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By F. M. Drake

IN this paper I propose to describe a plant which I lately erected for the Mount Stewart Lead and Silver Mining Company of Leadville, New South Wales, Australia. I am indebted to Mr. W. F. Burrow, of the Roads and Bridges Department of New South Wales, for the drawings which I present herewith, he having reduced and assembled them from my working drawings. The works are situated on a sloping hillside, but sufficient fall mas not available to have the furnace as high as desired, and at the same time have a good slag-tip and sufficient depth on the ore-floor to store a supply of ore and flux. The ore is delivered on the ore-floor (which is of the full width of the furnace-building, and extends back into the hill a distance of seventy feet) by two overhead tramways (not shown in the drawings), one on each side. From these the ore-mixture is made up on one side, while it is being taken to the furnace from the other. The fuel also is delivered by these tramways. The fluxes are stored on a floor directly at the back of the boiler and engine room. The boiler is a Cornish one, 6 feet 6 inches in diameter by 25 feet long. Its flue is 3 feet in diameter and does not contain Galloway tubes. The working-pressure is 40 pounds per square inch. The engine is an old-fashioned one, with a cylinder 16 inches in diameter by 30 inches stroke. Both the engine and boiler had been part of the plant of a defunct gold-mining company. They were very extravagant in their consumption of fuel. The blast is supplied by a rotary pressure-blower, of the Baker pattern, made in England. It has a capacity of 60 cubic feet per revolution. The furnace, which has several novel points, is shown in detail in Figs. 1, 2 and 3. The jackets are of cast-iron, and are higher than those generally used. They are different, too, in that the feedwater inlet is through a hand-hole plate near the bottom. This plate is fastened to the jacket by stud-bolts, and is provided with a baffleplate, which prevents the cooling water from impinging against the inner plate. The outlet is by a goose-neck at the top, which dis-

Jan 1, 1893

By H. J. Booss

THERE is a considerable lack of data on thermody-namic properties of hard-metal alloys. Only two papers 1,2 give mean values of specific heat in an unknown temperature range; more recently the author3 gave more detailed information on the mean specific heat in the temperature range from 20" to -190° C.. EXPERIMENTAL 1) Method—Experimental procedure followed a method suggested by Oelsen and coworkers.4 It requires samples of some 100 g of weight. The time of measurement is prolonged (30 min) but one run allows the measurement of a lot of positions on the heat-content curve. Results4 were found to be in good agreement with "best values" given by kelley,5 Kubaschewsky and Evans,6 etal.

Jan 1, 1960

By V. (“Viv”) R. Forbes

[ ] INTRODUCTION "The Liberal Party believes that the State's mineral resources belong to the people of Queensland and therefore it is essential that the state, on their behalf, obtains maximum benefit from the development of those resources." (Liberal Party, 1980 "Mineral Resources Policy" p.2). "Unless sound policies are implemented.... increased production of minerals will benefit only the already wealthy and the lucky few... A resources tax is essential to ensure that above normal profits are shared by the whole community.... Labor policies on taxation must ensure an adequate return from mining developments to the Australian people." (Labor Party Policy Paper, May 1982 "National Issues relating to rapid expansion of mineral production during the eighties". p.1, 37, 47). Judging from these statements, key political parties in Australia believe that mining profits are lightly taxed and that shareholders get most of the benefits. THE REALITY The Government Share of Income statement in Table 1 attempts to measure the government share of the income generated by the "Big Five" Australian mining companies - Broken Hill Pty Ltd. (BHP), CRA Limited, Utah International, MIM Holdings Ltd., and Western Mining Corporation ("WMC"). These companies control over 70% of Australia's mining industry including all the most profitable mining operations. They dominate the mining and processing of iron and steel, aluminum, manganese, coal, oil and gas, copper, lead, zinc, silver, nickel, gold, ferro-alloys and salt in Australia. They are also major world exporters of most of these products. They are thus a representative sample of the profit-earning side of the Australian mining industry. In their two most recent financial years these five companies had income before tax of A$6,980 million. The Government share of this income was A$6,285 million or 90%. The total market capitalisation of the Big Five is about A$8,480 million. (December, 1982). Governments thus consumed the equivalent of 82% of their total value in just two years. This was taking place while the politicians drafted the above statements calling for an "adequate" share for the people.

