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By F. Mallozzi

"With the requirements of air pollution codes becoming more and more critical, the role of the ventilation engineer in industry has become an extremely important one. The ventilation engineer must provide the expertise for the selection of equipment and the design of systems which will reduce pollution to an acceptable level. Because pollution control systems involve considerable capital and operating expense, he must have the ability to select, from a wide range of products, the collector which will provide the required performance with a minimum of down time for maintenance. This paper discusses the principles of wet collection and the forces used to separate particulate matter from a gas stream. Important factors in the selection of equipment, such as collection efficiency, horsepower requirements and water consumption for the various types of collectors, are discussed in detail, as well as the construction features normally incorporated to circumvent potential problems."

Jan 1, 1971

By V. G. Papangelakis, I. Guzman, S. J. Thorpe

Pressure oxidation (POX) leaching has been commercially applied to refractory gold ores since the 1980s. In the process, sulphide material is oxidized and dissolved in order to release the gold encapsulated in the sulphide matrix, making the gold available for further recovery. Redox potential is a significant parameter during POX, however it is only measured ex-situ at atmospheric conditions. In-situ redox potential measurement would provide instant information that allowing adjusting the process parameters accordingly. In the present study, redox potential measurements were performed during POX of sulphide material at high temperatures and high pressures. The working electrode was an Ir electrode combined with a flow-through reference electrode in a novel and robust configuration. Experiments were performed in a 2 L titanium autoclave, at 200–220°C, 50–110 psi oxygen overpressure, and 10–30% solids mass fraction. Temperature and oxygen pressure conditions were selected in order to reflect POX of refractory gold ores. Results show that the Ir electrode is consistent with theory and is able to perform well under these conditions, showing promise as a new in-situ sensor.

Jan 1, 2017

By R. B. Allen, E. E. N. Smith, B. C. MacDonald

"LocationTHE MAIN PROPERTY and townsite of Eldorado Mining & Refining, Limited, Beaverlodge Operation, is on the north shore. of Beaverlodge lake, •about seven miles north of the deserted town of Goldfields on lake Athabasca, and four hundred and ninety air-miles equidistant from Prince Albert, Saskatchewan, and Edmonton, Alberta. Both Eldorado Airlines •and Canadian Pacific Airlines maintain scheduled flights from Edmonton to the 4,000-foot airstrip on the north boundary of the Company property. Bi-weekly flights are made by Saskatchewan Government Airways from Prince Albert. Men and priority freight are flown by Company aircraft direct from Edmonton. From June to-September, when the waterways .are ice-free, freight is carried by barge from the •railhead at McMurray, Alberta, down the Athabasca river to Chipewyan, and across lake Athabasca to• Bushell, the terminal on Black bay, lake Athabasca, a total distance of 265 miles.The Eldorado Operation has been connected by a 12-mile, all-weather highway to Bushell. Where the highway crosses the park-land along the Fredette river, the townsite, Uranium, is under construction to provide services to the uranium-producing district. Spur roads have been built to the Ace, Fay, Eagle, •and Martin mines of the Eldorado operation, and to the properties of Pitch-Ore Uranium Mines, Ltd., and Rix Athabasca Uranium Mines, Ltd."

Jan 1, 1954

By Karl J. C. Harries

"The following is an outline or check sheet of some things that might be looked at when considering the concept of a royalty or royalty provisions. It is not intended as a guide to what is right or wrong (that is for the parties to decide). It does not replace the discussions in this work and is not intended to be a complete or exhaustive compilation of matters to be considered. It presumes that you are familiar with the basic concepts of royalties and has been included as an aide memoire only. Use it as such, with caution and common sense, and amend it to suit your needs. Remember also that brevity can result in inaccuracy or imprecision. This is a danger in many points below.The section references are for the sections where the matter is the main subject of the section – see the index for other sections.Royalty TypesNet Proceeds [§§ 5.8.; 11.4.]• No accepted or understood meaning within industry.• Literal meaning, if not defined, would imply “net operating profit”.• Often defined as net profit with revenue base of proceeds from sale of product.• Should not be used without detailed and clear definition. Net Operating Profit [§§ 5.8.; 11.4. – referred to only]• Net profit without recoupment of pre-production expenditures.• Revenue base: usually proceeds from sale of product.• Operating costs deductible.• Some reserves may be permitted.• Recoupment of some post-production capital expenditures may be permitted.• Not often seen as basis for royalties in Canada."

