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By L. J. Dijkstra

During the past several years an extensive study has been made of the internal friction of iron containing carbon or nitrogen in solid solution.1,2,3,4,5 It has been found that a sharp peak is observed in a plot of the internal friction vs. temperature of measurement. With a frequency of about one cycle per second, this peak occurs at 20°C for N and 36°C for C. The physical origin of this anomalous internal friction was traced by Snoekl to the local tetragonal symmetry of the interstitial positions in body centered cubic (bcc) iron. A change in stress induces a change in the equilibrium distribution of the interstitial atoms among the three types of interstitial positions, the three types corresponding to the three directions along which the tetragonal axis may lie. It is the continual striving of the interstitial atoms to maintain an equilibrium distribution that gives rise to a phase lag between strain and stress during oscillation, and hence, to internal friction. A similar type of internal friction has been observed in the bcc lattice of tantalum containing in interstitial solid solution C, N or 06. The theory of this anomalous internal friction leads to the prediction that its magnitude is strictly proportional to the concentration of either the carbon or the nitrogen in solid solution, provided the concentrations are small. This prediction has been verified3 for the range of solubilities ohtainable in alpha iron. This proportionality provides us with a new tool for the study of nitrogen and carbon in alpha iron. It thus enables one directly to follow the precipitation of either of these solute atoms from solid solution and also directly to measure their solubility as a function of temperature. A preliminary study of precipitation using this method has already been presented by the author.3 The purpose of this paper is to present more extensive observations upon the precipitation of N and C in alpha iron, as well as to present observations upon the solubility of N and C in alpha iron which, at least in the lower tempera- ture range, are believed more accurate than those previously obtained by other methods. The interpretation of the observations upon precipitation involves the use of the appropriate equilibrium diagrams. For ease of future reference in this paper the equilibriunl diagrams for the Fe-N and Fe-C systems are accordingly reproduced as Fig 1 and 2, respective]y. Experimental Procedure The specimens were made of West-inghouse Puron iron and consisted of wires 0.025 in. in diam and about 1 ft long. By a wet hydrogen treatment the specimens were purified of C and N. In

Jan 1, 1950

By D. W. Frommer

Processing nonmagnetic taconites by selective flocculation-desliming and flotation requires large volumes of water. If impounded without treatment, these off-process waters require excessively large areas for containment. To discharge the waste water into natural waterways would contribute to stream pollution and likely would not be permitted. In U.S. Bureau of Mines experiments conducted in the Twin Cities Metallurgy Research Center's 900-lb per hr pilot plant, approximately 85% of water requirements for the flotation-based treatment of a Michigan nonmagnetic taconite were met by reclaimed water. Water reclamation of the off-process streams from flotation was accomplished by controlled additions of lime, sodium carbonate, and a synthetic flocculant to reduce turbidities to 51000 ppm equivalent SiO*, while maintaining a Ca(II) content of =16 ppm in the finished effluent. Flotation concentrates of good quality were obtained using the reclaimed water. The cost of chemicals used in water reclamation was approximately equal to the savings in flotation reagents attributed to recycling of the water. Water quality is perhaps as important to flotation as are the reagents used. The character of water is extremely variable, depending on whether the source is a well, lake, or stream, upon the season and temperature, upon prior use, and upon the character of the watershed. All of these factors influence the water hardness and the quantity of other dissolved inorganic salts, turbidity, dissolved and suspended organic matter, dissolved gases, and pH. Frequently, the differences in water quality can measurably influence flotation selectivity, often to the point of spelling success or failure. Water hardness is particularly troublesome in flotation systems employing fatty acids, but other unrecognized constituents may also contribute to peculiarities in flotation behavior. Furthermore, a given water source may be entirely satisfactory in one flotation system, but entirely inappropriate in another. In recent years, society has given more attention than formerly to water use, even in areas where water is plentiful. However, both the demands of a growing population, with increased per capita needs, and also those of industry must be met. As a result, riparian rights must be negotiated with the appropriate government agency or agencies so that consumption of water is often allocated or otherwise controlled. Furthermore, the disposal of off-process industrial and domestic water is coming under the increasing scrutiny of these same governmental units. In these respects, the mineral industry is no exception, so that conservation, water reclamation, and reuse may be expected to assume increasing importance. In 1932, it was stated that, "water reclamation is generally more expensive than the economy in water and reagents resulting from its use. It is employed only if it is urgent to save water."' The economic aspects of this statement may still be true, but the unrestricted use of water is becoming less and less an option of the user. The purpose of this paper is to discuss an investigation conducted by the U.S. Bureau of Mines (USBM), in which water reclamation and treatment were undertaken to develop procedures for, and to assess the effects of, water reuse on the flotation treatment of nonmagnetic taconites. This investigation is a logical extension of previous work described by the author and associates at the Twin Cities Metallurgy Research Center employing selective flocculation-desliming and anionic flotation of silica from low-grade, nonmagnetic iron ores2,8 The requirements for the selective flocculation-de-sliming and the anionic flotation of silica processes determined the direction of the investigation. Previous studies had indicated that pH levels of about 11.0 and 11.8 were required for selective flocculation and flota-tlon, respectively. The calcium content of the water was believed to be of importance in both of these operations, and since Minneapolis tap water with a Ca(I1) content of about 16 ppm had been successfully used, an attempt was made to reclaim the process water at an equal level of dissolved calcium. Additionally, the objective was to nullify or effectively limit the effects of dispersants and fatty acid residuals from prior stages of processing. Last but not least, the system of water treatment had to have the capability of reducing turbidities to workable levels. With about half of the effluent being derived from the selective flocculation-desliming step and carrying about 25,000 ppm of highly dispersed, suspended fines, this last objective appeared formidable at the outset. Procedures for water reclamation were derived, in part, from well-known mineral dressing practices, from past observations and investigations, and from concepts contained in various pertinent publications on water treatment."c These procedures involved: 1) flocculation with lime and poly electrolyte-type flocculants, 2) lime-soda-ash softening, 3) chemical precipitation, and 4) mineral surface adsorption. Control was exercised at various stages of water reclamation by frequent measurements of Ca(11), pH, and turbidity.

