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By J. A. Finch

The use of pneumatic methods for measuring interface level in flotation columns is analysed. It is shown that errors can vary over a wide range (e.g. 40 cm. in one industrial application) depending on bubble size, particle size, bubble loading, liquid hold-up and bulk slurry density. Accurate measurements (+ 5 cm) can be obtained by measuring intensive properties of the froth near the interface. Methods based on the measurement of temperature and conductivity are described and analysed. Control of the interface level using these measurements subject to various disturbances is demonstrated.

Jan 1, 1988

By R. Sharma, S. Svoronos, S. Gupta, N. A. Abdel-Khalek

This paper presents the results of a column flotation study on a fine feed sample of Florida phosphate. A central composite design was used to study the main and interaction effects of parameters affecting column efficiency. The statistical design is also used to determine the optimum conditions. The results indicate that, depending on the experimental conditions, it is possible to produce a high-grade concentrate with a high recovery. A 15ppm concentration of nonionic frother was found sufficient to maximize the recovery at pH of 9.5.

Jan 1, 2002

By Thor Skjeggedal

The Norwegian contracting firm A/S Høyer-Ellefsen has for the last 2.5 years carried out substantial pregrouting in a 7600m long tunnel just outside the city of Oslo. This has been done to avoid groundwater lowering and consequent damage to surface structures. A WIRTH TBM is used in combination with a 2 boom Atlas Copco hydraulic jumbo for the drilling of groutholes in front of the machine. The detailed procedure for the drilling and grouting will be discussed, and the results obtained will be presented.

Jan 1, 1981

By S. Behl, S. Mathur

The viability of physico-chemical purification methods such as selective, flocculation for the separation of two components having similar chemical compositions and the same crystal structure but different surface properties is demonstrated. A loss of selectivity was observed in a dolomite-dolomite mixture when polyethylene oxide (PEO) (five million molecular weight) was used us the flocculant. Heteroflocculation was minimized using a low molecular weight PEO (MW =105) as a site blocking agent (SBA). Near complete separation was achieved at pH 10.5 at the 100% recovery level

Jan 1, 1996

By Paul D. Chamberlin

In the author's selenium report of 1994 (1), the literature was culled for selenium removal technologies that had a chance of producing effluents with <10 ppb of total selenium. Very few processes could achieve such low concentrations. The EPA dictates such low concentrations. Discharges to fishery waters must be <5 ppb total selenium. The federal max­imum contaminant level ( MCL) for drinking water is 50 ppb. The EPA regulations are for total selenium, not for the species, perhaps because an approved EPA assay method is atomic absorption and it cannot distinguish between species.

Jan 1, 1996

By F. Gray

The industrial diamond review arrived too late for the June issue. Since the economy is showing a resurgence of strength, diamond tool consumption is similarly on the rise. Tool makers indicate that sales are picking up. They expect that 1994 will see about an 8% increase. Fabricators expect this positive trend to continue through 1995. In the industrial sector, sales of grinding wheels, honing stones, cutting tools and the like are expected to grow 5% and 8% during the next year. Users still are buying only those products necessary for "just-in-time" production cycles.

Jan 1, 1994

By C. J. Barbee, P. J. Bethell, C. J. Wood

Heavy medium cyclones (HMCs) are commonly used to clean run-of-mine coal in the 50 x 0.5 mm size range. Unfortunately, field studies suggest that industrial HMC circuits are often poorly designed or improperly operated, resulting in significant losses of recoverable clean coal. To assist plant engineers in overcoming these problems, a spreadsheet-based simulation package has been developed to predict HMC performance based on mathematical expressions originally developed by Wood (1990). A case study is provided to illustrate the predictive capabilities of this simulation package for an industrial HMC.

Jan 1, 2004

By Thomas H. Holley

This presentation will describe the underground excavation at the Bad Creek Pump Storage project. It consists of approximately 3,650 m of various size tunnels ranging from 4.2 m to 9.0 m finished diameter, a powerhouse cavern 45 m high by 135 m long by 22.5 m wide, and 422 m of 9 m diameter shaft. A total of 350,000 m of rock will be excavated, hauled, and placed in a submerged wier. Temporary rock support consisting of shotcrete and/or rockbolts is installed as required.

