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By George P. Hulst

Prior to 1914, there was little demand for antimony in this country; its use was limited almost entirely to the manufacture of type and bearing metals. Practically no antimony ore was mined here, the market being supplied principally from China, and the alloy was produced by a direct mixing of lead and antimony. The great world war, with its demand for shrapnel in hitherto undreamed of quantities, precipitated a great boom in the price of antimony. Nominally quoted at 6 c. to 7 c. in 1914, the price increased by leaps and bounds to 45 c. in March, 1916. Under the stimulus of high prices many small mines were opened, for it became profitable to work ores containing as low as 20 per cent. antimony. High-grade sulfide ores (stibnite), containing 55 per cent. to 60 per cent. antimony, were received from Bolivia, China, and Alaska. Low-grade sulfide ores, running from 20 per cent. to 45 per cent., were produced in Nevada, California, Idaho, Utah, and Mexico; much of this ore was fairly rich in silver. The principal oxide ores came from Mexico and Oregon. The following analyses are representative of various types of antimony ore: Sb Pb Cu AG AU SIO2 Be S CaO Zn As Oxide ore.......... 25.09 46.02 0.95 0.40 10.98 0.30 0.22 Sulfide ore......... 55.20 15.12 1.05 19.87 0.40 0.25 Sulfide ore......... 37.16 11.30 0.10 34.0 7.60 0.50 17.05 7.40 0.36 0.10 Sulfide ore......... 41.55 18.00 0.20 76.0 0.04 5.60 0.25 18.17 0.85 0.25 0.10 The International Lead Refining Co. fortunately was equipped to handle these ores through residue and blast furnaces. The charge consisted of a variety of sulfide ores containing Sb 20 to 60 per cent. and SiO2 6 to 45 per cent., oxide ores containing Sb 20 to 40 per cent. and SiO2 10 to 45 per cent. Secondary materials, such as battery plates, battery mud, lead oxide, paint, etc., together with refinery skims, softener skims, and other refinery byproducts, were treated along with the antimony ore to furnish the lead required. All silver-bearing antimony ores were treated in the residue furnace, the sulfur, iron, and copper forming matte that carried part of the silver, the balance going into lead bullion. The antimony slag produced was sufficiently low in silver to warrant being smelted in the blast furnace to antimonial lead.

Jan 1, 1921

By E. L. Mann

Asbestos is the generic name given to a group of fibrous mineral silicates found in nature. They are all incombustible and can be separated by mechanical means into fibers of various lengths and cross sections, but each differs in chemical composition from the others. The commercial grades of asbestos consist of a spectrum of lengths and sizes and are not definitive as are most organic fibers. It is generally recognized that there are two main groups of asbestos. The first only contains the fibrous serpentine called chrysotile and comprises about 94% of the world production of asbestos. The second group contains five minerals in the amphibole series-crocidolite, amosite, anthophyllite, tremolite, and actinolite. The latter two have no commercial importance. Properties Physical The structures of asbestos fibers have been studied intensively over a number of years. X-ray diffraction patterns have been used as a means of identification and classification. Low angle X-ray scattering techniques have shown that chrysotile fibers are "hexagonally close packed" and parallel to each other, having cross-sectional diameters varying from 18Å to 300 Å units, while the amphibole fibers are many times larger in cross section, have considerably smaller surface areas, and the fibers are less symmetrical than those of chrysotile. Electron micrographs of the more common types of asbestos are shown in Figs. 1-4. Fig. 1 illustrates the softness and flexibility of chrysotile which make it suitable for textile-type products. For a number of years there was considerable controversy over the exact nature of chrysotile fiber, but the photographic evidence in Figs. 5-7 clearly show that it is tubular in structure. The curvature is the result of the mismatch in the dimensions of the silica [Si2O5] and brucite [Mg(OH)2] layers which are on the inside and outside, respectively, of a double sheet. This curvature had been predicted by Pauling (1930) years ago based on the molecular dimensions of the two component structures. In contrast to this, the structure of the amphiboles consists of two ribbon chains based on Si4O11 units separated by a band of seven cations with two hydroxyl groups attached to

