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By Byung-Su Kim

The recovery of elemental sulfur from sulfur dioxide has important implications for coal- burning power plants and nonferrous metal smelters. The thermodynamic analyses showed calcium sulfide to be suitable for recovering elemental sulfur from sulfur dioxide. At the temperatures of 1023 to 1153 K under the sulfur dioxide partial pressures of 5.0 - 60.1 kPa, elemental sulfur from sulfur dioxide was recovered by its reaction with calcium sulfide. The kinetics of this reaction were measured at the same conditions using a thermogravimetric analysis technique. This reaction scheme would be attractive if the coproduct calcium sulfate can be reduced to regenerate calcium sulfide. This is the topic of another paper. At 1073 K under the sulfur dioxide partial pressure of 25.8 kPa, 48 % of the original calcium sulfide was converted to calcium sulfate in an hour. Under the same conditions, 55 % of the regenerated calcium sulfide were converted to calcium sulfate in an hour even after third regeneration.

Jan 1, 1999

By Azizan A

The paper presents the results of chlorination of beneficiated rutile and struverite (a tantalum-niobium bearing, rutile type mineral in alluvial tin deposits) by carbon tetrachloride in the range 623¦K to 773¦K. The chlorination of rutile could reach near completion at 773¦K in about an hour. The process was found to follow a modified pore-blocking law, with the basic reaction constant K varying inversely with the particle size, so that the basic process may be a reaction under interface control (chemical control) as modified by the pore blocking effect. The pore blocking factors for the various experimental conditions were evaluated and the predicted curves showed good agreement with the experimental data. In the case of Struverite, a similar kinetic behaviour was observed but the chlorination process was relatively slower than that in rutile. It was also observed that in struverite, niobium got chlorinated faster, followed by tantalum and titanium, indicating a good possibility of separation by chlorination using carbon tetrachloride.

Jan 1, 1996

By M. J. Ferrante

High-temperature enthalpy and X-ray powder diffraction studies were con-ducted on aluminum sulfide (Al2S3) as part of the Bureau of Mines effort to provide new data for the advancement of mineral technology consistent with environmental preservation and energy conservation. Relative enthalpies were measured with a copper-block calorimeter from 298.15 to 879 K. Tabulated values are listed for heat capacity, entropy, Gibbs energy functions, and relative enthalpies between 298.15 and 900 K. Enthalpies were also given in the form of an equation and combined with published data to calculate standard enthalpies of formation and Gibbs energies of formation for reactions of crystalline aluminum with crystalline, liquid, and gaseous diatomic sulfur. The hexagonal, rhombohedral, and cubic phases of Al2S3 were identified by X-ray powder diffraction studies, and the indexed powder diffraction patterns of these phases are given. The hexagonal phase was stable to about 852 K and was partially converted to the rhombohedral form at 879 K.

Jan 1, 1981

In presenting the paper, Mr. Davey quoted subsequent work which yielded k, at miscibility gap composition, of about 50.We can confirm the authors' findings for k j 300 for nearly pure tin to k 2 about 50 (average 47) from our own experience and from repeat experiments on materials described by the authors, at 1150°C.During laboratory-scale trials we also found confirmation of the authors' view that it is more difficult to reach equilibrium from the high-iron (in slag) side than from the low-iron (high-tin) side. Working from the high-iron side tends to yield results that point to a high k, leading to optimistic values of iron/tin ratio in final slags. In practice the smelting of a tin concentrate, unless it is exceptionally low in tin and high in iron, must yield initially a tin-rich slag and therefore equilibrium can only be reached from the "tinny" side.

