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By Michael Harris, David Mosch

The numerous small high grade gold and silver deposits of the Front Range Mineral Belt have been uneconomical because of the high costs of mill construction, transportation, and smeltering. Processing of these ores must be simple, cheap, with satisfactory recoveries, and be environmentally acceptable. Vat leach with direct electrolytic recovery satisfies these parameters. A battery of testing was conducted utilizing bottle and bucket tests. A five ton vat leach with direct electrolytic recovery gave excellent results while a five ton heap leach proved impractical due to channeling.

Jan 1, 1987

By J. -C. Tricket, W. A. Düweke, S. Kock

The high-resolution three-dimensional (3-D) seismic reflection data acquired for Impala Platinum between 1998 and 2002 in the western Bushveld Igneous Complex form an integral part of Impala?s mine planning and development programme. These data are of sufficiently high quality to provide unprecedented 3-D models of the geological structure of the Merensky Reef and UG2Chromitite Reef, from which the mine deduces the size, geometry and distribution of mineable blocks. Conversely, the seismic data details zones of un-mineable ground caused by structural disturbances in the form of potholes, fault zones and pegmatoid intrusions. Furthermore, variations in dip and strike are well imaged, allowing the mine planners to optimize the mine design and layout. Current interpretation work is focused on enhancing the accuracy of these geological model seven further through the analysis of ?seismic trace attributes.? Preliminary testing on the Impala Platinum 3-D seismic volume has shown that attribute analysis has the potential to illuminate a number of geological reef disturbances that are otherwise ?invisible? in the original seismic time volume. Potential areas of improvement include inter alia: ?Improved definition of major fault zones ?Improved definition of zones of reef depression (?slumping?) ?Identification of reef terracing and/or reverse dip.. Keywords: 3D reflection seismics; Impala Platinum Ltd; seismic attribute analysis; Seismic Micro Technology (SMT); rock solid attributes; dip-azimuth; instantaneous phase and similarity.

Jan 1, 2004

By A. M. Keenan, R. H. Carpenter

Geologic faults have always been a plague to the mining industry, and have often reduced a mining venture from a profitable to a marginal operation, and even at times have forced companies to liquidation or bankruptcy. The use of geology as an economic tool in coal mining is not generally recognized, particularly in coal deposits which are horizontal or essentially flat-lying, in which faulting might not be prevalent. However in pitching coal seams, the existence of faults is considered an inherent challenge, as the inclination of the seam or seams per se, is a product of geologic disturbance. The problems of faults are a constant challenge to the management of the Thompson Creek Coal and Coke Corp. in their operations in Pitkin County, Colorado, where high grade medium volatile coking coal is mined. The background for this paper has been provided by the experiences of these operations, in overcoming the effects of faults in mining, and resolving the fault structures. The properties of the Thompson Creek Coal and Coke Corp. are located in central Colorado, 13 miles southwest of Carbondale, Colorado, a tow] on the main highway between Glenwood Springs and Aspen, Fig. 1. The elevation at the mine portals is approximately 8000 feet: there is a relief of some 2.000 feet to the adjacent ridges rising away from the [nine portals. 'The entire present production of these properties totaling 180.000 tons of clean coal per year is trucked to the railhead at Carbondale, Colorado, from whence it is shipped to the Geneva Steel Plant at Provo, Utah. The properties were initially opened in September 1952. A total of 125 men are currently employed with the production conting from the Number 1 and 3 Mines, and development workings of the Number 2 and 5 Mines respectively. While the No. 2 Vine was opened in 1953 prior to No, 3 Mine, the encountering of faults without sufficient development work accomplished to sustain production, necessitated the diversion of manpower to the other [nines for economic considerations. GEOLOGY '1-he productive coal seams occur in the Mesaverde formation of Late Cretaceous age which outcrop along the eastern edge of the Piance Creek Basin from the Grand Hogback, a part of the southwestern flank of the White River Uplift north of Glenwood Springs, southward through the 'Thompson Creek area and the