Jan 1, 1983

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 G. E. Steudel

The study of the possibility of effecting a lower cost in the manufacture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the consideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improvements, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph,5 whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality. Large slag volumes result in higher coke consumption and cost of disposal. Any attempt to lower slag quantity must be accomplished

Jan 1, 1940

By G. E. Steudel

The study of the possibility of effecting a lower cost in the manufacture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the consideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improvements, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph,5 whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality. Large slag volumes result in higher coke consumption and cost of disposal. Any attempt to lower slag quantity must be accomplished

Jan 1, 1940

By D. R. Shoemaker

BACKGROUND AND PROJECT DESCRIPTION A 16.5 kt per day underground molybdenum mine is being developed at Questa in northern New Mexico by Molycorp, Inc., a wholly-owned subsidiary of Union Oil Company of California. Small scale underground mining first began in 1920 and large scale open-pit mining was initiated in 1965, but because of the high stripping ratio, surface mining is being phased out. Construction of the underground mine began in 1979, and first production is scheduled for 1983. Full production of 9 Gg of molybdenum per year will be reached in 1984. Access to the mine will be through a service shaft, development shaft, and a conveyor decline. 'The service shaft, 7.3 m in diameter and 396 m deep, will be used for personnel, supplies, development muck, and fresh air intake. It will be equipped with a large, double-deck man and material cage. A double-drum hoist will be used for cage hoisting and will handle up to a 27-t load in balance with a counterweight. Development muck will be loaded into skips in balance and a single-drum auxiliary hoist will provide secondary access. The 396-m-deep, 4.3-m-diameter development shaft will be used for initial mine development, mine exhaust air, and a slick line pass for concrete entering the mine. The development shaft is located 400 m north of the service shaft. Run-of-mine ore will be transported from one of two 1.4-kt storage pockets by a 1.2-m-wide conveyor belt located in the 10°, 1980-m-long, 5.0-m-wide, and 4.6-m-high decline. The decline conveyor belt has a centerline length of 2229 m and a vertical rise of 363 m. It will be powered by four 448 kw motors driving a single drive pulley. There will be two 1.2 m transfer belts on the surface with centerline lengths of 322 m and 542 m. All three conveyor belts will be reinforced with steel cables. On the surface, in addition to the headframes and hoist houses for the two vertical shafts, a large maintenance shop, warehouse, compressor house, change house, main office building and timber framing building have been constructed.

Jan 1, 1981

By W. L. Batt

MODERN war means mechanization, and mechanization means raw materials, especially minerals-and lots of them. Let me recall a few events of recent history-events that constitute mile- stones down the road to the war now blazing on three continents and threatening to spread still farther. In 1932, Japan took Manchuria, largely upon the excuse that she needed raw materials; in 1935, Italy invaded Ethiopia, largely upon the excuse that Mussolini's growing nation needed raw materials-in this case oil; from 1933 to 1935, Germany cried out for a greater share of the world's natural wealth, and since 1938 has invaded or occupied ten other countries, frequently upon the excuse that she needed-and had a right to-more raw materials.