Jan 1, 1967

By D. G. Foot

The Bureau of Mines investigated energy-efficient methods for recovering potash values from process and waste brines. Laboratory pan evaporation of four chloride brines produced crude salts containing predominantly sylvite, halite, and carnallite. Six sulfate-chloride brines produced crude salts containing primarily schoenite, kainite, leonite, sylvite, carnallite, and halite. Potash grades ranged from 7.2 to 22.2 pct K20, and recoveries from 84 to 99 pct. Sylvite flotation from chloride evaporites, with amine collector, recovered 90 to 97 pct of the potash in a concentrate containing 54.3 to 60.3 pct K20. Fatty acid flotation of the high-sulfate evaporite recovered 78 pct of the sulfate minerals in a 27.8-pct-K20 concentrate. Flotation of the chloride minerals with amine collector recovered 80 pct of the potash in a 59.7-pct-K20 concentrate. Solar evaporation of 10,000 gal of brine recovered 99 pct of the potash in a crude evaporite containing 24.5 pct schoenite and 20 pct sylvite. Continuous flotation in a 100-lb/h process research unit recovered over 95 pct of the potash in schoenite and sylvite concentrates containing 28.0 and 62.3 pct K20, respectively. An economic evaluation suggested a rate of return of 3 pct for a new facility and 9 pct if. the process is adapted to fit an existing plant.

Jan 1, 1984

By Gregory B. Poole

Data collected with 16K sampling at a variety of blasting sites (construction, quarry, trench, mine) and monitoring locations was analyzed to effectively show the 16 different 1K readings all based on the microsecond that the unit went into monitoring.

Feb 1, 2020

By A. Lyndon

Combined loading tests are described to three 1.2m diameter cast-in-place bored piles, 12m long, formed through cohesionless material into stiff sandy clay. The maximum loading conditions imposed at the pile head were 2400 kN vertical, 160 kN lateral and 3120 kNm in bending, corresponding to twice the working load. Settlements, deflections and rotations of the piles were measured during the tests. An interesting finding which the test results suggest is that, in the short term, vertical movement had apparently no marked effect on lateral performance for piles working at design loadings.

Jan 1, 1989

By David Hoadley, Kenneth R. Maser, Ashok B. Boghani, James E. Billar, D. Randolph Berry, Mackenzie Burnett, Robert H. Trent

B 5 Typical Input - Ouput The input data and the printed output for case 8 of the simulations (Section A.4) are shown in the following. The branch, level, node, miner and hoist numbers mentioned in the input and output refer to the corresponding numbers in Figures A-3, A-4 and A -5, The plotted ouput, already shown in Figures A,11 and A-12 is not repeated. The input data shown does not include the following simulation parameters: DTIME = 0, 05 FTIME = 130 JPLT(K) = 1 for K = 1, 2, 3 0 for other K TPLSTR = 0.0 TPLST P = 130

Jan 1, 1976

By E. G. King

Enthalpy increments above 298 K were determined by the Bureau of Mines for magnesium hydroxide (brucite) between 298 and 900 K by the method of drop calorimetry. Enthalpy and entropy increments from this investigation were combined with other data from the literature to calculate absolute entropies, free energy functions, enthalpies of formation, Gibbs energies of formation, and equilibrium constants of formation from 298 to 900 K. By the use of published PVT data for water, a table of Gibbs energy change was given for the reaction MgO (periclase) + H2O = Mg(OH)2 (brucite) as a function of temperature and pressure.

Jan 1, 1975

By R. Schuhmann

Distribution of material in the fine sizes of a comminution product generally is well represented by the empirical equation [ ] in which y = cumulative percent finer, x = particle size, = distribution modulus, and k = size modulus. Charles found that the energy consumption in comminution is usefully expressed by another empirical relation, [ ] in which E = energy input per unit volume of material, A = a constant, k = size modulus based on Eq. 1, and n = a constant; ( 1-n) is the slope of a plot of log E vs log k. Holmes has presented energy equations similar to Eq. 2.

