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By C. Steve Albrecht

The primary objective of this study is to "evaluate the influences of different equipment systems on prime farmland subsoil compaction due to surface mining in the Midwest." The study is centered on four mine sites each utilizing a different combination of equipment for soil removal and replacement. Subsoil compaction is the primary focus of this research effort. Primarily owing to physical location, it is difficult to alleviate the increased compaction directly below the topsoil in reclaimed soils. Different equipment combinations cause varied effects when soils are replaced during surface mining activities. By collecting field data within a statistical framework, the effects of different equipment combinations on subsoil compaction can be evaluated and conclusions drawn. This research effort evaluates a comparison of natural and replaced soils by moist bulk density data. This comparison centers on changes in bulk density and texture to a depth of 48 inches. For each study site a complete case study is provided for use in data correlation and interpretation. Basis for this research effort centers on four (4) major experiment designs: o Comparison of soil bulk density values for three distinct equipment systems evaluated on the same general soil types (i.e. scrapers, bucket-wheel excavators and end-dump haulback); o Evaluation of the effects of soil replacement in repeated lifts using scrapers; o Presentation of bulk density values representing soil compaction for four (4) equipment systems (i.e. bucket-wheel excavators, draglines, scrapers and end-dump haul back) on four separate mine sites in the midwest; o A prediction analysis model is developed for subsoil bulk density following reclamation based on equipment characteristics and a statistical evaluation of texture. The soil data base of information includes data from the 1982 Bureau of Mines study entitled, "Impact of Surface Mining on Soil Compaction in the Midwest" as well as data from this research project. The final report is synthesized through an analysis of field data, statistical analysis and literature review. The final report focuses on the interpretation of results for practical application by government and industry.

Jan 1, 1984

By Erin N. Legault-Seguin

The Lac des Iles Mine (LDI) is located approximately 100 km northwest of Thunder Bay, ON. The mine has been in operation since 1993. The initial mine production was 2,300 mtpd and increased to 15,000 mtpd following a mine expansion project that was completed in 2001. Effective Oct. 29, 2008, LDI was shut-down under care and maintenance due to low metal prices. Prior to shutting down, LDI was the only primary PGM deposit in operation in Canada and of the few found globally. The complex geology and physical properties of the deposit causes several challenges for the concentrator. This paper will discuss the solutions to the main challenges of attaining nameplate tonnage and achieve satisfactory Pd recovery and concentrate grades.

May 1, 2009

By Pengfu Tan

A thermodynamic model of KIVCET lead smelting process has been developed. The model has the capability to predict the distribution behaviors of Pb, Zn, Fe, S, O, As, and heat balance in KIVCET direct lead smelting furnace. The compositions of lead bullion, matte, slag and gaseous phases, and heat balance are calculated. The effect of the ratio between oxygen and feed on the chemistry of the smelt and heat balance of the coke checker has been simulated and discussed. The thermodynamic model also helps to understand the tapping issue of the lead bullion. In this paper, some applications of this model in the industry have been presented.

Jan 1, 2006

By B. V. Johnson

The Federal Bureau of Mines has conducted research in support of the National Aeronautical and Space Administration's (NASA) Apollo program to acquire fundamental scientific and engineering knowledge of lunar-mineral resources for NASA's Office of Advanced Research and Technology. Equipment and techniques were developed to conduct soil shear tests in a simulated lunar environment. Shear tests performed with this equipment in ultrahigh vacuum and controlled laboratory atmosphere showed that simulated lunar soil increased in strength in the simulated lunar environment.

Jan 1, 1973

By J. A. Botín

In modern mining, the core of mine planning and design is the block model, a simplified representation wherein the ore deposit is divided in a number of virtual rock cubes, each characterized by technical and economics parameters. Mining engineers use this model to design the mine and develop production schedules. Traditionally, sustainability issues are not considered in the mine design, planning and engineering processes but rather are incorporated in the later stages of the feasibility evaluation as necessary expenditures and operating costs. This paper describes the results of a research project aiming to develop a methodology to incorporate sustainability parameters directly into the block model, so that they become an integral part of the design, planning and scheduling processes. It is considered that the incorporation of sustainable parameters into the block model would yield a more responsible mine design and improve mining projects sustainability performance

