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By Paul Worsey, Attila Szucs, Mark F. C. Schmidt

The use of water stemming with small explosive charges has been successfully developed at the University of Missouri-Rolla. Further study has been started to explore additional applications and uses for water stemming. Work has been focused in the areas of rock excavation and perimeter blasting techniques. This ongoing work is being conducted internally as graduate and undergraduate level research. A paper describing the general technique for use with boulder fragmentation was published in last year’s proceedings. The excavation work took place in the form of both small bench and trench shots while pre-split and trim shots comprised the perimeter-blasting portion. The excavation work was located in both the Roubidoux sandstone formation and the Jefferson City dolomite formation. Hole diameters ranged in size from 1.5” to 3” diameter and powder factors ranged from l/l 0 to l/3 of the conventional charge weight. The water was used to both couple the explosive charge with the surrounding borehole, as well as stemming. No other stemming material or products were used. The method of using water stemming in perimeter blasting was found to work well. However, when used for excavation purposes, adequate rock movement was not obtained. The rock was often split, but not fragmented. The airblast was surprisingly low, but the ground vibrations were excessive. This is not the case with trim shots, and therefore this method appears ideal for splitting purposes, especially in areas where people do not like to stem the tops of the holes, and where airblast and vibrations are of concern. Water coupling dramatically increases splitting and fracturing. However, fragmentation is not only fracturing but also the movement and swelling of the material. As the charge weights are reduced with this method, sufficient gas volume is not created to cause this necessary movement and swell, which is critical in rock excavation. This explains why its successful application was limited to perimeter blasting applications.

Jan 1, 2000

By Alvaro Fernández-Baca, Richard Chaiña

El proyecto Ayawilca, controlado al 100%por Tinka Resources y ubicado en el Perú central, contiene un expectante descubrimiento de zinc-estaño en un sistema de reemplazamiento en carbonatos (CRD), así como un recurso de plata cercano a la superficie. Basado en sondajes completados hasta finales de 2015, Tinka anunció un estimado de recursos minerales inferidos para la zona de zinc de 18.8 millones de toneladas con 8.2% de zinc equivalente, y un primer estimado de recursos minerales inferidos para la zona de estaño de 5.4 millones de toneladas con 0.89% de estaño.

May 1, 2008

DeLamar, the largest silver mine in the United States outside the Coeur d’Alene District in Idaho, came into production in April 1977. It is a joint venture of Earth Resources Company and Canadian Superior Oil Co., with Earth Resources being the operator. The property is a consolidation of several old operations dating back to the 1800s; the better known of these are the DeLamar and Sommer-camp mines. DeLamar is situated in southwestern Idaho near the Oregon border in the Owyhee Mountains, 88 km (55 miles) southwest of Boise, Idaho. Heavy snowfalls and very low temperatures are common in the wintertime. Planned production was 77,500 kg (2.5, million oz) of silver and 5 13 kg (16,500 oz) of gold per year. Expected life at a milling rate of 1,540 mtpd (1,700 stpd) was 20 years, with an ore grade of 144 g Ag and 1.58 g Au per mt (4.2 oz Ag and 0.046 oz Au per st). Present production is at the rate of 1,900 mtpd (2,100 .stpd). The total cost to bring the mine into production was reported to be approximately $22 million, and the mill was reported to have cost about $10,000 per daily ton of design capacity. At DeLamar, naurnanite (Ag2Se) is the principal silver .mineral occurring with acanthite (Ag2S) disseminated through quartz veins and a silicified rhyolitic ash-flow tuff of Miocene age. The mineralization is relatively iron and manganese free and is readily amenable to cyanidation giving recoveries of 85% silver and 92% gold, which occurs free. When the plant was started, considerable difficulties were encountered in handling the wet sticky ore because of the high clay content. Following primary crushing, the ore is stockpiled. Due to the sticky nature of the ore, the vibrating feeders initially installed under the stockpile were replaced by a reclaim hopper and an apron feeder which was fed 24 hours a day by a front-end loader. The minus 200 mm (8 in.) ore is fed to a 5.5 m by 2.75 m (18 ft by 9 ft) autogenous mill followed by a 2.75 m by 4.57 m (9 ft by 15 ft) ball mill and cyclones. The mills are equipped with 750 kw and 520 kw (1,000 hp and 700 hp) motors. Both mills discharge to the same cyclone feed sump and a portion of the cyclone underflow is returned to the autogenous mill. Cyclone overflow, which is 20% plus 74 microns (200 mesh), is leached in four agitators in series for 72 hours (45 hours at the 2,100

