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By R. A. Jankowski

The US Bureau of Mines (USBM) conducted a series of laboratory tests to investigate the effectiveness of using a flooded-bed scrubber with exhaust ventilation in deep-cut faces. An experimental test program to determine the impact on respirable dust levels resulting from changes in face airflow, curtain-setback distance, operator positioning and operating parameters of the external spray system on the miner was completed. Gravimetric sampling was conducted in the immediate return and at three sampling locations on the off-curtain side of the entry. Statistically significant differences in dust levels, on the order of 0.5 to 1.2 mg/m3, were observed between specific sampling locations and changes in several test parameters. Dust sampling results indicated that a flooded-bed scrubber and exhaust ventilation can be a viable dust-control system for extracting deep cuts up to 12.2 m (40 ft) in length. The impact on methane was not evaluated.

Jan 1, 1998

By Paul D. Dupree

Introduction Trapper mine is a dip-line strip coal mine, located about 10 km (6 miles) southwest of Craig, CO. The mine provides the Craig Generating Station, operated by Colorado-Ute, with about 2 Mt (2.3 million st) of coal annually. The Trapper mine is situated on a hillside that climbs in elevation from 2 to 2.5 km (6500 to 7400 ft) at average grade of 15%. The coal seams dip at about 16% down-grade and they outcrop near the top of the hillside. They exceed maximum stripping depth in 1 to 1.7 km (3500 to 5500 ft) of pit length. Stripping in the three existing pits is performed with three Page 752 23 m3 (30 cu yd) walking draglines. Two of the pits are currently being cast blasted. Until the active cut, the Derringer pit was cast blasted only where the lower mined seam exceeded stripping capabilities and the pit was a single seam operation. Now, a 9-m (30-ft) interburden bench has been incorporated into the stripping plan. This bench protects the first seam of coal from shear when the overburden is cast blasted. The cast blasted section of the 1.6 km (53,000 ft) long pit has been increased from the 457 m (1500 ft) casted previ¬ously to 1 km (3500 ft). High speed films show the coal losses due to shear of the blast are minimal. The pit width in Derringer pit is 61 m (200 ft). This is necessary for efficient spoil side stripping of a relatively thin interburden in the two-seam section of the pit. The Enfield pit is entirely cast blasted. The single seam pit is 46 m (150 ft) wide, 1.1 km (3600 ft) long, and varies from 15 and 46 m (50 and 150 ft) in overburden depth. The overburden strata at the mine consists of medium-hard layers of sandstone, shale, and

Jan 1, 1987

By Franz Deix, Bernd Braun

A sensitive portion of Section U 6/3 of the Vienna, Austria subway system required construction of two single-track tunnels under a telecommunications facility, where strict ground settlement limitations had been set. Local geology, hydrology, and site constraints led to the decision to use the New Austrian Tunneling Method (NATM) supplemented by compressed air and partial dewatering. In addition, ground freezing was used, as the tunnel would pass through alluvial layers beneath the facility, with a cover of as little as 1.5 m. Laboratory testing and two in situ test fields were used to collect frozen soil data so that the freezing scheme ensured minimal heaving magnitude.

Jan 1, 1987

By Sidney M. Cohen

This paper describes an economical method of upgrading a clay deposit using conventional separation processes into a suitable cement plant feed. The natural size difference of clay and silica in the clay deposit is utilized in the dry air classification device. A case study in which the raw material from a clay deposit is converted into a portland cement plant feed is presented.

Jan 1, 1983

A linear programming model with integer sequencing variables is developed to schedule limestone production for a quarry in Central Japan. The model is used to optimize raw material production from the quarry supplying limestone to cement, quick lime and aggregate plants. Typically, the mining in the quarry takes place in four to six different areas each area having 3 or 4 mining increments. Each mining increment contains different quality of limestone. The limestone production from the quarry is transported to the plants byway of 3 ore passes. Each plant requires a given quantity of limestone per period meeting certain quality constraints. This paper will describe development of the scheduling optimization model and how it is used to schedule production by period within the quarry to meet the capacity and blending requirements of the plants.

