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By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By A. V. Glembotsky, N. A. Greshnova, V. P. Seregin, V. A. Glinkin

Based on a new alternative depressant of zinc sulphide - dimethyldithiocarbamate- modern technology has been developed for the treatment of complex polymetallic ores. In addition to its general function (depression of sphalerite and iron sulphides to some extent), this double-acting reagent is a promoter for galena and silver minerals. Laboratory and plant tests, as well as industrial experience have all shown consistently higher silver recovery to the lead concentrate as compared with sodium cyanide, resulting in less pollution and a safer environment

Jan 1, 1995

By Bernie W. Martin, Duncan MacLennan

Building a new tunnel to accommodate CN North America's needs for double stacked container traffic for the next 100 years presented new and unprecedented challenges. Innovative approaches to contracting, especially the Negotiated Compressed Process (NCP), TBM design, project management, and the regulatory and approval process, were all required to deliver a new tunnel to meet the needs of CN North America within the approved completion schedule of 39 months compared to a traditional approach of 45 to 48 months. The NCP resulted in trimming 6 to 9 months from the project schedule, at a great saving of direct and indirect costs. The Negotiated Compressed Process is an innovative formalized approach to risk sharing for complex infrastructure projects. Ads were placed in the international trade journals and a short list of "approved" contractors and suppliers was developed. The approval process for this international project involved contacts with over two hundred agencies, departments and interested stakeholders.

Jan 1, 1995

By B. K. Cantrell, D. H. Carlson, W. F. Watts, K. L. Rubow

The use of diesel equipment in underground coal mines is controversial because the National Institute for Occupational Safety and Health (NIOSH) regards "whole diesel exhaust" as "a potential occupational carcinogen." Threshold limit values (TLVs) are recommended for the gaseous pollutants CO, CO2, NO, N02, SO2 and some hydrocarbons emitted in diesel exhaust. There is, however, no TLV recommended for diesel exhaust aerosols (DEA) nor is there a standard method for sampling these aerosols. The University of Minnesota and the USBM have collaborated to develop a personal diesel exhaust aerosol sampler (PDEAS) that utilizes size selective, inertial impaction and gravimetric analysis. During the field tests of this sampler, numerous air quality measurements were conducted in underground coal mines that used diesel equipment. The objective of this report is to present these data and to assess the impact of diesel face haulage equipment on mine air quality. Air quality data were collected in four underground coal mines. The mines used continuous miners and diesel haulage equipment. Daily samples were collected at the section intake, haulageway and on haulage vehicles. Both full-and partial-shift samples were collected using a variety of aerosol instruments. These included the PDEAS and the micro-orifice uniform deposit impactor (MOUDI) that sample the mine aerosol diesel exhaust fraction. Two of the mines were also monitored for the major gaseous constituents found in diesel exhaust CO, CO2, NO, NO2 and SO2. These measurements were made by either collecting bag samples for subsequent gas analysis or by Palmes dosimeters. Concentrations of CO, C02, NO, NO2 and SO2 were well below regulated levels. The mine average DEA concentration determined by the PDEAS in the haulageway for the three mines surveyed was 0.79 mg/m3, with a standard deviation of 0.19 mg/m3. DEA contributed 65% of the respirable aerosol, at this location. These results were the same as the mean DEA concentrations calculated from size distribution measurements obtained using the MOUDI. This was 0.78 mg/m3, with a standard deviation of 0.22 mg/m3.

Jan 1, 1992

By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By Anders Sellgren, K. Hansson, A. Jedborn

In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore. from a system's point of view, slurry transportation would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alternative examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasibiliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. 1. The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then temporarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe- line, and personnel are needed. In a first phase (Fig. 2). a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1990

By Jack S. Watson, Rasika Nimkar, Robert M. Counce, BRADLEY BENNETT, Patrick Zhang

The purpose of this work is to develop a conceptual recovery process for thorium dioxide (ThO2) from its primary source, monazite. The study-level design presented in this paper represents an economically viable method for conducting ThO2 extraction from monazite ore while also recovering byproducts that include P2O5 and other rare earth element oxides rather than producing unnecessary waste streams. Three key points to note in the development of this process design are: (1) the process economics relied heavily on recovery of rare earth byproducts, (2) the thorium handling portions of the process could effectively be segregated from the rare earth and phosphate handling portions of the process, and (3) the thorium and uranium content in waste streams should be carefully managed and eliminated where possible.

