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By Keith K. Griffiths

Computers are a valuable tool to the aggregate industry and should not be overlooked. The selection of proper mechanical components is necessary. Merely the addition of a computer to operate a mechanical system will not guarantee that the system will work. Material delivery by rail or truck for stock-piling and intra-plant material transfer can be accurately supervised by automation with simplified controls for the layman to use. There are mechanical problems that occur in material blending and load-out systems when such a system is automated. Such things as material flow, bridging, stockpiling methods, overloading of equipment and safety must be considered. How these problems can be overcome by proper selection of compatible mechanical and computer equipment will be discussed. Also, there are many individual material circuits in a plant that can be computerized at a nominal cost. This type of computer installation will not necessarily result in increased production, but it will produce greater efficiency in manpower. These small computers, binary counters, memory circuits and other devices and their application to the aggregate industry will also be presented. A careful evaluation of the mechanical and computer inter-relationship by someone understanding both "Rocks and Electronics" is essential to proper handling of material by computers. Computers for the mining industry may be either digital or analog computers. Unless the computer is to control some type of chemical process, such as a cement plant, the, analog computer is not required and the digital computer' will be quite adequate and less expensive.

Jan 1, 1967

By Albert J. Phillips

When brasses containing from 61 to 62.5 per cent. copper are very rapidly cooled from temperatures near their melting point to below 0" C., unusual results are obtained. These results are quite unexplainable from a consideration of the old copper-zinc constitutional diagram but they can be satisfactorily explained by means of the diagram revised by Genders and Bailey1 (Fig. 1). Experiments on a 60/40 Brass If an alloy containing 60 per cent. copper is heated to any temperature between approximately 760' C. and its melting point, it will consist entirely of beta. Upon slowly cooling, alpha will start to separate just below 760" C. Since this alpha will contain nearly 65 per cent, copper, the beta will be impoverished by its separation. As the temperature decreases the composition of the separating alpha will approach approximately 61 per cent. and the beta in contact with the alpha will maintain an equilibrium composition approaching 54.5 per cent. copper. If the rate of cooling is extremely slow, especially at the lower temperatures, stable equilibrium will be established by diffusion through the alpha and beta so that eventually 85 per cent. of the alloy will consist of alpha containing 61 per cent. copper, while 15 per cent. will be beta containing 54.5 per cent. copper. With a faster cooling rate diffusion cannot keep apace and the constituents will be cored, the alpha varying from 65 to 61 per cent. copper and the beta from 60 to 54.5 copper. In this case the amount of alpha will be less than 85 per cent. and the beta will be increased correspondingly. The general structure thus obtained is illustrated in Fig. 2. When a more rapid rate of cooling is effected by quenching in water, the alloy is retained largely in the beta form. However, the crystals of beta are usually outlined with dark fringes which at high magnifications are found to be very fine alpha-beta structures almost martensitic in appearance. The separation of this fringe can be greatly suppressed

Jan 1, 1930

By B. Steyn

AngloGold Ashanti?s Obuasi Mine, in Ashanti Province, Ghana, operates a plant utilizing bacterial oxidation technology. The mineralization consists predominately of arsenopyrite with gold associated in the sulphides. The plant was commissioned in 1994, at that time was the largest BIOX® plant in the world, and has been operated by AngloGold Ashanti since 2004. The constraints identified in the plantwere the flotation capacity, BIOX® feed size and disruptions to the BIOX® process. This paper focuses on the practical steps that were followed in order to increase the recovery of the plant. The original plant design was based on a head grade of 10 g/t and an overall recovery of 86.5%. In 2007 the plant was operating at an overall recovery of 79.2% at a head grade of 5.6 g/t and this has improved to 80.6% at a head grade of 5.8 g/t during the last quarter of 2008. Additional improvement to 81.5% are expected by the end of April 2009. Improvements that have been implemented include: ?Replacement of the regrind mill ?Installation of additional float capacity ?Installation of a scavenging circuit ?Installation of a complete water treatment plant. In addition we have focused on the effect of the water circuit on the BIOX® process and the steps involved in correcting this problem. The water circuit was originally designed as an open circuit and through environmental considerations the circuit was closed towards the end of 2007. The bacteria in the BIOX® are particularly sensitive to cyanide and cannot tolerate levels above 0.05ppm (free cyanide), this required extensive modifications to the water circuit. The plant improvement project is ongoing and the current scope of work should be complete by the end of April 2009. Further research and development work is being planned, which includes a ?white rot bacteria? that can consume the preg-robbing carbon and oxidize the sulphide in a single pass.