Jan 1, 2003

By W. J. Wuest, M. J. DeMarco

The US Bureau of Mines (USBM) recently developed a rock-mass classification system for evaluating the structural competency of bedded roof strata. The system is called the coal-mine roof rating (CMRR), When compiled with the companion Lotus 1-2-3 spreadsheet, CMRR data are a useful tool with applications to longwall gate-road pillar design, safety-hazard mapping and artificial roof-support analysis. This paper gives a review of some applications under actual conditions found in current US mining operations.

Jan 1, 1997

By Thomas D. Henderson, Wayne R. Hopkins, Arthur Lynch

Brought on stream in 1975, the 10,000-tpd copper oxide processing plant of Anamax Mining Co. at Twin Buttes, Ariz., is now at design capacity. Based on acid agitation leaching, solvent ion exchange, and electrowinning, the Anamax plant bridges an ore to metal gap by producing mine site copper cathodes from lime-bearing oxidized feed. Until recently, such material was considered submarginal because copper recovery was uneconomic using conventional methods. The new Twin Buttes oxide plant, the largest of its kind in North America, is one stage in the expansion of Anamax, a joint venture owned equally by The Anaconda Co. and Amax Inc. Other phases include an accelerated stripping program in the pit, an expansion of the sulfide concentrator from 32,000 to 40,000 tpd, and construction of a 57 MW diesel generating station for the oxide plant. Once the basic oxide flowsheet had been selected, Anamax awarded a contract to a joint venture of Arthur G. McKee and Co.' (Western Knapp Engineering Div.) and Davy Powergas' (a member of the Davy International group). The contract covered process design, engi¬neering, procurement, and construction. The property, identified by Banner Mining Co. in the 1950s and developed by Anaconda, was put into production in 1965 as an open pit mine equipped with a 32,000-tpd sulfide flotation concentrator. The Twin Buttes mine has since become one of the largest earth¬moving operations in the world, excavating some 10 tons of waste per ton of ore. At the original sulfide mill feed rate, the mining fleet moved 375,000 tpd of material (225,000 tpd alluvium and 150,000 tpd of rock). The most significant copper mineral in the ore body is chalcopyrite, which is scattered in pockets of weak mineralization. Cap oxidation has progressed to depths of several hundred feet, principally as chryso¬colla. Present estimates place minable reserves at 347 million tons of sulfide ore grading 0.6% copper and 55 million tons of oxide ore grading 0.9% recoverable copper from 1.25% total copper. The nonre¬coverable fraction is composed of native copper and insoluble copper oxides and sulfides. The oxide ore has been partially stripped, and 25 million tons having a copper content greater than 0.6% have been stockpiled. The stripping of oxide ore to expose sulfides will continue for several years. The copper oxide studies launched by Banner Mining in the 1950s were continued by Anaconda. Process routes investigated included: acid leaching; cyanide leaching; segregation roasting; chlorination, oxidation, and reduction; flotation in combination with leaching; cal¬cite flotation plus leaching; copper oxide flotation; and sulfidization flotation. By late 1968 test work favored a low temperature roast followed by an ammonia-ammonium carbonate leach of the ore. During this period, acid leaching had been discounted because of excessive con¬sumption of expensive acid by the high limestone ore matrix. Restric¬tive SO2 emission controls in 1971 suggested an increase in acid production and availability at a lower cost. Research then switched to acid leaching, resulting in development of the present flowsheet. Mechanically agitated leaching in a cascaded series of tanks was selected in preference to vat leaching after piloting a stirred-tank system. Preliminary testing of vat leaching indicated that inter-vat solution clarification problems could be caused by gas evolution from leaching of high limestone ore. Lack of research time for resolution of this problem, plus success of the pilot agitation system, dictated the final flowsheet selection. In 1973 Anamax decided to proceed with agitation