Jan 1, 1940

By F. Habashi

Conclusive evidence is given showing that the dissolution of gold in cyanide solution follows mainly the over-all equation A similar equation can also be written for the dissolution of silver. Theoretical derivation of the velocity equation for this reaction has been obtained, which describes quantitatively the experimental facts. The equation is as follows: where [CN-] and [O2] = the concentrations (in moles/ml) of cyanide and dissolved oxygen, respectively. DcN- and Do2 = the diffusion coefficients of cyanide and dissolved oxygen; 1.83 x 10-5 and 2.76 x 10-5 cm2 sec-1, respectively. A = the surface area of the metal in contact with the aqueous phase, in cm S = the thickness of the boundary layer; varies between 2 and 9 x 10-3 cm, depending on the speed and method of agitation. Rate being expressed in g. equiv. sec-1. INTRODUCTION During recent years a number of papers have been published on the kinetics of dissolution of gold and silver in cyanide solution.1,6 The authors agree that the rate of dissolution is diffusion-controlled, but disagree regarding the mechanism of the reaction. Some authors'-3 support the Bodlander equations: The apparently conflicting results, when critically reviewed, are concordant and lead to the conclusion that the proper equation for the dissolution reaction should be only Eq. 1. This conclusion is based on the following facts: 1) For every 2.2 equivalents of metal dissolved one mole O2 was consumed. 2) For every 1 equivalent of metal dissolved, 2 moles of cyanide were consumed.' 3) Hydrogen peroxide is formed during the dissolution of gold or silver, and for every 2.2 equivalents of metal dissolved, one mole H2O2 was produced.1,2 The fact that Kudryk and Kellogg4 did not detect H2O2 during the dissolution of gold should not throw doubt on the many affirmative reports of other investigators. Lund2 has shown that depending on the crystalline nature of the silver surface, (e.g. method of etching the sample) H2O2 may undergo catalytic decomposition heterogeneously according to: Further Kudryk and Kellogg4 carried out their experiments in the presence of 0.5% KC1 to achieve a greater electrolytic conductivity, and Matsudaira7 has shown that H2O2 is catalytically decomposed (homogeneously) in the presence of C1- ion in alkaline solution. 4) Experiments carried out by Boonstra8 and Lund2 showed that the dissolution of gold and silver respectively in KCN + H2O2 in the absence of oxygen was a slow process. Thus the reaction takes place to a minor extent only. Further kinetic studies showed that: 1) The rate of dissolution depends on the surface area of the metal in contact with the liquid phase, thus indicating that the process of dissolution is a heterogeneous process. 2) The rate of dissolution depends on the speed of stirring,5,6 thus indicating that the process is controlled by a physical phenomenon.

Jan 1, 1967

By C. Feng, I. R. Kramer

A study was conducted to determine the influence of the surface on the yield point of fcc metals and high-purity iron. For the high-purity fcc metals, the yield Point produced by restraining a specimen may be eliminated by removing a sufficient amount of the layer. A yield point similar to that in fcc metals was found when iron specimens were pre-strained after a prior deformation. This yield point is not due to an aging effect but is associated with the surface layer. The yield points in age-hardenable aluminum alloy and those connected with "work softening" are not affected by a surface-removal treatment and are, therefore, caused by bulk-type dislocation reactions. The yield point which appears in previously strained fcc metals has been the subject of numerous studies. Several theories have been advanced to explain this "unloading-reloading" yield point. Cottrell and stokes,' from their study on the effect of temperature on plastic properties of high-purity aluminum, considered that it was the result of an avalanche "breakdown" of Lomer-Cottrell sessile dislocations. Haasen and Kelly2 extended this concept of dislocation-dislocation interaction to account for the yield point when unloading and reloading was carried out at the same temperature. On the other hand, strain aging was found to be responsible for the yield point in A1-Mg alloy by Westwood and Broom,3 and in aluminum alloys containing various amounts of alloying elements by Smallman, Williamson, and Ardley.4 In these cases, the yield point is considered to be associated with the locking of dislocations by impurity atoms. Birn-baum and filer5,6 and Takamura and muira7 reported that, when a high-purity copper specimen was prestrained, then held at a lower stress, and finally reloaded at the prestrain temperature, a larger yield point was observed. Such an increase was attributed to the locking of dislocations by point defects. In addition, Bolling8 suggested that both mechanisms, i.e., dislocation-dislocation interaction and strain aging, are responsible for the yield phenomenon, at least in a brass. Thus, there appears to be experimental evidence for both the strain aging and dislocation-dislocation interaction theories to explain the yield point in the fcc metals and alloys. In a previous article,9 one of the authors observed that, if sufficient amounts of metal were electrolti-cally removed after deformation of a high-purity aluminum single crystal, the yield point did not appear upon reloading. Continued pursuit in this area revealed that the effect of the surface on the yield-point phenomenon is not limited to one metal but affects a variety of metals of different lattice structures. Nor is the effect limited to the yield points produced by simple unloading-reloading at constant temperature; the surface exerts an important role on the yield point even when the temperature is changed at some stage during the test. It is the intent of the present article to give additional information regarding the effect of the surface on the yield-point phenomenon and to show the association between the yield point in both fcc and bcc metals and the existence of a heavy concentration of dislocation in the region near the surface of a deformed specimen.10 EXPERIMENTAL Materials. The single crystals used in this study had a nominal cross section of 0.125 by 0.125 in. and a 3-in. gage length and were prepared by a modified Bridgman technique. The initial purity for the metals was 99.997, 99.999, and 99.999 pct for aluminum, copper, and gold, respectively. Bar stock of these same materials was used in the preparation of the polycrystalline specimens of aluminum and copper. The iron specimens were machined from rods of "ferrovac", reported to have a purity of 99.98 pct. In addition, a precipitation-hardening aluminum alloy, containing 1.5 pct Cu, 2.5 pct Mg, and 5.5 pct Zn, was used. The polycrystalline specimens of iron, copper, and the aluminum alloy had a diameter of 0.17 in. and a gage length of 2 in. while the high-purity aluminum specimen had a nominal rec-

