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By J. Wagner, P. Eger, G. Melchert

In 1991, LTV Steel Mining decided to close an open-pit taconite mine in northeastern Minnesota using a passive-treatment approach consisting of limiting infiltration into the stockpiles and wetland treatment to remove metals. More than 50 Aft (55 million st) of sulfide-containing waste had been stockpiled adjacent to the mine during its 30 years of operation. Drainage from the stockpiles contained elevated levels of copper, nickel, cobalt and zinc. Nickel is the major trace metal in the drainages. Before the closure, the annual median concentrations ranged from 1.5 to 50 mg/L. Copper, cobalt and zinc are also present but they are generally less than 5% of the nickel values. Median pH levels range from 5 to 7.5, but most of the stockpile drainages have pH levels greater than 6.5. Based on the chemical composition of each stockpile, a cover material was selected. The higher the potential that a stockpile had to produce acid drainage, the lower the permeability of the capping material required Covers ranged from overburden soil removed at the mine to a flexible plastic liner. Predictions of the reduction in infiltration ranged from 40% for the native soil to more than 90% for the plastic liner. Five constructed wetlands have been installed since 1992. They have removed 60% to 90% of the nickel in the drainages. Total capital costs for all the infiltration reduction and wetlands exceeded $6.5 million, but maintenance costs are less than 1 % of those for an active treatment plant. Because mine-drainage problems can continue for more than 100 years, the lower annual operating costs should pay for the construction of the wetland-treatment systems within seven years.

Jan 1, 2000

By Edward Judd

CONSTRUCTION of the long projected Moffat tunnel, on the Denver & Salt Lake R. R., between Tolland and Irving, Colorado, is now actually and actively progressing. This 6.1-mile bore through the Rocky Mountains, at an elevation of about 9200 ft., will eliminate 33 1/3 miles of probably the most tortuous standard-gage track in the world, and at the same time will avoid the necessity for crawling over the snow-bound continental divide at an elevation of 11,660 ft Completion of the tunnel within the stipu-lated four years, is expected to nave an immediate effect on the develop-ment of the known coal fields in north-western Colorado, already penetrated at Craig, the present terminus of the road, while it will also enable the railroad to carry out the original designs of its founder, David H. Moffat, for through traffic from farther west. The intention now is to accomplish the latter object by an extension from McCoy following the canon of the Grand River down to a connection with the Denver & Rio Grande at Dotsero. Legal impediments to the financ-ing of the tunnel having been removed by action of the Colorado legislature in April, 1922, the Moffat Tunnel Improvement District issued bonds to the amount of $6,720,000 which have been widely and quickly distrib-uted through a New York banking house. Responsi-bility for the execution of the project has been vested in the Moffat Tunnel Commission, to which a board of consulting engineers has been appointed for carrying out the details of administration. This board includes D. W. Brunton and L. D. Blauvelt, of Colorado, and J. Vipond Davies and J. Waldo Smith, of New York; R. H. Keays, who has managed large tunnel projects in New York, is chief engineer for the Commission.

Jan 11, 1923

By Hoskings R

Conceptual studies of the run of mine stockpiling and reclaiming requirements com- menced in 1975. Final selection of equipment was made in 1978 and the equipment was commissioned in June, 1979. Equipment selection was made after extensive investigation and visits to existing facilities throughout the world. The equipment selected is described and com- pared with the alternatives. Foundations were designed to minimise clean up effort and to ensure stability of equipment under design wind loading without storm bays. Program- mable logic control enables the stacking and reclaiming facility to provide a blended feed to the coal preparation plant.

