed by the reduction of cuprous oxide by wustite as shown in Eq.2. Here X represents element X in the alloy phase while (X) represents the element in the slag phase. The copper and iron transfer behaviour obtained in this study are explained by a ?Stepwise equilibrium model.

Jan 1, 1985

By Z. Yang, X. Bian

"The synthesis of 3-hydroxy-2-naphthyl hydroxamic acid (H2O5) as a collector for bastnaesite is described. The flotation performance of H2O5 with respect to bastnaesite was investigated by flotation tests, measurements of the surface adsorption capacity, and polarizing microscopy. The results indicated that H2O5 exhibits superior collecting performance compared with direct flotation recovery of bastnaesite, with recoveries above 90% at pH 8–9. The adsorption mechanism of H2O5 on bastnaesite was studied by solution chemistry analysis and ?-potential tests. The results showed that the adsorption of H2O5 is associated with the types and concentrations of hydrolysis products of rare earth cations on the bastnaesite surface. At pH 8–9, the hydrolysis-dominant components RE(OH)2+ and RE(OH)2 + of the rare earth cations were adsorbed onto the bastnaesite surfaces, which were considered to be the major positive active points and were beneficial to chemical adsorption of the reagents. The pH range of 7–10 was optimum for bastnaesite floatability, and the adsorption of H205 on bastnaesite was by means of chemical adsorption through formation of a five-membered ring chelate. IntroductionThe rare earth elements (REEs) have long been known as ‘industrial gold’ and are widely used in the petroleum, chemical, ceramic, metallurgical, and permanent magnet material industries because of their special physicochemical properties (Chen, 2011; Krishnamurthy and Guptac, 2002; Esquivel, Bohe, and Pasquevich, 2002; Abrahami, Xiao, and Yang, 2015). Demand for REEs is growing in response to continuous advances in science and technology. Currently, most REEs are extracted from bastnaesite ((Ce,La)CO3F) (Chelgani et al., 2015; Yosry, 1990), which is recovered from rare earth ores by flotation. Typically, hydroxamic acids are used as bastnaesite collectors (Xu, 2015; Ren, 1997; Mousumi and Venugopal, 2016). Research into the mechanism of interaction between hydroxamic acids and bastnaesite is important for improving the recoveries of bastnaesite. However, there have been few studies that examine this mechanism."

Jan 1, 2017

By X. Nie

A new one-step process was developed to completely eliminate iron and directly obtain pure Ni, Co, and Cu chloride catholyte from nickel rich matte by low temperature (400 0c) chlorination roasting of Ni, Co, and Cu with HCl and leaching with water (PH 7). The effects of roasting temperature, time, oxygen potential and HCI addition and leaching temperature, pH, H2S04 addition and residue sulphur on the recovery of metals and purity of leaching solution were investigated. More than 99 % of Ni and Co and 95 % of Cu in the matte was directly extracted into the solution while 99.995 % of Fe remained in the residue. The iron concentration in the leaching solution was less than 3.6 X 10-5 mol/L.