Jan 1, 1965

By J. J. Predali, J. M. Cases

J. J. Predali and J. M. Cases (Attachè de Recherche and Charge de Recherche, respectively, Centre de Recherches sur la Valorisation des Minerais, Institut Polytechnique de Nancy, Nancy, France)-The role of CO, in carbonate flotation is one of the parameters of a complex phenomenon: the adsorption of salts of fatty acids onto minerals. The authors therefore must be congratulated for their work. Their experimental results suggest some questions: 1) Recovery with C02, benzoic acid, or HCl, look very similar [(Fig. 1)] to that with the use of air; however, the recovery decreases very slightly (<10%). But in these two kinds of experiments, two parameters are different: the pH which is near 5.8 (with C02) becomes 9.1(without C02) and the concentration of Ca* I x 10-3 mole per 1 (without C02) becomes 1.9 X 10&apos; mole per 1 (without COO. 2) Adsorption experiments (Fig. 2) permit computing the surface coverage 9 which is 7.4% (without C02) and 8% (without CO_.) for an initial concentration of 12 X 10-6 mole per 1 in sodium oleate. It seems doubtful to us that this small variation can produce large increases of the recovery in flotation. 3) It is difficult to correlate adsoption and flotation experiments because they do not seem to have been done with the same size of particles. The concentration needed to obtain 95% recovery (Fig. 1) is about 3.0 X 10_" mole per l, that is, 1.13 X 10-6 mole for about

Jan 1, 1973

By G. R. Palmer

The occurrence of mercury minerals in precious metal-bearing ores has long been noted. During cyanide leaching of these gold-silver ores, some mercury forms a soluble cyanide complex, which is present in the pregnant leach solution and passes through the process operations of the cyanidation mill. As a result, mercury is found on the electrowinning cell cathodes, on the stripped carbon before It is volatilized during carbon regeneration, or in theme tailings pond. The U.S. Department of the Interior&apos;s Bureau of Mines is examining the effectiveness of two ion-exchange resins to adsorb gold and/or mercury from a cyanide leach solution. The resins, Duolite A-7 and Ionac A-305, were contacted with leach solutions containing varying concentrations of mercury, gold, and silver. The resin A-7 reached a maximum mercury loading of 0.0165 lb Hg/ft3, and A-305 reached a maximum mercury capacity of greater than 0.44 lb Hg/ft3. Adsorption of gold by both A-7 and A-305 is influenced by the presence of additional ions in the solution.