Jan 1, 1987

By J. A. Burger

Research in the eastern U.S. has shown that productive mine soils and forests can be restored using the following guidelines: (1) cooperation among the mine operator, landowner, and regulatory authority in developing the mining permit; (2) replacing topsoil or creating a desirable medium in which trees are planted; (3) loosely grading non-compacted topsoil; (4) using native and non-competitive domestic ground covers that quickly protect the site; and (5) planting a proper silvicultural mix of crop trees for their commercial value along with nurse trees for wildlife and soil improvement.

Jan 1, 2002

By T. D. Rourke, Fred H. Kulhawy, O&apos

From mid-1976 to early 1978, a 5.64 m diameter tunnel, 2491 m long, was excavated by a modified Robbins full-face TBM under shallow cover in highly stressed sedimentary rock in Rochester, NY. In-situ stress measurements were made, and tunnel behavior was monitored with MPBXs, inclinometers, tape extensometers, and instrumented steel ribs. Detailed geologic mapping was also done. In this paper, the as-monitored tunnel behavior is described, and the TBM performance is evaluated from field records. Recommendations are also made for future tunneling in similar geologic environments.

Jan 1, 1981

By Andreas Eberhorn

The evaluation criteria for cost comparison of the Gearless Drive, also called Ringmotor or wrap-around motor, as variable-speed drive with a Twin Drive System using synchronous motors are reviewed. Both mechanical and electrical components of the drive system are considered. The evaluation criteria are described and for the cost comparison the applicable method is demonstrated. The cost comparison is based on a medium and large sized mill system. Each drive package will display cost details with regard to capital cost, maintenance, power losses, overall efficiency, overhead and loss of production. As conclusion the benefits and disadvantages of the two drive/mill packages will be presented together with a payback diagram.

Jan 1, 1996

By Andrew J. Gaber

Beginning in 1961 the Coal Research Bureau of the West Virginia University School of Mines began a program of research to investigate the potential of the magnetic separation process to remove sulfur of the pyritic form from pulverized coal. Using a so-called isodynamic magnetic separator (1) it was demonstrated that significant quantities of pyritic sulfur could be removed from coal. Pulverized coal samples representing several different coal seams were individually screened into sizes in the 48 mesh by 200 mesh (297 to 74 microns) range, and untreated and steam-air treated samples of each were processed by the magnetic separator. The results of the investigations to magnetically remove pyrite from coal were published in two separate reports issued by the Coal Research Bureau (2,3), The isodynamic magnetic separator was designed about 30 years ago for use in geological and mineralogical laboratories (1,4,5) to isolate one or more minerals exhibiting only small differences in their magnetic propertiese Specifically, to accomplish this, the air gap of the unit was designed to provide a nearly constant magnetic gradient both laterally and longitudinally in the separating zone. It is reported that mineral concentrates of 99.5 percent purity can-be produced on closely-sized fractions after several cleanings through the unit. The isodynamic magnetic separator, therefore, was not expected to process coal containing sizes smaller than 200 mesh, or to demonstrate commercial capability to process pulverized coal. The primary purpose of the Coal Research Bureau in using the unit was that of determining if the magnetic process could remove pyrite from coal, and thereby encourage further research directed toward development of acceptable full-scale processing equipment.

Jan 1, 1969

By J. E. Lonergan

The in-pit placement of crushing and conveying equipment results in a flattening of the pit slope of a multiple-bench open pit mine. To maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must minimize the cost of the flatter slope which is expressed in terms of additional stripping or tied up ore. A computer program that compares the tonnage aspects of various conveyor layouts will be discussed.