Jan 1, 1980

By Harry Perry, Mary S. Esfandiary

This is a revision of a directory of coal research organizations throughout the world that was first published in 1961 as Bureau of Mines Information Circular 8008. Because of the wide acceptance of the first edition , a revision which would supply up- to - date information on coal research and which would facilitate communication among organizations of similar interest seemed important . This compilation includes the names and activities of organizations of various types -government , university , private , and industry . The areas of investigation that each organization is engaged in are indicated in broad terms : Analysis and constitution , mining and geology , preparation , briquetting , combustion , high- and low-temperature carbonization , gasification , conversion to solid and liquid fuels , and health and safety . Publications issued by and describing the work of each organization are cited ; almost all of the organizations also publish results of their investigations in the technical press . Selected technical and trade publications that publish information on the science and technology of coal and are not published directly by any of the organizations listed are cited in the pages following the list of organizations in each country . Information that would improve coverage and accuracy of the next revision of the publication will be welcomed by the Bureau of Mines .

Jan 1, 1966

By Tao Long

Recently, the great amount of funds has been invested in the reconstruction of damaged infrastructure and house of subsidence area of coal mine and mine reclamation. This is a benevolent and popular work, which is highly praised. However, because of complicated collapse area problems in China, especially the less integrated control research in the collapse area of metal mine, and the technological weakness, how to effectively control the subsidence area, restore landscape and ecological environment, and utilize the abandoned mine land, are major issue related to the mining area of sustainable development. The research situation about the formation of metal mine collapse area at home and abroad is reviewed, and then the effect of the collapse area on the mine safety and ecological environment, the existing technical problems and development direction are analyzed. Lastly, the technical measures of the collapse area control and cooperative underground mining is put forward.

Oct 1, 2010

By Chunhua Liao, Chao Peng, Guangsheng Huo

To get the separation of tungsten and molybdenum, we adjusted the pH of the solution to 7, at which tungstate (WO4 2-) will be polymerized into HW6O21 5-, while molybdate will still exist in the form of MoO4 2-. Two strongly basic anion exchange resins (D201, D213) were experimented for the separation of tungsten and molybdenum. The results of condition experiments indicate that D213 resin performs better in the treatment of lower concentration solution A (WO3 10g/L, Mo 4.12g/L) with a separation factor as high as 10.64. While D201 resin is more suitable for the high concentration solution B (WO3 70g/L, Mo 28.97g/L) with a separation factor of 6.9. The results of column testing performed on D213 resin show that the solution after loading process with a mass ratio of Mo/WO3 = 50 and the desorption solution with a mass ratio of WO3/Mo = 27 can be obtained

Jan 1, 2014

By G. E. Karantzavelos

Introduction Pure alunite (KAl3 (SO4 )3 (OH)6 ) contains 37% A1203. In the alunite ore found on Milos Island, Greece, opal is the principal gangue while some aluminum, iron, and titanium oxides are also present. Processing of alunite ore leads to the recovery of potassium sulfate, alusil (alumina-silica mixture), and sulfur dioxide. An interesting process has been developed in Greece (Svoronos, 1977) consisting mainly of roasting the alunite ore mixed with KCl to disrupt the mineral lattice in preparation for separating its components. The thermal decomposition of alunite at approximately 600°C occurs by the reaction: KAl3 (SO4)3 (OH)6 + nSiO2 + 6KC1 - 4 K2SO4 + 3 A1203 . nSiO2 + 6 HCl + 3H20 According to the process pure alunite mixed with 543 kg KCl, 840 kg K2SO4, 370 kg alusil, 264 kg HCl, and 65 kg H2O are produced. The roasted calcine is leached in hot water. The potassium sulfate leach slurry is then thickened and washed to obtain a solid residue (alusil) plus a solution rich in potassium sulfate. Assuming the alunite ore feed consists of 50% to 60% alunite, the solid residue might consist of approximately 22% to 27% A1203, silica, iron, and titanium oxides. This leach residue could be digested in a hot sodium hydroxide solution to dissolve alumina. An amount of the alumina (approximately 10%) would not dissolve. In addition, part of the silica would dissolve requiring removal from the solution in a subsequent desilication operation. The same leach residue could be attacked by mineral acids such as a diluted hydrochloric acid to extract aluminum chloride (Sawyer et al., 1983). Alumina and aluminum are presently produced from bauxite of which Greece has important reserves. If this aluminous material (alusil) could be beneficiated by any physical process, then it would be considered as a raw material for the production of fire-clay refractories instead of an alternate source of alumina. Experimental procedure Extensive laboratory work (mineralogical and chemical analyses) indicates that alusil is a mixture of liberated (mostly) alumina and silica particles, not a natural silicate mineral. Table 1 shows the chemical assays of two alusil samples that were obtained from a roasted alunite ore leaching pilot plant. The samples used for the experiments were a very fine alusil powder (K80 = 400 mesh, 38 µm). These were pulped in tap water. Desliming of the pulp before each flotation test was ineffective. Flotation tests were performed in a 2000 ml Denver cell. A standard 500-g sample was first conditioned at a pulp density 50% solids for 5 min with hydrofluoric (HF) acid solution (40%). The temperature of the conditioned pulp was measured and kept nearly constant at 50°C. The pH of the conditioned pulp was also regulated to 2.5-3 with HF solution. Then ARMAC C solution (2%) as well as two drops of fuel oil and one drop of AEROFROTH 65 were added and the pulp was reconditioned for 2 min. After conditioning, the viscous pulp was transfered to the flotation cell. It was then diluted with warm water to about 15% solids and floated to separate silica from alumina particles. A rougher froth product that was cleaned and recleaned was obtained from the flotation tests.