acid leaching to produce liquor for a solvent ion exchange (SIX) and electrowinning (EW) installation. At that time, such systems were proving their eco¬nomic viability, and they allowed direct recovery of high quality cath odes while bypassing the smelting and refining of lower grade cement copper precipitates. The Anamax oxide plant processes a grade of about 1% recovera¬ble copper from stockpiled sources and new mine production. Unit operations include size reduction, agitation leaching, countercurrent decantation washing of residue, pH adjustment of solution, clarifica¬tion, solvent ion exchange, and electrowinning of copper, with a mini¬mum of environmental impact. The feed to the primary crushers is composed of: 1) -10-in. stockpiled oxide ore, which has passed through the mine crusher and has been grade-segregated to allow blending for a consistent feed material. 2) -4-ft ROM oxide ore stockpiled in the early mining operation. 3) -10-in. ore currently being mined. Ore is delivered in 100-ton mine trucks to a dump pocket, where an apron feeder transfers the ore to an inclined 6-in. fixed grizzly. The grizzly oversize passes through a 48 x 60-in. primary jaw crusher, where it is reduced to -6 in. The fixed grizzly undersize and the crusher product are conveyed to a 4-in. vibrating grizzly. Oversize from this grizzly reports to a 7-ft Standard cone crusher, where it is reduced to -1 in. The crusher product and the vibrating grizzly undersize are conveyed to two 8 x 20-ft double-deck vibrating screens that operate in closed circuit with two 7-ft short head cone crushers. Screen undersize (-1/2 in.) is delivered to a covered conical fine ore stockpile with a designed live capacity of 15,000 st. Beneath the stockpile are ten Pioche belt feeders feeding two conveyor belts that supply the milling circuits. Each group of five feeders has two variable speed and three fixed speed drives. Grinding to 95% -48 mesh is done in two parallel lines, each consisting of an 111/1 x 18S -ft rod mill and a 12' x 30-ft ball mill in open circuit. Water is added to maintain a pulp density of 72% solids in the mills. Further water additions are made to the ball mill discharge sumps to create a 60% solids slurry for transfer to the leaching reactors. The nominal grinding rate is 440 tph solids. A splitter box on the product from the mills separates the incoming slurry into two streams, one for the leaching reactors and the other for the pH adjustment reactors. This split is determined by the lime content of the feed material and the pH of solution entering the pH adjustment reactors. Five 30-ft diam by 31-ft high mechanically agitated, rubber-lined leach tanks are arranged in cascade, with gravity transfer from one to another by enclosed launders. Concentrated sulfuric acid (93.2%) from storage is added to the pulp principally at the first leach tank, and the reacted slurry leaving the cascade after a 5-hr residence time at 50% solids is pumped to countercurrent decantation (CCD) wash thickeners.' At this point, most of the soluble copper has been leached from the ore. Up to 250 lb of sulfuric acid is consumed per ton of ore, the acid being supplied in tank cars from a nearby smelter. Leached slurry at 50% solids is washed countercurrently with return raffinate from the SIX plant in a series of four Doff-Oliver center caisson, 400-ft diam thickeners. Solids advance from thickeners No. I through No. 4 to the tailings dam. SIX raffinate advances from thickener No. 4 to thickener No. l, where a pregnant solution overflows to pH adjustment. To optimize solvent extraction, the pH of the pregnant solution from CCD thickener No. I is adjusted from 1.5 to 2.5 by adding unreacted ground slurry (split from the grinding circuit) to the preg¬nant overflow of No. I CCD thickener. The resulting slurry is retained for 45 min in three mechanically agitated leach tanks. CO, generated during leach and pH adjustment is ducted to a demister, where entrained acid droplets are removed. The cleaned CO, and the No. 4 CCD underflow are the only effluents from the plant. The underflow of No. 4 thickener mixes with tailings from the sulfide concentrator and is neutralized. Slurry exits the pH adjustment vessels at about 10% solids and