Jan 1, 1985

By Jae-Chun Lee, Sadia Ilyas

"The present study was carried out to develop an economically feasible and environmentally friendly technique to extract metals such as Al, Fe, Cu, Zn and Ni from electronic scrap by employing Penicillium chrysogenum strain KBS3. Studies were carried out in various leaching modes, including one-step leaching with combined growth and leaching phases (Mode 1), two-step leaching with alternative growth and leaching phases (Mode 2), spent medium leaching with separated growth and leaching phases (Mode 3), leaching with fresh growth medium without fungal culture (Mode 4), leaching with deionized water (Mode 5) and leaching with commercial organic acids (Mode 6). The main lixiviants produced in leaching Mode 1 were citric, tartaric, oxalic and gluconic acids in concentrations of 12, 2.5, 1.8 and 152 mM, whereas in leaching Mode 2, the concentrations were 15, 0.5, 1.0 and 1,162 mM, respectively. In leaching Mode 3, 63 mM citric acid, 23 mM tartaric acid and 29 mM oxalic acid was observed. At 5% pulp density, both leaching Modes 1 and 2 showed similar metal extraction yields from the electronic scrap at 35º C. Approximately 96% Al, 98% Zn, 48% Cu, 25% Fe, 66% Pb and 73% Ni were extracted in leaching Mode 1 in 14 days. In leaching Mode 3, 85% Al, 81% Zn, 97% Cu, 22% Fe, 50% Pb and 63% Ni were extracted in 14 days. The extraction yield of Cu was much lower in leaching Modes 1 and 2 (48- 50%) than in leaching Mode 3 (97%), as well as in leaching Mode 6 (98%) with organic acids, including citric, tartaric, oxalic and gluconic acids.IntroductionTechnical innovations in electrical and electronic equipment with new functionalities and designs stimulate consumer purchasing and lead a rapid increase in the production of waste materials. Terazono et al. (2006) have reviewed the data on electronic waste generation in different parts of the world. Studies by Bertram et al. (2002) have also affirmed that waste electrical and electronic components are the fastest growing waste category, which emphasizes the need for efficient recycling strategies. The recycling of electronic waste is an important subject, not only from the perspective of waste treatment, but also from the perspective of the recovery of valuable materials. However, the recycling of electronic scrap is still limited, due to the heterogeneity of the materials present in the products (Veit et al., 2005).Bioleaching-based processes offer a number of advantages over conventional pyro or hydrometallurgical methods (Lee et al. 2007; Lee and Pandey, 2012). These advantages include simple, inexpensive (in some cases) and ecofriendly approaches and the generation of less active leach residue relative to other metallurgical processes. Although biohydrometallurgical processes have been successfully applied to the leaching of metals from ores (Olson et al., 2003; Rawlings, 2002), data pertaining to the application of these processes for the extraction of electronic scrap are still scarce. Recently, a few studies have been undertaken for the extraction of metals from electronic scrap/ printed circuit boards (Brandl et al., 2001; Faramarzi et al., 2004; Ilyas et al., 2007; Ilyas et al., 2010). These studies were conducted with mesophilic chemolithotrophic bacteria (Acidithiobacillus ferooxidans and Acidithiobacillus thiooxidans), cyanogenic bacteria (Chromobacterium violaceum) or moderate thermophilic bacteria (Sulfobacillus thermosulfidooxidans). However, compared to bacterial leaching, fungal leaching has several advantages, including (i) the ability to grow at a higher pH, thus making it more suitable for the bioleaching of alkaline solid waste; (ii) a generally faster leaching process with a shorter lag phase and (iii) the chelation of excreted metabolites (e.g., organic acids) with metal ions, thus reducing toxicity to the biomass (Burgstaller and Schinner, 1993; Aung and Ting, 2005)."