Jan 1, 2003

By John A. Meech, Irene E. Cristoveanu

"The separation of hematite fines (100% - 38 um) from quartz fines (100% -75 um) was investigated using coarse hematite (-212 + 150 um) as a carrier in an oleic acid collector system. Oleic acid additions, per cent coarse additions, combined conditioning time and ferric chloride additions were examined.It was found that carrier flotation is a dominant mechanism for fine hematite recovery at oleic acid levels greater than 200 g/t and coarse additions of 10 per cent. Hetero-aggregation is favoured under conditions of lower levels of oleic acid in the presence of ferric chloride and at conditioning times greater than ten minutes.Ferric chloride acts as a depressant for pre-reagentized coarse hematite at oleic acid levels below 150 g/t. At concentration levels above 500 g/t, ferric chloride interferes significantly with the success of carrier flotation of fine hematite.The sensitivity of carrier flotation to changes in oleic acid addition, ratios of fine and coarse material, conditioning time and the presence of an inorganic electrolyte emphasizes that its application is only suitable when flotation conditions can be stabilized.IntroductionThe production of fine material is unavoidable during the comminution of iron ore . Due to the difficulty of recovering iron from this fine material by various gravity separation methods, flotation can often be a more attractive process producing high-grade material at high recovery. One variation of the flotation process, that can be useful for recovering ultra fine minerals (<10 um) , is known as carrier flotation. This process involves the utilizat ion of coarse particles ( > 150 p.m) which act as flocculation centres to which the fine hematite particles (10 - 40 u.m) become attached and are carried or "" piggy-backed"" into the froth phase."

Jan 1, 1985

By M. J. MONEA

Canada's only economic high-swelling bentonite deposit is mined in Wilcox, Saskatchewan, by A vonlea Mineral Industries Limited. The bentonite resulted from devitrified and altered volcanic ash released during the Laramide Orogeny. Volcanic ash was deposited in a pre-Cretaceous basin that was formed by salt solution collapse structures. Sedimentary rocks in the Wilcox area belong to the Snakebite Member of the Upper Cretaceous Bearpaw Formation. Bentonite beds are 61 cm, 46 cm and 30 cm thick; these are separated by 365 cm and 182 cm of shale respectively. Saskatchewan bentonite is used primarily in the oil well industry.

Jan 1, 1984

By Joseph W. Mills

"The Mic Mac property comprises 994 acres held under mining concession and approximately 2,000 acres held under development licenses in Bousquet township, Abitibi county, western Quebec.Gold was first discovered in this locality by Tom Duval in June, 1936. In February, 1937, Cassels Duval Mines, Limited, was incorporated under Quebec charter to take over the property. In September, 1937, the property was examined by Agaura Explorations, Limited, and an option agreement was signed on October 6th, 1937. In June, 1938, an option for complete ownership was taken on adjoining claims belonging to Norgold Mines, Limited, and in September, 1938, a mining plant was installed and the sinking of a 3-compartment shaft begun. In June, 1939, an option was taken by Mic Mac Explorations, Limited, on seven claims of Mooshla Gold Mines Company, Limited, that adjoined the Norgold block to the east. In December, 1939, a new company, Mic Mac Mines, Limited, was formed, amalgamating the Cassels Duval, Norgold, and Mooshla claims, as well as two claim groups to the west of the mine known as the Vaughan and West Cassels groups."