Jan 1, 1994

By H. R. Nelson, Martin S. Maier

The diffusion of metals plays an important role in many metallurgical processes,lJ as, for example, the formation of alloys by the annealing of mixed powders, the homogenization of segregated alloys, and the carburization of steel. Other equally important phenomena, such as recrystalli-zation and creep, are based on the movement of like atoms. Although much is known concerning the effects of diffusion in metals, very little is known concerning the details of the physical processes involved. 1,3-5 The data that should give most information concerning the processes involved in diffusion are those on the diffusion of like atoms—self-diffusion. Very little is known concerning the rates of self-diffusion of metals because before the discovery of artificial radioactivity in 1934 there was no means of detecting the movement of the diffusing atoms through the solvent metal except in the elements possessing natural radioactive isotopes—lead, thallium and bismuth. Using the induced radioactivity of the element as an indicator, atoms can now be "tagged" so that the rate of diffusion of the radioactive isotopes of an element (tagged atoms) into the stable isotopes of the same element can be measured. As the radioactive and stable isotopes of an element differ only in nuclear structure, it is quite certain that all isotopes of an element diffuse at the same rate. Two techniques have been used to measure self-diffusion. In one, a thin layer of radioactive metal is deposited on the face of the solvent metal, and the decrease in the surface activity as the radioactive atoms diffuse into the solvent metal is measured. This method has been used to determine the coefficients of self-diffusion of lead6-l0 and bismuth10 using natural radioactive isotopes, and of gold1&apos; and copper12 using artificial radioactive isotopes. In the other technique a thin layer of the radioactive metal is deposited oil one face of the solvent metal, and after diffusion has occurred the specimen is cut in thin sections to obtain the concentration-penetration curve. This method has been used to determine the coefficients of self-diffusion of gold,13 zinc,14-l6 silver," and copper.17 The second method was used for the work described in this paper. Materials The coefficients of self-diffusion were determined in two grades of polycrystalline copper and in single crystals of copper. Commercial OFHC copper was used in the form of disks I in. in diameter by % in. thick. Samples of high-purity copper described by Phillips and Skinnerl9 were supplied through the courtesy of the American Smelting and Refining Co. The high-purity copper rod was used without remelting by cross-rolling to a thickness of 0.075 in., from which disks 7/8 in. in diameter were cut. A good grade of bus-bar copper was used to make the single crystals.

Jan 1, 1942

By Martin S. Maier, H. R. Nelson

The diffusion of metals plays an important role in many metallurgical processes,lJ as, for example, the formation of alloys by the annealing of mixed powders, the homogenization of segregated alloys, and the carburization of steel. Other equally important phenomena, such as recrystalli-zation and creep, are based on the movement of like atoms. Although much is known concerning the effects of diffusion in metals, very little is known concerning the details of the physical processes involved. 1,3-5 The data that should give most information concerning the processes involved in diffusion are those on the diffusion of like atoms—self-diffusion. Very little is known concerning the rates of self-diffusion of metals because before the discovery of artificial radioactivity in 1934 there was no means of detecting the movement of the diffusing atoms through the solvent metal except in the elements possessing natural radioactive isotopes—lead, thallium and bismuth. Using the induced radioactivity of the element as an indicator, atoms can now be "tagged" so that the rate of diffusion of the radioactive isotopes of an element (tagged atoms) into the stable isotopes of the same element can be measured. As the radioactive and stable isotopes of an element differ only in nuclear structure, it is quite certain that all isotopes of an element diffuse at the same rate. Two techniques have been used to measure self-diffusion. In one, a thin layer of radioactive metal is deposited on the face of the solvent metal, and the decrease in the surface activity as the radioactive atoms diffuse into the solvent metal is measured. This method has been used to determine the coefficients of self-diffusion of lead6-l0 and bismuth10 using natural radioactive isotopes, and of gold1&apos; and copper12 using artificial radioactive isotopes. In the other technique a thin layer of the radioactive metal is deposited oil one face of the solvent metal, and after diffusion has occurred the specimen is cut in thin sections to obtain the concentration-penetration curve. This method has been used to determine the coefficients of self-diffusion of gold,13 zinc,14-l6 silver," and copper.17 The second method was used for the work described in this paper. Materials The coefficients of self-diffusion were determined in two grades of polycrystalline copper and in single crystals of copper. Commercial OFHC copper was used in the form of disks I in. in diameter by % in. thick. Samples of high-purity copper described by Phillips and Skinnerl9 were supplied through the courtesy of the American Smelting and Refining Co. The high-purity copper rod was used without remelting by cross-rolling to a thickness of 0.075 in., from which disks 7/8 in. in diameter were cut. A good grade of bus-bar copper was used to make the single crystals.

Jan 1, 1942

By R. Barany

The heat of formation at 298.15 ° K of muscovite was determined by solution calorimetry . The value obtained for the heat of formation from the elements was -1,421.2 ± 1.3 kcal / mole , and the value for the heat of formation from the oxides was -55.9 ± 1.3 kcal /mole .

Jan 1, 1964