Feb 27, 2013

By K. S. Liddell, L. A. Smith

"Kell is a hydrometallurgical treatment option for the recovery of platinum group metals (PGMs), gold, silver, and base metals from flotation concentrates to refined products on site. The process is cyanidefree, while eliminating emissions of sulphur dioxide, arsenic trioxide, and other toxic species often emitted by smelters and roasters. Kell has been tested on a range of concentrates, including PGMs (UG2, Merensky, Platreef, Great Dyke, Great Lakes, Lake Superior and others), refractory gold-silver, coppergold, and polymetallic concentrates. High extraction efficiencies (>95%) are achieved for value metals (Pt, Pd, Rh, Au, Ag, Ni, Cu, Co) and secondary metals (Sb, Zn, Pb, and others). Studies comparing Kell with smelter-refining show economic benefits: capex 18–33% of smelting; opex 51–66%; electricity consumption 13–46%; and environmental benefits: greenhouse gas emissions 57–61% of smelting, acidification 5–27%, human toxicity 37–62%, and freshwater ecotoxicity 43–55% of smelting. Synergies for Kell plants alongside smelter-refineries include utilization of recovered acid from sulphur abatement systems and excess capacity in precious metal refineries. Kell is unconstrained by concentrate grade and impurities, allowing the co-location of Kell plants for treatment of non-smeltable concentrates alongside existing smelters.Kell has been demonstrated in a nine-week integrated pilot campaign at 1:1000 scale, supporting a bankable feasibility study for a 110 000 t/a plant treating a UG2-Merensky concentrate at Pilanesberg Platinum Mine (PPM), owned by Sedibelo Platinum Ltd, which has along with South African Industrial Development Corporation (IDC), invested in Kell. Zimbabwean Mining Development Corporation (ZMDC) has signed a Memorandum of Agreement (MOA) with KellTech for a centralized PGM concentrate processing plant."

Jun 1, 2019

By E. Roman, M. Brissette

"For its end of life, Kemess Mine processed its difficult-to-treat hypogene ore from the east pit, which is a lower grade copper and higher pyrite content gold ore. Metallurgical testwork showed that improvement was possible with finer grinding. In July 2010, small grinding media were introduced into the regrind mill resulting to a nearly 50% decrease of the power draw with a finer grind, an increase of 1.1% in the final copper grade, an increase of 1.2% in copper and 2.9% in gold recoveries at the cleaners. The overall recoveries of copper and gold increased by 0.6% and 1.3% respectively.INTRODUCTION Kemess Mine was located at 430 km North West of Prince George in North Central British Colombia. Its production started in May 1998 by Royal Oak Mines and was acquired by Northgate Exploration in 2000. Kemess reached its end of life production with its last of production on March 9th, 2011. The concentrator nominal capacity was 56,000 TPD consisting of a primary gyratory crusher feeding two identical conventional SAB circuits totalizing 35,840 kW available for grinding. Rougher flotation consists of four parallel rows of eight 130 m³ tank cells to produce a concentrate of 2% Cu and 6 gr/t Au which was pumped to a 5 m x 10.8 m regrind ball mill. The regrind installed power was 4,480 kW to generate a final grind of 80% passing 40 microns. This product fed the two stage cleaner flotation circuit, first stage being two parallel rows conventional 17 m³ Denver cells and second stage, two 3.3 m x 11.0 m flotation columns in series, designed and constructed by Kemess, to produce a 22% copper concentrate grade. The cleaner flotation middlings were returned back to the regrind mill (Froehling, Mohns, Roman & Grady, 2005; Froehling, Mohns, Rajwani & Fullam, 2007). The large copper-gold porphyry deposit was composed of 15% of oxidized supergene and leach cap ore and 85% of hypogene ore having three major alteration zones, mainly characterized by their sulphur/sulphide content. Chalcopyrite is the main copper mineral species with minor bornite. Gold was mainly present as free gold or electrum. The first alteration zone consisted of a phyllic alteration characterized by moderate and intense sericite alteration which had the highest mill throughput and twice the mine grade when in contact with the supergene horizon. The second zone consisted of a potassic alteration characterized by more competent groundmass and moderate to intense quartz stockwork which had a slightly lower mill throughput. Finally, a third alteration found beneath and on the east side of the deposit having almost no quartz stockwork and hardened groundmass, sulphide content is two to four times more. This zone had the lower mill throughput with significant drop in the mine average gold and copper grade. Through its mine life, the average life of mine copper head grade was 0.21% Cu and 0.64 gr/t Au. Head grades began to decrease from 2008 to be substantially lower in 2011 as shown in Table 1."