Jan 1, 1985

By William J. Gregg

Mechanical testing results on oriented Devonian gas shale core throughout the Appalachian basin show a preferred fracture direction perpendicular to bedding. The preferred direction is generally parallel to the Allegheny Front, but is only coincidentally parallel with in-situ stresses in some areas of the basin. The strength of the preferred orientation of induced fractures decreases with distance from the Front. It is suggested that strain related to Paleozoic deformation is responsible for this mechanical fabric and thus the direction of induced fractures in laboratory tests may not be of use as indicators of in-situ stress orientations as previously suggested.

Jan 1, 1986

By James A. Brierley, Rong Yu Wan

Bacterial pretreatment was evaluated as a process for enhancing gold recovery from a refractory ore from Newmont Cold's Cold Quarry deposit. This ore is refractory due to sulfide locking of the gold and preg-robbing carbon. The microbe Thiobacillus ferrooxidans effectively oxidized the sulfide component of the ore, but the preg-robbing carbon continued to be active. Thiourea was examined as an alternative to cyanide for extraction of the gold from the biooxidized ore. The thiourea, coupled with CIL, was effective for gold extraction resulting in up to 89% gold recovery for the sulfidic-carbonaceous refractory ore. The thiourea gold complex appears not to be bound by the preg- robbing carbon of the ore when in the presence of recovery carbon (CIL)