Jan 1, 2015

By R. Q. Phillips

INTRODUCTION Since Canada commenced sulphur exports in the early 1960&apos;s, the volume of exports has grown steadily and by 1984 Canadian sulphur accounted for 45% of internationally traded sulphur. In 1984 a total of 7.6 million tonnes of Canadian sulphur was shipped to offshore markets and the U.S.A. Total shipments, including domestic sales, were 8.3 million tonnes. With a production of some 5.6 million tonnes, inventory was drawn down by nearly 3 million tonnes in order to keep pace with demand. Since 1979, world consumption has exceeded total world production and Canadian sulphur inventory has played a truly crucial role in balancing world supply and demand. Consequently, by the end of 1984 the Canadian inventory, which stood at 21 million tonnes in 1978, had been almost halved to 12 million tonnes. This inventory, however, is still by far the largest in the world and will continue to play a vital role in the world supply until expected depletion around 1990. Canada is the world&apos;s second largest producer of sulphur following the U.S.A.. The Current production level of between 5 and 6 million tonnes per annum is forecast to be maintained to 1990 with only a marginal fall in that production level in the following decade to the year 2000. This will ensure that Canada retains its position as one of the largest producers and most important suppliers of sulphur in the world market well into the future. CANADA AS WORLD SUPPLIER OF SULPHUR Since Canada commenced sulphur exports in the early 19601s, exports have grown steadily. (Figure 1) The rate of growth accelerated in recent years and in 1984 a total of 7.6 million tonnes was sold for export, compared to 3.0 million tonnes in 1975 and less than 1 million tonnes in 1964. World trade in sulphur also increased during this period, from 11 million tonnes in 1975 to 17.0 million tonnes in 1984 (Figure 2). Most of this increase in volume was accounted for by Canada and Poland. As a result, the importance of Canada as world supplier expanded from 28% in 1975 to 45% in 1984. The markets for Canadian sulphur can be divided into two broad categories, i.e., solid and liquid (Figure 3). The solid market is the largest, taking about 70% of Canadian output, virtually all of which is sold offshore. The U.S.A. and the Canadian domestic market take about 20% and 10% respectively, in liquid form. In 1984, (figure 4) total Canadian sulphur sales exceeded 8 million tonnes with offshore export sales at a record 5.8 million tonnes and sales to the U.S. of 1.8 million tonnes. Domestic sales accounted for only 0.7 million tonnes. The general trend for offshore export sales has been upward whereas there has been relative stability in domestic sales. The exceptional growth in offshore sales in the 1970&apos;s necessitated considerable capital investments by the industry in order to establish a logistics system capable of handling increasing offshore tonnages. Over 50 sulphur plants are currently operating in Alberta and British Columbia. Some of these are connected by rail, (Figure 5) the balance being generally connected by road. There are currently fourteen of the larger rail-connected plants designated as export plants, and these are equipped with sulphur forming facilities. The long distances between these plants and tidewater terminals in Vancouver is a very significant factor in complex logistics of Canadian offshore sulphur exports (Figure 6). Alberta and B.C. plants are located an average of 1,400 kilometers from tidewater.

Jan 1, 1986

By Subhash B. Bhagwat

No economic activity would take place without human ingenuity, driven by our needs and wants. Economic activity is made possible by the energy from fossil and other energy sources, our technological hardware, the buildings and infrastructure made from materials such as metals, plastics, wood, water, and numerous industrial minerals. Industrial minerals are difficult to define; however, they generally include utilitarian minerals other than metals and fuels. Some of the industrial minerals form an exception to this categorization and, as explained elsewhere in this book, belong to more than one category because of the manner in which they are used. Industrial minerals are so numerous, and their applications so ubiquitous that they often do not receive the attention given to metals or fuels. Nevertheless, without industrial minerals, many of our most fundamental economic activities such as construction of our cities and towns, control of floods, transportation of our people, the search for minerals, the smelting of metals, and the delivery of supplies of clean water would be impossible. How then do we assess the economic importance of such a variety of materials lumped together under the general designation of industrial minerals? The price per ton or the total amount consumed do not satisfactorily explain the relative value of these materials to the economy. For example, the current price of limestone is about $5/t and the price of fluorspar is about $190/t. However, nearly 1.1 Gt of limestone, valued at about $5.5 billion, are consumed in the United States annually, compared with 590 kt of fluorspar, valued at about $1 12 million. It would not, however, be accurate to conclude that the mineral with the higher total value is always more economically important. Many other questions must be answered before making judgments of this type. These questions include how many people are employed in the production, processing, and distribution of the mineral? How is the mineral used? Is it partly or entirely imported? Where do the imports come from? How is the mineral transported? Is the mineral considered critical to national security? Is there a substitute material available in case the mineral supply is interrupted? Where would the substitute be produced? What would be the basic material from which the substitute will be produced? Even when these questions have been answered satisfactorily, the economic importance of the mineral is not completely known until the consuming sectors are studied as well. For example, even though the manufacture of fighter aircraft is vital to the national security, it employs only a fraction of the number of people employed in the housing and construction industry or the automobile industry or the appliance industry. Mineral commodities consumed in these latter sectors of the economy are more significant than those consumed in the production of fighter aircraft, because nearly two-thirds of the Gross Domestic Product (GDP) depends on the purchasing power of individual consumers. In comparison, the entire defense budget accounts for only 6% of the GDP. There is no easy way to present the importance of industrial minerals to the US economy because most minerals have diverse applications and products made from industrial minerals often con- tribute to the production of other goods and services. One way of highlighting the value of a mineral would be to relate it to the GDP reported by categories such as construction, machinery, stone, clay and glass products, electrical and electronic equipment, paper and allied products, and chemicals and allied products. These GDP categories represent industries fundamental to the economy totaling $644 billion, or about 12% of the GDP in 1990. However, these categories are not convenient for discussion of the many minerals that are used in our complex economy. In a general way, minerals are the 5% of our economy that make the other 95% possible. This chapter, therefore, discusses a selected few industrial minerals to illustrate their importance to various sectors of the economy.