Jan 1, 1954

By D. L. Mckay

"Texasgulf&apos;s 25-foot-diameter Kidd Creek No. 2 Shaft is situated next to the company&apos;s No .1 Shaft, so that both shafts can share common surface and underground facilities. Shaft construction, covering a span of five years from collaring to commissioning , started by freezing the 50 feet of grey varved clay covering the bedrock. After first completing the concrete headframe, sinking to the 5/00-foot horizon utilized two different shaft construction methods. The top 2800 feet was excavated by slashing a 6-foot-diameter borehole and hoisting the broken material up the 3OO0-foot-deep No.1 Shaft; the remaining 2300 feet was excavated by conventional benching methods. Prior to installing the 5100 tons of galvanized shaft steel on the completion of the sinking operation, the sinker was able to install all thirteen armoured electrical cables in one pass with his modified galloway stage. All loading and lip pockets were installed along with the steel guides and bun tons during the equipping operation.IntroductionOn November 21,1978, Texasgulf completed the No.2 Shaft at Kidd Creek Mine. This shaft is the largest shaft (volumewise) in North America and is unique in many respects . The purpose of this paper is to outline some of the unique features of this shaft as well as to record how this facility was constructed.Soon after sinking No.1 Shaft, Texasgulf started a program of exploratory diamond drilling from the 2800 level in No. 1 Mine. The drilling outlined sufficient ore below that level to justify a second mine. Feasibility studies demonstrated that production could be increased by one-third exclusively from this deep ore . The study called for a 25-foot-diameter shaft to the 5100-foot horizon to service a mine producing 5,000 tons of ore per working day.The site chosen for No.2 Shaft and head frame is 350 feet southwest of No.1 Shaft. This, along with No.1 Shaft, will be away from the influence of ground subsidence when mining at depth. The collar and head frame foundations are located on a flat rock ledge some 70 feet below surface. This flat ledge was found by test holing through the overburden by diamond drilling. It is fortunate that this location best satisfied the surface layout and logistics for both surface and underground ore and waste handling."

Jan 1, 1981

By N. Dave

"IntroductionIn the mineral processing industry, certain minerals such as uranium, thorium and vanadium, etc., are selectively extracted with high degree of purity from solutions containing variety of other impurities by solvent extraction process. This separation process consists of bringing two immiscible or partially soluble liquid phases in contact, where mass transfer of one or more components takes place by ion exchange under certain controlling parameters such as pH, salt concentration and temperature. The processes taking place are primarily physical because the solutes being transferred are ordinarily recovered without chemical change. For efficient extraction more than one stage is required and the solutions are recycled counter currently through other stages. A multi-stage counter current solvent extraction plants essentially consist of four circuits namely: extraction, water wash, stripping, and solvent scrubbing. In extraction circuit the desired mineral is extracted at a certain pH from an aqueous solution onto an organic solvent by bringing them in contact with mixing in large mixer-settler units. The mixed solutions are allowed to settle and separate into component organic and aqueous phases in horizontal settlers by providing sufficient residence time. The solutions are then moved counter currently to other units- the aqueous from the underflow and the solvent organics from the overflow of the settler. The solution streams coming out of extraction circuit are loaded organics and barren aqueous or raffinate solution. The loaded organics are washed with water in a water wash circuit to remove transported impurities from extraction circuit. In the stripping circuit, the desired mineral is stripped selectively from the solvent phase back into an aqueous phase with a stripping agent at a selected pH. The solution stream leaving, stripping circuit are stripped organics and pregnant strip (aqueous) solution. The stripped organics are scrubbed off other loaded impurities with a conditioning solution in, scrubbing circuit and the cleaned solvent is re-used in the extraction circuit.A typical solvent extraction plant for uranium extraction and purification at Agnew Lake Mine is described in this chapter."