Jan 1, 1961

By Mark Kennedy

New technology projects, particularly those involving first-of-a-kind metallurgical facilities, have a poor track record of achieving their overall financial targets, as they incorporate technological risks in addition to normal project risks. Traditional project management tools focus on risk from concept development through implementation, but typically fail to address the additional risks associated with a new technology project including slow start-up, slow ramp-up, and failure to achieve design performance (quality or throughput). These additional risks affect future cash flows produced by the project and hence its value, typically measured as net present value (NPV) or internal rate of return (IRR). Common accounting practice is to reduce the present value of future cash flows by applying a chosen discount rate, often reflecting the corporation?s cost of capital. This discount rate does not contain an allowance for technological risk, which is often underestimated or misunderstood by both the financial and engineering project managers. This paper will focus on the topic of project risk and new technology, and how these play a role in the probability of achieving the predicted future cash flows. By properly addressing technological risk in the form of risk-weighted returns, marginal projects can be prevented from proceeding, while sound projects can be given the additional time and resources required to achieve the optimal level of front-end-loading (the level which returns the maximum risk-weighted net present value). A simple set of mathematical tools should enable engineers and financial professionals to establish a common level of understanding, thereby leading to a more accurate assessment whether a new technology project is ready to pass through its next level of approvals.

Jan 1, 2013

By P Jones

Fosterville Gold Mine is an underground operation located approximately 20 km northeast of Bendigo and 170 km north of Melbourne. Fosterville is the most productive gold mine in Victoria, producing over 100 000 ounces of gold annually. Access to the underground is via two portals cut into previously mined open pits. Fosterville is 100 per cent owned and operated by Northgate Minerals, a gold and copper producer with mining operations, development projects and exploration properties in Canada and Australia. (Northgate Minerals Corporation, 2010).Underground operations commenced in 2006 with ore production focused on three orebodies – Phoenix, Ellesmere and Falcon. Primary ventilation to the existing operation is provided by wall mounted exhaust fans located underground. This system provides approximately 220 m3/s of primary ventilation to the operations with intake achieved through the portals. Exploration drilling during recent times has uncovered another lode, the Harrier orebody, located 2.2 km to the south of the existing workings. Due to the location of the new Harrier orebody, vent analysis revealed the requirement for a separate source of primary ventilation to service Harrier. Figure 1 shows a current underground long section and the location of Harrier in relation to the existing workings. A number of options were considered, including a parallel decline from one of the existing two portals, however, a 4 m diameter, 350 m deep surface ventilation shaft was selected as the most suitable and feasible option for the mine.In December 2009, geotechnical drilling commenced to determine the most suitable location for the shaft. Site locations were constrained by a number of factors including interaction with current and planned surface infrastructure as well as proximity to nearby community members at the southern end of the mining lease. Results from the geotechnical drilling indicated that the majority of the proposed shaft would be suitable for a 4 m raise bored shaft. However, the top 40 m of the proposed shaft were deemed unsuitable for raiseboring due to zones of weakness that would likely influence the stability of a 4 m diameter shaft (Robinson, 2010). Following independent geotechnical advice and input from experienced raiseboring contractors, it was decided to construct a 40 m presink prior to raiseboring activities.A crane mounted auger was chosen as the preferred method to construct the presink. This method differs substantially from traditional blind sinking methods which are highly labour intensive. The auger, or drill, is designed and built to be attached to a crane base using engineered pivoting cradles. The auger is suspended from the crane jib by pennant ropes attached to hydraulic rams. The auger applies rotation and axial loads to the drilling tools via a telescopic steel bar (Kelly Bar). Drill tools are attached to the bottom of the Kelly Bar and comprise an assortment of drilling buckets and reaming barrels. Drilling buckets cut in situ material and collect the spoil for removal and are used for the excavation of the ‘pilot hole’. Reaming barrels are used to expand the diameter of the excavation by directing cuttings into the pilot hole which are then removed by the drilling bucket. In June 2010, the presink was constructed utilising a crane mounted auger with support to the shaft walls applied via a remote fibrecrete unit. The excavation method involved cutting a series of smaller diameter holes (2.4 m wide) before reaming out to the final diameter (4.7 m wide) in 4 - 5 m vertical increments. Fibrecreting activities were carried out on the backshifts, to ensure the stability of the excavation and minimise interaction between the two units. The presink was completed on time and offered many advantages compared to traditional blind sinking methods, including safety, costs and timing. This paper will examine the benefits of this method as well as lessons learnt and future applications in the mining industry.