Jan 1, 2012

Project-finance risk assessment is mature and approached systematically. Mining industry due diligence remains fragmented with specialists addressing different aspects of the project such as reserves/JORC, environmental impact, commodity market, even financial analysis/ modelling. Each professional has a natural tendency to believe that the risk factors they have examined are the key ones necessary to be reviewed, as if their review will stand in isolation from others in the way a project materialises. By first laying out the project risk categories and risk matrices, this paper examines the due diligence process with a keen eye to the risk assessment and overlap/interaction. Some cross fertilisation of due diligence techniques from other industry sectors is placed into the mining context. Six case studies are then presented with regards to due diligence features in each projectÆs feasibility review or financial information/ offering memorandum. A power case first shows the differences between the due diligence expectations of banks versus bond financings. Then some Australian and Indonesian mining development cases are analysed to see the due diligence learning points. Some of these were project/due diligence failures. The conclusion focuses on the value of a systematic risk assessment for project evaluation, all from the project-finance perspective.

Jan 1, 2007

By Costa

The fly-in-fly-out (FIFO) mining model has significant implications for the industry as it plays an important role in fulfilling the economic, social and workforce needs of the contemporary mining industry. It has increasingly been adopted in countries such as Canada and Australia, but is the current nature of the FIFO model challenging the place women have achieved in the mining workforce? Research in the mining and oil and gas industries reveals that major sources of stress for fly-in-fly-out mining workers are the times of parting and of reunion, the challenges in defining roles within the family, parenting, and conflict between spouses over the use of the leave time and money. So how are women in the industry coping with the FIFO schedules and consequent intermittent relationship with families and partners? How is the mining industry adapting to recruit and maintain female employees in FIFO operations? This paper describes an exploratory study conducted in Canada, which intended to provide some initial insights that might help us answer the questions above. There are several inferences made from this exploratory study. It is clear that FIFO poses significant challenges for both female and male workers. It is largely seen as rigid, extremely challenging for employees with young families and workers who are single mothers, and it is generally incompatible with starting a family and caring for small children. FIFO creates severe stress on relationships and on workers psychological well-being. It seems to be indeed harder for women, since they are often in lower paying positions and therefore have less financial reward for the FIFO lifestyle. On the other hand, FIFO creates the opportunity for young women, including Aboriginal women, to jump-start their careers, as well as opportunities for mature women to work while pursuing other personal interests during their time off. This paper concludes with recommendations to improve the FIFO system in order to increase the ability of mining companies to hire and retain women

May 1, 2006

By A. Dogan Paktunc

Abstract - The Mechanic Settlement pluton is a small body of mafic and ultramafic rocks of tholeiitic affinity, occurring in the Late Neoproterozoic volcanosedimentary rocks of the Avalon terrane of the Appalachian Orogen. The lenticular-shaped intrusion is layered and consists of ultramafic and mafic lithologies in at least six cyclic units that are relatively thin or short-lived. The lowermost exposed portion of the intrusion is characterized by the cumulate sequences of olivine and olivineclinopyroxene and display relatively flat rare earth element profiles. In the overlying lithologies, plagioclase is the dominant cumulus phase following olivine in the crystallization sequence. Olivine ranges in composition from Fo73 to Fo85 and plagioclase from An62 to An89. Variations in the compositions of orthopyroxene and clinopyroxene are within Wo0-4En52-87Fs11-45 and Wo39-50En23-50Fs3-19. These mineral compositional variations in part reflect cryptic patterns and correspond to normal fractionation trends in cyclic units. Elevated Pt and Pd concentrations of up to 2.6 ppm occur in ten zones within the central cyclic series in association with both the peridotitic and gabbronoritic rock types. The mineralized zones are about several meters thick and appear to have lateral continuities over several hundred meters, suggesting that they are stratiform. Platinum-group elements (PGE) occur as discrete platinum-group minerals (PGM) including vysotskite, merenskyite, michenerite, sperrylite, stillwaterite, and hollingworthite that are associated with magmatic sulfides composed of pentlandite, pyrrhotite, and chalcopyrite. PGM grains range in size from a few micrometers to well above 20 µm. Sulfur isotope ratios and the Se/S ratios of the mineralized rocks are typically higher than the mantle values. Sulfide saturation probably occurred as a result of assimilation of crustal rocks and magma mixing during emplacement of subsequent pulses of magma. Repetition of the PGE mineralization in rather thin cyclic units suggests that the replenishing magma probably had high PGE levels and was saturated with respect to sulfide during emplacement. Ingression of the replenishing magma into the chamber in a turbulent plume and its mixing with the resident magma was probably responsible for the formation of the PGE mineralizations in the Mechanic Settlement pluton.