Jan 1, 1971

By I. N. Zavarine

THE present paper is a continuation of the work on the relationship between the magnetic and the phase transformations in carbon steels during quenching. An account was given by the author in a previous paper It of an investigation of this relationship during a continuous quench. The experimental difficulties then encountered prevented a conclusive proof of the existence or absence of such a relationship. The conclusions of the previous paper were derived indirectly. The present paper deals with the same problem, but the method of attack is modified. The technique developed by Davenport and Bain2 in their well-known study of austenite transformation at subcritical temperature was adopted in the present investigation. They have shown that the austenite to martensite transformation can be caused to proceed at a relatively slow rate by quenching steels into a hot metal bath. Slowing down of the rate of transformation, together with an observation by Lewis3 that a eutectoid steel can be preserved in the nonmagnetic state by such quench-ing, pointed out a possibility of a study of both transformations under more favorable conditions than previously described. The preliminary steps were connected with the technique of the hot bath quench. To study this phase of the present work a modification of the apparatus previously describe1 was used. The modification con-sisted in the use of a hot quenching bath instead of brine, and in fixing the secondary and the primary coils of the apparatus in a stationary position in and around the bath. Tubular specimens, as in the previous investiga-tion, were heated in a suitable furnace to a temperature in the austenitic.

Jan 1, 1935

By Alan J. Strauss

The Group IV elements germanium, tin, and lead form nine 1:1 compounds with the Group VI elements sulfur, selenium, and tellurium. This paper reviews the properties of the pseudobinary solid solutions formed by these compounds, including the extent of mutual solid solubility, temperature-composition phase diagrams, transport properties, deviations from stoichiometry, optical properties, and energy band structure. Particular emphasis is placed on the Pbl-xSnxTe and Pb1-xSn,Se alloys with rocksalt structure, because of current interest in these malerials for generating and detecting infrared ,radiation. ThE Group IV elements germanium, tin, and lead form nine 1:1 compounds with the Group VI elements sulfur, selenium, and tellurium. Some of the physical and electronic properties of these compounds, including their melting points1-8 and energy gaps,4'9-12 are listed in Table I. Four compounds (SnTe, PbS, PbSe, and PbTe) have the cubic rocksalt (Bl) structure. At room temperature GeTe has a rhombohedra1 structure closely related to the B1 structure, into which it is transformed at about 400°C. Four compounds (GeS, GeSe, SnS, and SnSe) have the orthorhombic B29 structure. In samples which have not been intentionally doped with impurities, the electrical conductivity is due primarily to electrons or holes produced by the ionization of donor or acceptor lattice defects associated with deviations from stoichiometry. Undoped samples of PbS, PbSe, and PbTe may be either n type or p type, depending on whether they contain excess lead or an excess of the Group VI element, respectively, but only p-type samples of the other compounds have been reported. This paper will review the properties of the pseudo-binary solid solutions formed by the nine 1:l compounds. The topics to be considered include the extent of mutual solid solubility, temperature-composition phase diagrams, transport properties, deviations from stoichiometry, optical properties, and band structure. Particular emphasis will be placed on the Pb]-xSnxTe and Pbl-xSnxSe alloys with B1 structure, which are promising materials for generating and detecting infrared radiation in the 8 to 14 µm atmospheric window and beyond. MUTUAL SOLID SOLUBILITY The extent of mutual solid solubility in the pseudo-binary systems has been investigated for fourteen of the eighteen ternary systems (in which the two terminal compounds have a common element) and for nine of the eighteen quaternary systems. In most cases, X-ray diffraction measurements made at room temperature were used to determine the structure and lattice parameter(s) of the phase(s) present in samples of various compositions prepared by freezing from the melt. In many investigations, including the extensive studies of Krebs and co-workers,13'14 the samples were annealed at elevated temperatures before X-ray measurements were made. In some cases, metallographic examination and thermal analysis have also been employed. The results for the ternary and quaternary systems are summarized in Tables II and III, respectively. (The original references should be consulted for the annealing temperatures.) The solubility of one compound in the other is at least 5 mol pct in all cases, and is often much larger. Complete solid solubility has been observed in all systems so far investigated in which the terminal compounds have the same structure, although in the PbS-PbTe system complete solubility is limited to elevated temperatures. Thus complete solid solubility occurs in five systems where both compounds have the B1 structure (SnTe-PbTe, PbS-PbSe, PbS-PbTe at temperatures above 805°C, PbSe-PbTe, and SnTe-PbSe) and in three where both compounds have the B29 structure (GeS-SnS. GeSe-~n~e-, and SnS-SnSe), as well as in two involving GeTe and a compound with B1 structure (GeTe-SnTe and GeTe-PbTe). On the basis of thermal analysis data, complete solid solubility at sufficiently high temperatures has also been reported for the GeSe-GeTe2 and SnS- pbS22 systems. This seems unlikely, however,