Jan 1, 2009

The four major contracts for the construction of Museum Station are: 1) Installation of Soldier Piles, 2) Structural Contract including the temporary support, 3) Architectural Finishes which include Electrical and Mechanical Services and 4) Supply and Installation of lifts and escalators. Specific areas of interest include the fabrication, installation and removal of temporary support steelwork, the monitoring of ground loading, and the interface between the structure and. the temporary support. The information collated during construction confirmed the design assumptions and criteria. No major changes'were required in the design of the construction methods. The open cut excavation and construction of this major structure in an inner city area was significant by any standard and was successfully completed with minimum disruption to the public and without damage to adjoining or surrounding property.

Jan 1, 1981

By Lyndon L. Foley

The proper spacing of oil wells is a problem of vital importance to the oil industry. Conservation demands a maximum recovery, while economic considerations attach primary importance to profitable extraction of oil. Formulas have been evolved,1-4 which attempt to solve the spacing problem mathematically, using reservoir characteristics as factors. Comparisons have been made of fields producing from supposedly similar sands, which have been drilled to varying densitiesl5,6 and also of different leases in the same field. 5-8 Intermediate wells have been drilled in old, depleted fields to test the efficiency of drainage by the original operation.9-11 Intermediate wells have been drilled also to obtain cores of the depleted sand for analysis, from which to learn the degree of efficiency of extraction. The problem is being attacked by many students, and along several different lines. Unfortunately, the fields now depleted were produced generally without restriction, and conclusions drawn from them may not apply to fields operated under restricted production. Reservoir conditions in a restricted field, especially reservoir pressures and pressure gradients within the reservoir, are quite different from those in a field produced to capacity. Many people have held the idea that the closer the spacing, the greater the recovery. This general statement is not always true, as other factors than well spacing enter into the recovery problem. Cutler,5 who has been quoted frequently by advocates of close spacing, limited his conclusions to gas-drive fields, and to the particular fields and spac-ings that he investigated. Hill and Sutton7 analyzed four tracts in the Powell field; the tract most closely drilled had the third largest recovery per acre-foot, and the tract with the widest spacing had the second largest recovery per acre-foot. A search of the references to well spacing in the publications of the U. S. Bureau of Mines5,13-19 discloses repeated condemnations of the wastefulness of the practice of town-lot drilling.

Jan 1, 1938

By S. Hamm

Elevated total dissolved solids (TDS) content in waters downstream from surface coal mines in Central Appalachia is an increasing concern, as unnaturally high TDS and/or associated specific conductivity (SC) may contribute to deteriorated stream ecosystem health. The aim of this preliminary study was to explore the efficacy of accelerated leaching as a means of minimizing long-term TDS/SC generation by problematic overburden materials. Five spoil samples were collected and leached using simulated rainwater solutions under ambient and acidified pH conditions. SC and pH were monitored during each leach cycle, and ICP-MS was used to determine how leaching conditions affected release of specific ions. This paper highlights the results of both jar and column leaching tests. Among the most significant findings was that, for certain spoil materials, SC in leach column effluent was effectively reduced when spoil was leached first with acidic rainwater followed by ambient rainwater (vs. ambient rainwater leaching alone). This suggests that accelerated leaching might aid in long-term stabilization of some materials.

Feb 23, 2014

By Franklin D. Slagle

Porosity and bulk-and sand-grain-density values were determined for 2,040 core samples of 30 formations from 47 wells in New York, Ohio, Pennsylvania, and West Virginia, and are compiled for petroleum industry use in well-log evaluation and geologic studies.