Jan 1, 2000

By L. J. Barnes

The process for bio-remediation of sulphidic slag (BROSS) operates at 45 degrees C in a neutral (pH 7), aerobic aqueous environment in which sulphide in the slag is converted to sulphate. A two-stage version has been developed that involves quantitative chemical pre-oxidation of total sulphide to thiosulphate, followed by complete biooxidation of thiosulphate to sulphate. Ferric hydroxide, also produced, acts as a flocculant in the bioreactor effluent, concentrating heavy metal precipitates in the thickener as a sludge which can be returned to the smelter. The process has a high sulphate production rate and low acid and nutrient consumptions. Microorganism growth is the rate-limiting stage, requiring a minimum liquid residence time in the bioreactor of about 7 h. Although large quantities of oxygen are needed, sufficient gas mass transfer can be attained in an air-lift reactor. Products are an aqueous stream containing sodium sulphate with only trace levels of toxic heavy metals and a benign, toxic metal-containing sludge

Jun 18, 1905

By J. J. Carlson

Studies were carried out to determine whether filtration rate changes observed in an iron ore concentrator were due to variations in particle shape. It was confirmed that ores mined from different locations differed in filtration rates by as much as 60%. However, when image analysis was used to quantify the shapes of the individual particles in the filtration feed, it was found that the particle shapes showed almost no variation, and the shapes did not correlate with the filtration rate. It was also determined that the filtration rate variations were not caused by changes in particle size distribution. These results suggest that the differences in filtration rate are due to factors other than size and shape, such as surface chemistry.

Jan 1, 2008

By Lon A. Williams

This report, authorized under Contract Number H0220065, Amendment Number Six, documents the results of an investigation of lighting system modifications and designs for working place illumination in high and low coal entries of underground coal mines. The portable and machine mounted system designs, both modified and prototype, represent a continuation of work performed under Contract Number H0111403, Amendment Number One. See the Final Technical Report dated May 26, 1972 for detailed information concerning this prior Contract. All modified and prototype systems met the performance requirements of Bureau of Mines Specification 2G, as proved by test. In addition, the systems met the requirement for illumination, as specified in the Contract and in the proposed standards which were generated in response to Public Law 91-173.

Jan 1, 1976

By B. J. Vorster

Introduction The South African Iron and Steel Co. Ltd., in Sishen, operates a large iron ore mine in the Cape Province of the Republic of South Africa. The mine is about 600 km (370 miles) west of Johannesburg. Its annual capacity is 100 Mt (110 million st) of ore and waste. Present production is 60 Mt/a (66 million stpy). The mine operates continuously six days a week. Primary equipment includes 55 154 t (170 st) Wabco and Unit Rig trucks and 14 P&H shovels. During 1979, a feasibility study into the application of a trolley assist system was launched. This was due to the escalation in diesel fuel costs and the appeal by the South African government to reduce the use of fossil fuels. Tests and economic evaluations proved the system technically feasible. The estimated discounted cash flow return exceeded 50%. Diesel fuel consumption would be reduced by 120 ML (31.7 million gal) in the next 10 years. The construction of 7.7 km (4.8 miles) of line and the conversion of 66 154 t (170 st) trucks were completed in March 1982. System description The conventional 154 t (170 st) diesel electric rear-dump haul-truck is equipped with two direct current wheelmotors in the rear axle. They drive the rear wheels through planetary gearboxes. A 1.2-MV (1600-hp) diesel engine and an alternator is used to power the wheelmotors. The trolley assist system consists of an overhead power line, a current collector mounted on the truck, and add-on control equipment where external electric power is routed by way of the current collector to the motorized wheels. The diesel engine is used to propel the truck on level sections. On inclines, the wheel-motors of the truck are coupled by way of a current collector to the externally generated electrical power supply for propulsion. The diesel engine, having lower output than the wheelmotors, limits the truck performance. Previously, maximum output was required from an engine on inclines. By supplying the wheelmotors from a virtually unlimited outside electric energy source, the truck performance is enhanced to the extent that speed is increased by 46% and fuel consumption reduced by 75%. System design The unique operational requirements of this mine rendered the already proven trolley pole system useless. Minimum loss in truck mobility and flexibility was required. Overhead lines and substations were required to be relatively light and portable. This enables fast repositioning of a line with no loss in material. These unique specifications set the base for the design of the Sishen system. Steel I-beams, 203 x 203 mm x 52 kg/m (8 x 8 in. x 35 lb per ft) with lengths varying from 8.3 to 11.3 m (27 to 37 ft), mounted on base plates about 4 x 2.6 m (13 x 8.6 ft) were used as masts. The masts were erected at about 50 m (165 ft) and the base plates covered with about 12 t (13 st) of iron ore to keep them absolutely stable. Results from the pilot installations indicated that using single cantilever cross arms for insulator and overhead conductor suspension was not practical. The tension in the overhead varied too much with temperature changes, causing too much sag in the lines. There had to be a positive and negative conductor. It was decided to suspend each polarity on an independent cantilever and keep the conductors under constant tension by weights acting through a two-to-one pulley tensioning arrangement. For each overhead conductor, a standard 160-mm (6.5-sq in.) grooved copper railway conductor was used. To obtain the necessary current carrying capacity, two of these conductors are used in parallel per pole. Insulators are rated at 4 kV and are of the composite type. An 11-kV three-phase line, supported by the pantograph line masts on the opposite side of the dc line, feeds the rectifier stations. The rectifier stations consist of an 11-kV to 860-V, three-phase 1.46 MVA transformer and a three-phase diode bridge rectifier. The output voltage is 1200 V do earthed through a current limiting resistance network so that the line po-