Mar 21, 2011

By E. S. Machlin

THE function of this paper is to present certain results based on the fact that the strain energy arising from the solution of out-of-size solute atoms into the solid matrix is free energy and not internal energy, and that this strain energy is almost wholly shear strain energy as Zener1 pointed out. One natural consequence of this fact is an expression for the excess entropy of mixing due to the strain energy. Also, a derivation of the integral molar free energy of mixing partitioned to strain energy will be made using the model of Lawson2 and thermodynamics. This derivation has the advantage of showing in a simple way the assumptions involved in Lawson's treatment. The excess entropy of mixing partitioned to strain energy can be derived in two ways. The first way makes use of the fact that the strain energy is almost wholly in the form of shear strain energy. In this case, as Zener1 pointed out, the change in entropy due to the addition of shear strain energy is to a first approximation given by ?S/?? p,T = - ?lnµ/?T p,? Hence, the excess entropy of mixing partitioned to strain energy will be given by Sx ? = ?S/?? p,T Gx ? = - ?lnµ/?T p,? Gx ? where, Gx ? is the relative integral molar free en- ergy partitioned to strain energy; S is the entropy; is the shear strain energy; µ is the shear modulus; T is the absolute temperature; p is the pressure; and ? is the shear strain. Zener,1 as well as others, has pointed out that the shear strain energy stored in a compressible isotropic solid due to the insertion of an off-size incompres-sible sphere into a spherical hole in the solid, is approximately given by 2/3 µ/O (O°I - O)2 where a is the volume of spherical hole and no , is the volume of sphere. If we now consider the introduction of one mol of such spheres into an infinite volume of solid, then the relative partial molar free energy excess due to strain energy alone is GM ? = 2/3 µ/v (v ° 1 - v)2 where v is the molar volume of solvent and v°1, is the molar volume of pure solute type 1. Let us consider that these spheres are the two types of atoms in a binary solution, then there will be a similar relation for solute 2. The total relative integral molar free energy excess due to strain is then, to the approximations stated above GM ? = GM ? x1 + GM2 ? x2 = 2/3 µ/v [(v ° 1 - v)2 x1 + (v ° 1 - v)2 x2]. Lawson's relation for the strain energy can be de-

Jan 1, 1955

By D. A. Reynolds

THIS PAPER gives results of an investigation of the composition and carbonizing properties of coals from the Beckley bed, Caretta No. 5 mine, Susanna, McDowell County, and Glen Rogers No. 2 mine, Glen Rogers, Wyoming County, W. Va. The investigation included BM-AGA carbonization tests at. 800° and 900° C. and various chemical and physical tests designed to evaluate agglutinating, plastic, and expanding properties of coal. Both coals were of low-volatile bituminous rank, and each was blended with 70 and 80 percent high-volatile A coal from the Pittsburgh bed, Warden mine, Allegheny County, Pa. The Beckley bed was 39 to 58 inches thick at three points in the Caretta No. 5 mine, and the coal was 39 to 49 inches thick. The carbonizing sample contained 83.0 percent fixed carbon on the dry, mineral-matter-free basis. It contained 3.1 percent moisture, 15.8 percent volatile matter, 72.4 percent fixed carbon, 0.7 percent sulfur, and 8.7 percent ash, as carbonized. The ash softened at 2,680° F. The free-swelling index was 7;i, and the agglutinating value was 7.7. This sample fused weakly in the Gieseler plastometer; its blends with 70 and 80 percent Pittsburgh coal attained maximum fluidities of 830 and 600 dial divisions per minute, respectively. The single coal yielded 83.0 percent coke at 900° C. Yields of other carbonization products per ton of coal at 900° C. were: Gas, 9,950 cubic feet; tar, 3.5 gallons; light oil, 1.36 gallons; and ammonium sulfate, 16.3 pounds. Blending with Pittsburgh coal lowered the yield of coke and raised the yields of other products significantly. Caretta coal yielded strong, blocky coke at 900° C., 89 percent of which was plus-2-inch size. Strength indexes of 900° C. coke from 30:70 blends of Caretta No. 5 and Pittsburgh and of Pocahontas No. 3 and Pittsburgh coals were, respectively: 2-inch shatter, 68 and 53; 1%-inch shatter, 90 and 85; 1-inch tumbler, 50 and 53; and %-inch tumbler, 62 and 65. The specific gravity of the 900° C. gas from Caretta No. 5 coal was 0.272, and the heating values were 491 B. t. u. per cubic foot and 2,440 B. t. u. per pound of coal. The blends yielded much richer gas. Yields of tar acids and bases (0.09 and 0.04 gallon per ton, respectively) were within the normal ranges for low-volatile coals. Caretta No. 5 coal expanded 7.5 percent in the sole-heated oven at a charge density of 55.5 pounds per cubic foot. The blends containing 70 and 80 percent Pittsburgh coal contracted 4.8 and 7.5 percent, respectively.