Jan 1, 1985

YEAR OF ELECTION 1913. FRANK DAWSON ADAMS, Montreal, Canada. 1921. WILLIAM CUTHBERT BLACKETT Sacriston, Durham, England. 1935. GUILLAUME DANIEL DELPRAT Melbourne, Vic., Aust. 1920. HENRY STURGIS DRINKER Merion Station, Pa. 1933. KARL EIKERS NEW YORK, N. Y. 1922. FEDERICO GIOLITTI Torino, Italy. 1906. SIR ROBERT A. HADFIELD London, England. 1928. JOHN HAYS HAMMOND New York, N. Y. 1921. FRANK WILLIAM HARBORD London, England. 1917. HERBERT HOOVER Washington, D. C. 1905. HENRI LECHATELIER Paris, France. 1931. WALDEMAR LINDGREN Cambridge, Mass. 1921. C. McDERMID London, England. 1913. EZEQUIEL ORDONEZ Mexico City, Mexico. 1935. ROBERT PEELE New York, N. Y. 1911. ROBERT H. RICHARDS Cambridge, Mass. 1934. T.A. RICKARD Victoria, B.C., Canada 1933. ALBERT SAUVER Cambridge, Mass.

Jan 1, 1936

By Gerald G. Griswold, James O. Greenslade, Cherie Tilley, Richard Reseigh

INTRODUCTION The Harrison Western Corporation, a leading Denver based mine contracting and engineering concern, is presently engaged in sinking two 1,006 m (3,300 ft) shafts for the Phillips Uranium Corporation at their Nose Rock Project, approximately 13 miles northeast of the small community of Crownpoint in McKinley County, New Mexico. The Nose Rock Project is the first attempt by the Phillips Uranium Corporation to tap the deep uranium reserves in what has become known as the Grants Mineral Belt. (See Figure 1) Project Description Phillips Uranium Corporation's plan for the large 2700 metric ton per day (2950 tons/day) mining facility calls for a series of deep access and ventilation shafts ranging from 4.27 m (14 ft) to 5.49 m (18 ft) in diameter to approximate depths of 1,006 m (3,300 ft). The initial pair of shafts consists of one production shaft and one ventilation shaft separated by a distance of 91 m (300 ft). The interbedded layers of sedimentary sandstones and shales to be penetrated by the shafts contain several major water producing aquifers, the deepest being the mineralized zone called the Westwater Canyon Member of the Morrison Formation. A series of temporary and permanent water pumping stations is planned. Generally, the temporary stations are located above major aquifers to facilitate water removal while sinking through the aquifer and the permanent stations constitute what, in the final mode, will be the mine dewatering system. In addition, Phillips Uranium Corporation has In- stalled and is maintaining a system of depressurization wells that temporarily pump the major aquifers and considerably reduce the water inflows during shaft construction. The major aquifers are also chemically grouted prior to sinking. Site work for the project was initiated in the fall of 1976. Full scale shaft sinking commenced in November of 1977 on the 4.88 m (16 ft) diameter ventilation shaft and on the 5.49 m (18 ft) production shaft by another contractor. Harrison Western Corporation began work on the project on November 4, 1979 wlth the production shaft at a depth of 633 m (2,076 ft) and the ventilation shaft at a depth of 474 m (1,554 ft). This paper will address only the portion of shaft sinking completed by Harrison Western. Geology and Hydrology The Nose Rock Project is located on the Chaco Slope in the southern extreme of the San Juan Basin in northwest New Mexico. The southern San Juan Basin is generally bound by the Defiance Uplift to the west, the Zuni Uplift to the south and the Nacimiento Uplift to the east. The Grants Mineral Belt occupies most of the southern portion of the San Juan Basin and this project is located on the northern extreme of the Grants Mineral Belt. The San Juan Basin is comprised of sedimentary rock of continental, marginal-marine and marine origin, that dip northward from the Zuni Uplift and Chaco Slope into the interior of the San Juan Basin. (See Figure 2) Major Formations The major geologic formations to be penetrated range from mudstone and shales to siltstones and sandstones. The sandstones are highly productive aquifers which, under static conditions, would flow artesian. Water temperatures in the aquifers are high and therefore are a force to be dealt with. Temperatures range from 18ºC (65OF) in the upper aquifers to 48ºC (118º'F) in the lower. The mudstones are generally bentonitlc. (See Figure 2) GALLUP SANDSTONE: The Gallup Sandstone is the first major regressive wedge in the San Juan Basin of the Cretaceous era and at this project is about 35 m (115 ft) thick. It is medium grained and prior to depressurization carried a hydrostatic water pressure of over 72.5 kg per cm2 (1,031 psi) which 1s artesian. Compressive strength ranges from 334.7 kg per cm2 (4,760 psi) to 421.8 kg per cm2 (6,000 psi). Water temperature is approximately 30°C (86°F).