Jan 2, 1960

By Wei Li

Carbothermic reduction kinetics of ilmenite concentrates catalyzed by sodium chloride and the microwaveabsorbing characteristics of the resulting reductive products were investigated; the reduction degree of the reaction was deduced to be . Results show that the activation energy is 112.03 kJ/ mol when NaCl is used as catalyst, lowering to 23.11 kJ/mol for the final stage activation energy using sodium silicate as a catalyst. There is a sharp change in the reduction rate constant at temperature ranges of 1,223 K - 1,273 K and 1,373 K - 1,423 K. Microwave-absorbing characteristics of reductive products were measured by the method of microwave cavity perturbation. It was found that microwave-absorbing characteristics of reductive products obtained at temperatures of 850° C, 1,100° C and 1,150° C show sharp changes. Through X-ray diffraction (XRD) characterization, it was found that the formation of Fe occurs at a temperature of 900° C during the reduction process of ilmenite concentrate. The decrease of FeTiO3 content results in the decrease of microwave-absorbing characteristics, while the increase of Fe content gives rise to the increase of microwaveabsorbing characteristics. The decreasing trend is larger than the increasing trend, which is the main reason for the sharp changes in the microwave-absorbing characteristics of reductive products at the temperature range of 850° C to 900° C. When the decreasing trend is smaller than the increasing trend, it becomes the main reason for the sharp changes in microwave-absorbing characteristics of reductive products at the temperature range of 1,100° C to 1,150° C. Kinetics results show that sodium chloride can catalyze the carbothermic reduction of ilmenite, and microwave-absorbing characteristic investigations indicate that microwave heating can be used in the carbothermic reduction of ilmenite.

By D. Duff, R. A. Cameron, R. E. S. Whitehead

"Although CO2 content indicates the quantity of carbonate in altered basaltic rocks, it does not measure the intensity of carbonate alteration. CO/CaO molar ratios, which correlate with the relative proportions of individual carbonate species, measure the intensity of carbonatization. The Timmins gold area is characterized by pervasive carbonatization, with large parts of the area having CO2/Ca0 ratios greater than 1, representing C02 supersaturation relative to CaO. The mean CO/CaO in the Timmins area is0.67, compared to 0.13 in two unmineralized control areas. Regional arsenic abundance in the gold-bearing area averages 4 ppm, compared to about 0.35 ppm for the non-mineralized areas. The average K20 and Rb abundances are similar in both mineralized and non-mineralized areas (about 0.30% K20 and 7 ppm Rb). Although there appears to have been no wholesale addition of K20 or Rb to the mineralized area from external sources, these alkalies have been redistributed internally, resulting in patterns of local enrichments and depletions which are characteristic of the mineralized region.Local target anomalies associated with known mineralization consist of CO2/Ca0 > 1.5, As > 10 ppm and K2O > 0. 75%. Not all three values are required to be above these thresholds in order to identify local targets. Combinations such as C02/ Ca0 > 1.5 and As < 10 ppm, CO/CaO 1-1.5 and K20 > 0. 75 or K20 > 0. 75 and As > 10 ppm are equally indicative of proximity to gold occurrences. Gold-bearing veins and associated anomalies occur throughout the entire volcanic succession and are not restricted to any particular stratigraphic horizons. The levels and patterns of alteration and occurrences of gold appear to be structurally controlled."

Jan 1, 1982

By Jie Li, Shu Yang, Zhongliang Tian, Xun Hu, Yanqing Lai, Shaolong Ye

The electrolyte with high solubility for Sc2O3 is very important to realize the preparation of Al-Sc master alloy by the molten salt electrolysis. The solubility of Sc2O3 in K3AlF6- Na3AlF6-AlF3 melts is determined with an isothermal saturation method. The effects of KR (ratio of K3AlF6 to K3AlF6+Na3AlF6), the concentration of Al2O3 and the temperature on the solubility of Sc2O3 are also investigated. The results indicate that the solubility of Sc2O3 in K3AlF6-Na3AlF6-AlF3 melts increases and then decreases with KR increasing. The solubility of Sc2O3 is 8.698 wt.% in (K3AlF6-Na3AlF6)(KR25%)-18 wt.%AlF3 melts at 940℃. The effect of temperature on the solubility is little, and the values are 8.575 wt.% and 8.762 wt.% respectively when the temperatures are 930℃ and 960℃. The solubility decreases from 8.809 wt.% to 7.058 wt.% as the concentration of Al2O3 increases from 1.5 wt.% to 5.0 wt.% at 950℃.