Jan 1, 1969

By R. S. Busk

TWO groups of binary alloys were prepared. The first group consisted of those elements relatively soluble in magnesium: Li, Al, Zn, Ga, Ag, Cd, In, Sn, Hg, T1, Pb, and Bi. These are predominately Group B elements. The second group consisted of most of the remaining metallic elements, all of which are relatively insoluble in magnesium: As, Au, Ba, Ca, Ce, Cu, Ir, La, Mn, Ni, Pd, Pt, Rh, Sb, Si, Te, Ti, W, and Zr. These are all Group A, transition, or rare earth elements. Alloys of the first group were prepared by melting sublimed magnesium and elements as pure as obtainable in a graphite crucible, using a standard flux1 for protection. The melts were poured into a small graphite mold to produce 3xY4 in. diam slugs. Spectroscopic analyses were made of each slug. In all cases, the total impurity content was less than 0.05 pct. The slugs were then heat treated 5°C below the eutectic temperature until complete solution, as judged metallographically, was obtained. Filings from the center of the slugs were compacted at room temperature into 1/2 in. long x 1/4 in. diam cylinders, sealed in pyrex under 1/2 atm of purified argon, annealed at the heat-treating temperature for 1 hr, and quenched in water. After the X-ray diffraction patterns were made, each compact was analyzed chemically for the solute. It was found that the Mg-Li alloys reacted during filing to reduce the Li in solid solution. For this series of alloys, the slug was cold worked, annealed in purified argon, the surface dissolved in dilute HC1 to a depth of about 0.005 in., and the X-ray exposure made on the fresh surface. This procedure resulted in sharp, continuous lines and in true values for the lattice parameters. The alloys in the second group were prepared in the same way except that about 1 wt pct of each solute was added. Since the solubility limit was exceeded, not all the solute dissolved. These alloys were analyzed spectroscopically only to confirm the presence of the solute. The diffraction patterns were of the back-reflec-tion type using the characteristic copper radiation* and a flat cassette. Both the film and the specimen were rotated. The specimen temperature was maintained at 25" ± 1°C. The films had sharp lines; their diameters were measured directly to 0.01 cm. The parameters were calculated from the line diameters using a modified version of Cohen's extrapolation method.2 his method effectively eliminates systematic errors such as film shrinkage, inaccurate alignment, and non-precise film-to-speci-men distance. The random errors were minimized

Jan 1, 1951

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE ST. Louis, Mo. PAST PRESIDENTS WILLIAM L. SAUNDERS, NEW YORK, N: Y. L. D. RICKETTS NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 1, 1918

By I. N. Zavarine

THE present paper is a continuation of the work on the relationship between the magnetic and the phase transformations in carbon steels during quenching. An account was given by the author in a previous paper1 of an investigation of this relationship during a continuous quench. The experimental difficulties then encountered prevented a conclusive proof of the existence or absence of such a relationship. The conclusions of the previous paper were derived indirectly. The present paper deals with the same problem, but the method of attack is modified. The technique developed by Davenport and Bain2 in their well-known study of austenite transformation at subcritical temperature was adopted in the present investigation.. They have shown that the austenite to martensite transformation can be caused to proceed at a relatively slow rate by quenching steels into a hot metal bath. Slowing down of the rate of transformation, together with an observation by Lewis3 that a eutectoid steel can be preserved in the nonmagnetic state by such quenching, pointed out a possibility of a study of both transformations .under more favorable conditions than previously described. The preliminary steps were connected with the technique of the hot bath quench. To study this phase of the present work a modification of the apparatus previously described1 was used. The modification consisted in the use of a hot quenching bath instead of brine, and in fixing the secondary and the primary coils of the apparatus in a stationary position in and around the bath. Tubular specimens, as in the previous investigation, were heated in a suitable furnace to a temperature in the austenitic