Jan 1, 1990

By Robert B. Fulton, Gill Montgomery

Fluorspar is the commercial name for fluorite, a mineral that is calcium fluoride, CaF2. The name, derived from the Latin word fluere (to flow), refers to its low melting point and its early use in metallurgy as a flux. It is the principal industrial source of the element fluorine. Two other minerals, cryolite and fluorapatite, have significant fluorine content. Cryolite, sodium aluminum fluoride, Na3AlF,, is a rare mineral that has been found in commercial quantities only in Greenland. The natural material has been supplanted by synthetic cryolite for its principal industrial use in the manufacture of aluminum. Fluorapatite, Ca5F(PO3)2, is a source of phosphate for fertilizer manufacture, containing a small percentage of fluorine. Commercially mined deposits of apatite have varying amounts of fluorine, chlorine, hydroxyl, and carbonate. HISTORY Fluorspar was used by the early Greeks and Romans for ornamental purposes as vases, drinking cups, and table tops. Various peoples, including the Chinese and the American Indians, carved ornaments and figurines from large crystals. Its usefulness as a flux was known to Agricola in 16th century Europe. Fluorspar mining began in England about 1775 and at various places in the United States between 1820 and 1840. Production grew substantially following the development of basic open hearth steelmaking, wherein it is used as a flux. Use was stimulated by growth of the steel, aluminum, chemical, and ceramic industries, particularly during World Wars I and 11. Fluorocarbons entered the picture in 1931. The use of anhydrous hydrogen fluoride (HF) as a catalyst in the manufacture of alkylate for high octane fuel began in 1942. Differential flotation for separating fluorspar from galena, sphalerite, and common gangue minerals in the 1930s and the application of heavy media concentrating methods to the treatment of low grade ores in the 1940s were outstanding technological advances that facilitated increased production. Pelletizing and briquetting of flotation concentrates for use in steel furnaces and the development of flotation schemes for beneficiating ores containing abundant dolomite and barite have been major improvements in the industry. USES OF FLUORITE Fluorspar is used to make hydrogen fluoride (HF), also called hydrofluoric acid, an intermediate for fluorocarbons, aluminum fluoride, and synthetic cryolite. It is used as a flux in the steel and ceramic industries, in iron foundry and ferroalloy practice, and has many minor specialized uses. Hydrogen fluoride is produced by reacting acid grade (97% CaF,) fluorspar with sulfuric acid in a heated kiln or retort to produce HF gas and calcium sulfate. After purification by scrubbing, condensing, and distillation; the HF is marketed as anhydrous HF, a colorless fuming liquid, or it may be absorbed in water to form the aqueous acid, usually 70% HF. Synthetic cryolite, organic and inorganic fluoride chemicals, and elemental fluorine are made from hydrofluoric acid. The acid itself is important in catalysis in the manufacture of alkylate, an ingredient in high-octane fuel for aircraft and automobiles; in steel pickling, enamel stripping, and glass etching and polishing; and in various electroplating operations. The manufacture of one ton of virgin aluminum requires about 12 to 29 kg of fluorine content in synthetic cryolite and aluminum fluoride. This quantity, through improved technology and recovery practices, is being lowered significantly in countries with the most advanced technology (i.e., Australia and Sweden), while others (i.e., Surinam and South Africa), remain at the high end. Elemental fluorine is prepared from anhydrous hydrofluoric acid by electrolysis. Gaseous at room temperature and pressure, fluorine is compressed to a liquid for shipment in cylinders or in tank trucks. Elemental fluorine is used to make uranium hexafluoride, sulfur hexafluoride, and halogen fluorides. Gaseous uranium hexafluoride is used in separating U235 from U233 by the diffusion process. Sulfur hexafluoride is a stable high dielectric gas used in coaxial cables, transformers, and radar wave guides. Halogen fluorides have important applications, mostly as substitutes for elemental fluorine, which is more difficult to handle. Emulsified perfluorochemicals, organic compounds in which all hydrogen atoms have been replaced by fluorine, are undergoing investigation as promising blood substitutes. They transport oxygen and, in conjunction with a simulated blood serum, perform many functions of whole blood. With further development, these organic compounds may ultimately, in emergencies, be useful in saving lives of animals and humans during periods of acute shortages of natural blood. Inorganic fluorides are used as insecticides, preservatives, antiseptics, ceramic additives, and fluxes and in electroplating solutions, antioxidants, and many other products. Boron trifluoride is an important catalyst. Organic fluorides are volume leaders in the fluorine chemical industry. Fluorinated chlorocarbons and fluorocarbons are prepared by the interaction of anhydrous HF with chloroform, perchlorethylene and carbon tetrachloride, and are characterized by low toxicity and notable chemical stability. They perform outstandingly as refrigerants, aerosol propellants, solvents, and cleaning agents and as intermediates for polymers such as fluorocarbon resins and elastomers. Fluorocarbon resins are inert compounds that have unusually low coefficients of friction and have found a number of applications as lubricants for parts that cannot be oiled, e.g., bearings for window raising equipment located inside of automobile doors, in small electronic equipment, for the manufacture of chem-

Jan 1, 1994

By Peter M. Turcic

A Mine Safety and Health Administration (MSHA), National Institute for Occupational Safety and Health (NIOSH) and Bureau of Mines (BOM) interagency task group was established to assess the health and safety implications of the use of diesel-powered equipment in underground coal mines. The objective of the group was to conduct an assessment to identify areas of concern, to evaluate the adequacy of existing methods for addressing the identified concerns, and to recommend agency approaches for addressing inadequacies. The interagency group concluded its work in April 1986 and presented a report to the Assistant Secretary for Mine Safety and Health. The report included the conclusions reached by a group and recommendations for regulatory action by MSHA. This paper will present the methods and approaches used by the interagency task group and will discuss the findings, conclusions and recommendations of the group. The paper will attempt to meld together the technical and policy aspects of the issues involved with diesel equipment in coal mining.