Jan 1, 1987

The Kwademen Au deposit was discovered in 1977- 1978 following a regional geochemical survey carried out as part of a UNDP (United Nations Development Programme) and BUMIGEB (Bureau des Mines et de la Geologie du Burkina) mineral exploration programme. The climate in Burkina Faso is rather dry, with markedly seasonal rainfall averaging 800 mm pa in the Kwademen project area. The relief is low and the drainage poorly developed, so that soil sampling was preferred to stream sediment sampling. Rocks have been subjected to deep lateritic weathering, so that they are transformed into a mixture of Fe-oxides and kaolinite over depths varying between 30 and 40 m. At the regional stage, soil samples (30-40 cm depth) were collected a a 300 x 500 m grid. The Kwademen deposit was indicated by Cu, Zn, Mo and Au anomalies. Subsequently, an area of 9.2 km2 was covered by a detailed geochemical survey by soil (100 x 50 m) and auger sampling. The regional anomaly was fully confirmed, and subsequent drilling intersected mineralized ore shoots, justifying the establishment of an evaluation project led by the U.N.R.F. (United Nations Revolving Fund) in 1983- 1985.

Jan 1, 1988

By W J. Rankin, D R. Swinbourne

There are large tonnages of low-grade manganese ores in India fromwhich it is uneconomic to produce ferroalloys by conventionalsubmerged arc furnace smelting. A possible way to utilise these ores is to smelt them with carbon to produce pig iron and a manganese-rich slag, then to metallothermically reduce the molten slag to silicomanganese. Thermodynamic modelling of the metallothermic reduction stage using aluminium, silicon or aluminium-silicon alloy as the reductant at 1450¦C showed that an acceptable grade of silicomanganese could be produced. Manganese recoveries were predicted to be about 80 per cent. Heat balances indicated that silicothermic reduction may not produce enough heat to maintain reaction temperature. Small-scale laboratory experiments showed that the manganese grade of the alloys, and the manganese recoveries, were close to those predicted, but those for silicon were muchlower. Aluminothermic reduction produced the highest grade manganesealloy. The slag-metal equilibrium involving manganese, silicon and their oxides was investigated and it was found that equilibrium was not attained in any of the experiments, due probably to the very low rate of reactions involving silicon. This could be used to advantage in practice to achieve higher manganese levels in the alloys.