Jan 1, 1984

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970

By N. J. Grant, Per Knudsen, J. T. Blucher

The fatigue behavior of three SAP alloys was studied in ternzs of strain rate and temperature, at high strains. The k values in the modified Manson-Coffin equation, Nk4 = C, were less than 0.5 under all test conditions, and change with strain amplitude for the lower-oxide alloys at about 2 pct strain. Lowest k values were near 0.25. Strain rate had no effect on life at 80 F, but had an increasingly greater effect with increasing temperature above 500". Life decreased with decreasing strain rate, above 500"F, and with increasing temperature. Ductility at fracture in a tension test was indicated to be an important factor in determining 1ife in these big+-strain tests with the SAP alloys. INEVITABLY, in the course of mechanical tests at elevated temperatures, particularly if significant time at temperature is involved, there are large changes in structure; these changes make it difficult to relate behavior patterns over ranges of temperature or strain rates at high temperatures. Such changes are to be expetted in low cycle fatigue at low strain rates and high temperatures. Accordingly, it was of great interest to examine the low cycle fatigue behavior of SAP / an aluminum oxide dispersion-strengthened aluminum, a type of alloy which had shown unusual structure stability to temperatures as high as 1000" to 1150°F and resisted recrys-tallization essentially to the melting temperature.&apos;j3 Since the matrix is pure aluminum, there are no complications of averaging, agglomeration, or phase solution. It was also desirable to check the Manson-Coffin equation4?&apos; for the SAP alloys, namely N~E~ = , where ep is the total plastic strain amplitude, k and C are constants, and N is the number of cycles to failure. Here, too, was an opportunity to check the roles of temperature and strain rate with a very stable material. Tavernelli and coffin6 had concluded that k had a value of about 0.5 for many alloys and C was equal to ~/2, where E is the fracture ductility determined from a static tension test. The results were obtained from low-temperature tests where creep and diffusion processes are unimportant. Manson7 found k = 0.6 fitted his data reasonably well; however, in later analyses of a large amount of low cycle fatigue data generated at room temperat~re@"~ he found k to vary from 0.6 for short lives to 0.21 for long-life fatigue tests. In the latter studies,89g Manson separated the total strain range into elastic and plastic components when he found that k was influenced by the nature of the strain. The use of EL (total strain) instead of EP (total plastic strain)4&apos;5 makes a difference in the resultant k value. The ratio of changes with temperature, strain rate, and strain; further, there are the problems in the determination of the elastic strain. Based on these considerations, and the improved fit of points in a plot of by Wells and Sullivan,&apos; is also utilized in these studies. Anderson and wahl,14 using commercial 1100 aluminum, and Blucher and Grant,15 using 99.99 pct pure aluminum, found an increase in life with increasing test temperature. Anderson and Wahl were the first to report low cycle fatigue results from SAP materials. With increasing temperature, the role of strain rate becomes more important. In this regard, care must be exercised to differentiate between frequency (wherein strain rate may vary from zero to a maximum in each cycle, sinusoidally, for example), and constant strain rate, as used in the present study, in a saw-tooth type cycle; in the latter case, the frequency is not specified but can easily be calculated from the strain and strain rate data. It has generally been found that life in low cycle fatigue tests decreases with decreasing frequency16 or with decreasing strain rate at elevated temperatures.15 Coffin,17 reviewing Eckel&apos;s work,16 also reported that k increased with decreasing frequency for acid lead, yielding values from 4.0 at a frequency fo 6.6 cycles Per day to 1-46 at a frequency of 7440 cycles per day; the value of k decreased to 0.58 at a frequency of 2.38 x lo6 cycles per day. EXPERIMENTAL PROCEDURE Three SAP alloys, of two nominal compositions, were tested. Alcoa supplied XAP 005 as 2-in.-diam extruded bar, of nominal composition A1-7 wt pct A1203. The Danish Atomic Energy Commission supplied SAP 930 (A1-7 wt ~ct Ala3) and SAP 865 (A1-13 wt pct Al&) manufactured by Swiss Aluminium Ltd., in the form Of $-in.-diam extruded rod. Metallographic comparison of the structures of XAP 005 and SAP 930 showed the former to have a more uniform oxide distribution. Button-head specimens were machined in the longitudinal direction of the bar with 0.4 in. gage length by 0.2 in. diameter, with a fillet radius of j-B in. After machining, the specimens were electropolished in a 1 to 4 mixture of perchloric acid to methanol to remove all machining marks. All test bars were in the as-extruded condition. The fatigue tests were performed on a hydraulically activated, axial strain machine, with complete reversal of strain.15 Test conditions were:

Jan 1, 1969

P. H. ARGALL, Magdalena, N. Mex. (written discussion *).-The Chilean mill has generally been considered a fine-crushing machine and as such has had many advocates; it has also. had quite a number of denouncers; hence, the action of the Portland engineers in using the Chilean mill for the preparatory work usually accomplished in, rolls or rock breakers and adding a ball mill to do the fine grinding is novel, though of doubtful utility in the long run, even though a fair efficiency is obtained for the combination machines, as measured by Chilean mill practice. In order to follow the development of Chilean mill practice on Cripple Creek ores, I might say that two 6-ft., Chilean mills were placed in operation in Stratton&apos;s Independence mill in April, 1908, treating ore crushed by rolls to pass Y4 in. screen aperture. Screening was, however, found to be an unnecessary refinement, as was also the use of plows in the mills; both screening and plowing were abandoned before the Portland engineers finished the experiments in stamp milling upon which they were then engaged. Two additional 6-ft. Chilean mills were added in the spring of 1911, three of them giving the required 10,000 tons per month capacity, one being always in reserve. The Portland company began using Chilean mills in June, 1910, at its Victor mill, and the same company purchased the Stratton&apos;s Independence mill in June, 1915, and remodeled it; it is now known as the Independence mill. A description of Stratton&apos;s Independence milling practice, is given in the November, 191.1, issue of Mining Majazine (London). These Chilean mills were used as fine grinders and that they accomplished their purpose is shown by the screen analyses, in which 62 per cent. of the pulp passed 150-mesh screens. To make coarse crushers of these Chilean mills, the Portland engineers have increased the speed, substituted steel in various parts for cast iron, provided new and enlarged thrust bearings to take care of the centrifugal force incident to higher speed, and have, thereby, greatly increased the cost of the machines. They have also reduced the width of the tires and dies from 8 in., as used in the original mill, to 7 in., a reduction of 12 ½ per cent. in the crushing surfaces, to say nothing of the loss of abrasion or attrition incident to the larger surface. At the same time, they have apparently increased the weight of the rollers by using tires 5 7/8 in. thick as against tires varying between 3 and 4 in. on the original mills, thereby increasing

Jan 4, 1920

By J. R. Long, T. R. Graham, R. S. Dean

FOLLOWING the development of elec¬trolytic manganese production by the Bureau of Mines, an extensive program was planned to study the character of this high-purity product and its possible utilization in various alloy combinations. Previous work reported in the literature concerning alloys containing substantial amounts of manganese dealt largely with material prepared from relatively impure grades of the metal, and in many instances the impurities thus introduced did not permit true evaluation of manganese as an alloying element. One phase of the Bureau&apos;s program included a general review of the system and the establishment of the physical properties of copper-manganese-zinc alloys, using electrolytic manganese, and with particular emphasis directed to com¬positions within the alpha solid solution area. A number of reports of the work on alloys in this system, giving specific data concerning the equilibrium system and some of its alloys, have already been published. One study1 established the alpha solid solution area at various temperature levels and composition ranges up to 50 pct manganese. Other investigation reports presented the properties of selected alloys; namely, a 65 pct copper, 10 pct manganese, 25 pct zinc alloys&apos; a sensibly white manganese brass composed of 70 pct copper, 20 pct manganese, 10 pct zinc,3 and a series of alloys with a constant copper content of 60 pct and 5 to 25 pct manganese, balance zinc.4 In these reports tensile properties as influenced by cold-working and subsequent annealing were given in detail. Additional data are now available on other alloys of the alpha region up to 50 pct manganese, and these, together with previously published information, are sufficiently complete to establish the relationship between properties and composition for this portion of the ternary system. The present report is a summary of the tensile properties of wrought alloys in the form of sheet and rod stock. The summary is necessarily limited to data for only one condition of cold-work and a single annealing temperature, since a complete exposition of all the data available would be too extensive for this presentation. Nevertheless, it is possible to make a reasonable approximation of the tensile properties in other conditions of cold-working and annealing by a correlation of the data previously presented on specific alloys and the data given here.