Jan 1, 1935

The Carol concentrator of the Iron Ore Company of Canada was completed in 1962 and through expansions .has become one of the largest iron ore beneficiation plants in the world with a new feed capacity of 147,000 It (149,350 mt) per day. The Carol ore contains specular hematite and magnetite having a bulk density of 166 lb/cu ft (2660 kg/m3) It is crushed to 7 in. (178 mm) for feeding to autogenous mills for subsequent concentration by Humphrey’s s spirals. Mine ore is loaded into 150 It (152 mt) trucks by 10 cu yd (7.65 m3) shovels and then, at a transfer station, is loaded into side-dump rail cars of 100 It (101.6 mt) capacity. Each train is composed of twenty cars and operates completely automatically as it hauls the ore to the primary crushing station, dumps its loads, and returns to the loading station, a 11.5 mile (18.5 km) round trip. The primary crushing station is composed of two 60 in. by 89 in. gyratories installed in 1962 and 1972. Neither crusher has a grizzly ahead of it. Rail cars are dumped into 300 It (305 mt) pockets which contain splitters for spider protection. What few oversize boulders which do not go through the crushers are drilled and blasted. Liners in the dump pocket and the crushed ore pokcet under the crusher are manganese steel. Both crushers are direct driven by 700 hp (5 15 kw) motors, and the hardness of the ore causes them to pull almost their full power. Crushed ore drops into a 160 It (163 mt) pocket equipped with radioactive level control units for low-intermediate, and high level indications. The low level unit stops and starts the apron feeder, the intermediate unit controls the feeder speed, and a high-level reading stops the car dumping. Apron feeders are 72 in. (1. 82 m) wide and are variable speed from 12 to 52 fpm (3. 7 to 15. 8 mpm). They discharge to 60 in. (1.52 m) wide takeaway belts. Crushing capacity, nominal and maximum for each crusher, is 4, 000 ltph (4. 064 mtph) and 4,600 ltph (4,674 mtph). The crusher eccentric is removed by a conventional cart which travels on removable rails. Mainshaft removal and maintenance is done by 65 and 75/10 ton cranes. Mainshaft replacement requires eight hours, a usual time, but concave replacement requires only sixteen hours, a result of a fine maintenance organization. Mantle relining is done at grade next to a repair track. Zinc is used for backing both mantles and concaves. Both the electrical rooms and the dust collectors are located below grade,

Jan 1, 1978

By G. Meyrick, H. W. Paxton

Observations on the tensile deformation of single crystals of austenitic stainless steels as a function of composition, orientation, and temperature are described and compared with relevant data for other alloys. Crystals free from second phases show sharp yield points followed by Lüders' extensions of up to 20 pct. The critical resolved shear stress is strongly temperature-dependent increasing fourfold between 423O and 77°K. A temperature-independent low rate of hardening follows the Luders' extension and is frequently associated with overshoot. Unlike pure metals where Proficse cross slip usually occurs at the end of linear hardening, this stage is terminated either by cross slip or by slip on the conjugate system, the particular node apparently depending upon initial orientation and composition. The presence of 6 ferrite or martensite has a profound effect upon the magnitude of the yield stress and the shape of the stress-strain curve. In experiments on transgranular-stress corrosion cracking of austenitic stainless steels in boiling magnesium chloride, several workers1'2 have shown that increasing amounts of nickel tend to reduce the probability of specimen failure. Reed and paxton3 have observed in single crystals that the macroscopic crack plane changes as the nickel content is increased. For 18 pct Cr-8 pct Ni, the crack plane is sensibly perpendicular to the applied tensile stress; for 20 pct Cr-20 pct Ni the crack plane is (100). The average crack-propagation speed is much lower in the 20 pct Ni crystals. Furthermore, the possible importance of stack-ing-fault energy in nucleation and growth of cracks has been suggested by a number of workers. While the general effect of variations in stack ing-fault energy appears to be quite well-correlated with observed variations in cracking characteristics, each of the theories assigns rather different detailed reasons for these effects. Because the stack ing-fault energy of austenitic steels is believed to be increased with increasing nickel content, and because some theories of trans-granular cracking postulate a mechanical stage in the crack propagation (rather than relying entirely on electrochemical solution), it appeared to be of some interest to investigate the mechanical properties of a series of different compositions of aus-tenite single crystals. These experiments were considered of intrinsic value in view of the relative scarcity of such data on fcc alloys. This paper gives the results of some experiments carried out with this aim in mind, and suggests various 'other experiments which may be helpful in future studies. The experiments are also of interest in view of the recent interest in obtaining higher-strength steels by control of the imperfection content of austenite; although the compositions are not directly comparable, the results are still probably valuable. Comparisons with experiments on other fcc alloy crystals are made where possible to see how well the general pattern of behavior is understood. EXPERIMENTAL PROCEDURE The alloys used in this investigation were vacuum-melted heats of nominally 20 pet Cr-20 pet Ni-60 pet Fe, 20pctCr-16 pct Ni-64 pct Fe, and 20 pctCr-12 pet Ni-68 pet Fe.* Analyses of the material are

Jan 1, 1964

By H. R. Reed

THE cost of hoisting with electricity in deep mines is a matter about which little information has been published. This paper makes a detailed study of such costs and presents the information in a series of curves. In particular the curves show the cost of hoisting ore from depths varying from 500 to 7000 ft. vertical with the shaft at 70º incline for the double cylindroconical and double cylindrical drums, the hoists operating balanced, using Ward Leonard and Ilgner-Ward Leonard control, and designed for most economical operation from each depth. Operation with the double cylindrical drum is expanded to cover hoisting from 2000 to 5000 ft. using equipment designed for 5000 ft. and from 4000 to 7000 ft. using equipment designed for 7000 ft. Additional curves are included to show the ratings of the necessary hoist motors, energy input per hoisting cycle and size of flywheel in tons for complete equalization when the Ilgner-Ward Leonard system is used, also hoisting efficiency. The duty imposed on a mine-hoist motor fluctuates between wide limits during a cycle of operation. The demand on the motor may vary from a large horsepower value near the beginning of the cycle to a small or a negative demand near the end of the cycle. Such factors as tonnage to be raised per hour, depth of shaft, and balanced or unbalanced operation have such a marked effect on the operating cycle and vary to such an extent for different mines that no set of rules can be formulated for quickly determining the size of the motor required. The duty cycle from which the motor rating is derived must be determined for each case according to the conditions of the proposed hoisting operations. The motor torque required at the cardinal points of the duty cycle is determined from the moment or torque in pound-feet necessary to raise and lower the load and ropes, accelerate and retard the load, ropes, drum, etc. The torque diagram depends largely upon the type of drum. Two different types of drums, if used on exactly the same duty cycles, may have widely different peak demands and capacities of hoist motors. The essential feature in the design of hoisting drums and in determining