Jan 1, 1989

By Hiwa Salimi, Foley. Stephen R, Loghman Moradi

Common water-miscible organic solvents including DMSO, DMF, methanol, ethanol, acetone and THF in combination with HCl and low concentrations of H2O2 were investigated for dissolution of gold. THF showed the best result demonstrating that gold powder could be completely dissolved in 1.5 hours in 1.0 mol/L HCl and 0.2 mol/L H2O2. No complete gold dissolution was observed for the rest of solvents even after 24 hours. The rate of gold dissolution was measured by dissolving gold wire in THF/HCl/H2O2 mixture. The rate of gold dissolution in optimised condition was 75 g m-2 h-1 at room temperature and significantly increased to 590 g m-2 h-1 by raising the temperature up to 60 oC. Although a THF/HCl/H2O2 mixture showed a slower dissolution rate in comparison with aqua regia, however the applied conditions are considerably milder which is highly promising for gold leaching in an industrial scale.

Jan 1, 2015

By H. C. Cooke

"The Eustis mine, which was discovered in 1865, lies about 6 miles southwest of the city of Sherbrooke, Quebec, in lots 2, 3, and 4, range 9, Ascot tp. The mine was operated almost continuously from 1865 to 1937, under vanous owners.The ore was pyrite with varying amounts of chalcopyrite, a small amount of zinc blende and galena, and a low content in gold and silver. Parts of it ran quite high in copper, but the mine average would probably have been between 3 and 4 per cent. The high percentage of pyrite made it profitable to separate the ore into a copper concentrate and a pyrite concentrate. The latter was shipped away for manufacturing sulphuric acid, or was treated in a plant erected in 1887 at Capelton by the Nichols Chemical Company.The ore is a replacement of mica schist, which in turn was formed by the shearing of rhyolite flows. Ore also replaces the 'green rock', a term locally applied to carbonatized mica schists. The ore contained numerous 'horses' and smaller fragments of country rock, in all stages of replacement, ranging from those whose character was quite distinct to those whose ancestry could only be detected by an indefinite banding preserved in the sulphides."

Jan 1, 1954

By P. Müller, W. J. Przybylowicz, V. M. Prozesky, D. de Bruin

"Abstract -The common style of gold mineralization in the Basal Reef and other auriferous conglomerates of the Witwatersrand Supergroup comprises clusters of gold grains occurring interstitially to quartz and pyrite pebbles as well as gold accumulations attached to the surface of and infilling hairline cracks and fractures in round, compact, detrital pyrite. The latter textural association is generally interpreted as a result of remobilization of originally detrital gold during regional metamorphism and related hydrothermal fluid circulation. There is general agreement that, although gold solubility was Iow, ore modification occurred during the metahydrothermal event, but uncertainty exists about the magnitude and scale of gold remobilization .Micro-textural and chemical investigation of compact round pyrite and associated gold accumulation has established that gold is preferentially deposited on chemically heterogeneous pyrite surfaces. Combinations of alternating high- and low-As zones in pyrite result in mixed np-type conductivity accompanied by electron exchange and the reduction-driven deposition of gold at the cathode. This mode] does not explain gold accumulations which are not immediately linked to chemical heterogeneities and it is inferred that an additional control on the deposition of gold was exerted by physical defects which lead to zones of increased charge density and/or areas of focussed conductivity. The inferred precipitation mechanisms suggest that submicrocopic physical and chemical controls are the important factors that determined the sites of gold nucleation and that precipitation mechanisms based on physicochemical changes of the hydrothermal system were not involved. The Basal Reef contains on average close to 5% of round, detrital, chemically zoned pyrite which provides a large number of electrochemically active pyrite surfaces. The extremely effective electrochemical precipitation mechanism together with the low solubility of gold in Witwatersrand fluids is the main reason for the spatially restricted remobilization of gold in the conglomerate reefs."