Jan 1, 1982

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 E. W. Enqelmann

INCREASED metal consumption throughout the world in the past three years has brought greater activity in the concentrators and mills that treat the ores.' Comparatively low prices have made greater economies necessary, with a continuing search for more efficient flow sheets, equipment, processes, and reagents. Comminution GRINDING has long been too ex- pensive and even any small increment of saving has been eagerly sought. Although not recently new the use of rod mills to cover the gap between intermediate crushing and coarse grinding has made further headway. They eliminate dusting and the closed- circuit operation of fine crushers, either roll or Symons cone, and the rod-mill product is suitable for secondary mills which can be operated with smaller balls and iron liners instead of coarse balls and steel liners. The newer practice is radically different from the older application of rod mills where they at- tempted to use them instead of ball mills for fine grinders. For that purpose they were not well adapted, as they do not have sufficient grinding surface available per cubic foot of mill volume.

Jan 1, 1940

Jan 1, 1879

By J. S. Ll. Leach, J. B. Cohen, M. B. Bever

IN the work reported here, the heats of formation of a copper-rich Cu-Sn (a) solid solution and of the Cu.Sn (0 phase were measured by tin solution calorimetry. An approximate determination of the heat effect attending the precipitation of 6 from a supersaturated a solid solution was also made.

Jan 1, 1955

By D. R. Irving

Prior to this time, ample supplies of high grade mineral fillers, such as clay, have been available close to consuming centers. Now depletion of these accessible deposits, coupled with other factors such as increasing demand due to population expansion and standard of living rise, has made development of upgrading techniques attractive to producers. Various investigations being carried out by the U.S. Bureau of Mines are described. Factors making the development of improved methods for upgrading clays attractive to producers include: depletion of accessible deposits of high-grade clays, rising transportation and labor costs, more exacting specifications, and increased demand resulting from an expanding population and a constantly rising standard of living. Industry recognizes the need for upgrading presently unusable materials and has cooperated with the U.S. Bureau of Mines by providing samples from submarginal deposits and by evaluating the products that USBM upgrades. Following are some examples of recent clay beneficiation work by the Bureau. FIRE CLAY BENEFICIATION At the Rolla Metallurgical Center, Rolla, Mo., the Bureau is conducting research to develop technically and economically feasible mineral-dressing methods to remove quartz, pyrite, and other impurities from submarginal fire clays. The research has been spurred by the growing shortage of high-grade bur-ley, flint, and plastic clays in the Missouri fire-clay district resulting from increased consumption and the low rate of discovery of new deposits of adequate quality. There apparently is ample tonnage of submarginal flint and plastic clays to maintain the industry for many years if such material can be upgraded to commercial quality. Samples have been submitted to USBM by several major producers of refractories, including A. P. Green Fire Brick Co.; Harbison-Walker Refractories Co.; Mexico Refractories Div., Kaiser Aluminum and Chemical Corp.; and Wellsville Fire Brick Co. Research on some samples has been completed and still is in progress on others. The Bureau has found that specimens containing quartz or pyrite in grains coarser than the accompanying clay generally can be upgraded either with a gravity table or hydraulic cyclone. The improvement in quality is notable not only in chemical analysis but in an increase in pyrometric cone equivalent (PCE) of one or more cones. As an example, consider the results attained on a sandy flint—a fine-grained mixture of kaolin and halloysite with quartz and minor quantities of iron and manganese oxides. Chemical analysis was: 35.4 pct alumina (A12O3) and 47.8 pct silica (SiO2). The sample contained 9.2 pct free silica as quartz. The PCE was 33. The sample was crushed to -10-mesh, blunged with water, and then stage-ground to -48-mesh. The ground sample was tabled to yield a clay fraction containing 37.6 pct A12O3, and 45.3 pct SiO2 of which 2.4 pct was free silica. The PCE was 34. Recovery of alumina by this process was 73.6 pct. Concentration of the same clay in the hydraulic cyclone was even more effective in respect to quartz removal. The clay fraction analyzed 38.6 pct A12O3, 45.0 pct SiO2 of which 0.5 pct was free silica as quartz, and had a PCE of 34. Recovery of alumina was 71.9 pct. The hydraulic cyclone was also effective in reducing the pyrite content of a plastic clay which likewise was a mixture of kaolin and halloysite, plus some pyrite, quartz, and carbonaceous material. It analyzed 30.6 pct Al2O3, 52.3 pct SiO2, 2.3 pct Fe2O3, and 0.76 pct S. The sample was blunged and screened through 10-mesh to remove some coarse pyrite, then ground to -65-mesh and cycloned to remove most of the remainder. The final product recovered 81.8 pct of the alumina at a grade of 32.3 pct A12O3, 50.9 pct SiO2, 1.4 pct Fe2O3, and 0.1 pct S. The PCE was 32-1/2, compared with a PCE of 31-1/2 for the crude clay. The fine beneficiated clays, of course, must be pelletized and dried or fired to meet commercial acceptance. This operation has already been solved by industry. Little progress has been made on two remaining problems. One is that of fine-grained impurities, not separable by table or cyclone. Flotation is being investigated to solve this. Although fine pyrite is readily rejected, a satisfactory quartz-clay separation has not been attained. The second problem, a stubborn one, is that of alkali. Samples containing as much as 5 pct potas-