Jan 1, 1953

When the Organic Act of 1910 established the Bureau of Mines within the Department of the Interior, the Bureau's programs involved only health and safety research and education, and fuels testing. As the Federal Government's primary agency for research in minerals, the Bureau has expanded its scientific and technological investigations to mining, processing mineral substances, eliminating occupational health and safety hazards in the mineral industries, economic development, and resource conservation. Today, Bureau activities cover a broad spectrum of programs for meeting the diverse (and sometimes conflicting) needs of mineral industries, government, and the public. In the 1960's, the Bureau initiated programs to resolve environmental problems associated with past mining and mineral processing. Dependence on foreign sources of raw materials critical to our national economy and competition from other countries for these world supplies stimulated the need to conserve and use with intelligence the Nation's minerals, metals, and materials.

Jan 1, 1978

By V. Panayotov, M. Panayotova

"The combined use of semiconductor properties and specific flotation characteristics of minerals under different conditions with external influences ensures a possibility for process control and management. This work presents the relations between the volumetric and surface characteristics of minerals and the minerals' flotation ability (flotability), as well as the possibility for the flotability purposeful change. Results from proprietary laboratory and industrial research are presented showing how controlled impacts on minerals and flotation pulp improve flotation technology. Such an approach reveals prospects for the creation of new technologies for the minerals' beneficiation.INTRODUCTIONThe decrease of the useful components content in the ores, the increasingly complex nature of the raw materials, the increasing energy consumption per the unit of production, and the need to solve the problems related to the environment pollution require solutions different than the existing classical ones. One of the ways to find such solutions is the detailed consideration of the fundamental properties of minerals and, in particular, of minerals having semiconductor properties. The studies of Plaksin et al. (1968), Chanturia and Shafeev (1977), Chanturia and Vigdergauz (2008), Carta et al. (1973), Hoberg and Shneider (1975), Rao et al. (1992) have given an idea of the energy state of the minerals, their flotation ability, and the ways in which this ability can be changed in one or the other direction. The main line in these works is the study and use of the relationship between bulk and flotation properties of minerals. To improve the efficiency of the process, it is also necessary to consider the relationship between the surface properties of the minerals, especially minerals-semiconductors, and their flotation ability.THEORETICAL BACKGROUND IN BRIEFA quantitative estimation of energy changes in a solid phase in the process of flotation may be made by taking in account the changes in the Fermi level. Its position in the forbidden band is an indicator about the direction and the intensity of electron exchange between solid and liquid phases. The Fermi level energy EF enables the estimation of flotability, intensity of needed external impact before flotation, as well as efficiency of the process going on. Quantitative estimation of these processes can be made by means of the flotability coefficient Kf."