Jan 1, 1981

By M. R. Geer, H. F. Yancey

With the exception of pneumatic processes and a few special beneficiation methods of comparatively limited application, all mechanical coal-cleaning and mineral-dressing processes involve the admixture of solid particles with a liquid, generally water. Upon completion of the beneficia-tion process the solids must be reclaimed from the water. With coarse particles separation of solids from water is accomplished readily by the simple expedient of screening, but recovery of fine solids presents a much more difficult problem. Thickening and dewatering are practiced so generally, both within the mineral industry and elsewhere, that they have been the subject of study through the years and a number of processes for making the solid-liquid separation have been developed. Dorr-type thickeners have been standard equipment lor many years, and more recently driers employing centrifugal force have gained in popularity.' A cyclone device employed to thicken the loess suspension is a novel feature of the loess heavy-medium coal-cleaning process developed by the Netherlands State Mines in the thirties. The cyclone thickener was first mentioned by Driessen2 in 1939 in a paper describing the loess process and has been the subject of two more recent reports by the same author.3 The cyclone does not differ greatly from cyclone dust collectors used in this country for years, except that it is operated wet rather than dry. Advantages of the cyclone over other thickening devices are its simplicity of construction, high capacity, and the fact that, having no moving parts, it is simple and inexpensive to operate. The Bureau of Mines became interested in the cyclone after one of the present authors observed it in operation in the Netherlands in 1945. The first experimental work with the cyclone conducted by the Bureau of Mines was an investigation of the unit functioning as a heavy-medium cleaner for fine coal—a use for the cyclone that was conceived long after it had been perfected as a thickener. The results of this investigation were published last year.4 The present investigation deals exclusively with operation of the cyclone as a thickener. Object and Scope of Investigation Although the cyclone thickener has been dealt with in three reports by Driessen, its use as a thickener was not the principal subject of any of these reports and therefore neither the cyclone itself nor its operation was described completely. The design of the cyclone was described only in a very general way, only limited operating results were given, and virtually nothing was disclosed about the influence of the various possible operating variables. Moreover, the Netherlanders apparently had employed the cyclone exclusively for thickening loess and had not employed it On coal 'lurries. One English colliery has used a cyclone to thicken coal slurry,5 and as it appeared that

Jan 1, 1949

By Francis Perron

The Directive 019 on mineral industry, which was officially adopted in 1989, is the main tool used by the Quebec Ministry of Environment for the environmental evaluation and control of mineral exploitation projects presented for approval. This document contain the environmental guidelines which must be met by all the mineral companies during and after exploitation. The Directive 019 also contain all the information required by the Quebec Ministry of Environment in order to fully analyse the environmental impacts caused by the mineral projects proposed to the Ministry for approval. The proposed modifications of the Directive 019 cover many new topics that were not addressed in prior editions, these include : water recirculation, groundwater protection, tailings management and post exploitation environmental controls.

Nov 1, 2002

By David Norval

The increased tendency to use hydrometallurgy as an extraction step towards final refining following a pyrometallurgical smelting process requires melts which have to be cast and or comminuted which can then be taken up into solution or leached. To generate surface area for optimal leaching, a common process is to granulate and in many cases mill, which is a general practice with friable material with properties associated with higher sulphur, carbon and silicon levels. However, in the case of metallic melts milling can be unfeasible due to the ductility of the granulated particles. Water solidification is capable of producing either fine particles by means of granulation or atomisation. Atomisation however, can produce particles as fine as 40 microns without a milling circuit.

Jan 1, 2005

By Andrés Gonzalo Reyes Unda

Este trabalho de investigação foi realizado com a intenção de obter uma rota alternativa de lixiviação de ouro para minérios de baixo teor para uma planta industrial que adsorve ouro atualmente pelo processo CIL (Carbon in Leaching) na região da Ilha Cristalina de Rivera (ICR). Foram realizados testes em garrafões de 20 litros para estabelecer critérios de operação e dimensionamento de uma planta de lixiviação em pilhas permanentes para tratar 1.000.000 toneladas de minério por ano de minério de baixo teor (0,45 - 0,65 g/t) com granulometria menor que 12mm e com produção anual podendo alcançar de cerca de 455 kg de ouro.