Jan 1, 1987

By E. G. Joe

Introduction Uranium mining has been actively carried on in Canada for over twenty-five years mainly in Ontario at Elliot Lake in Bancroft, and in Northern Saskatchewan at the operations of Eldorado Nuclear Ltd. During this period, approximately one hundred million tons of mill tailings have been created and lie mainly on the surface close to the mining and milling installations. If all of the present reserves of uranium ores in Canada are mined, a further eight hundred million tons of mine tailings could be created in Ontario, Saskatchewan, the Northwest Territories, Labrador and British Columbia. The present control of effluents from existing tailings is managed by the operating mines under the regulations of both federal and provincial control agencies. It is the long-term effect of these tailings on the environment that has not been clearly defined or measured. Several years ago, the Canada Centre for Mineral and Energy Technology of the Department of Energy, Mines and Resources, embarked upon a program of research directed at the long-term effects of uranium mine tailings. Much of the research to date has been on tailings in the Elliot Lake, Ontario area where some tailings piles have existed for over twenty years. The program consists of two projects, one to address the problem of existing tailings and the other to examine alternatives in the management of current and future tailings. TAILINGS CHARACTERISTICS Uranium tailings are generally the result of sulphuric acid leaching in which the uranium and alkaline minerals such as carbonates and chlorites are leached out. If the ore contains sulphides, these minerals potentially present an additional hazard of acid generation over time and could increase the mobility of heavy metals and some radionuclides such as thorium-230. Except for the extracted uranium, most of the radionuclides from the ore remain in the tailings pile. Uranium ores in Canada span a wide variety of mineralization and associated gangue composition. Uranium minerals include pitchblende, uraninite, uranothorite, brannerite and the phosphate uranium minerals autunite and saleeite. Gangue minerals vary from acid consuming carbonates, and chlorites to acid producing sulphides and inert silicates. With such a wide variety in ore composition, the characteristics of the resultant tailings in Canada are in many cases site specific and the solutions to the long-term effects of these tailings may also be site specific.

Jan 1, 1980

By Larry Whittle

Gold milling activities generate waste streams containing dissolved metals and cyanide, among other contaminants, that require some form of treatment prior to discharging to the environment. To date, this treatment has involved the destruction of cyanide through natural degradation, chemical oxidation, or a combination of both followed by the precipitation of metal hydroxides. Although these processes have been proven to be capable of achieving. regulatory levels for metals and cyanide, they often generate significant levels of cyanate and, at times, ammonia, that may be toxic to rainbow trout and thus undesirable for regulatory compliance. Tallon Metal Technologies Inc. (Tallon) has developed Vitrokele™, a chelating resin that has the ability to capture metals (including gold), both total and weak acid dissociable cyanide, as well as thiocyanate from both slurries and solutions. The loaded resin is eluted and returned to the recovery circuit while metals and cyanide are recovered from the eluent stream. This process was demonstrated at pilot scale at B.P. Resources' Hope Brook Gold Mine in Newfoundland, and more recently at Canamax's Bell Creek Mine near Timmins, Ontario. The results from these pilot programs showed metal and cyanide recoveries of 99% or more with the concentration of cyanide and metals in the effluent reduced to levels with minimal or no toxicity to rainbow trout. Furthermore, neither cyanate nor ammonia were generated from the process. The loaded resin was effectively stripped and re-introduced to the recovery circuit; cyanide was captured in the eluent stream and made available for recycling to the mill leaching circuit; and metals were captured for re-use, sale or disposal. This paper summarizes the results obtained from the pilot tests at the Bell Creek and Hope Brook Gold mine sites.

Jan 1, 1992

By G. A. Campbell

Mining companies have been very interested in the presence of a number of small but high-grade ore bodies of base metals in northern Wisconsin and the west/central section of the Upper Peninsula (UP) of Michigan. The region is rural with outdoor recreation/tourism and lumbering being the key industries. The economy of the region is based on the use of its natural resources, and mining is not a new activity to the region, especially the UP. The western UP was historically a key mining district for copper and iron, and today the central UP has a significant iron-mining district. However, all efforts to develop the base metal ore bodies have been met with a storm of controversy involving the local community, regional activists, Native Americans, and the state government, and only one mine has gone into operation. There is concern by some stakeholders about the sustainability of mining in communities that depend on the environment for tourism, recreation, and other commercial activities. There have been four attempts to get the necessary approvals to mine in this region: the successful Flambeau Mine effort (Wisconsin), the failed Crandon project (Wisconsin) and Lynne project (Wisconsin) efforts, and the continuing effort to get the Eagle Project (UP of Michigan) approved. The cases are presented to illustrate the project situation, the history of the permitting effort, and the opposing parties and their concerns. The case of the Eagle Project is looked at in detail by using the method of discourse communities and discourse analysis. Results of the analysis provide insight into the communication efforts to address community concerns about sustainability and what more can be done to improve this dialogue.