Jan 1, 2018

By Walter R. Storms

"The New Mexico potash exploration described herein was begun in February- 1944 by the Federal Bureau of Mines with the object of determining the amount of potash ore in the United States Potash Reserve in Eddy County, N. Mex., by exploring the deposit from the surface with vertical drill holes. The work was completed in October of the same year after 16 holes had been drilled on the reserve and in the immediate vicinity.The Potash Reserve is located in secs. 26, 27, 34, and 35, T. 20 S., R. 29 E., Eddy County, N. Mex., about 14 miles northeast of the city of Carlsbad. Drilling was done in secs. 26 and 35 of the reserve, in sec. 14 to the north, and secs. 25 and 36 to the east. All work was done on Government-owned, unleased land.The reserve may be reached from Carlsbad by traveling northeast by automobile for about 14 miles on paved U. S. Highway 62. This highway passes through the southern part of the reserve. Other parts may be reached by a State highway which passes along the eastern side of the area, then by poor dirt roads to various drilling sites. A branch of the Atchison, Topeka and Santa Fe Railroad also passes across the southern part of the reserve en route from Carlsbad to the plant and mine of the Potash Co. of America, a few miles northeast of the Project and 22 miles from Carlsbad.Work on the project was greatly facilitated by the help of Ralph H. King, associate geologist, R. K. Bailey, geochemist, H. C. Spicer, geophysicist, Hope Bunker, junior geologist, and R. H Allport, district mining supervisor, all of the Federal Geological Survey; and Charles H. Johnson, who at that time was district engineer for New Mexico for the Mining Branch of the Bureau of Mines."

Apr 1, 1946

By Jake Hutchison, Wade Hutchison

The opposition to the project has risen as a result of some political pressures and from public perceptions about the project’s necessity and its impact on the neighborhood. As a result of this controversy, the Underground Conservation Hall was quickly termed “Bredesen’s Bunker” by the Media. Close exposures for blasting, required by the Conservation Hall construction project, have been regulated by specifications that regulate seismic monitoring, pre-blast and post-blast inspections. Planned blasting for the project included blasting to a depth of 30 feet (9.14 meters) within 35 feet (10.67 meters) of the front of the newly renovated Tennessee Executive Residence. Excavation for the project found more competent rock at higher elevations than identified in test drilling which required blasting within 2 feet (0.61 meters) of the side of the Residence.

Jan 1, 2009

By Carlos Rabanal, Levi Guzmán

"The concern regarding grinding media consumption rates is as old as the invention of the tumbling mill. Over the years many materials and grinding media shapes have been tested. Through all this, steel grinding balls have proved to be the most effective media for comminution in tumbling mills. There have been several approaches to be able to the wear of the ball charge in any given mill, all of them based on limited empirical evidence. The mining industry still utilizes the Bond Abrasion test that was developed in the 1960´s. As shown later in this publication, this approach shows error greater than 60%. More recently, in 2007, Radziszewski from Mc Gill University has proposed a different total grinding media wear model; this model is based on decoupling the effects of the abrasion, corrosion and impact wear mechanisms. Also in 2007, Benavente, at the time with Moly-Cop Peru, presented an empirical model relating both ore properties and operational conditions that would affect the wear performance of the grinding media. For the database referred in this article, Benavente’s model shows an average error of ± 12% which represents a higher degree of improvement in comparison with previous approaches. The current paper presents recent research conducted by Moly-Cop Peru to further improve the capability of the empirical model originally developed by Benavente to estimate grinding media wear rates. The work performed included extensive determinations of the Bond Abrasion Index for different types of ores, collected at industrial grinding applications where the corresponding operational information was made available. The results of the research enabled us to build a more robust and confident model for estimating the grinding media wear rate. The average error of this new model is in the order of ± 9% which is a significant improvement over the models previously described."