Jan 1, 1994

By Niels Jonassen

INTRODUCTION The occurrence of the radioactive element radium, [2gg~a], often in trace amounts, in the crust of the earth gives rise to a flux of the daughter product radon, [&apos;ii~n], across the earth-air interface. This so-called exhalation process has been studied by many investigators [KO 45, Kra 64, Pe 66, Wi 741] over several decades. In the earlier investigations the exhalation process was studied in a purely geophysical context, but over the last 20 years it has been realized, that under certain conditions the exhalation may be intense enough to also have radiological implications. This is for example the case where relatively radium-rich wastes or by-products, like uranium mill tailings and phospho-gypsum, are stored in larger quantities. In order to be able to predict the radiological impact on the environnent it is necessary to know the rate of exhalation of radon from the surface of the exhaling material. Usually the determination of these exhalation rates has been done by measurements in situ directly upon the accumulated material [~le 74, Ha 791. Although such field measurements have the obvious merit of dealing directly with the full scale problem, the results are often lacking in generality, making predictions for other situations difficult or even impossible. The present paper describes some results of an investigation where it is attempted from laboratory measurements on samples of limited size to obtain material parameters, which are suited for extrapolating and predicting the radiational behaviour of larger accumulations of the material. THEORY Let us consider a sample of a radon exhaling material enclosed in a container with a dead space volume, Vd, and initially containing radon free air. If the total exhalation rate (number of atoms per unit time) is c, the radon concentration (number of atoms per unit volume) N will initially grow with the time t according to the equation [ ] where [R] is the decay constant of radon. As the concentration (or activity) in the container grows, the net exhalation rate will slow down because of back diffusion, and, furthermore, radon atoms may be removed from the container, not only by radioactive decay, but also by diffusion leakage [JO 81 ]. If, however,only the initial. part of the activity growth curve is considered, the total exhalation rate, c, is given by [c = r•Vd(2) where r is the initial activity growth rate in the container. Let the container be cylindrical with an inner cross-section S and filled to a height L with a radon exhaling material with the specific mass p, the porosity E, the radon pore production rate f and the diffusion length i. When the container is left open,i.e. the radon exhalation takes place into a virtually radon free room, the radon area exhalation rate, E, can be written [Jo 80] E = Ef?t tanh L.R (3) assuming the exhalation only takes place in one direction, towards the surface of the material. It appears that for values of L > 2•ithe area exhalation rate becomes E a = Efi (4) independently of the sample dimensions and the total exhalation rate c is then c = E •S (5) For L ;~ 0.5•i, tanh P _ k and equation (3) may be written E ~ EfL (6) then c = EfL•S = p•m= Em •m (7) where m is the mass of the sample and Em = P (8) is the mass exhalation rate. It thus appears that the radon exhalation rate from a given amount of material may often be conveniently expressed by c = E •Sfor L z 2•I (5) a orc = Em •mfor L0.5•i (7) where S is the total exhaling area and m the total exhaling mass, E and E the area and mass exhalation rates respectively defined by equations (4) and (8).]