Jan 1, 1967

By M. Zamalloa, P. J. Mackey, P. Coursol, Y. Prévost

"Over the last decade or so, the availability of commercial thermodynamic modelling packages, such as FactSageTM, have been utilized to help the plant optimize slag chemistry and to minimize slag handling issues that can arise, such as those due to a process upset, for example, by undue feed variability. This paper illustrates an extension of the application of this software in examining a large range of NCV operating parameters, such as the %Fe/SiO2 ratio in the slag and the level of minor gangue components in the slag including CaO, Al2O3, ZnO, MgO, and PbO. The conditions influencing the slag liquidus range of NCV slag are discussed with respect to the overall process chemistry.Given that a sufficiently low %Fe/SiO2 ratio is used in NCV slag, most components such as CaO-ZnOPbO- Al2O3-MgO, in small concentration, all contribute to lowering the NCV slag liquidus. Considering the present level of these minor components, the NCV can operate consistently at 1200°C if the %Fe/SiO2 is between 0.6 and 0.8 in the slag. Using the type of phase diagrams presented in this paper, it is possible to develop different Noranda reactor/Noranda converter operating strategies adapted to the feedstocks available at Xstrata mines/plants or from other sources. The NCV is unique in that a fayalite-based slag is employed for high-efficiency continuous converting with excellent operating results."

Jan 1, 2007

By Alexander L. Holley

THE 400-ton testing machine, ordered in June, 1875, by the United States Board appointed to test "iron, steel, and other metals," has lately been completed at the Watertown Arsenal, thoroughly proved and accepted by the Board. The excellence of the machine in every respect is more than satisfactory, and its accuracy is at first sight astonishing, although an investigation of its principles must show that if the weighing apparatus will weigh at all, it must do so with perfect accuracy, because all its movements are absolutely without friction. The proof experiments were numerous, and the effects of recoil after sudden ruptures at maximum loads, were watched with great care, but without much anxiety, because the weighing parts affected are by no means delicate in structure, and their motion is almost infinitely small. Among the tests were the following : A forged link of hard, wrought iron, 5 inches in diameter between the eyes, was slowly strained in tension, and broke short off with a loud report at 722,800 pounds. The diameter before breaking at the point of fracture was 5.04 inches; after breaking, 4.98 inches. In order to see if the weighing parts had been disturbed by the recoil, which was obviously near the greatest recoil the machine will ever suffer, a horse-hair was next tested ; it was 7-1000ths of an inch in diameter; it stretched 30 per cent., and broke at 1 pound. Other horse-hairs vary in tenacity between 1 and 2 pounds. Of course, the accuracy of the machine on such delicate specimens, and indeed on specimens having some hundreds of pounds tenacity, has been checked and proved by other weighing machines. A 5-inch round bar, turned down to 3 5/8 inches diameter along the centre, was pulled apart at 430,200 pounds tension. Then some more horse-hairs were tested ; also copper wires 19 1/2-1000ths of an inch in diameter, which averaged 25 pounds tenacity..

Jan 1, 1879

By Tabbetha A. Dobbins

We report studies concerning the time frame for catalytic reactions between NaAlH4 (sodium alanate) and TiCl3 during high energy milling. Spectral features in x-ray absorption spectra were analyzed for samples milled for various times (0 minutes, 1 minute, 5 minutes, 25 minutes, and 125 minutes). A structural transition from Ti3+ to Ti0 is observed within the first 5 minutes of milling. The Ti0 structure persists for samples milled for times longer than 5 minutes?even after those samples underwent a subsequent single hydrogen desorption/absorption cycle. Samples milled for less than 5 minutes retain Ti3+. For samples milled for 1 minute, the Ti-K edge position shifts from Ti3+ to Ti0 after a single desorption/absorption cycle. However, the first coordination sphere around the Ti0 absorber occurs at longer distances for 1 minute milled sample. These results hint at the formation of Ti0 pure metallic clusters upon desorption/adsorption cycling of 1 minute milled sample. Previously reported x-ray absorption spectroscopy studies have demonstrated the structural transition from Ti3+ (in TiCl3) to Ti0(determined to be TiAl3 found at the surface of the alanate powder). Aluminum deficiency on the NaAlH4 lattice is responsible for controlling hydrogen absorption capacity. These time-resolved spectroscopy studies suggest the possibility for controlling aluminum deficiency, while still maintaining titanium activation of the alanate powder, by manipulating the coordination environment of the titanium dopant during the first five minutes of high energy milling. Specifically, a possible approach to mitigating aluminum deficiency during doping would be to introduce an alternative species to which the to Ti3+ catalyst would form a metallic bond after the TiCl3 has attached at the surface of the sodium alanate powder and before the Ti3+ forms the TiAl3 intermetallic species.