Jan 1, 1994

By Jun Shan

The preparation of poly(N-isopropylacrylamide)-monolayer-protected clusters (PNIPAM-MPC) of gold nanoparticles was carried out in a homogeneous phase using three ways, in which three types of PNIPAM ligands were employed. The first type comprised of PNIPAMs with narrow molar mass distributions, synthesized by reversible addition-fragmentation chain transfer (RAFT) polymerization and thus, bearing a dithiobenzoate at the chain end. These polymers were used directly to passivate the gold nanoparticles upon the Schiffrin reaction in a one-pot synthesis. The second type of ligands was derived from the first one through hydrazinolysis, and they therefore contained a thiol end group. The third type of ligands was PNIPAMs obtained through conventional radical polymerization, post-modified to contain thiol end groups. The PNIPAM-MPCs were characterized by high-resolution transmission electron microscopy, and by dynamic light scattering. The one-pot synthesis utilizing the ligands of the first type turned out to be a simple and facile method, compared with the other two ways, with which the size of the gold nanoparticles can be easily manipulated mainly by adjusting the molar ratios of PNIPAM / HAuCl4. PNIPAM is a more efficient ligand to stabilize the gold nanoparticles in water and in organic solvents than alkanethiols. The surface density of PNIPAM chains ranged from 1.8 to 3.6 chain/nm2, which is much lower than that typical for alkanethiols. The thickness of a PNIPAM monolayer bound to the gold core is somewhat larger than the size of the random coil of corresponding free PNIPAM in aqueous solution, which suggests that the conformation of a PNIPAM chain bound to the gold core is to extended.

Oct 1, 2003

By Takahiko Okura, Chiharu Tokoro, Shunta Shibuya

Copper concentrates bearing arsenic as enargite (Cu3AsS4) or tennantite (Cu12As4S13) have disadvantage on copper smelting environmentally and economically, so those concentrates should be treated decently. In this paper, some experiments of arsenic removal from several kinds of copper concentrates by heat treatment in an inert atmosphere were discussed. In this study, several samples in which arsenic contents varied 0.18 to 8.3 wt % were heated to the temperature range of 400 °C to 700 °C under argon gas flow during 1 hour. The heat residues and volatilized materials (V.M.) were analyzed by XRD, XRF, ICP-AES, XAFS and MLA to analyze residual concentration of arsenic and residual mineral composition. Experimental results showed that arsenic removal from copper concentrates advanced in 2 steps; at the temperature range of 500 ° C to 550 °C and of 550 °C to 650 °C, respectively. MLA analysis showed that a few of arsenic in enargite was removed by 550 °C heating and a majority of it was still combined with copper and sulfur as tennantite. On the other hand, arsenic in tennantite was removed almost completely by 650 °C heating. Furthermore, a particular type of arsenic mineral, such as Cobaltite (CoAsS) still remained after heating at 700 °C.

Jan 1, 2014

By D. H. Whiting, A. J. Sinclair

"The San Antonio gold mine, in the Archean Rice Lake Greenstone Belt, was in continuous production from 1932 to 1968, during which period the mine yielded 1.35 million ounces (41 985 kg) of gold and 195,000 ounces (6065 kg) of silver from 4.88 million tons (4.44 Mt) of ore for average grades of 0.28 ounces per ton (9.0 g/t) gold and 0.04 ounces per ton (1.3 g/t) silver. After the 1968 closure, several publications appeared implying that gold grade of ore decreased with increasing depth.Recent observations by the authors show that the relatively low grades in the middle levels of the mine are controlled geologically. However, the reduced average are grade in the lower levels of the mine was due to inadequate mining control that led to a high level of dilution. The similarity of geological factors such as vein styles within structural domains, variations in mining control, recent ore muck sampling results and lithogeo chemical profiles all indicate the similarity of ore in both upper and lower levels.Strict mining control should yield improved grades of gold for future production from the lower levels at San Antonio, in contrast with the production grades of the late 19605that represented extensively diluted ore. High-quality production records are an essential component in re-evaluating a past producer, particularly where workings are accessible. Poorly founded conclusions which cannot be checked by access to workings can have a negative impact on the evaluation procedure."