Jan 1, 1961

By Roshan B. Bhappu, Thomas J. Lien

The DYNA WHIRLPOOL Processor (DWP) is a cylindrical dynamic heavy media vessel. It is used to beneficiate fine ore sizes from one and one-half inches in diameter down to 65-mesh. Process units available treat up to 80 tons per hour. Units can be manifolded for higher capacity plants. Finely ground magnetite and ferrosilicon are used for media. , Installed capital costs are from $4,000 to $6,000 per ton hour of feed. Operating costs are from $0.30 to $0.50 per ton of feed. The DWP is used to beneficiate metallic and non-metallic ores and coal. The OCC Vessel is a static heavy media separator used to beneficiate coarse ore up to six inches in diameter and auto scrap. Process units available treat up to 300 tons per hour.

Jan 1, 1978

By Ray W. Amstutz

The shaft was drilled on US government land m the Piceance Creek basin and was funded by the US Bureau of Mines (USBM). It was drilled to 723 m to provide access to rich deposits of oil shale, nahcolite, and dawsonite for geotechnical test work and environmental mining research. Rotary drilling and dual-string, reverse-air, jet-assist circulation methods were used. Drilling required eight months. Site preparation, running, and cementing the casing required another four months. A 2.44-m ID steel welded casing, 508 mm thick was installed and 2.7 kt of cement were slurried and pumped into the annulus.