Jan 1, 2006

By H. E. Miard

"IntroductionIT HAS BEEN SAID, with good reason, that the future of the coal mining industry, in our mountainous regions at least, depends on ability to extend operations successfully to depths greater than those reached until now. However, at many points, Nature has given definite hints that trespassing there could be fraught with very un-pleasant possibilities, unless the intruder be governed by a series of rigidly observed principles. In the coal mines of British Columbia, so far, bumps have taken place only at the Coal Creek colliery, but there they gave ample evidence of the extent to which they can hamper underground operations. Some of these manifestations have also been experienced at the McGillivray Creek colliery, in Alberta, at Carbonado, in Washington, and at Sunnyside, in Utah.Cognate phenomena, known as 'air blasts', have been experienced also in some deep metalliferous mines in British Columbia. Such occurrences were responsible for three fatal accidents at the Le Roi and War Eagle mines, near Rossland, in 1916 and 1917.In more or less disturbed coal fields, another form of stress release, equally detrimental, has been still more generally experienced than have been the bumps. Coal outbursts have taken place at Morrissey and Coal Creek, in the Crow's Nest Pass district, and at Cassidy, South Wellington, and the Nanaimo Reserve colliery, on Vancouver island. Manifestations of this kind have been made the subject of intensive study in European countries where they occurred on a really impressive scale, which must not be construed as meaning that some of our own outbursts did not approach a comparable magnitude."

Jan 1, 1953

By Thomas O. Glover

Economic growth in Cameroon was restricted during 1987 owing to a decrease in oil prices during 1986 and the falling value of the U.S. dollar in 1987. The yearly budget reflected a significant drop in revenues and expenditures. Taxes on imports and some exports were increased to raise revenue. The shortage of revenue and the lack of liquidity had restricted the Government's operation and investment programs. External public debt reached an estimated $3.2 billion2 representing 20.8% of the estimated gross domestic product (GDP) of $12.3 billion. The Government operated at an 8.7% deficit in 1987. The GDP growth rate had fallen to 0.6% in 1987 and was projected to remain at that level in 1988. Economic planning was managed through a series of 5-year plans. The country was under the sixth 5-year plan, which would end in 1991. The plan stressed balanced development as one method to achieve the desired results. Sources of revenue for the year included direct and allied taxes (31.2%). customs collections (31.0%), and oil royalties (28.0%). The oil royalties in 1987 represented a larger part of the budget than in 1986. In January 1987, eight African oil exporting countries, including Cameroon, formed the African Petroleum Producers' Association (APPA). The eight countries represented about 80% of the total African oil production of 5.2 million barrels per day in 1987. Cameroon produced approximately 3% of total APPA oil production. Cameroon produced approximately 68.5 million barrels of crude oil in 1987, slightly less than that produced in 1986. Earnings also fell in 1987, not only from falling production, but also from declining oil prices and the fall in value of the U.S. dollar on world markets. Oil had been the main stimulus of Cameroon's growth throughout the past decade, and a key generator of export revenues and financial resources for the economy.

Jan 1, 1989

By I. Orford, V. Rozon

"The Sä Dena Hes Joint Venture Mill was designed to process 450,000 tonnes per year of ore grading 11.3% lead and 13.2% zinc.Run of mine ore is crushed in a jaw crusher and then ground in a closed circuit SAG mill and ball mill. A standard differential float in conventional flotation cells produces quality lead and zinc concentrates. Dewatering of concentrates is achieved using high rate thickeners, lamella clarifiers and pressure filters. Process water is 100% recycle.Plant design was based on metallurgical test work performed on drill core samples. The plant is operating at better than design specifications.INTRODUCTIONSä Dena Hes Joint Venture was formed in 1989 to develop the Sa Dena Hes lead/zinc sulphide mine, located approximately 45 km north of the Town of Watson Lake, Yukon. The mine is owned by a Joint Venture partnership between the operator, Curragh Inc. (87.5%) and Hillsborough Resources Limited (12.5%).The project was completed in a record ten months, working throughout the severe Yukon winter and through an abnormally wet spring and summer. The first truckloads of concentrate were shipped from the mine on August 1, 1991.The mill currently processes 550,000 tonnes of ore per year grading 6.9% lead and 11.5% zinc. Ore is crushed in a jaw crusher and ground to 60% passing 75 microns in a closed circuit operation of one SAG mill and ball mill. Using conventional flotation cells a 150,000 tonnes of concentrate is produced in a year with lead concentrate grading 78% lead and a zinc concentrate grading 60% zinc at overall recoveries of 89% and 90% respectively.Concentrates dewatering is accomplished using high rate thickeners, clarifiers and pressure filters. The lead concentrate and zinc concentrate contain 5.4% water and 7.8% water respectively. Concentrates are trucked to Skagway, Alaska for subsequent shipping to overseas smelters."