Jan 1, 2008

By Hans von Michaelis

It is appropriate that this opening session of HAZPAC '91 and Randol Gold Forum CAIRNS '91 should address the question of what the two industries can learn from each other. The panel is made up of members with experience in both industry. Several are experts on managing both hazardous waste and the mine/mineral processing environment. There is a marked difference in the regulatory needs with respect to treating and disposing of mining and mineral processing wastes, and "hazardous" or industrial waste materials. The topic of this paper is on the things that the two industries have in common: Many of the processes and technologies that are being used or are currently being developed for treatment of hazardous waste are the same or similar to the processes that have been used for decades or k,nger in the mineral processing industry. Hazardous waste treatment and remediation of contaminated solid and/or groundwater is a relatively new industry. Many of the practitioners in this new industry are relatively inexperienced in most of the "new" technologies that are being applied or developed.

Jan 1, 1991

By Michael J. Virnig, Keith E. Weerts

The development of heap leach technology for the treatment of low grade gold and silver ores has led to a boom in gold production in the Western United States in the last 15 years. The heap leaching process consists of stacking the mined ore on a specially prepared leach pad followed by leaching with dilute alkaline cyanide solution. Depending on the nature of the ore, it may require crushing and/or agglomeration prior to stacking on the leach pad. As the dilute alkaline cyanide solution migrates down through the stacked ore, the precious metal values along with a portion of the base metal values present in the ore are dissolved to give a pregnant leach solution which is collected at the base of the stacked ore. The precious metal values are then recovered by zinc precipitation or by adsorption on activated carbon. Once all of the economic metal values have been leached from the ore, the spent heap leach pad, the leached materials, and associated process solutions must be decomissioned. Decommissioning requires that the spent heap leach pad along with the leached materials and process solution be treated to reduce the concentration of the residual cyanide, base metal cyanide complexes, and other anionic metal species to levels where they no longer represent a long tenn liability for the environment and the operator. The levels to which these components must be reduced is dependent upon the applicable water quality standards as determined by the responsible federal, state, and local agencies. As environmental awareness has grown, these standards have become increasingly stringent. A number of treatment options have been developed for the de-toxification of the process residues. These include natural degradation which incorporates bio-degradation, fresh water rinsing, alkaline chlorination, hydrogen peroxide, and sulfur dioxide/air. Each of these options has been reviewed with regard to relative effectiveness, generation of hazardous by­products, reagent handling considerations, relative costs, and time requirements by Denton, Iverson, and Gosling I. As an outgrowth of the program to develop Henkel's guanidine based gold extractants, COGNIS, Inc., the environmental ann of Henkel, has developed the CyanoMet R process for the extraction and concentration of the various cyanide and anionic metal species from the dilute aqueous alkaline solutions generated during the process of detoxifying spent heap leach pads. This paper describes the CyanoMet R process, its application to three different feed solutions, and treatment of the concentrate generated by the process.