Jan 1, 2015

By Mehmet Guney

Acceptable working environments in production areas and service openings of Mindola Mine on the Zambian Copperbelt are currently maintained with an air volume of 1000 cu m/s and 5500 kW of installed refrigeration plant. It is planned to open up additional deeper areas of the mine for future production. Temperature problems, due to the increased heat generation from various underground sources are recognized as potential hinderances to efficient production. Findings of research work indicate that the present total heat output of 58,000 kW will inevitably increase by the second five-year period of the deepening program to 52,900 kW and by the fifth five-year period to 63,200 k\V. Estimated WB and DB temperatures, representing the total heat generation in stope faces and development headings during these same periods are predicted to increase to 53.8/40.2 °C and 35.5/42.1 °C respectively. It is assessed that the nominal refrigeration capacity needs to be progressively increased to 9,000 k\V This paper deals with the thermal problems in Mindola Mine, current heat output calculations, future prediction of temperature and heat load in various working areas of the mine with the application of computer programs and interpretation of data in order to find the effective solution for long-term planning of ventilation and air-conditioning.

Jan 1, 1982

By James A. McGurty, Walter J. Koshuba, Samuel G. Gordon, Gilbert E. Klein

ONE of the problems encountered in working with metals at elevated temperatures is the instability resulting from solid-solid diffusion at a -common interface. A determination of the nature and magnitude of this phenomenon is important as related to structural properties of materials potentially useful in high temperature applications. This paper is concerned with the solid state reaction between metallic molybdenum and beryllium at elevated temperatures. In the older literature,&apos; there were no known inter-metallic compounds in the system Be-Mo, and the list of beryllium intermetallics was rather small. New compilations have been made by Smithells&apos; and by Taylor," listing all the then known beryllium binary intermetallic compounds. In addition to these, some new phases have just been reported&apos; in the system Zr-Be and some further compounds of the type MBe,3 have been currently described. The new compound MoBe,, which is reported upon in this paper differs considerably from the above mentioned MBe13 group, and will be discussed in detail. Diffusion Study The nature of the solid diffusion at a Mo-Be interface was determined by sealing a beryllium rod inside a molybdenum capsule and heating at a temperature of 2000°F&apos; (1100°C) for 100 hr. The molybdenum capsule was fabricated from % in. swaged molybdenum rod obtained from the Fansteel Metallurgical Corp. The beryllium metal was obtained from Brush Beryllium Corp. as % in. cold-drawn rod. This material assayed 99.05 pct Be and contained as impurities 0.84 pct Be0 and 0.19 pct Be2C. The molybdenum capsule was % in. in OD x 2 in. long. To prepare it, a concentric hole 0.250 in. + 0.001 in. — 0.000 in. in diam was drilled and reamed in one end of a rod to a depth of 1 &apos;/2 in. A 1 in. long beryllium rod 0.250 in. + 0.000 in. — 0.001 in. in diam was pressed into the molybdenum capsule. A molybdenum plug 0.250 in. & 0.001 in. in diam x 9/16 in. long was then used to seal the beryllium in place. The capsules were placed in a molybdenum resistance furnace and heated at 2000°F (1100°C) for 100 hr. On cooling, the specimens were cut and examined metallographically. Fig. 1 is a micrograph at X250 of a portion of the interface from a transverse section. Four phases are clearly visible, with sharp boundaries between them. Phase A is a section of the molybdenum capsule; phase B is the molybdenum-rich intermetallic composition MoBe,; phase C is the beryllium-rich compound MoBe13; phase D is a section of the beryllium rod. All phases were readily polished to a high metallic luster. The molybdenum-rich phase has a pronounced purple tinge. The beryllium-rich phase (C) is only slightly darker than the pure beryllium. In polarized reflected light, however, it shows striking birefringence between crossed Nicols. The black areas represent flaws or pores in the material. A small representative amount of powdered material from each of the phases was carefully removed under the microscope with a diamond tool. With this material, powder X-ray diffraction patterns were made for the identification of each phase. Table I lists the X-ray diffraction spacings and corresponding indexes characteristic of these intermetallic compounds. Synthesis To establish the formula of the new intermetallic compound, X-ray diffraction patterns were made from powders obtained from solidified melts of known composition. The syntheses were made by