Jan 1, 1993

By H. R. MacMichael

THE rate, at which zinc ore can be sintered varies widely, depending on the characteristics- of the ore. A coarse table concentrate already roasted to a very, low sulfur content may sinter at the rate of-8.6 tons per hour per 100 sq.ft., of wind-box area. At the other extreme is a finely ground flotation, concentrate, with considerable sulfur remaining in the ore as delivered to the sinter plant, for which corresponding rate may be only 3.8 tons per hour, or 44 per cent of the first rate. A usual, size-of sinter machine in zinc plants is one with wind-box 3.5 by 33 ft., making 115.5 sq. ft. of area, and using a 100-hp. fan. Such a machine with the coarse concentrate described would produce 10 tons of sinter per hour; and with the fine, flotation concentrate would produce only 4.4 tons per hour. Ore of the latter class is -sintered at the plant of the Mexican Zinc Co. (subsidiary of the American Smelting & Refining Co.) at Rosita, Mexico. ; Consequently, an unusually large sintering capacity per ton of ore to be treated is required, and provision was made for treatment of 9000 tons concentrates per month, preroasted to. 5 to 9 per. cent sulfur. For this duty three of the usual 33-ft., 100-hp. sintering machines would have been needed.. Considering the auxiliary equipment. necessary for proportioning, handling, and mixing charge, as well as means for disposition of the sinter, a three-unit plant becomes-very complicated, especially if the capacity should be doubled later. After much study it was decided to build a single sintering unit of ample capacity for the work, the plant thus being much simpler and cheaper in construction as well as in operation (Figs. 1 and 2). The sintering machine was built with a wind-box 5.25 by 60 ft. (315 sq. ft.) and provided with a 300-hp. fan.

Jan 1, 1932

By Eero Pokki

The Päijänne Tunnel is being constructed to conduct raw water from Lake Päijänne to the Helsinki Metropolitan Area. The length of the tunnel will be 120 km and the cross-sectional area is 15.5 sq m. Excavation started in 1973 and will be completed in 1982. 94 km of the tunnel have been constructed from Lake Päijänne towards Helsinki. The construction has advanced according to plan, and due to Finland's hard bedrock there have not been any great surprises. When completed, the tunnel will meet the water requirements of the Helsinki Metropolitan Area for a long time in the future because, with a free fall of 36 m, 9.6 cu m/s of water can be conducted through the tunnel from Lake Päijänne to the Helsinki Area. With pumping it will be possible to obtain a flow of 15 cu m/s, although at first the discharge will only be 5 cu m/s. The tunnel was partly taken into service as early as March 1979 when it was possible to conduct water to the municipalities north of the Helsinki Area. It is already possible now to conduct water to Helsinki, too, but from the Korpimäki pumping station onwards it must be done through the River Vantaa, which is polluted. When the tunnel is completed in 1982, good quality water can be conducted from Lake Päijänne to the Helsinki Metropolitan Area through a continuous rock tunnel. The cost estimate in 1973 was $ 88 million while it is now $ 90 million notwithstanding high inflation. AREAS NEEDED FOR CONSTRUCTION The areas needed for construction were bought through voluntary deals. Areas of 4 - 5 hectares were bought along the tunnel line at about 5 km intervals. The excavation of access tunnels was started from these places, and the crushed rock from the main tunnel was also transported there. However, it was not necessary to buy the area

Jan 1, 1979

By Gavin Beer

The Chelopech Copper/Gold Mine, located in west-central Bulgaria, has been in operation since 1953 producing a low-grade copper/gold concentrate containing up to six percent arsenic by weight. The penalties and restrictions imposed on smelting this concentrate, have seen the past operators of the mine strive to find a balance between achieving acceptable metal returns from the smelters and installing novel or unproven technologies in the face of changing political and economic climates. This paper documents the colourful history of the concentrate processing alternatives that have been investigated and implemented over the 45 year life of the Chelopech Mine, culminating in the high pressure oxidation option that is now the focus of a Definitive Feasibility Study (DFS) presently being undertaken by GRD Minproc Limited. The fundamental chemistry of the preceding alternatives is outlined followed by a discussion as to why each was ultimately not employed. A detailed description of the technology proposed in the DFS is given including: pressure oxidation, arsenic fixation, SX/EW copper recovery, CIL gold recovery and AVR/SART cyanide/copper recovery processes. Keywords: Arsenic, Scorodite, HPOX, Chelopech, Dundee