Jan 1, 1990

By W. P. Clement, M. C. Bell

"PT International Nickel Indonesia, a subsidiary of Inco Limited, mines and processes nickel laterite ores for the supply of nickel matte, primarily to the Japanese market. It has large ore reserves and a substantial portion of its energy requirement is supplied from its own hydro-electric plant. Since the early 1980s, the company has increased its production level and has dramatically reduced its unit costs. Projects are in progress to further decrease costs and recently a major expansion - 31% increase in nominal capacity - has been undertaken, consistent with marketing opportunity in Japan. The expansion project, which has a low capital cost per unit of production, is outlined in this paper.IntroductionPT Inco is a fully integrated mining and smelting facility located in Indonesia on the island of Sulawesi, and is based on a very large, world class nickel laterite orebody located 60 km inland from the Gulf of Bone. Construction of a single furnace line plant started in 1973. Complete supporting infrastructure of townsite, hospital, schools, shopping centres, road systems, and an airport, all had to be built. Prior to completing the first phase, the decision to expand the plant to a three kiln-furnace line was made. A major factor in the decision was the ability to develop the hydroelectric potential of the nearby Larona River. This involved construction of a 30 m high dam, and a 7 km surface canal to the generating station. The penstocks drop 150m to three 55 Mw turbines. The power generated is sufficient to supply the entire electrical load for the complete smelting complex and town site. The unique capability to supply the high smelting energy requirements for a lateritic nickel ore by fully captive hydro power, has, over the past decade insulated PT Inco from the large upward swings in oil prices suffered by other lateritic nickel producers.Wet ore is first partially dried to 20% moisture in two co-currently fired rotary driers, screened on an internal trommel, and delivered to two dry ore storage buildings. The partially dried ore is calcined at 750°C and partially reduced in three large 100 m by 5.5 m kilns. Reduction is accomplished by lancing bunker C oil onto the ore bed at the discharge end of the kiln. The kilns are equipped with five air pipes with onboard blowers to gradually bum the combustible gases generated by the lance oil. The kilns were also originally equipped with two side burners to supply additional heat close to the feed end of the kiln. Calcined ore is transferred by 20 ton refractory lined containers to the furnace feed bins at the top of the smelter. Smelting is carried out in three 18 m diameter furnaces powered by 45 Mw transformers. Slag is skimmed at 1550°C through water cooled breast plates and removed to the slag dump by Haulmaster Carriers. Matte of 25% Ni, 1O% S, 65% Fe is tapped at 1350°C into the converter aisle. The iron is slagged in converters to produce a nickel sulphide matte at 80% Ni. Three rotary Kaldo converters were initially installed. The product matte is granulated, dried, and shipped in returnable rubber bags to Japan. Most of the matte is fluid-bed roasted to nickel oxide and then directly reduced in a second reducing fluid bed reactor at our facilities in Japan.The smelting complex was designed to smelt two million tonnes of dry ore per year, equivalent to 34 000 tonne nickel production at an ore grade of 2% Ni. The cost of the entire project including working capital was $900 million."

Jan 1, 1990

By C. A. Godden, T. N. Atkinson

Metallographic techniques developed from studies of corrosion films and electroplated coatings on aluminum alloys, are discussed. These methods, the result of work at the Naval Research Establishment, give extremely flat fields at the interface despite the large difference in hardness, in cases where electrolytic polishing could not be used. Techniques to overcome the similarity in refractive index of immersion oil, aluminum ox1d~, and mounting plastic are also considered, as are methods for improving visual and photographic contrast in areas of interest. A device for marking specimens is also de-scribed. ELE•CTROLYTIC polishing is inapplicable in the study of surface films on aluminum, whether corrosion films or electrodeposits. Mechanical procedures must there-fore be followed. Sample flatness and edge retention are of paramount importance. Special measures are required in the study of oxide films, because of the close proximity of the refractive indices of aluminum oxide, mounting plastic, and immersion oil.