Jan 1, 1982

By D. F. Kelsall

The hydraulic cyclone, in simple or modified form, is finding increasing application in metallurgical processing. In this article, the author considers several aspects of conventional applications, less conventional applications and some potential uses of modified cyclones. Information concerning the mechanism of separation within a cyclone and the effect of the major variables is summarized. During the past twenty years, and especially in the last decade, hydraulic cyclones have found increasing application in metallurgical processing as classifiers and, under special circumstances, as thickeners. To a more limited extent cyclones have been used to extend heavy, or dense, medium separation into the finer size ranges, and several attempts have been made to separate dispersions of immiscible liquids with a view to use in the field of liquid-liquid extraction. A large proportion of published work reports incomplete data, applicable only to particular systems. The more useful papers, covering data of relatively wide significance, have been reviewed recently in considerable detail1,2 and no attempt will be made to cover the same material in the present paper. However, before considering several aspects of conventional application, less conventional applications and some potential uses of modified cyclones, it is considered pertinent to summarize information relating to the mechanism of separation within a cyclone and the effect of the major variables. LIQUID FLOW PATTERNS At any point within a simple cyclone the liquid velocity may be resolved into three components, i.e., the tangential or swirling velocity (v), the vertical velocity (w), and the radial velocity (u). The relative values of these components at selected positions within a typical cyclone have been established for water. 3 The extent to which similar flow patterns exist in systems of considerably higher viscosity, or apparent viscosity, operating under conditions of reasonable pressure drop, is not known with precision, but indirect evidence indicates that up to approximately five centipoise there is no marked change. Beneath the bottom of the vortex finder, envelopes of constant tangential velocity are cylinders coaxial with the cyclone, and, at any horizontal level, starting from near the conical wall, this component of velocity increases with decrease in radius (r) according to the relationship vrn = constant (where 0 < n < 1) This holds to regions near the center of the cyclone where v reaches a maximum value at a radius somewhat smaller than that of the inside wall of the vortex finder. At this point, the centrifugal acceleration, v2/r reaches a maximum value. As the radius is further decreased the tangential velocity decreases so that v = r. Above the bottom of the vortex finder there is a tendency for the first relationship to persist to near the outside wall of the vortex finder, until v reaches a maximum and then decreases rapidly as the wall is approached. The two most significant features of the vertical velocity distribution are an approximately conical Table I. List of Symbols v = liquid tangential velocity at any point in the cyclone w = liquid vertical velocity at any point in the cyclone u = liquid radial velocity at any point in the cyclone r = radius from cyclone axis to any point ? = half angle of conical portion of the cyclone ?= fluid viscosity d = particle diameter s = liquid specific gravity p = particle specific gravity d50 = the particle diameter which distributes 5090 to overflow, 50% to underflow e = overflow diameter, in. b = feed diameter, in. G = gpm of feed slurry

Jan 1, 1963

By N. F. Schulz, H. A. Lex

Representative samples of typical manganiferous iron ores from the Cuyuna Range, Minn., were reduced to the metallic iron state at temperatures just short of fusion with an excess of coke in the manner prescribed by the R-N Direct Reduction Process. The magnetic separation characteristics of the reduced ores were determined at several grinds by means of the Davis magnetic tube concentrator. The efficiency of metallic iron concentration by magnetic separation was a direct function of the degree of metallization of the ore sample and of the degree of liberation at-tained in grinding. Some improvement of liberation characteristics was observed if a small amount of sodium chloride was present in the ore during reduction. Despite the extremely fine-grained structure of the crude ores, iron concentrates containing more than 90% Fe at iron unit recoveries of 90% or more were obtained from well-metallized samples of three of the four ores tested. The rejection of manganese to the tailings corresponding to such separations varied from 80% to 95%. The manganiferous iron ores of the Cuyuna Range in Minnesota constitute a major potential source of manganese in the United States provided that a technically and economically feasible process can be devised for its recovery. The iron and manganese of these ores occur chiefly as oxides or carbonates in the form of very small, irregular, discrete grains embedded in a matrix of quartz or other siliceous gangue. The present investigation was undertaken as a sequel to an earlier one conducted jointly by the U.S. Bureau of Mines and the R-N Corp. Those results indicated that the R-N Direct Reduction Process might be a means for recovering both a metallic iron concentrate and a marketable manganese by-product from Cuyuna manganiferous iron ores.&apos; The R-N Process is one of many proposed schemes for the treatment of crude iron ores involving the di- rect reduction of the iron content to metallic iron followed by physical extraction of the iron. In general, it consists of the following stepsz2 1) Iron oxide in crushed natural ores or iron concentrates is about 95% reduced to metallic iron at elevated temperatures but without fusion in a rotary kiln in the presence of a solid carbonaceous fuel (usually coke), limestone, and combustible gases. 2) The kiln product, consisting of metallized ore, excess coke, lime, and coke ash is cooled. The excess coke is recovered and recycled to the kiln feed. 3) The metallized ore is ground for liberation and magnetically separated to yield a metallic iron concentrate. 4) The metallic iron concentrate is agglomerated by a briquetting process. The reduction kiln temperatures must be high enough for the desirable chemical reactions to proceed at practical rates without causing fusion of any of the ore constituents. The time of exposure of the ore to the reducing conditions must be ample for 90% to 95% metallization of the iron oxides present. If more than 95% of the iron is metallized, however, difficulties may arise in the briquetting step, presumably because of accompanying higher carbon content of the metallic iron. Although not always necessary, powdered limestone or dolomite may be added at some time during reduction to aid in eliminating sulfur and phosphorus from the final product.3,4 Other additives, notably sodium chloride, when present during reduction, result in improved liberation characteristics of the metallic iron from certain reduced ores.&apos;v6 Under the reducing conditions normally employed, P(co + H2)= 1 atmosphere, and 1830&apos; to 2150°F (1000° to 1170°C), the R-N Process is claimed to be adaptable to a wide variety of low, medium, and high grade iron and manganiferous iron ores. The manganese oxides and carbonates are reduced mostly to nonmagnetic MnO which is rejected to the nonmagnetic tailings. Manganese-bearing pyroxenes of the original ore, however, may be rejected poorly, if at all, apparently because these minerals occur within the iron lattice or are entrapped in the metal particles.7 The investigation was limited to studies of the Cuyuna ores listed in Table I, which were reduced to the metallic iron state under various conditions