Jan 1, 1947

By M. B. Standing, D. L. Katz

Equilibrium data are reported on the composition and densities of coexisting vapor and liquid phases as a function of Pressure and temperature for four hydrocarbon systems prepared from crude oil and natural gas. The data were observed over a range of pressures from 1000 to 8220 Ib. per sq. in. and at temperatures ranging from 35&apos; to 250°F. The compositions of the hydrocarbon systems were such that the critical temperatures of the mixtures were lower than the range of the investigation. Under these conditions, it is shown that the composition of the systems has a marked effect upon both the absolute value of the equilibrium constants and the change of the constants with pressure for the several components comprising the system. From weight-balance considerations, it has been possible to construct the partial phase diagram of a mixture of a 49.9" A.P.I. crude oil and natural gas having a gas-oil ratio of 3660 cu. ft. per barrel. Descriptions of the apparatus used, the analytical procedure, the material used, and the equilibrium measurements are included. A knowledge of the physical behavior of naturally occurring hydrocarbon mixtures is fundamental to the treatment of many problems in the production, transportation and rehing of petroleum. The recent discoveries of pools having formation pressures up to 7500 lb. per sq. in. and higher, and the development of full-scale equipment to operate at pressures of 5000 lb. per sq. in. or more have greatly increased the need for knowledge of the composition, temperature, pressure and density relationships of coexisting vapor and liquid hydrocarbon phases. The following paper presents the results of investigations upon 22 mixtures of natural gas and crude oil of known com~o~itio~ that had been brought to equilibrium under controlled conditions of temperature and pressure. Data are presented upon the densities and compositions of the coexisting vapor and liquid phases for four different systems containing hydrocarbon material having boiling points from minus 2 5S°F. to well over I000 F. The data were observed Over a pressure range from Iooo Ib. per Sq. in to 8220 lb. Per sq. in. and a temperature range from 35&apos; to 2 50°F. The accepted method of computing the physical relationship between two coexisting phases of a hydrocarbon mixture is by use of the "equilibrium constant" K as proposed by Souders, Selheimer, and Brown. definition,

Jan 1, 1944

By M. B. Standing, D. L. Katz

Equilibrium data are reported on the composition and densities of coexisting vapor and liquid phases as a function of Pressure and temperature for four hydrocarbon systems prepared from crude oil and natural gas. The data were observed over a range of pressures from 1000 to 8220 Ib. per sq. in. and at temperatures ranging from 35&apos; to 250°F. The compositions of the hydrocarbon systems were such that the critical temperatures of the mixtures were lower than the range of the investigation. Under these conditions, it is shown that the composition of the systems has a marked effect upon both the absolute value of the equilibrium constants and the change of the constants with pressure for the several components comprising the system. From weight-balance considerations, it has been possible to construct the partial phase diagram of a mixture of a 49.9" A.P.I. crude oil and natural gas having a gas-oil ratio of 3660 cu. ft. per barrel. Descriptions of the apparatus used, the analytical procedure, the material used, and the equilibrium measurements are included. A knowledge of the physical behavior of naturally occurring hydrocarbon mixtures is fundamental to the treatment of many problems in the production, transportation and rehing of petroleum. The recent discoveries of pools having formation pressures up to 7500 lb. per sq. in. and higher, and the development of full-scale equipment to operate at pressures of 5000 lb. per sq. in. or more have greatly increased the need for knowledge of the composition, temperature, pressure and density relationships of coexisting vapor and liquid hydrocarbon phases. The following paper presents the results of investigations upon 22 mixtures of natural gas and crude oil of known com~o~itio~ that had been brought to equilibrium under controlled conditions of temperature and pressure. Data are presented upon the densities and compositions of the coexisting vapor and liquid phases for four different systems containing hydrocarbon material having boiling points from minus 2 5S°F. to well over I000 F. The data were observed Over a pressure range from Iooo Ib. per Sq. in to 8220 lb. Per sq. in. and a temperature range from 35&apos; to 2 50°F. The accepted method of computing the physical relationship between two coexisting phases of a hydrocarbon mixture is by use of the "equilibrium constant" K as proposed by Souders, Selheimer, and Brown. definition,