Jan 1, 1962

Over five decades ago Kitchener et al. in England and La Mer and Healy in the US were advancing our understanding and application of flocculants, primarily related to water treatment and filtration of sludge. This paper identifies the beneficial uses of flocculants in the mining industry and revisits some less common applications. Flocculants are used in unique ways in the mining industry: (a) to increase the effectiveness of sediment ponds to achieve discharge compliance; (b) to eliminate air emissions and costs associated with concentrate drying; (c) to reduce soil loss by applying a flocculant to erodible soils, and reducing wind and water erosion; (d) to remove/thicken metal precipitates in acid rock drainage treatment; and (e) to enhance clarification and thickening in mineral processing. The mechanism of how flocculants, coagulations and potentially-determining ions function is described and how their use can be controlled and optimized. The potential toxic effects of flocculants and coagulants when used to achieve effluent permit discharge compliance for suspended solids is discussed. A review of numerous best management practice (BMP) guidelines for the design and operation of sediment ponds treating mine site runoff to remove total suspended solids (TSS) and turbidity revealed an absence of a predictive methodology to address regulatory discharge compliance for TSS concentration, turbidity and flocculant-induced toxicity. An essential component of pond design is the need to determine whether flocculants and/or other settling aids are required based on the site-specific soil particle size distributions upslope of the pond. Procedures are provided to predict the need for settling aids and estimate pond discharge quality and downstream impact for TSS, turbidity, toxicity and sub-lethal toxicity. The need to develop BMP guidelines are discussed for (a) the application of flocculant to surface soils to reduce erosion, (b) the application of flocblocks to increase the ?sediment control? aspects of erosion and sediment control, as part of the mine site erosion control plan strategy and (c) the application of flocculant to silt socks to treat mine site runoff.

Aug 1, 2013

By M. M. Papic, J. C. Whitehead

"A sample of Hat Creek coal has been extracted using supercritical aromatic solvents in a laboratory-scale extraction unit. Extract yields in the range 21 to 27 per cent of dry-ash-free coal were obtained.These yields are significantly lower than yields available from high-volatile bituminous coal and some lignites. Thus in a combined extraction/hydrocracking scheme for the production of liquid fuels from Hat Creek coal the char product would exceed that required to supply the energy and hydrogen requirements of the process. Excess char could be used for the production of power or gas for sale.IntroductionThe Supercritical Gas Extraction (SGE) process developed by the U.K. National Coal Board operates in two stages: an extraction stage in which a coal extract is separated from mineral matter and undissolved coal and a subsequent hydrogenation stage.The SGE process is based on the ability of a compressed supercritical fluid to dissolve relatively high molecular weight substrates. It is operated under conditions where a hydrogenrich portion of the coal is selectively extracted in a short residence time extraction stage leaving a hydrogen-deficient char. The process schematic shown in Figure I illustrates that the extract produced is separated from the solvent and passed to a conventional hydroprocessing stage while the char product is utilized in the production of heat, power and hydrogen for the process. An extract yield in the order of 50 per cent dry-ash- free coal represents a ""balanced"" case for such a process in which there is just sufficient char for the provision of these commodities."

Jan 1, 1985

By R. Lepsoe

THE basic products of the Company are lead and zinc. Around 950 tons of these metals are produced each day. In the last decade, the production of by-products has increased to such an extent that they are now nearly as large an operation as that of the base-metals. Of the many by-products, the most objectionable, and up to a few years ago the most useless, was the sulphur. The sulphur comes in the ore from the Sullivan mine. It is in the ore chiefly as lead, zinc, and iron sulphide. By selective flotation, these three sulphides are separated into the lead concentrates, the zinc concentrates, and the mill tailings. These mill tailings contain the bulk of the iron and the sulphur. Of the 6,5CO tons of ore treated per day, about 5,000 tons are wasted to the mill tailings dump, and, while now valueless, will in future years undoubtedly be used as a source of iron and sulphur. There are about 1,500 tons of sulphur and 2,200 tons of iron per day in the tailings, and about 400 tons of sulphur left in the total concentrates. In the recovery of lead and zinc in the pure form, the first step is the roasting operation. Here, most of the sulphur is driven off as sulphur dioxide in the gases. It has always been apparent that this sulphur should be. recovered and utilized if possible. At the same time, this would cure the smoke nuisance. Because, in spite of the low concentration of sulphur dioxide and the fact that it was discharged from stacks at an elevation of 400 feet above the plant, some damage to vegetation close to the plant had been experienced.

Jan 1, 1937