Jan 1, 1937

By F. D. Delve

RECENT studies by the author on the reduction of titanium from TiO2-bearing slags have been accompanied by difficulties that are attributed to

Jan 1, 1959

By P. E. Hildebrand, S. W. Mao, C. N. Crain

Hydrologic and economic criteria figure in many obvious ways in water resources development, but they are rarely linked quantitatively, and most of the applications are pertinent only to the case in point. In the preparation of a regional plan for the massive program of groundwater development in the Punjab, situations have arisen where the choice among two or more alternatives was made by analyzing the economic implications of critical hydrologic parameters. Two of these situations which are of more or less general interest are described herein. The first involves basic development policy, i.e., the question of whether groundwater development should be based on the demand of the economy, or the conservation concept of safe yield. The analysis demonstrated that mining of groundwater is clearly indicated for areas where alternative supplies are unavailable. The second problem involves integrating hydrologic data into an economic model to determine optimum well design. From an analysis of all pertinent factors, a simple nomograph was derived for determining well specifications on site during construction. The nomograph can be modified for use in any alluvial terrain. In the Northern Zone of the Indus Plains —the region known as the Punjab- Pakistan is proceeding with integrated development of water and land resources on a scale never before contemplated. The heart of the development plan is a massive program of groundwater development to supplement canal irrigation supplies and to provide drainage relief to the irrigated lands. By the close of 1966 about 2500 high-capacity wells were in operation and furnishing the water supply and drainage requirements for nearly two million irrigated acres. The development plan calls for the construction of a total of nearly 30,000 wells with an aggregate capacity of 120,000 cfs by 1980. This is part of an integrated program eventually affecting over 20 million acres of irrigated lands. [This development program is administered by the West Pakistan Water and Power Development Authority (WAPDA) for which the firm of Tipton and Kalm-bach, Inc. serves as regional planning consultant for the Northern Zone. The material presented herein is derived from planning studies which have been made in connection with the preparation of a regional development plan for the Northern zone.] Pakistan is proceeding with this large-scale development of land and water resources at a time when she is at the threshold of an almost calamitous population growth. Furthermore, she is attempting this development at a time when most of her usable land is already occupied and her food demands exceed current production. Here in the Northern Zone is one of the largest continuous areas of irrigated land in the world, easily irrigated, with a warm, dry climate and ample potential supplies of surface water and groundwater. Except for an abundance of land and labor, other primary resources are lacking or in very limited supply. The agriculture sector so dominates the economy that only by means of an effective development of the agricultural resources can other sectors grow and contribute to a viable economy. Because of inadequate irrigation, farming here has not effectively participated in the twentieth century revolution in agriculture, and only by means of massive development of potential supplies can Pakistan catch up. Capital for any kind of mass development must come from the outside. But the overriding restraint to development of the land and labor resources is the availability of water. ECONOMICS OF DEVELOPMENT The decision-making process in planning regional development becomes increasingly complex as the various sectors of the economy become more highly developed. Where, as in this case, one sector dominates the economy and its development can be considered the key to development of other sectors, the decision-making process is simplified and it is easier to identify alternatives and to trace the prob-