Jan 1, 1952

By Franz Xaver Spachtholz

K+S Potash Canada GP (KSPC) is part of the K+S Group, one of the world&apos;s leading suppliers of standard and speciality fertilizers and the world?s leading salt producer. Globally, the K+S Group employs more than 14,000 people. The K+S share ? the only commodities stock on the German DAX index ? is listed on all German stock exchanges. Legacy, K+S Group?s most recent project, is a greenfield solution potash mine and production plant in the south of the Canadian province of Saskatchewan. Therefore K+S will invest 4.1 billion CAD. With the expected commissioning in summer 2016 the company projects to reach the two-million-tonnesmark of production capacity at the new location at the end of 2017. The gradual extension of the annual capacity to 2.86 million tonnes of potash products will occur afterwards. The Legacy Project will be Saskatchewan?s first new potash mine in nearly 40 years. It is one of the world?s most economically attractive potash projects. The project will contribute to Saskatchewan?s growth through employment, tax revenues, utilities, and local partnerships. Located in the southern part of Saskatchewan?s potash belt, the deposit is too deep to be mined conventionally due to rock mechanics and high temperatures. Legacy will therefore be a solution mine. Proximity to the urban centers of Regina and Moose Jaw is imperative, with their workforces and economy, as well as access to highways, railway carriers and, most importantly in the prairies, water. With this unique project setting and mining technique, we are convinced that Legacy will be one of the most efficient, economic, sustainable and safe potash facilities in the world.

Aug 1, 2013

By W. D. Robertson, H. Conrad

Incremental loading creep tests were conducted on magnesium single crystals in tension and direct shear at 78" to 364°K. The creep curves were transient in character and exhibited waves or periodic behavior, especially at small strain rates. A sigmoidal curve was frequently observed following the first occurrence of creep. Repeating of segments of a master curve was observed at all temperatures indicating that identical conditions for flow existed after various stress and strain histories. A logarithmic time law with a independent of temperature whs observed at 78°and 203°K, while a parabolic (power) time law was approximated at 298" and 364°K. The departure from logarithmic creep was attributed. to recovery, which was observed at 298" and 364°K. The temperature independence of a represents a serious objection to the theory of low-temperature creep based on the intersection of dislocations. IN a previous paper1 it was found that the initial creep rate of magnesium single crystals, y, following the addition of a stress increment in the range of 78° to 364°K, was given by the expression ? = C exp - {?H/kT} exp {B(T- h ?)} [1] where B and C are constants, AH is an activation energy of 10,300 cal per mole, T is the applied stress, ? is the total strain at the time of the stress increment, and h is the strain-hardening coefficient. The present paper is concerned with certain other features of creep behavior observed during this investigation; namely, the form of the creep curves, time laws for creep, recovery, and microscopic and X-ray observations of the structural manifestations of the creep flow. To the authors&apos; knowledge the temperature range of the present investigation (0.08 Tm to 0.40 Tm) has not been covered previ-