Jan 1, 1981

By Gaither H. Kitchell

Introduction Fire assaying is an ancient method of determining the amount of gold and silver in geological materials. Today&apos;s commercial, mining, school, and private fire assaying laboratories still use this basic assaying method. One of the primary reagents used to collect gold and silver from the ore is litharge (PbO), or lead oxide. The Federal Occupational Safety and Health Administration&apos;s (OSHA) and the Nevada Division of Occupational Safety and Health&apos;s (OSH) have a lead standard. It states that a person exposed to airborne lead at concentrations greater than 50 µg/m3 (1.4 µg per cu ft) of air over an eight-hour period has gone beyond the permissible limit for health safety. It is not uncommon to get airborne lead concentrations of 1400 µg/m3 (40 µg per cu ft) in the flux/sampling area or 900 µg/m3 (25 µg per cu ft) around the gas/ electric furnaces during the fusion and cupelling phases. Therefore, any individual working in these areas for more than 30 days per year without proper protection is a likely candidate for lead poisoning. Lead In A Human Lead enters the body as a toxic substance by breathing it into the lungs and upper respiratory tract and by getting lead particles into the digestive system through ingestion. Smoking, eating, drinking, chewing tobacco and gum, or wiping a lead-contaminated hand across the mouth are all potential ways forgetting lead into the body. Lead taken in large quantities can actually kill a person in a matter of days. This is rare, however. The real hazard of lead overexposure is for the unsuspecting person who works in an unsafe lead environment for months or years. This individual can slowly receive damage, sometimes permanent, to the central nervous system, gastrointestinal tract, kidneys, blood, gums, and reproductive system. Safe Blood-Lead Level The Lead Standard states that the maximum safe blood-lead level is 40 µg/100 g (40 µg per 3.5 oz) of whole blood. It recommends 30 µg/100 g (30 µg per 3.5 oz) if a person is of child-bearing age. People not working around lead products normally have a blood level of less than 10 µg/100 g (10 µg per 3.5 oz). A person not adequately protected can exceed 40 µg/100 g (40 µg per 3.5 oz) in a few weeks by fire assaying only 100 samples per day per week. Studies show that some individuals have developed lead-related diseases with blood-lead levels less than 80 µg/m3 (2.3 µg per cu ft). Lead Poisoning Symptoms There is a long list of symptoms associated with lead poisoning. The most common are: a prolonged, sweet, metallic taste in the mouth; loss of appetite; anxiety and irritability; weakness and dizziness; nausea and tremors; and insomnia. Preventive Measures Based on the lead standard, a number of preventive measures can be taken to ensure that fire assay laboratory personnel frequently exposed to lead are protected against overexposure. First, the actual lead concentration within the assaying area should be determined. In Nevada, one can contact the Department of Industrial Relations, Division of Occupational Safety and Health for assistance in monitoring any laboratory. The agency will also recommend ways to protect against lead. This agency does the monitoring and provides an educational class on lead safety. This type of agency can be found in other states. Commercial consulting firms are also available, but they are expensive. Regardless of the technique used to determine lead concentrations, it is most likely that the average fire assay area exceeds the 50 µg/m3 (1.4 µg per cu ft) lead concentration limit. This is considered unsafe.

Jan 8, 1985

By J. E. Moyle

Introduction In October 1983, Rio Algom Ltd. decided to develop the East Kemptville tin deposit. This decision was the result of intensive geological and metallurgical evaluation work, as well as careful studies of environmental and marketing aspects. The property is located about 54 km (34 miles) northeast of Yarmouth, Nova Scotia, near the village of East Kemptville. Geology The geological setting is a granitic intrusive at the contact with metasediment. The tin mineralization is in the form of cassiterite (SnO2), which is about 80% tin. Sulfides are associated with the deposit containing silver, copper, and zinc, which are recoverable as byproducts. The mineralization is associated with quartz veins and greisen alteration. In the center of the deposit, the zones are so close together they have the appearance of a massive or porphyry type deposit. Ore Reserves The mine plan is based on 56 Mt (62 million st) at 0.165% tin. Up to the ninth year, mill feed will average about 0.2% and lower grade material will be stockpiled. In later years, the average grade will be 0.16% tin. A 0.08% cutoff grade is planned for the entire 17 years of production. East Kemptville is a relatively low grade tin deposit, and only a combination of favorable circumstances has made it economic. • The ore body is situated at, or very close to the surface, permitting mining by open-pit methods at a very low stripping ratio. • The tin occurs in relatively large crystals that allow for good recovery by gravity techniques. • It is ideally located in a well-settled area, with all essential infrastructural requirements already in place. • The positive and cooperative response from various governmental agencies involved are encouraging. Mine Plan The mine plan incorporates an initial pit, followed by three expansions to the ultimate pit. Mining will be at 14 kt/d (15,000 stpd) and will include 9 kt (10,000 st) of ore plus 5 kt (5,500 st) of waste and low grade. After five years, the mining rate will increase to 21 kt/d (27,500 stpd). It is planned to drill vertical 165-mm (6.5-in.) blast holes on a 6-x 6-m (20- x 20-ft) pattern. Bulk slurry explosives will be used to break the rock. Loading will be with hydraulic shovels into 70 t (77 st) haulage trucks. The trucks will deliver the ore to a 13 x 20 m (42 x 65 ft) gyratory crusher, where it will be crushed to 150 mm (6 in.). Secondary and tertiary cone crushers will be used to further crush the ore to 13 mm (0.5 in.). Process Design Tin extraction technology is practically unknown in North America, but Rio Tinto Zinc, which has a 53% ownership in Rio Algom, is involved in tin production in the UK. Rio Algom has used this experience and expertise in metallurgical evaluation work and in tin-related aspects of plant design. Mineralogical and metallurgical test work have been performed, both on laboratory and pilot plant scales. Mineralogical work was done largely in the UK, the laboratory test program was carried out at the Lakefield research facility in Ontario, and pilot plant work was conducted in Australia. This program has resulted in significant improvements in recovery and concentrate grades, over best projections that could be made before this work was carried out. Concentrating Process The concentrator is being designed to process about 9 kt/d (10,000 stpd). The first step is to reduce the size of the ore to -700 µm (-24 mesh) in a rod mill/ball mill grinding circuit. The ball mill will operate in closed circuit with high capacity banana-type screens. The screened product will be deslimed to 30 µm (560 mesh) with cyclones, then passed over the pri-