By D. Gu

Breakage of rock by machine is one of the major methods of excavation. With the development of modern engineering techniques, the advantages of using this method have become even greater. Improving efficiency and reducing the cost of rock breakage is the main goal of mechanizing the rock-breaking operation. For a given stratum, the efficiency of a machine excavation depends, to a large extent, on the design and method of application of the tools used. In a planetary cutter head, the subcutter heads move like planets around the main drive shaft. The general structure of the cutter head is shown in Fig. 1. As shown in this illustration, the main drive shaft (1-1) is connected to the tie bar (1) and, thereby, to a gear(5) that meshes with a fixed-gear ring(2). The subcutter head (4) is rigidly connected to the gear (5) by the subcutter-head axis (2-2). When the main drive shaft (1-1) rotates, the subcutter head (4) rotates around the axis (2-2), under the action of the tie bar and with the help of sleeve pipe (3).At the same time, axis (2-2) also rotates around the main shaft (1-1). As shown in Fig 1, is the angle between the main shaft and the subcutter-head axis.

Jan 1, 1995

By H. Keith Henson

The purpose of this paper is to describe the possible application of two relatively new technologies: flat load cells, and improvement seeking computer programs (referred to in the literature as heuristic programs); to a very old problem, inventors control of ball mills. The combination of these two technologies seems most applicable to grate discharge mills operated closed circuit. It is possible that other applications such as controlling the inventory of autogeneous mills might also be found. Direct control of ball mill inventory has been the goal of a number of investigators over many years. Many patents and articles on the subject have claimed success, but operating installations are hard to locate. An excellent review of the advantages and problems of direct inventory control is contained in an article by Oswald and Ziegler (1). Figure 1 is a conventional representation of throughput vs. inventory with other factors such as slurry consistency and rock twee held constant. There are common sense, physical arguments? and experimental evidence supporting the existence of a unimodel curve of this twee? though direct measurement of the shape is difficult and how it moves as rock type changes is riot well known. Figure 2 represents the problem faced by all mass measurements of inventory, a tare weight of about 25 time the inventory mass. Worse yet, the tare is variable due to ball and liner wear and the Periodic addition of balls. The problem then breaks down into two parts: 1) obtaining enough mass

Jan 1, 1977

By C. R. Tatman

Failure to achieve the planned production rate at new mines and mine expansions has a major adverse impact on a project's economics. A study of 60 steeply dipping tabular deposit-type underground mines showed that 35% of the mines did not achieve their production goals. An empirical formula based on the geometry of the deposit and expressed as a rate factor is applicable to select a production rate for this type of deposit. The formula is: Annual production rate = Rate factor x Rate multiplier. The appropriate rate multiplier for each category of deposit thickness and the associated risk can be used to select an appropriate production rate.

Jan 1, 2002

By Marc E. Orman, Mark E. Smith

Regulatory and environmental concerns continue to broaden the range of mine wastes that must be capped with low permeabili­ty liner systems at closure. The mining industry has been a leader in developing liner technology for leach pads and tailings impoundments, but it is a newcomer to cap designs. Closure of waste piles, leach heaps and tailings require special design con­siderations. Factos such as differential settlement, surface ero­sion, veneer stability, vegetation and burrowing animals must be considered.

Jan 1, 1996

By Raimo J. Leimala, Heimo U. A. Saarinen, Matti Seilo

Processes for hydrometallurgical treatment of crude nickel hydroxy carbonate and impure high grade ferronickel to produce cathode nickel and cobalt hydroxide are described. Outokumpu nickel refining process originally developed for nickel matte has been adapted to the mentioned new raw materials. Both new process routes contain the Outokumpu cobalt removal and nickel electrowinning stages but leaching and solution purification stages differ from the industrial process in operation in Harjavalta.