Jan 1, 1970

By H. Majima, E. Peters

The oxidation rates of pyrite, pyrrhotite, chalcopy-rite, chalcocite, covellite, bortzite, galem, sphalerite, and stibnite have beet2 carefully compared at 120 oC, using aqueous phosphate solutions buffered at pH 2.7, 7.1, and 11.2 as zvell as in 1.0 M NaOH and in am~onia-containing solutions. Under conditions of these studies, variations in rates of more than a factor of two per unit of mineral surface area were not observed for different minerals in phosphate buffered solutions, except that pyrrhotite showed unusually rapid oxidation rates In acid solutions, and stibnite showed very rapid oxidation rates in basic solutions. In caustic solutions, oxidation vates varied considerably, the most rapid being for pyrite and the slowest for sphalerite. In the presence of ammonia, only copper minerals showed more vapid oxidation rates in neutral solutions. Oxidation rates were measured on the basis of oxygen-consunzption rates per unit surface area of exposed mineral, using equivalent surface areas, mineral mesh sizes, pulp densities, agitation, and oxygen pressure. OXIDIZING pressure leaching of sulfide minerals has been a commercial method of treating concentrates of nickel, cobalt, and copper for over a decade, and recently the treatment of lead and zinc concentrates,' and even the recovery of sulfur from pyrrhotite by such processes, have been shown to be technically feasible.' Furthermore, the leaching of uranium ores by solutions acidified through the oxidation of pyrite and other sulfides present in the ore is commercially practicable.3 5 The mechanisms by which sulfide minerals oxidize in aqueous environments are not yet understood. However, certain generalities can be made for systems containing no unusual reagents, in which oxygen is the oxidizing agent. In acid solutions the oxidation of many sulfide minerals leads to the formation of elemental sulfur.8'7 The attack of some sulfides, notably galena (PbS), sphalerite (ZnS), and pyrrhotite (FeS), at temperatures below about 110°C leads to an almost quantitative conversion of the sulfide mineral to elemental sulfur according to the equation: On the other hand pyrite (FeS2) and copper minerals do not yield elemental sulfur quantitatively. Not more than 50 pct of the sulfur in pyrite is converted to the elemental form in acid oxidation, and this percentage is lowered by higher temperatures and possibly higher oxygen pressures.8'9 Copper minerals sometimes yield elemental sulfurlo'" but, since sulfur can reduce cupric ions with the precipitation of CuS or Cu2S, it is not likely that elemental sulfur is obtained in good yield. In acid solutions sulfur that is not converted to the elemental form always appears as sulfate salts or sulfuric acid. Sulfide minerals that yield elemental sulfur nearly quantitatively on oxidation in acid solutions tend to generate H2S when the conditions are not oxidizing. However, although it has been suggested that H2S is an essential intermediate in the mechanism of oxidation," this is doubtful because it is oxidized very slowly by oxygen. In basic solutions elemental sulfur is never formed, as predicted by thermodynamical considerations, summarized in the form of a potential-pH diagram in Fig. 1. However in neutral and basic solutions, unstable forms of sulfur not shown in the diagram appear during oxidation, notably thiosulfate (SOT-) and polythi-onates (S,O;-, x = 2 to 6).l3,l4 On prolonged oxidation only hexavalent sulfur as sulfate (SOi-) or sulfamate (NHzSO;) is found to remain, the latter appearing only in the presence of ammonia. The oxidation kinetics for thiosulfate by oxygen are the subject of a separate paper,'5 and were studied on the presumption that thiosulfate may, in fact, be a primary oxidation product for sulfide minerals, at least in neutral and basic solutions. A survey of the literature indicated that comparative oxidation rates for different sulfide minerals were needed to establish the role of the mineral structure and its thermodynamic properties on the oxidation rates and mechanisms.

Jan 1, 1967

By P. A. Boeck

The thermal insulation of high-temperature equipment for industrial purposes is a subject that has not received from engineers and designers the attention its importance deserves. This may be attributed to a number of causes: In the first place, heat flow is a rather difficult factor to measure under the working conditions of an industrial furnace, and until the comparatively recent introduction into certain operations of the heat balance sheet, with a systematic attempt to account for the discrepancies in the totals, the various causes of heat losses and the reasons for their existence were not brought to light. Moreover, such rapid advancement and radical changes in the design of equipment, especially in metallurgical lines, have occurred that practice has apparently outdistanced the theory of design. In the modern tendency toward greatly increased size of units, effort has seemingly been made to utilize the heat of fuel gases without proper attention having been paid to conserving heat energy and confining it strictly to the points of maximum activity. The systematic study of the thermal properties of structural materials, which is being carried out by various government bureaus and also by a number of technical societies, is forming the basis for more intelligent and consistent work in the economically important subject of prevention of fuel waste. Advantages of Insulation It is not necessary in this practical and utilitarian age to enlarge upon the need of prevention of heat losses, but attention may advantageously be called to the other benefits derived simultaneously with the saving in fuel and the increase in thermal capacity of furnaces when properly insulated. In most industrial furnaces—which are here spoken of in the broadest sense, indicating general high-temperature equipment for any purpose and heated in any manner—the operations are generally carried on through a temperature range which has been found by practice to give the most ef-

Jan 1, 1916

The great Bingham Canyon mine and mill complex of Magna, Arthur, and Bonneville had their beginning in a 300 ton per day operation which was commenced by Daniel C. Jackling and his associates in 1904 as the Utah Copper Company. This small plant was quickly expanded to 500 tons per day and eventually Kennecott, the Utah Copper successor, was processing about 45,000 tons per day of ore in each of its mills, Magna and Arthur. Finally, in 1966, the Bonneville crushing and grinding plant was added to produce a slurry of ground ore for flotation concentration at the expanded Magna and Arthur mills. Bonneville is one of the few railroad dump primary gyratory crushing plants constructed in recent years; another which is described in this Monograph is Cuajone in Peru, started in 1976. Two others are the Climax Henderson unit and the Phelps Dodge Tyrone crusher, which has been fed by trucks since it commenced operation. The Bonneville, Henderson, and Cuajone crushers are all located considerable distances from their mines, hence the use of rail haulage. Shovels in the Bingham pit are 8 to 15 yd capacity; and ore is sent to the Bonneville crusher in 100 ton rail cars, where they are dumped, two at a time, by a hydraulically driven rotary dumper. Ahead of the 54 in. by 74 in. Allis-Chalmers gyratory are two grizzlies, each 27 ft 9 in. (8.46 m) long with bar spacings of 4 ½ in. (114 mm) at the feed end and 6 in. (152 mm) at the discharge end. Grizzlies are sloped at 38. Oversize rocks are re- moved from the crusher pocket by an overhead crane, and the crusher produces a nominal 6 in. (152 mm) product. Rail cars are dumped in the direction away from the crusher, and water sprays at the dumper level and in the rock boxes are used for dust suppression. There are two 72 in. by 19 ft (1.83 m by 5. 8 m) apron feeders under the 500 st cr'ushed ore pocket and two more under the grizzly undersize pocket. Feeder speeds are variable up to 72 fpm. Grizzly undersize is dropped to the feeder pocket by three shelves forming a rock ladder attached to the pocket walls. Each pair of apron feeders, one for crushed ore and one for grizzly undersize, discharge to a 72 in. (1. 83 m) picking belt for tramp removal; this belt feeds one of two 54 in. (1.37 m) feeder conveyors which in turn discharge to a 60 in. (1. 52 m) takeaway belt operating at 175-300 fpm (53.3 - 97.4 mpm). Both picking belts are equipped with metal detectors and tramp iron magnets, and a separate 30 in. (760 mm) conveyor is used to remove tramp material.