Jan 1, 1998

By J. A. Lombardi

Acid-Minesite-Solutions (AMS) are defined as copper-containing pit lake, tailing pond, acid-mine-drainage, and heap and waste dump irrigation solutions. All of these solutions are characterized by copper, acid and water contents that have value as individual commodities. HW Process Technologies, Inc. (HWPT) has developed nano- and hyper-filtration membrane methods for the separation of copper and other base metals, acid and water from AMS. These treatments are collectively called HWPT?s Engineered Membrane Separation® (EMS®) systems. The EMS® method of copper recovery produces small volumes of higher-concentration copper solution from large-volume, copper-dilute AMS. A case study of the 4,000 gpm Cananea EMS®-AMS plant for the recovery of copper from an acid-pit-lake resource is described. The EMS® method of copper removal from AMD and tails pond waters has been bench-scale demonstrated and are described, complete with projected budgetary capital and operating costs. Additional treatments for the recovery of sulfuric acid and high-quality water from the permeate (membrane passing) waters for the above-described AMS treatments are presented, with budgetary capital and operating cost estimates. [ ]

Jan 1, 2009

By Michael B. Tessieri

With the increasing environmental demands and higher fuel efficiency standards on auto manufacturers, the use of aluminum in automotive applications is expected to show significant growth for the balance of this decade. Cast and wrought aluminum usage is reviewed by auto part, and alloy family for the year 2000 automobile. The analysis addresses the effect of this increased volume on the recycling infrastructure and identifies issues and opportunities, which will face the recycling industry.

Jan 1, 1995

By F. Habashi

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

Jan 1, 1967

By Louise Reber

THE town of Jerome is located in Yavapai County in north central Arizona. It has a population of over 6000 people and the two important mines of the district, the United Verde and the United Verde Extension, lie almost within its limits. Perched high on the southwest side of the Verde Valley, at elevations of 4800 and 5500 ft.. (1463 and 1676 m.), both mines have long tunnels from their lower levels through which ore is hauled to their respective smelters. The two smelter towns of Clarkdale and Verde, out in. the center of the valley and about 2000 ft. (609 m.) below, form part of the wide panorama that meets the eye at Jerome. This panorama and the topographic and surface features and the geologic formations are shown in a separate pamphlet. In the vicinity of Jerome, the great southern escarpment of the Arizona plateau forms the northeast side of the Verde valley. The general level of this escarpment, or "the rim" as it is called, is about 4000 ft. above the Verde River. A corresponding elevation is reached only by the, highest parts of the Black Hills, which bound the valley to the southwest. The two lava-capped mesas, known as Mingus and Woodschute Mountains, form the culmination of the Black Hills. That part of the old Black Hills mining district which lies between their crests -and the Verde River is now known as the Jerome or Verde copper district. However, if the smelters are not included, a strip about 7 mi. long and 3 mi. wide (11 by 5 km.) embraces practically all the properties belonging to the Jerome district. This strip lies parallel to and along the side of the Verde valley from about First View, 2 mi. northwest of Jerome, to the neighborhood of Allen Spring, near the south end of Mingus. Mountain. The rank of the district in copper production, as well as features of peculiar geologic interest in connection with the ore occurrences, entitle the district to much more consideration than it has received in the literature of copper mining. Arizona stands preeminent as a copper-producing state. In 1918, it produced four-tenths of the total amount produced in the United States, and nearly equaled the, combined production of its three chief competitors, Montana, Michigan, and Utah. The production of the Jerome district places it in the sixth place among the copper