Jan 1, 1979

By W. R. Reed

This study evaluates the roof-bolter canopy air curtain (CAC) system in an exhausting face ventilation system, demonstrating its effectiveness and illustrating the CAC protection zones. Two scenarios are considered: (1) a roof-bolting machine in the center of the entry for installation of the fifth row of bolts from the face, and (2) a roof-bolting machine positioned close to the face for the installation of the last row of bolts. In both scenarios, the bolting machine is placed in an environment that contains 6.0 mg/m3 of respirable dust and is ventilated by an exhaust curtain with 9,000 cfm (4.25 m3/s) of air to simulate the roof-bolter machine operating downstream of a continuous mining machine. Two operation positions are simulated at the same bolting location: dual drill heads in the inward position for two inside bolts and in the outward position for two outside bolts. The influence of the CAC on airflows and dust dispersion is evaluated with the CAC operating at 250 cfm (0.12 m3/s) with dust reductions ranging from 39.5 to 82.8 percent. When the roof bolter is operated close to the face, increasing CAC airflow was required for adequate protection, as the dust reductions can be as low as 39.5 percent. Increasing CAC airflows to 350 cfm (0.17 m3/s) and 450 cfm (0.21 m3/s) were evaluated, demonstrating that dust reductions increased to 59.7 percent (350 cfm) and 72.0 percent (450 cfm). These results revealed improvement in respirable dust reductions compared to the CAC airflows at 250 cfm when the bolter is operated close to the face, showing the CAC can provide protection against high respirable dust concentrations for roof-bolter operators.

By Peter J. Alderson, M. P. Evans

The application of conveyor belting in bulk materials handling is crucial to the success of most mining enterprise selection of the cover rubber grade and thickness has a direct bearing on the cost per ton conveyed. Current and future demands placed on belt covers by higher speeds, increase capacities and greater lifts means today's cover will not be adequate for tomorrow's applications. The simple selections of RMA Grade I and I1 and MSHA above and below ground need to be expanded upon. This paper describes various new approaches for selecting conveyor belt covers for specific applications. It also presents examples of material and compounding techniques available to the belt manufacturer to improve the performance of their covers.

Jan 1, 1996

By Francesco Chiappalone, Marco Invernizzi, Franco Gasperin, Brad Yoder, Fabrizio Fara

The Ship Canal Water Quality Project includes a 2.7-mile, 18-ft and 10-in internal-diameter tunnel that extends from Ballard to Wallingford, two popular neighborhoods in the West of Seattle. During the TBM Main drive a 12' × 12' boulder with variable strength up to 47,000psi was encountered resulting on signicant damages to the Cutter ead. The rst part of this paper describes the design and construction of an underground recess in front of the TBM to create a safe working environment for the TBM repairs. The second part describes an injection campaign performed to seal an unforeseen water inow appeared during the excavation of the 11 Avenue Diversion Structure.

Jun 23, 2024

By K. Hansson, A. Jedborn

Background In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore, from a system's point of view, slurry transporta¬tion would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alterna¬tive examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasi¬biliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. l . The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then tempo¬rarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situ¬ations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe¬line, and personnel are needed. In a first phase (Fig. 2), a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1989

By Ian Corrans, Gary Johnson, John Angove

The Activox™ process is a new method for the pre­treatment of refractory gold ore concentrates prior to cyanidation. The process has been developed in Western Australia, and relies on a combination of fine grinding and pressure oxidation to achieve the destruction of sulphide minerals and the liberation of encapsulated gold. The process operates under much more moderate conditions than those that would typically be used for conventional pressure oxidation of similar concentrates. This results in significantly lower capital and operating costs being achievable.

Jan 1, 1993

By E. M. R. Santos

Acid mine drainage (AMD) is a serious environmental problem in mining activities. Prediction is the first step for evaluate the potential of acid generation and there are some possible methods to estimate it. They are described as static tests, which examine the balance between acid producing and acid consuming materials, and kinetic tests, which attempt to predict drainage quality over time. Static methods are very simple and give a qualitative classification of the ores based on their potential to generate acid drainage. The characterization of ores to determine their acid producing potential is vital for the prevention of long term environmental liabilities. Laboratory tests were conducted under different methodologies to predict the acid producing potential of sulphide overburden, ore materials and tailings from a deposit of copper ore.

Jan 1, 1999

By B. M. Davis

The problem of assessing uncertainty in grade estimates has been a topic of research and debate since kriging was first introduced. Use of the kriging variance was originally touted as all that was necessary in order to produce confidence limits for grade estimates. Since then the prudence of using the kriging variance for confidence interval estimation has been questioned, and techniques such as the jackknife and conditional simulation have been proposed to develop "true" interval estimates. This paper compares the differences between interval estimates produced by the three methods and makes conclusions about the performance of confidence intervals produced by the different methods.

Jan 1, 1997

By T. W. West

A description of a coal washing spiral is presented, together with a discussion of the parameters which define the envelope within which the spiral can be successfully applied. The correct process engineering, installation, operating and maintenance parameters are addressed as well as actual results for new and retrofit installations on USA coal capital and operating costs of spiral circuits are compared to those of traditional fine coal circuitry.

Jan 1, 1988