Aug 16, 2017

The underwater explosions have been used for several decades for the determination of energy content of explosives. This test method is the only one by which shock, fly rock and total energy can be measured separately. It enables to develop the energy content of explosives for different purposes; the military high explosives should contain more shock energy whereas the fly rock energy is very important feature for the blasting explosives. The purpose of this study was to establish a test place for the underwater experiments in an old water filled open pit, to estimate the repeatability of the energy measurements, to calculate the standard deviations and to compare the total energy content of commercial explosives to the theoretical values calculated by the thermo dynamical program. The test method seems to work quite well especially for the homogeneous explosives. The difference between the energy content of theoretically calculated and measured values for the homogeneous emulsion explosives was about -3% and for the dynamite about -5%. The energy difference for the unhomogeneous ANFO was detected to be -19%. The diameter of the charge was 15 cm and it was obviously too small for ANFO to detonate ideally. On the basis of the test results this test method seems to fit best for the cap sensitive, completely homogeneous, spherical charges whose masses are big enough and which are initiated from the centre of the charge. The underwater explosions as a test method is the only one which can separate shock, bubble and total energies from each other and therefore makes it possible to develop the properties of explosives for different purposes which is very important feature of this method.

Jan 1, 2005

By C. J. Coulson, G. A. McKay, A. M. Cormie

"The Leitch mine, in the Beardmore area, Thunder Bay district, Ontario, is a part of the Leitch estate. It includes a group of patented claims, staked by P. A. Leitch of Port Arthur early in the century, along bands of jasper, hematite, iron-bearing slates, carbonatized schists, and greywacke, commonly known as 'iron formation'. The folding and shearing in these variably interstratified beds is a striking structural feature. The vein pattern is very definitely related to these structures.ROCKSThe rocks in the area are typical Keewatin lavas-greenstone-overlain by a series of Timiskaming sediments consisting of arkosic greywacke, iron formation, slate, greywacke, conglomerate, and greywacke, in this order from the contact. The sedimentary beds are generally conformable with the older rock contact except on local folds.Small intrusions of diorite occur as nearly vertical dykes and are believed to be of Algoman age. To the northwest of the mine there are larger intrusions of granodiorite belonging to the same period.Several dykes of quartz diabase up to 100 feet in width, striking generally north and south, with a vertical dip, cross the property. These have been classified by W. C. Laird as pre-Keweenawan in age. One of these dykes in particular is of importance for it cuts through the No. 2 vein and serves as a pillar when stoping. It is approximately 6 feet wide.The youngest rock type represented is a sheet of olivine diabase that extends over an area of thousands of square miles in and around lake Nipigon. It underlies most of this area and its outcrops form some of the most interesting topographical features, with fault scarps rising hundreds of feet. In the vicinity of the mine shaft the sheet is about 600 feet thick and extends from a depth of 1,871 feet to approximately 2,470 feet. It has a dense, hard, chilled margin and a sharp, tight contact. It intruded the sediments with very little lateral disturbance. The chief effect was baking, which darkened and hardened the greywacke near the contact. The sheet dips flatly to the cast."