Jan 1, 1978

By F. F. Sharpless

IT IS to be. read in Western newspapers that Hon. S. S. Arentz who introduced H. R. 7736, 'has repudiated it, that certain mining men in Utah and Nevada say that they will fight it "until hell freezes over." It is to be read in more temperate expressions that "the eminent engineers who drafted the bill exhibited a woeful misunderstanding of Western sentiment and opinion." Furthermore, it is averred that there is no need for revising the mining laws, that we have got along pretty well under the existing laws, etc. Such statements are so surprising in view of, the facts of record that it is hard to believe that the oppon-ents of the bill, who make them, are well informed, or that they represent general opinion. It will be useful therefore to review the record of the case. Our GOOD OLD MINING LAWS We have indeed got on well in our mining industry, but that has not been by virtue; of our good old mining laws. Rather has it been in spite of them. Ever since the Richmond-Eureka case in the '70s there has been litigation arising from them. Early in the '80s the apex law was discarded by common consent in Leadville, and it has been ignored in every district of flat-lying ore deposits ever since. The great mines of the present time are the porphyry copper deposits. At Bingham, Ely, Ray, Chino, Miami, Clifton and Ajo the apex law is simply inapplicable. A law that is inapplicable to the modern conditions of mining may be an old law, but it can not properly be characterized as a "good" law. For such reasons, and many others of weight, there has been a demand for revision of the mining laws; dat-ing ,back at least three decades. President Roosevelt appointed a commission to advise respecting the question, said commission comprising. Dr. James Douglas (chairman), John Hays Hammond, D: M. Barringer, Walter B: Devereux, and William B. Mather. This committee was unanimous on the following points:

Jan 1, 1922

By D. W. Brunton

The maps of our large mines are usually prepared with the greatest care; and it is somewhat singular that, in compari~on with the great amount of time and money spent in surveying and platting, 80 little actual use is made of them. Almost the only purpose for which a completed survey-map is afterwards consulted, is the determination of the relative position of the different workings to each other, and to the boundary-lines of the property. After the completion of such a map, it should be made the beginning of another, and in most cases a far more important undertaking, namely, its utilization as a starting-point for a complete inventory of the company's underground possessions. The ordinary mine-survey map, being nothing but a record of what has been done, is, in one sense, only ancient history. To increase its value, such additions should be made as will render it a complete statement of the amount and value of ore in sight at any particular time, and a guide for future development. Comparatively little extra labor is involved in this undertaking, since the larger and most expensive part of the work has already been completed when the mine has been surveyed and mapped. The necessary additions consist in working out and platting on the maps, the geology of the mine as exposed in the workings, in such a manner that the geological survey may be of daily use in the development and operation of the mine. The first step towards the production of a geological map consists in tracing individual level-sheets from the general or composite map. The area to be included, outside of the property in question, will depend very much on local conditions; but, for geological, legal and commercial reasons, it should be extended as far as reasonably practicable. The scale to be

Jan 1, 1906

M. A. SCHEIL (presented in written form).-Messrs. Franks, Binder and Brown are to be congratulated for the splendid paper that they have presented. We have not made stress-corrosion tests with the cantilever type of specimen which they describe, but it is well known that a uniform stress across a certain portion of the gage length can be obtained with such a specimen. This type of specimen would appear quite expensive to make. The writer was interested in the authors' remarks concerning the erratic results obtained from specimens with cold sheared edges. It might be of general interest to record here some results we obtained with annealed sheet N4 in. thick of type 304 stainless steel. After specimen No. 4 had been exposed 65 days without any signs of cracking we applied a strain gage and measured the stress as 25,810 psi. This value is considerably higher than we find for some of the type 347 stainless alloys and suggests that 18-8 might have a different threshold value for stress-corrosion cracking than 18-8 columbium. The test data in Table I show that strain caused by cold work or polishing has a pronounced effect in lowering the time for failure. In my paper I showed values of 8 hr. and 17 hr. to stress crack cold-rolled type 301 stainless steel stressed beyond the yield point. This steel contained 8.31