Jan 1, 1944

By Stanley J. Gasior, Albert J. Forney, Joseph H. Field

Most bituminous coal mined near Eastern industrial areas requiring high-Btu pipeline gas is caking and therefore unsuitable for fixed-bed pres- sure gasification by present techniques. If the caking property of such coal can be destroyed, the resulting fuel can be gasified in a fixed bed at pressure with steam and oxygen as a preliminary step in making a high-Btu gas. To find a practical way to convert strongly caking coals of granular or small lump size to noncaking fuel was the object of this study. In previous work, strongly caking coals of 1/8- to 1½ -in. size were converted to a noncaking coal char by treatment in a static bed in about 3 hours1 and coal of 18-100 Tyler mesh in a fluidized bed in about 5 min.2 An inert gas such as steam, nitrogen, carbon dioxide or a mixture of all three with a small amount of oxygen was used in both studies. In our work a strongly caking bituminous coal from the Pittsburgh seam of 4-8 Tyler mesh size was converted to a noncaking form in free-fall in about 2 sec by dropping it through a countercurrent flow of steam containing 4% to 7% oxygen at 1150º-1200°F and 300 psig. This method should also be applicable to smaller-sized material.

Jan 3, 1965

By George Stern

The purpose of this investigation has been to study one method for making ordinary carbon steel by the powder metallurgy technique. This method con- sists of pressing and subsequently sintering an intimate mixture of iron powder and gaphite in a suitable atmosphere to Produce a steel. Two types of iron Powder have been studied, electrolytic and reduced. The sinkring of mixtures of loose Car-bonyf iron and carbon black for several hours at I000 to 1100 C. (1830&apos; to 2010°F.), followed by forging of the resultant ingots to bars, has been studied by Offermann.&apos; TrailZ has reported on the properties of hydrogen-reduced and electrolytic iron powders, both containing 0.5 per cent carbon. Methods for making steel bodies by powder metallurgy have also been sug- gested by Goetzel13 who recommends the addition of steel powder of well-defined carbon content to pure iron powder, and Volterra,~ who suggests compaction and subsequent sintering of a mixture of steel powder, iron oxide and iron powder. Additional methods for making steel bodies by powder metallurgy have been reported in the literature by Kohering Allen,6 Balke,7 Wulff,8 Boegehold,9 and others.,0,11,12 It is felt that the method of pressing iron powder and graphite, as described herein, lends itself best to production be-Cause of the lubricating action of the

Jan 1, 1946

By George Stern

The purpose of this investigation has been to study one method for making ordinary carbon steel by the powder metallurgy technique. This method con- sists of pressing and subsequently sintering an intimate mixture of iron powder and gaphite in a suitable atmosphere to Produce a steel. Two types of iron Powder have been studied, electrolytic and reduced. The sinkring of mixtures of loose Car-bonyf iron and carbon black for several hours at I000 to 1100 C. (1830&apos; to 2010°F.), followed by forging of the resultant ingots to bars, has been studied by Offermann.&apos; TrailZ has reported on the properties of hydrogen-reduced and electrolytic iron powders, both containing 0.5 per cent carbon. Methods for making steel bodies by powder metallurgy have also been sug- gested by Goetzel13 who recommends the addition of steel powder of well-defined carbon content to pure iron powder, and Volterra,~ who suggests compaction and subsequent sintering of a mixture of steel powder, iron oxide and iron powder. Additional methods for making steel bodies by powder metallurgy have been reported in the literature by Kohering Allen,6 Balke,7 Wulff,8 Boegehold,9 and others.,0,11,12 It is felt that the method of pressing iron powder and graphite, as described herein, lends itself best to production be-Cause of the lubricating action of the

Jan 1, 1946