Jan 1, 1993

By S. K. Kawatra, K. J. Shoop

An analysis of an iron-ore processing plant showed that its performance was affected by seasonal temperature variations, resulting in reduced grinding efficiency during the winter months. Laboratory and pilot plant studies were conducted to determine the cause of the observed loss of efficiency at lower temperatures. Two types of experiments were performed: drop tests and wet grinding tests. Drop testing was performed to determine the effect of brittleness and frozen pore water on the breakage characteristics of the ore. The wet grinding tests were divided into laboratory tumble tests and pilot-plant tests. The combined results of drop testing and wet grinding were used to make predictions on the performance of plant-scale autogenous grinding. From the drop tests, it was determined that dry iron ore, basalt and sandstone did not change in brittleness over the temperature range studied (-25º to 25ºC). However, when the rocks were water-saturated, the highly porous sandstone became considerably more fracture-resistant upon freezing, due to "grouting" of the pores and preexisting cracks by ice. This effect was not seen for either the iron ore or the basalt, because they were considerably less porous than the sandstone. Tumble tests on both laboratory and pilot scales showed that both the addition of fines to the test charge and a reduction in temperature caused a decrease in the amount of new fines produced. This reduced fines production was the result of increased slurry viscosity when either the fines content increased or the temperature decreased.

Jan 1, 1998

By Etienne Simode

When a fire breaks out in a mine access road, the hot smoke, due to heat transfer with the surrounding strata, progressively cools as it moves forward. Depending upon the resulting temperature distribution, the density of the fluid changes. This change will bring about: Firstly, a variation in the resistance of the roads, a static factor which will affect ventilation flow rates; Secondly, the appearance of extensive aerodynamic driving loads in the vertical and angled roads, a dynamic factor which can have an extensive impact on ventilation. Because of the appearance of combustion and distillation products, the vaporization of water, which wets the timbering and strata, and the expansion of gases, fluid rates can at times be increased tenfold in the area of the fire. These disturbances may cause explosions and ventilation inversions, which fill other working areas with smoke toxic to workers, and which can come up upon rescue teams from behind. Due to the progressive heating of the surrounding rocks, which will slow down heat transfer, these dangerous conditions change over time. At the Lorraine Basin Coal Mines, programs have been set up and employed for the computer simulation of these phenomena over time. They serve to anticipate or to understand such phenomena, the input data being the characteristic parameters of the mine ventilation network, the operating curve of the main fans, and measurements or estimates of some temperatures. Two programs, employed together, have been developed for these studies: One program, called RESO, calculates the distribution of flow rates and pressures throughout the entire network; The other, called FEUMIN, determines smoke temperatures over their entire path.

Jan 1, 1980

By Z. Stanojevic-Simsic

The aim of this paper was to study sludge dewatering. The sludge which was examined in laboratory scale was produced by neutralization of a mixture of metallurgical wastewaters. Experiments were carried out using synthetic wastewaters with chemical composition similar to that of metallurgical wastewaters, resulting from electrolyte regeneration, copper refining and precious metals recovery. During investigations gravitational and vacuum techniques were examined. Parameters such as water volume, height of sludge and time of dewatering were monitored. In this paper the results obtained are presented and discussed. The application of vacuum filtration enables better dewatering and compaction of sludge.

Jan 1, 2006

By Ümit Yasar Onder, Tevfik Prof. Dr. Güyagüler

This paper describes the results of an investigation done to determine thermal properties and heat emission of coal measure rocks in Karadon district of Zonguldak Coal Colliery. In Karadon District, depth of working level is about -600 m where the heat problem arises. In the near future, workings will be at deeper levels and working conditions will be more difficult due to high heat emission. At the beginning of the paper the importance of wall rock heat emission in underground mine have been explained and then the amount of heat emission to the ventilating air has been calculated. It has been found out that the strata is the main source of heat in the mine. Therefore, geothermal gradient of the region has also been determined simply by measuring the virgin rock temperatures in different levels of the mine. All other data directly related with heat emission from different sources has been collected and has been analyzed by using a special computer package. As a conclusion, total heat and moisture content in ventilating air and heat balance has been determined for Karadon Coal Mine. In addition, some recommendations and suggestions have been made for the future works when the mine workings go to the deeper levels.

Jan 1, 1995