Jan 1, 1954

By Simon D. Strauss

This paper discusses the ways in which the metals industries dealt with surplus capacity in the past, primarily the twenty years, 1950 - 1970. During this period, the bulk of metals production in the market-economy countries was controlled by private sector companies. Since then, public sector companies have played an increasingly important role in the metals industries of the market-economy countries. Of course, the change was not overnight. It occurred gradually over a period of many years. Indeed, by 1965, several developing nations had adopted policies of government control of metal operations previously in private hands. The metals industries are so many, so large and so diverse that the constraints of time make it impossible to deal with all segments. To illustrate the problems of surplus capacity in the metals industries during the period, this paper focusses primarily on three metals. One experienced rapid market expansion between 1950 and 1970; one had a growth more or less in line with the expansion of the world economy; and one lagged well behind the pace of industrial expansion. No one will be surprised if aluminum is identified as the metal with rapid growth; copper is identified as the metal that kept pace with the world economy; and tin is identified as the laggard. Individual computations of the rate of growth for these metals show some variations, but most estimates place the increase in aluminum consumption in the market economy countries between 1950 and 1970 at a compound annual rate of about 8%. The corresponding figure for copper is a compound rate of 4%. Tin consumption showed no growth. A widely accepted point of view is that sharply rising demand will force the price of a commodity to increase substantially in order to attract a necessary expansion of supply. The corollary to this is that stagnant demand will hold prices down, as there is no need to stimulate new supply and it may be necessary to force marginal producers out of the market. The experience with aluminum, copper and tin in the years 1950-1970 does not support this theory. In 1950, the average prices in the United States market as reported by the American Bureau of Metal Statistics were 17.74 a pound for aluminum, 21.24 a pound for copper, and 95.54 a pound for tin. The corresponding figures in 1970 were 28.74 a pound for aluminum, 57.74 a pound for copper and 174.24 a pound for tin. Aluminum, with the greatest market growth, rose by 60%, whereas tin rose 82% and copper by 172%. Clearly, other factors were a more important influence on price than was the trend of demand To revert now to the capacity issue, during the period under review aluminum demand was rising so sharply that that industry had no sustained problem with surplus capacity. Instead, if anything, capacity may at times have been a constraint on demand. The industry succeeded in attaining explosive market growth through imaginative research and promotion programs. Some of this growth was the result of substitution of aluminum for other commodities – steel, copper, wood and glass as examples. But much of the growth was the consequence of developing entirely new products through technological innovation. Prior to World War 11, the number of aluminum producers was extremely limited. Alcoa and Alcan in North American and Pechiney and Alusuisse in Europe were the dominant firms. The developing countries had virtually no production and in Japan and Germany production was only a small fraction of the North American output. During the War years the United States had financed the construction of many new facilities. After the war these plants were sold by the government to Reynolds and Kaiser in order to foster more competition in the market. Subsequently many other concerns became aluminum producers, both in the United States and elsewhere and competition for markets intensified.

Jan 1, 1989

By Mohab Sabry, Emre Biringen

"Over the past decades, there has been an increasing recognition of the strategic use of pile-supported rafts in design of heavily-loaded structures to reduce total and differential settlements. However, such a hybrid foundation construction method has not been widely utilized in many countries, including the U.S., due to code limitations. In current practice, the foundation design is based on either (i) the bearing capacity of the raft supported by the subgrade soils, or (ii) solely the pile capacity. This paper addresses the potential use of such hybrid piled-raft systems, where the structural load is shared by both by the piles and the subgrade soils directly beneath the raft, in foundation design of power plant structures. The study assessed some of the characteristics of piled-raft behavior by undertaking three-dimensional finite element analyses of two raft sizes, and various pile layout patterns, including the rafts with piles distributed evenly and only in the central area of the raft. The computer software PLAXIS 3D with the Mohr-Coulomb constitutive soil model was used to facilitate the modeling of such cases. The results presented in this paper indicate how strategically locating the piles could reduce the differential and total settlements.INTRODUCTIONUnder the high applied loads coming from the superstructure, the addition of piles is primarily required to improve the factor of safety against failure when a raft does not provide adequate bearing capacity. If the bearing capacity of the raft is sufficient to carry the total load with a reasonable safety margin, then the addition of piles is usually intended to reduce the settlements to an acceptable amount. In current practice, the design process for pile-supported raft foundations conservatively concentrates on providing solely the pile capacity to carry the total structural load without taking the contribution of the raft into account. This paper addresses the potential use of hybrid piled-rafts in controlling foundation settlement of power plant structures by considering the support of both the piles and the subgrade soils beneath the raft to carry the load."

Jan 1, 2017

By Joshua Hoffman, Braden Lusk, Kyle Perry

This study explores how different geometries affect the shock tunnel’s performance and is broken down into three phases. Phase 1: Open arena testing in which theoretical energies produced by the detonation of small (<1.0 kg) charges are compared to actual experimental values. Phase 2: Shock Tunnel testing in which realized energies are experimentally recorded for the UK and S&T shock tunnels. Phase 3: Analysis of the data to begin building a pressure-forecasting tool for the UK shock tunnel given the explosive characteristics, sample size and desired pressure and impulse.

Jan 1, 2009

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By Herbert H. Kellogg

DEPOSITS of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine (-325-mesh) sand assaying 95.5 per cent Si02, and a very fine (-3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. EXPERIMENTAL PROCEDURE The Sample.-The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotation Feed.-By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine fraction was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper.