Jan 1, 1969

By N. A. D. Parlee, E. M. Sacris

OVER 55 years ago, Sieverts and Hagenacker1 and Donnan and shaw2 made determinations of the solubility of oxygen in liquid silver, over a rather short range of temperatures (973° to 1125°C), using quartz bulbs, mainly at 1 atm. From the results of their investigations, Mizikar, Grace, and parlee3 plotted % O vs 1/T and obtained the following straight-line equation for solubility at 1 atm: log % O = 723.7/T - 1.0949 Recent work on the kinetics of the Ag-O2 system394 indicated there might have been some error due to reaction with the direct heated silica bulbs. Thus it became important to redetermine the solubility with the silver contained in a better refractory such as alumina, and by induction heating where the walls of the container were not so strongly heated. It was also desirable to extend the temperature range. A modified Sieverts apparatus of fairly standard type was used except that a special type of reaction tube was employed, Fig. 1. Dried argon with a guaranteed purity of 99.9 pet, and dried medical-grade oxygen with a guaranteed purity of 99.6 pet containing a little nitrogen, were the gases used. The measuring system consisted essentially of a manometer and a water-jacketed 100-cc gas buret with an open-end vertical tube parallel to the buret. This tube has the same diameter as the buret in order to give the same capillary effect on the mercury. A screw leveling device was used with the buret leveling bulb to help provide accurate pressure readings. A 2-mm ID capillary line connected the buret to the reaction chamber, helping to reduce the dead volume, which was determined with argon. The heating medium was a 2.5-kw output induction furnace. The induction coil was made of a 3/16-in. copper tubing with eight turns and an inside diameter of 4 cm. The reaction bulb was made of a quartz tube 18 cm long and 3 cm in inside diameter, Fig. 1. Inside the quartz tube were two alumina crucibles (Mor-ganite Triangle RR, XN10 and XN5) one inside the other. Below these two was an alumina disk. The disk prevented the corners of the outer crucible from expanding against the quartz tube. The solid sample (99.994 + pet Ag), which had been previously cleaned by air oxidation and repeated removal of scum, was placed inside the inner crucible. An alumina disk with a hole at the center covered the two crucibles. In order to decrease the dead volume, an evacuated quartz "filler" tube was placed on top of the crucible cover. The part of the reaction tube above the spherical joint was made up of Pyrex glass with a flat glass top. It was here that a prism was placed to facilitate the use of an optical pyrometer, previously calibrated at the freezing point of silver. This type of reaction bulb and furnace assembly was specially designed for this work, and used in preference to the relatively new water-