Jan 1, 2005

By K. Horikawa, S. Kitahata, H. Kobayashi

"It has been clarified that the Al-5%Mg alloys containing 200 mass ppm of sodium showed grain boundary (GB) fracture, even when tested at room temperature, as well as at high temperatures (~500°C). SEM/EDX study revealed that the GB fracture was associated with the segregation of sodium on grain boundaries of the Al-Mg base alloys. Differential scanning calorimetry (DSC) revealed that the GB fracture at room temperature of Al-Mg alloys was caused by sodium segregation in the solid state. It was also shown that the tendency of the GB fracture strongly depended on the heat treatment; the alloys after the solution heat treatment showed severe GB fracture compared with the alloys after the homogenizing treatment. Moreover, it was shown that the GB fracture at room temperature by sodium segregation was suppressed by the addition of bismuth or indium. The suppression of the GB fracture by the additional elements was believed to be the scavenging effects by forming the sodium bearing compounds in grains.INTRODUCTIONIt is widely known that Al-Mg base alloys containing high amount of magnesium(~5 mass%)show better workability at room temperature, while that the alloys show grain boundary (GB) embrittlement at higher temperatures ranging from 200 to 500°C, depending on grain size(Otsuka, 1984), sodium impurity level(Ransley, 1959, Okada, 1997, Horikawa, 2001)and testing strain rate(Yamada, 2012). In the case of coarse-grained Al-5%Mg alloys (~300 ?m), 2 mass ppm of sodium causes severe GB fracture around 300°C, however, the Al-5%Mg-2ppm alloy does not show GB fracture at room temperature. It has also been shown that the GB embrittlement caused by 2ppm of sodium was suppressed by the additional elements, such as bismuth (Talbot, 1995), antimony (Ueda, 1997) and indium (Horikawa, 2012). Basedon these facts, we have concluded that the GB embrittlement at higher temperatures of Al-Mg alloys is attributed to the synergy effects of GB segregation of atomic sodium and GB sliding. For the GB embrittlement mechanism of Al-Mg alloys, involvement of low melting point metallic phases has been pointed out (Lynch, 2002),since the GB embrittlement appeared only at high temperatures. Up to date, the effects of high amount of sodium addition more than 2 ppm on GB fracture were not fully investigated in Al-Mg alloys. In the present study, the effect of 100 ppm ordered sodium impurity on GB fracture of Al-5%Mg alloys was examined, together with the effect of another additional elements."

Jan 1, 2017

Upgrading of MgO in Dolomite Ore by Froth Flotation

Sep 13, 2010

By Jorge Vargas Fernández

Con el objetivo de incrementar la productividad de la minería, se plantea la realización de una positiva acción nacional de formación y capacitación en todos los niveles de la industria minera a través de un análisis de la situación actual respecto de los mineros, ingenierios y técnicos. Por ello, el presente texto señala la descripción del sistema "B" en funcionamiento de la productividad minera y la tecnología adecuada a la realidad.

Nov 1, 1972

By Stanley A. Bunk, Desmond J. Bruyns

The purpose of roasting gold ores is to release for extraction, typically by cyanidation, the small particles of gold which are surrounded by refractory stone. The roasting simply opens up passages for the penetration of solution into the interior of the ore particles. This is accomplished by the removal by volatilization or formation of volatile oxides of certain constituents such as sulfur, arsenic or antimony. Temperature is an important parameter. High temperatures tend to form a dense iron particle rather than a "spongy" calcine. The dense iron particles result in lower gold recoveries. High temperature can cause melting of some components which also results in gold encapsulation. Roasting of refractory gold ore concentrates has been practiced for decades. Multiple hearth, rotary kiln and muffle reactors were first used for roasting. Fluid bed roasting provided a low capital cost, low maintenance technology with better process control and soon became the favored technology. The first fluid bed concentrate roasters were commissioned in the late 1940's. Early fluid beds were "bubbling" type. Environmental considerations did not significantly impact on the design. Feedstocks were highly exothermic and reaction rates were relatively rapid. Roasting today must compete with other technologies for treatment of refractory ores. Ore bodies which are not

Jan 1, 1993