Jan 4, 1984

By Donald E. Ralston, A. T. Yu, Richard P. Killmeyer, Peter Phillips, R. W. Vander Laan, John N. Murphy, John R. Messineo, John Peters

Overview Coal production in the first half of 1982 remained near the high levels begun in mid-1981. Production then fell sharply during the latter part of the year due to recession-weakened demand. Even so, 1982 coal production was estimated at 756 Mt (833.4 million st), a record high that exceeded 1980&apos;s previous production record of 752.7 Mt (829.7 million st). Though 1982 coal production was substantially below industry expectations, the attained production represented a modest achievement in view of the recession&apos;s severity. In 1982, the gross national product declined 1.8%, its worst performance in 36 years. The recession, along with mild weather, improved energy efficiency, and conservation efforts caused the nation&apos;s energy use to drop 4.1%. With the iron and steel, construction, and auto industries experiencing production declines, industrial energy use fell 10.8%. The extent to which the nation&apos;s economic slump affected coal markets in 1982 varied: • total domestic coal use declined 3.4%; • exports remained strong for most of the year; • electric utility coal use declined for the first time in 25 years; • coal use at coke plants declined by one-third; and • coal use in other industrial sectors stayed surprisingly near the 1981 level. Coal Use Domestic coal use totaled 641.8 Mt (707.5 million st) in 1982,22 .9 Mt (25.2 million st) below the 1981 level. Nearly all the decline occurred at coke plants where use dropped to 37.1 Mt (40.9 million st) in 1982 from 55.3 Mt (61 million st) in 1981. The decline in all other industrial and electric utility coal markets was only 5.3 Mt (5.8 million st). Residential and commercial coal use increased 635 kt (700,000 st) to 7.3 Mt (8.1 million st) from 6.7 Mt (7.4 million st). Power generation decreased 2.3% in 1982, the first year since 1945 that output failed to increase. The decline was due to a drop in industrial demand. Electricity sales to residential, commercial, and other nonindustrial users increased slightly. But coal&apos;s competitive position stayed strong. Power generation from oil and natural gas fell 28.9% and 11.7%, respectively. Power generation from coal declined only 0.9%. Among the primary energy sources, coal fueled 53.2% of US electric power generated in 1982, compared with 52.4% in 1981. In 1982, US steel production declined to 66.1 Mt (72.9 million st), the lowest level in 50 years and down from 109.6 Mt (120.8 million st) in 1981. The reduced steel demand forced 13 coke plants to shut down. Coal Exports Coal exports in 1982 totaled 96.4 Mt (106.3 million st), 5.7 Mt (6.3 million st) less than in 1981. Despite the worldwide recession and above normal coal stockpiles in many countries, exports were higher than many expected. This showed the relatively strong US position in international coal trade. The US continued to export near-record coal tonnages despite increasingly competitive influences. These included an improvement in Australia&apos;s labor difficulties, the return of Polish coal to Western European markets, and lower oil prices. In 1982, US port facilities were improved and expanded, and more efficient methods for scheduling foreign vessels were implemented. As a result, high volume coal shipments to US ports were routinely loaded and dispatched unhampered by bottlenecks that plagued East Coast ports in 1981. Appalachian coalfield production declined 1.4% from 1981 output and 3.9% from 1980 output. The interior region of the US showed an increase of 7.8% in production. This was primarily because of a

Jan 5, 1983