Jan 1, 1945

The Bowen Basin of Queensland and the Hunter Valley of New South Wales are the two largest black coal producing areas in Australia. Mining activities require the availability of large on-lease storages for water and tailings. The legislative framework for the regulation of large dams is slightly different between the states. The paper provides a broad outline of the regulatory framework in each state, with particular reference to Queensland and New South Wales. In New South Wales, water storages and tailings dams are treated similarly by the regulator. Under the Queensland Water Act 2000, now proclaimed, referable dams are restricted to water supply dams the failure of which involves population at risk downstream. Tailings dams are excluded, being deemed hazardous waste dams regulated under the Environment Protection Act 1994. These storage facilities are required to include a ædesign storage allowanceÆ so that the possibility of spillway discharges is minimised. New South Wales is moving toward this requirement. A curious anomaly is emerging in Queensland: the legislative definition of hazardous waste û material resulting from the processing of minerals û does not extend to coal tailings since coal is not included in the definition of a mineral. The paper explores this and other ramifications for mining operations of the new legislation as well as general trends in criteria affecting the cost of providing water and tailings storages.

Jan 1, 2003

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 H. Do, R. Yoon

Mineral flotation is effective over a relatively narrow particle size range of approximately 10 to 100 µm. If the upper particle size limit can be increased, however, a run-of-mine ore may be coarse ground, and only the middlings are ground further to achieve required grades. This will result in significant energy savings and increased throughput. In this communication, methods of increasing the upper particle size limit will be discussed on the basis of a flotation model that can predict flotation from both hydrodynamic and surface chemical parameters. An example of increasing flotation recovery by control of a chemistry parameter is given.

Jan 1, 2007

By L. Ya Agapova

Emissions from pyrometallurgical processing of copper concentrates are usually subjected to wet scrubbing. The washing sulphuric acid so produced contains up to 0.2 g/L of rhenium, which is recovered by solvent extraction in the form of crude ammonia permeate (NH4ReO4). The mother liquor, obtained from the sulfuric acid wash, contains up to 35 g/L of rhenium and significant amounts of impurities. Crude rhenium salt is commonly processed by double recrystallization to produce marketable salt. A ton of high grade NH4ReO4 can be produced using 3.0-3.5 m3 of mother liquor. We propose an efficient technology of mother liquor pre-treatment aimed at removal of water-soluble organics with subsequent electrodialysis. Electrodialysis of mother liquor in a multichamber electrodialyzer resulted in the production of pure rhenium acid, from which a marketable NH4ReO4 is precipitated using ammonia solution. Multiple steps of recycling consisting of dissolution of NH4ReO4 salt, rhenium acid production from NH4ReO4 salt by electrodialysis and precipitation of NH4ReO4 from rhenium acid made it possible to upgrade the purity of obtained rhenium compounds by 1-2 orders of magnitude. A rhenium metal powder of 99.997% purity was produced by high-temperature reduction of high-purity NH4ReO4 with gaseous hydrogen. Rhenium metal powder of high purity can be produced from sulfate solutions of NH4ReO4 by non-platinum anode-based membrane electrolysis processes also. Keywords: rhenium, mother liquors, membranes, electrodialysis, rhenium acid, ammonia perrhenate, reduction, powder, electrolysis

Sep 1, 2012

By Howard M. Graff, Sidney C. Bouwer

The economics of mining has traditionally been viewed quite apart from the economics of blast furnace operations. It was realized, of course, that blast furnaces would operate best with good raw materials, but the shipping grade of ore, for example, was decided by a sort of collective bargaining process between the ore producers and the blast furnace operators. The mine operators had to keep their costs down, and were usually successful in convincing management that the ore in any particular deposit was endowed by nature with a certain iron content and that the mine grade could not be raised beyond a certain economic limit. The blast furnace operators would have liked a better grade, but as long as they were able to make a fairly good showing it was more or less accepted that they had to take the ore as it came. The mine operators were able to ship to the nation&apos;s blast furnaces thousands of tons of iron ore that would be completely unacceptable by present-day standards. As late as the early 1950&apos;s, the major steel companies were witnessing an actual decline in the iron content of their incoming ore supply as the mining companies mined progressively leaner material and upgraded it only enough to pass competitive market requirements

Jan 8, 1965