Jan 8, 1920

By B. H. Johnston, B. N. Doyle

Introduction In 1983-84 Kidd Creek Mines Ltd. in Timmins, Ontario, started up the second commercial zinc pressure leaching plant based on the Sherritt Gordon Mines Ltd. direct pressure leach process. The first plant using this process was that of Cominco Ltd. at Trail, B.C. (Parker, 1981; Martin and Jankola, 1985). The Kidd Creek plant was constructed to take advantage of the extra capacity that the existing purification, cell house, and casting sections had in comparison to the roasters and acid plant. In addition, pressure leaching offered flexibility in other areas such as treatment of low grade concentrates. The process design package and the detailed design of the autoclave system was carried out by Sherritt. The detailed design for the rest of the plant was done by Cambrian Engineering Ltd. under the direction of Kidd Creek. The construction and commissioning of the plant has been described earlier (Johnston, 1983). This paper describes, briefly, the application of the direct pressure leaching process for zinc concentrates and the integration of the pressure leaching plant into the main zinc plant at Timmins. It discusses some of the startup problems and changes made to the plant. The damage following a fire inside the autoclave is also described. The subsequent repair and recommissioning of the autoclave has enabled Kidd Creek to demonstrate the originally sought advantages of pressure leaching and to take advantage of their extra cell house and casting capacity to produce more zinc. Process principles The development of the direct pressure sulphuric acid leach of zinc sulphide concentrate at temperatures above the melting point of sulphur and at elevated oxygen pressure has been described in detail in a number of papers (Martin and Jankola, 1985; Johnston, 1983; Veltman and Bolton, 1980; Kawulka et al., 1975; Veltman et al., 1977). The most significant aspect of the process is that sulphidic sulphur is converted to elemental sulphur in reactions that can be simply described as follows. ZnS + H2SO4 + 1/2 O2 ZnSO4 + S° + H2O (1) FeS + H2SO4 + 1/2 O2 FeSO4 + S° + H2O (2) The rates of these reactions increase with temperature. However, at temperatures above the melting point of sulphur, 12°C, the presence of a surface active agent is necessary to prevent the occlusion of the sulphide particles by elemental sulphur and the consequent termination of the leach. The zinc that is dissolved according to reaction (1), after purification of the zinc sulphate solution, is recovered by electrowinning. The iron that is put into solution according to reaction (2) is further oxidized to the ferric state and subsequently, as the acidity decreases, precipitates as hematite or as complex basic sulphates. MINERALS AND METALLURGICAL PROCESSING Examples of this are shown below. 2FeSO4 + H2SO4 +1/2O2 - Fe2(SO4)3 + H2O (3) Fe2(SO4)3 + 6H2O - Fe2O3.3H2O + 3H2SO4 (4) 3Fe2(SO4)3 + 14H2O - 2(H3O)Fe3(SO4)2(OH)6 + 5H2SO4 (5) (hydronium jarosite) The conventional zinc process involves concentrate roasting to produce zinc oxide and sulphur dioxide. The zinc oxide is leached in spent electrolyte and the sulphur dioxide is converted to sulphuric acid. The pressure leaching process can, therefore, be used to replace conventional concentrate roasting, sulphuric acid manufacture, and atmospheric leaching with one unit operation. Integration of pressure leaching into the Kidd Creek flowsheet A block flowsheet of the Kidd Creek zinc plant, with the new pressure leach plant integrated, is shown in Fig. 1. The original plant, which was constructed in 1972, is based on standard electrolytic zinc plant technology. Concentrates are dead roasted to remove sulphur. Sulphur dioxide is treated to produce sulphuric acid. Calcine is leached in two stages using cell-house electrolyte. Iron is precipitated as jarosite and solutions are purified to remove copper, cobalt, nickel, and cadmium to provide feed for a partially mechanized cell house. Zinc is plated from solution, and the resulting cathode is melted, cast, and sold (Johnston, 1983). The integration of the pressure leaching plant into the main zinc plant is relatively simple. As shown in Fig. 1, a fraction of the zinc concentrate and spent electrolyte are fed to the pressure leach autoclave, while part of the calcine from the roaster is used as a neutralizing agent in the oxidation step along with. the jarosite leach solution. The main exit stream from the pressure leaching plant is the slurry discharged from the oxidation step, consisting of neutral zinc sulphate solution, unreacted calcine, and precipitated iron oxides. This slurry is transferred to the neutral leach step in the main zinc plant. The leach residue from the pressure leaching step, consisting of basic iron sulphate and elemental sulphur, is washed and impounded along with the jarosite residue produced in the main plant.