Jan 1, 2014

By Chen-Luh Lin, Jan D. Miller, Sindhoora Puvvada

The depletion of good-quality phosphate resources, coupled by a need to supply acceptable products at a rapid rate, while meeting the specifications of fertilizer chemical plants, is of significant concern to the phosphate industry, particularly the production of concentrates with tolerable dolomite content. Tremendous efforts have been made in the past to develop an efficient and economical process that can effectively reject dolomite. However, dolomite removal continues to plague the phosphate industry due to several process-specific challenges being encountered with various beneficiation technologies. In this regard, detailed characterization of phosphate reserves is crucial to understand textural limitations and help design appropriate strategies to achieve quality products. In this paper, the development of procedures to establish dolomite rejection/francolite recovery curves for gravity preconcentration of pebble phosphate from Central Florida is described using high-speed X-ray computed tomography (HSXCT). The developed procedure is illustrated with the analysis of high MgO pebble phosphate samples (10.39 mm × 1 mm) received from South Pasture and Four Corners locations operated by the Mosaic company. The rejection/recovery curves were calculated for each sample for five different particle size classes to observe the trend in rejection/recovery with variation in particle size. In order to evaluate the reliability of the HSXCT data, analysis using the high-resolution X-ray computed tomography (HRXCT) technique was performed on one particular sample (FC-3-1). Finally, to assess the separation efficiency, the HSXCT and HRXCT rejection/recovery curves obtained were compared to experimental gravity preconcentration results reported in the literature. In this way, liberation limitations for gravity preconcentration in the processing of pebble phosphate have been established based on X-ray computed tomography analysis.

By John F. McElhiney, Hosseim Kazemi

This chapter is a review of fluid flow and mass trans- port in fractured rocks. The topics include: single-phase and multiple-phase flow theory, formation productivity or injectivity improvement by artificial fracturing, and mass transport associated with fracture acidizing. Possible extension and application of such ideas to in situ mining in fractured rocks are presented.

Jan 1, 1974

By J. Dvorácek

Similar to the case of any other business venture, the activities of mining industries are profit-oriented. Nothing will stop successful businesses in their quest for new localities to be prospected for minerals of their choice, because the useful economic life of any mine is limited. Before starting the process of land acquisition, it is important for any mining company to be able to forecast the profitability of the venture. The focus of this paper has been on the prospects of success, concerning the case of the opencast extraction of aggregates for construction. A set of 70 successful and loss-making quarries was subjected to investigation. An international company owns the quarries, which are located in seven European countries. The method of discriminate analysis provided the criteria for assessing future economic performance. The same criteria served the purpose of classifying other quarries belonging to the company.

Jan 1, 2012

By T. Hudaverdi

Control of the particle size distribution of a muckpile after blasting is always an important subject for mining industry. Blasting has a significant impact on downstream processes such as loading, crushing and grinding. Improvement of blasting results provides increase in loader and excavator productivity due to increased diggability capacity, and increased bucket and truck fill factors. Suitable and uniform particle size distribution results increase in crusher and mill throughput and decrease in energy consumption in size reduction process. Digital image analysis technique has been used for many years in mining industry in order to determine blast fragmentation. The images of muckpile surface can be used to determine the mean fragment size (x50). The source of digital images can also be haul truck, leach pile, waste dump, stockpile, and conveyor belt. The screening that is conventional method to measure blast fragmentation is time consuming and expensive. Digital image analysis is a cost effective and faster method (1).

Jan 1, 2012

By T. Güyagüler, A. Güngör

Labour protection, that is prevention of occupational disease and reduction in the frequency of accident, has always been a matter of major concern of mining industry. Management and the government should promote and maintain high safety standards through some measures to reduce accident frequency and occupational disease. This can be achieved by using the science and technology for improving the working conditions, educating workers and managers about accidents and prevention techniques. The detailed analysis of an accident requires knowledge of many factors such as location, time, type, cost of accident, victim, nature of the injury, personal and environmental factors etc. that can be obtained from the standard coded accident report form. There is a need to prepare safety data in a common format to produce a common safety reporting system and ‘Occupational Health and Safety Management Tool’ is being designed to manage this by using database and Internet technology. Precautions should be taken according to results from database queries by providing comprehensive and accurate national data. To have an accident is unfortunate, however, to have an accident and learn nothing from it is unforgivable. Keywords: Occupational health and safety, Accident report form, Accident database, Occupational diseases and accidents, Safety reporting system.