Jan 1, 1965

"The method of conveying ore from the ore-pocket at the shaft collar to the railroad ore bins, whence it is transported to the reduction works, is by cars with electric haulage.The cut shows the car coupled to the electric locomotive at the moment of dumping into the railroad ore bins.The locomotives used are of several makes, usually having two 12 ½ h-p. motors of a rated draw-bar pull of about 1,200 lbs., using 500 volts D. C. Track gauge is 30 inches, and speed is six miles per hour.The car holds two skip loads-10 tons—and has sides opening outwards the full length of the car, for rapid side-dumping. One man operates car and motor, and loads and unloads the car. The trolley pole runs in both directions without reversing.Loading is clone by operating a valve connected to a compressed air lift, which raises and lowers a gate in the ore-pocket chute, the ore running by gravity into the ore car spotted under the gate of the chute. Dogs hold the sides of the car in place at the bottom, the tops of the sides being hinged, and a hand lever releases the dogs, permitting the sides by the pressure of the ore against them to open at the bottom.With little manual effort one man does the work which was formerly done, when hoisting ore with cars and cage, by 10 men under laborious conditions."

Jan 1, 1913

By H. L. Talbot, R. S. Young, A. Golledge

The differentiation of oxidized forms of lead from lead sulphide in complex products by chemical analysis is of considerable importance to certain mining and metallurgical companies. A method for the separation of "oxide" lead, (non-sulphide), from sulphide lead, based on the solubility of the former and virtual insolubility of the latter in concentrated ammonium acetate solution, has been employed for many years in the industry.1,2 Unfortunately this solvent, while satisfactory for lead sulphate and carbonate, is without action on lead phosphates or vanadates, with the result that at .Rhodesia Broken Hill Development Co. where these latter minerals are present, they have counted as sulphides when using this procedure. For the control of operations in the new concentrator it became necessary to have a closer check on sulphide lead than could be obtained by the only current method of differentiating this constituent from oxidized lead. For some time this problem has engaged our attention, and after a large number of solvents had been tried, the following procedure was evolved. Like all methods depending on selective solution in a com- plex material, it will not give 100 pct separation into oxide and sulphide portions. Because of the ease of oxidation of galena this is almost an impossibility. Our procedure will, however, give a recovery of 99-100 pct of the oxidized lead, contaminated with very small quantities of lead sulphide, and a corresponding recovery of 97-98 pct of the sulphide lead. The procedure is simple and reproducible, and gives results quite .satisfactory for routine control purposes. Experimental For this work standard samples of the following pure lead minerals were obtained from Broken Hill, Northern Rhodesia. These were hand-picked specimens, and were carefully ground to —200 mesh. Since it was realized that a fair-sized sample would be necessary for this work to thoroughly test the large number of procedures we had in mind, a compromise had to be effected between a few grams of very pure minerals and a large sample of lower grade material. For this work the presence of gangue was not deleterious, so long as the oxide lead minerals did not contain lead sulphide, and vice versa. Table I shows the lead minerals used, and their total lead content, both theoretical and actual. Apart from the pyro-morphite the minerals are seen by analysis to be quite pure. Many solutions were tried in an effort to dissolve the oxidized lead minerals and

Jan 1, 1949

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who registered at Institute headquarters during the period Feb. 10 to Mar. 10, 1915: W. R. Van Liew, Batoum, Russia. Edward S. Jones, Scranton, Pa. Olof Wenstrom, Boston, Mass. Scott Turner, Tromso, Norway. D. G. Miller, Denver, Colo. A. H. Boyd, Denver, Colo. E. E. Olcott, New York, N. Y. M. M. Thompson, Morsemere, N. J. Francis R. Pyne, Elizabeth, N. J. F. C. Merry, Brooklyn, N. Y. Charles W. Goodale, Butte, Mont. Charles A. Filteau has been appointed General Manager for the Inca Mining & Development Co., Province of Carabaya, Department of Puno, Peru. Charles T. Kruse has been appointed Superintendent of the Jones & r, Loughlin mines in the Marquette and Menominee ranges.

Jan 4, 1915

By John J. Cioston

THE steadily rising flood of gold production from all parts of the world has created an avalanche of rumors regarding the stability of the present price of this metal. Markets have been unsettled from Johannesburg to London from New York and Toronto to Melbourne, despite repeated denials from, American and British authorities of any immediate intention to alter the price. Various economists and financial writers have blossomed forth with predictions that the price would be set at the old level of $20.67 per oz.; some have even been so brash as to state That a price of $15 would be necessary to bring production within control. These effusions, based entirely upon monetary theories and with no regard to production costs, serve to confuse rather than clarify the situation.

Jan 1, 1937

Jan 1, 1938