Jan 1, 1945

By John K. Berry

The handling and disposal of mine water is a much larger problem than is apparent at first glance. Many more tons of water are removed from underground coal mines in the United States each year than tons of coal. In the old anthracite region of Pennsylvania the average was around 36 tons of water for each ton of coal produced, while bituminous mines average about 5 or 6 tons. Water encountered in the mines may come from several different sources. In mines with light cover and without a firm solid roof, rain may seep directly into the workings. This is also true of deeper mines where pillar falls have broken the roof to the surface. The top under creek beds is generally faulty and broken and water comes through the roof at this point. Deeper mines are frequently overlain by water-bearing strata and pillar breaks and faulty top can admit water, while places running to the dip often accumulate water which comes up through the bottom. In all cases, except where passing under a year-round stream, the amount of water made is usually pretty closely connected with the weather-more water during the rainy season, less in the dry, although the deeper a mine is, the longer it is before changes in rainfall are felt. Another source of water, not connected with seasonal changes, is old water, accumulated over a period of time in abandoned workings of the mine, or from flooded adjacent mines. Here the water filters through the coal, the parting, or comes up through cracks in the bottom. Similarly, core holes, drilled to prospect the property, can be a source of unexpected water and modern safety requirements for dust-allaying water at the working faces can turn a pleasant working area into a soggy mess over the weekend if the system is not properly maintained and operated. Why water must be disposed is obvious. If allowed to accumulate, sooner or later the mine will fill up, or at least it will become "roofed" in low places, blocking ventilation and traffic. But it must also be controlled long before it reaches this stage. Water standing in the working faces is unpleasant for the men who have to work there and also cuts down on the amount of work they can do, as they must watch where they step all the time. It is especially unpleasant in low coal.

Jan 1, 1973

By Raymond L. Orr

Calorimetric measurements have been made of the heats of solution of gold axd indium in a number of liquid tin-rich alloys at a temperature of 705°K. Relative partial molar enthalpies of gold were determined as a function of dilute solute concentrations in the binary Au-Sn system and in the ternary Ag-Au-Sn and Au-In-Sn systems. Relative partial molar enthalpies of indium were determined for dilute solute concentrations in the In-Sn, Ag-In-Sn, and Au-In-Sn systems. Through application of the interaction coefficient concept introduced by Wagner and more recently extended by Lupis and Elliott, ualues have been obtained for the Au-Au, Ag-Au, Au-In, and Ag-In enthalpy interaction coefficients in liquid tin. The results are inte@reted in terms of solute atom distributions through comparisons with the predictions of dilute solution models. INTEREST in the thermodynamic behavior of dilute liquid alloys stems from two primary sources. From a practical viewpoint, it is often of importance to know or to be able to predict the effect that one solute will have on the thermodynamic properties of the other solutes in a multicomponent system. Studies of dilute solutions may also be rewarding from the theoretical point of view. Some of the difficulties arising from the more complex interactions possible in concentrated solutions are avoided, leading to easier interpretations in terms of solutibn models and bonding energies. For example, in a binary alloy the limiting values of the partial molar properties of the solute represent the case for which each solute atom is completely surrounded by atoms of the solvent, and no other interactions are possible. Solute - solute interactions in dilute solutions are conveniently treated by the interaction coefficient concept of ~agner.' Using a Taylor series expansion for the logarithm of the activity coefficient, In yi, of a component, i, in a solution consisting of dilute solutes with atomic fractions xi, xj, xk, and so forth, in a solvent, s, and neglecting the second- and higher-order terms, Wagner obtained the expression: Where Wheyz is the limiting value of yi in the pure solvent, s. The interaction coefficients, e:, e:, and so forth, are defined by: Since yi is related to the excess partial molar Gibbs energy of component i by GfS = RT In yi, E: is re- ferred to as the "Gibbs energy interaction coefficient". Lupis and Elliott~ extended Wagner's treatment to the relative partial molar enthalpy, aHi, and the excess partial molar entropy, qS. They defined the enthalpy interaction coefficient as: and the excess entropy interaction coefficient as: leading to expressions similar to Eq. [I]: The three interaction coefficients are related by? The self-interaction coefficients, E:,qf,and at, represent the effects of interactions between atoms of component i in solvent s; €3, q{, and 03 represent the effects of interactions between i and j atoms in solvent s; and so forth. From the Maxwell-type relationships between partial molar quantities, it may be shown1'2 that e{ = ej, r}\ =tj), and ai - crj. Experimental determination of the interaction coefficients requires extremely precise measurements of the appropriate properties as functions of solute concentrations in very dilute regions. Few such data exist for intermetallic alloys, even for binary systems, because of experimental difficulties. This is especially true with respect to enthalpy data for dilute alloys. Examination of the compilation of data for binary alloys by Hultgren, Orr, Anderson, and ~elle~~ reveals that determination of limiting values of &fii(Xi =0) often requires extrapolations to infinite dilution from xi = 0.05 to 0.10. Direct measurements of the enthalpy interaction coefficient, qi, for multicomponent systems are virtually nonexistent. This quantity is of course subject to direct calorimetric measurement, but its determination must be limited to cases where high experimental precision is possible. The heats of solution of gold and indium in liquid tin can be measured with relatively high precision, which makes determinations of 73 involving these metals as the measured solutes experimentally attractive. This paper presents the results of such determinations of the Au-Au, Ag-Au, Au-In, and Ag-In enthalpy interaction coefficients in liquid tin. MODELS FOR DILUTE-SOLUTION BEHAVIOR Random-Solution Behavior. In the quasi-chemical treatment of solutions, only nearest-neighbor bonds are considered, and a composition independent value

Jan 1, 1967