Jan 1, 1959

By Yang Cui

The partition ratios of lead and antimony between liquid copper and MgO saturated Cu2O-CaO-SiO2 flux and their removal rate by using CuCl-based flux were measured between 1423 K and 1573 K. For Cu2O-CaO-MgOsatd. system, the partition ratio of Pb was not affected by the CaO content at 1573 K, while that of Sb increased with increasing CaO content. The maximum value was 160 at the CaO content of 15.4 mass%. For Cu2O-CaO-SiO2-MgOsatd. system, two liquid phases of calcium silicate rich phase and Cu2O rich phase were observed at 1573 K. The partition ratios of Pb and Sb between calcium silicate phase and liquid copper depended on the (mass%CaO)/(mass%SiO2) ratio. When the (mass%CaO)/(mass%SiO2) ratio increased, the partition ratio of Pb decreased and that of Sb increased. For CuCl flux, the removal of Pb from liquid copper at 1473 K was possible, and a relatively low concentration could be reached. Moreover, most of lead as metal chloride vaporized from flux. For the removal of antimony from liquid copper at 1423 K, it was difficult by CuCl flux, and the removal ratio was approximately 20% for 75 minutes. However, the addition of CaO or Na2CO3 to CuCl flux was relatively effective to remove antimony, and the concentration of antimony reached a relatively low value after 15 minutes by using CuCl-CaO or CuCl-Na2CO3fluxes. Moreover, a small part of antimony as chloride vaporized from flux.

Jan 1, 2006

By Kenji Monden

"The creep properties of tin-based lead-free solders Sn3.0AgO.5Cu and SnO.7CuO.1Ni and lead solders 10Sn90Pb and 50Sn50Pb were investigated at temperatures between 298 K and 398 K. The creep-rupture time decreases with increasing initial stress and temperature. The analysis of the solder creep curves uses the Omega method. The creep rate E is expressed by following formula: 1n ? = 1n ?0 + O?, where ?0 and O are experimentally determined. The parameter Eo increases with increasing initial stress and temperature. The activation energy for ?0 is 108 kJ/mol for Sn3.0Ag0.5Cu and 85 kJ/mol for SnO.7CuO.INi. These energies suggest that the lattice diffusion of tin is dominant. The energy of 1OSn90Pb is 37 kJ/mol and that of 50Sn50Pb is 67 kJ/mol. The creep-rupture time is calculated using the parameters, ?0 and O. The calculated creep-rupture time is in good agreement with the measured creep-rupture time.IntroductionLead-free solders are replacing traditional tin-lead solders as part of a major trend toward protecting the earth's environment. However, tin-lead solders are currently needed for a high temperature range. Since creep is the main deformation mechanism in solder joints, understanding a constitutive model for creep and examining creep data for solders are important for reliability analysis of solder joints. A steady-state creep equation is generally used for solders [1,2], and it supposes that the secondary-, or minimum-, creep rate occupies the entire creep life. However, secondary creep accounts for only about 15 % to 40 % of the entire creep life for solders. Thus, the equation may not represent the entire creep life."

Jan 1, 2012

By J. D. Mooney, D. L. Darnell

The accompanying chart has been drawn for the convenicncc of engineers as&apos; a means of quickly determining the correct number of plies of conveyor belts operating under specific conditions. The calculations are based on the average safe strength (factor of safety, 15) of the various standard rubber conveyor belts. The calculations assume maximum loading conditions; that is, the belt is considered as carrying the greatest load that it will handle without spillage at ordinary belt speeds. This not only produces the most economical operating conditions, but also the maximum tension in the belt. The chart is a graphical representation of the formula: Where, p = kg W(L + 10H) p = the correct number of plies k = a constant, depending on the type of drive g = the weight &apos;in pounds per cubic foot of material handled W = the width of the belt in inches L = the length of the belt in feet (approximately twice the center distance). H = the difference in elevation between the head and tail pulleys, in feet. For a simple drive, with a bare pulley, k = ----1 For a simple drive, with a rubber-lagged pulley, k = --- 1 For a, tandem drivej with bare pulleys, k = 1/375.000 For a tandem drive, with rubber-lagged pulleys, k = ---1 The chart is drawn for a simple drive, with a bare pulley (k — =e)1 ; therefore, the number of plies obtained from the chart should be multi- &apos; plied by the factor 0.83 or 5/6 for simple, lagged drive; the factor 0.67 or 2/3 for tandem, bare; and the factor 0.55 or 11/20, for tandem, lagged. The formula p = kgW(L + 10H) has been developed mathematically

Jan 1, 1916