Jan 1, 2003

By J. R. Thoenen

This circular is part 2 of the third subject (entitled "Excavation") of a series summarizing; the technical problems involved in the production and preparation of sand and gravel. Part 1 discussed the use of power shovels, draglines, and excavator cranes. This paper discusses the use of the power scraper, slackline cableway excavator, and hydraulic monitor, and part 3 will discuss the use of various types of dredges. Subsequent papers in the series will discuss the use of haulage equipment, the application of both excavating and haulage equipment to various types of deposits, and operating costs. THE POWER SCRAPER The power scraper is a younger member of the various forms of gravel excavators. It is a mechanized enlargement of the slip scraper and other scrapers which depend upon animals for power and has a greater range of operation from a single set-up than any other excavation unit except the slackline cableway excavator.

Jan 1, 1935

By C. H. Lee, K. Adham, D. Lemieux, J. P. Mai, A. M. S. Hussein, S. Francey

Copper concentrates contain an increasing amount of arsenic, as the older sources of cleaner ores are depleted. At the same time, the market-determined price penalties on arsenic rich concentrates are rising due to limited arsenic treatment capacity at smelters and increasingly stringent environmental emissions limits, necessitating arsenic removal from the copper concentrates. Heating the concentrate to about 700°C can remove significant amounts of arsenic (as elemental and/or sulfide); however, it is important to achieve that removal economically and with due consideration to post-roasting arsenic (toxic waste) treatment. A comparison is made among different roasting options (kilns, multi-hearths and fluid beds) to qualify their potential application to arsenic roasting of copper concentrates. Both direct and indirect methods of heating as well as inert, air and oxygen atmosphere options are considered. Technology comparisons are made with respect to the roaster’s operation, its off-gas treatment and arsenic stabilization. INTRODUCTION In the last twenty years, the arsenic content in globally produced copper concentrates have risen steadily due to the depletion of cleaner ores and the increase in complex concentrate production. Smelter contracts typically limit arsenic to less than 0.5wt%, with a penalty paid per unit of additional arsenic and other contaminants such as antimony. While penalties can vary, they can be on the order of $70/dry metric tonne of concentrate per 1% of arsenic. Strategies used by miners to avoid the price penalties have been to blend clean and complex concentrates, and to remove arsenic from complex concentrates to achieve smelter limits. Arsenic can be removed from copper concentrates by the heat-induced decomposition of the predominant enargite mineral (Cu3AsS4) at ~700°C. Roasting at such elevated temperature is a commercially proven option, but still suffers from process, cost and environmental challenges. After roasting, the released arsenic sulfide fumes can be separated from most of the entrained concentrate dust in cyclones. Then, the As2S3 laden off-gas can be oxidized to generate a still volatile oxide (As2O3), which can be further separated from the residual concentrate dust in a hot electrostatic separator. Next, the off-gas can be cooled with air or water, to either deposit the arsenic oxide in a dry or dissolved state. Dissolved arsenic acid can be oxidized to As(V) and reacted with iron to yield a stabilized ferrihydrite precipitate or scorodite that can be safely disposed. In the case of dry arsenic trioxide, it can be mixed with reactants (e.g. iron) and fluxes before thermal stabilization in a glass matrix, thus eliminating the need for costly oxidation.

Jan 1, 2019

During the second quarter of 1983, the economic recovery showed further strength abetted by the performances of consumer spending, residential investment, and the slowing down of inventory reductions. Real gross national product (GNP) increased from the first to second quarter at an annual rate of 9.2 percent, compared with 2.6 percent from the fourth quarter of 1982 to the first quarter of 1983 and an overall 1.9- percent decrease in 1982. Total industrial production increased 1.8 percent in July, and productivity as measured by output per man-hour increased 1.3 percent between 1982 and the first quarter of 1983 (table 1). Especially noteworthy .was the performance of real consumer durable goods expenditures, which increased at an annual rate of 32.0 percent from the first to the second quarter (about half of the increase consisted of automobile sales) and real consumer spending, which increased between these periods at an annual rate of 9.7 percent. Residential investment increased at an annual rate of 76.1 percent between the first and the second quarters,1/ and private housing starts by about 38,000 units in both March and May.

Jan 1, 1983