Jan 1, 1987

By Subrahmanyam T. V, Rao V. D

GOLD, the most important and highly coveted valuable metal, is being sought and exploited in almost all the geological environ- ments by the developed, developing and underdeveloped coun- tries of the world. Brazil, one of the five largest countries of the world, has similar geological and lithological environments as of Africa, Australia and Canada and also produces this noble metal. With the recent gold rush in Serra Pelada Hill in Para State of Brazil and the consequent migration of a large number of enter- prises to the Amazon region in search of geological conditions similar to that of Serra Pelada, the status of Brazil in gold produc- tion has been considerably elevated to be among the first five of the world. The principal geological environments or the lithologies where gold can be explored are the greenstone belts, banded iron for- mations, quartz lodes and placers; and the tectonic controls like shear zones in these horizons, or the tectonic relics or the annular- structures that limit the quartz vein systems in felsic as well as volcanogenic systems and epithermal models. The greenstone, BIF associations are present in all the Archean-Proterozoic ter- rains in Central and Northern Brazil. The quartz boudins in Al rich schistose rocks and the weathered and chemically enriched

Jan 1, 1988

By Martin Rubenstein

CdS crystals have been grown from a number of metallic solvents such as bismuth, tin, lead, and cadmium. Etching studies have shown that plastic deformation occurs if the crystals are not removed from the solvent prior to the solidification of the solvent, on cooling. The deformed crystals show a umique exciton fluorescence as a function of edge dislocation density. If one grows the CdS in the eutectic alloy of the above four metals (commonly called Wood's metal) the crystals can be removed from the solvent with hot water and no plastic deformation occurs. In this paper, the liquidus solubility measurements of CdS, as a function of temperature, are presented. The data were obtained using a high -temperature filtration technique. CADMIUM-SULFIDE crystals have been grown from a number of metallic solvents1 such as cadmium, bismuth, tin, and lead. Liquidus solubilities of CdS in cadmium,2 bismuth,3 and tin4 have already been measured. Crystals of CdS, in all four metals, have been grown by solution growth: 1) by cooling a saturated solution and 2) by a solution transport method.1'"1 CdS crystals grown in these four solvents have a few characteristics in common: 1) 1.8°K photolumines-cent emission consisted mainly of the radiative recombination of the bound exciton commonly known as I,, 2) slip lines which could easily be seen by the naked eye, and 3) edge dislocation densities in the order of l05 per cu cm.1 It was decided that these slip lines and the high edge dislocation densities were caused by a plastic deformation of the CdS crystals. It was felt that this plastic deformation did not occur during the growth of the crystals nor during the cooling of the solution, but did occur when the solvent which was in contact with the crystals froze. If these assumptions were valid, the slip lines and the high number of dislocations could be reduced or eliminated by removing the crystals from the solvent before the solvent froze. Since crystals of CdS had already been grown separately in such solvents as bismuth, lead, tin, and cadmium, it was felt that crystals could be grown in a eutectic mixture of these four metals. In this work a eutectic (or near eutectic) mixture of bismuth, lead, tin and cadmium in the proportion 50, 26.5, 13.5, and 10 wt pct, respectively, was used to grow CdS crystals. Such a mixture has a melting point of about 70°C and is close in composition to the alloy commonly known as Wood's metals. If the crystals could be grown from this mixture of solvents, and if hot water (>75°C) could be used to separate the crystals of CdS from the metallic solvent, it was hoped that CdS crystals could be grown with little or no plastic deformation which had been ob- served when crystals were grown from these solvents uncombined. CdS crystals were grown from this low melting eutectic mixture of bismuth, lead, tin, and cadmium using the solvent transport method. CdS powder and the appropriate amount of metals were sealed in a quartz tube under a pressure of about 5 X 10-6 torr. This ampule was then placed in a vertical position in a furnace. The temperature was raised to about 900°C. The furnace was designed so that the top of the liquid column within the ampule was between 10° to 40°C higher than the bottom of the liquid column. These temperatures were measured on the outside of the quartz ampule. The ampule was maintained at temperature for 7 to 14 days (depending on the temperature at which transport was taking place) and then the furnace temperature was lowered until the temperature was about 125°C. The ampule was then removed from the furnace, placed in water maintained at about 90°C, and opened in this 90°C environment. The crystals could then be removed from this two-phase liquid (Wood's metal and water) by mechanically picking them out. Alternatively, the crystals could be quantitatively removed by adding an excess of mercury to the mixture of metals, crystals, and hot water. The hot solution of metals and the hot water could be evacuated using a small diameter tube connected to a vacuum. Small amounts of mercury and water could be removed by heating the crystals in vacuum. Crystals prepared using this technique showed no evidence of slip. However, some of these crystals did show edge dislocation densities as high as l04 per cu cm. Some few selected crystals showed no dislocations. Single crystals of CdS were grown as large as 5 by 5 by 0.5 mm. The ampules for the growth of these crystals were 13 mm O.D., 11 mm I.D., 150 mm! LIQUIDUS SOLUBILITY MEASUREMENTS The CdS starting materials was G.E. 118-8-2 powder which was fired in H2S at 1000°C, and then a vapor transport technique5 was applied to produce a "sound" mass of CdS. The Wood's metal was prepared by weighing out bismuth, lead, tin, and cadmium in the proportions of 50, 26.5, 13.5, and 10 wt pct, respectively. The bismuth, cadmium, and lead were from the American Smelting and Refining Co. (ASARCO) and all had purities of 99.999+ pct. The tin was 99.9999 pct spectroscopic grade from the Vulcan Materials Co. The appropriate mixture was placed in a quartz tube, evacuated to a pressure of 5 X 10-6 torr, melted to a liquid, cooled to room temperature under this same vacuum. This ingot was then placed in another quartz tube, evacuated to 5 x l0-6 torr, and sealed off under vacuum. The ampule was then horizontally placed in a furnace. The temperature was raised to 600°C, and over a period of several hours the ampule was vigorously shaken several times. The ampule was then removed from the furnace, and the metallic liquid was

Jan 1, 1970

By M. A. Rabah

This paper focuses on the separation and extraction of lead and zinc from Egyptian ore at the Eastern Desert close to Red Sea. The ore sample contains cerussite, galena, hemimorphite and other impurities. Experimental program includes flotationof galena using xanthate collector and cerussite using oleic acid as collector and sodium silicate to depress hemimorphite tailing. Galena, and hemi-morphite are converted to carbonate with sodium carbonate prior to metal extraction. Lead and zinc are extracted pyrometallurgically at 1050°C and750°C or electrochemically respectively. The overall performance of the applied method amounts to 99.2%. Extraction efficiency of lead and zinc amounts to 98% and 94.5 % respectively. A model assumed that galena and cerussite in the ore are partially oxidized and can be liberated by grinding. Oleic acid selectively reacts with cerussite in a multistep sequence. Preliminary cost estimate shows that the price of the products is competitive to their local market prices.

Jan 1, 2000

By J. A. Botín

Identifying, quantifying and minimizing technical risks associated to investment decisions is a key challenge for the minerals industry decision takers and investors. However, risk analysis in most bankable mine feasibility studies are based on stochastic modelling of project NPV which, in most cases, fail to provide decision makers with a truly comprehensive analysis of risks associated to technical and management uncertainty and, as the result, are of little use for risk management and project optimization. This paper presents a value-chain risk management approach where project risk is evaluated for each step of the project lifecycle, from exploration to mine closure, and risk management is performed as part of a stepwise value added optimization process.

Jan 1, 2011

Jan 1, 1969

Jan 1, 1969