Jan 4, 1986

By Horace R. Graham

CHEMICAL nitrogen and the "nitrates" of commercial significance are derived mainly from three basic sources: (1) the natural deposits in the form of nitrate-bearing earth and clay, which, being largely water-soluble, can exist only in the most arid portions of the earth; (2) coal, which yields not only nitrogen but also the additional by-product hydrogen required for the synthesis of ammonia and its compounds; (3) the atmospheric air, which contains 75.5 pct by weight of pure nitrogen gas. NATURAL NITRATE DEPOSITS Occurrence Although natural nitrate deposits consisting principally of the salts of sodium and potassium have been reported in Egypt, South Africa, Mexico, Argentina, Colombia, Peru and the USA, the only known natural nitrate resources of commercial importance are the vast deposits found in the Atacama, Tarapaca and Antofagasta regions of northern Chile, lying in a narrow strip 10 to 50 miles wide and roughly 450 miles long, between latitudes [19O] and [26OS]. In addition to yielding nitrates of sodium and potassium, these deposits are the source of the major portion of the world's supply of iodine. The nitrate-bearing ores known in Chile as caliche are variable in composition and may contain from 5 to 30 pct each of nitrate, chloride and sulphate. While the deposits are present chiefly in the form of the sodium salt, potassium, magnesium and calcium also occur in percent- ages from 0.1 to 5; usually as double salts with sulphate, which are but slightly soluble in nitrate leach solutions. These double salts are bloedite(Na2S04.MgS04.4H,0), glauberite (Na2S04.CaS04), polyhalite (K2SO4.- MgS04.2CaS04.2H20). Some deposits contain up to30 pct of their ni-

Jan 1, 1949

Recovery of Pentlandite and Chalcopyrite from Tailings by Bi-Cyclone Powerful Flotation Column

Sep 13, 2010

By H. H. Richardson, E. H. Dix

Of all the alloying elements used in commercial aluminum alloys, copper stands out as by far the most important, and it is perhaps for this reason that the constitution of the aluminum-copper system has received the attention of a number of experimenters covering a period of approximately 20 years. It would be expected that during this time the equilibrium relations would have been thoroughly and exactly worked out. In a general way this is true, but there is a surprising lack of agreement in the details of the high aluminum end of the diagram, particularly in regard to the solid solubility of copper in aluminum, which has been given variously as 2 per cent. to 5 per cent. at the eutectic temperature. The importance of this is much greater than might at first appear to one unfamiliar with modern commercial aluminum alloys. This is true because, with one exception, the susceptibility of the strong alloys to heat treatment depends, in part at least, upon the solid solubility relations of copper in aluminum. In 1919, Merica, Waltenberg, and Freeman1 published the first detailed information on these solubility relations. They found that the solubility of copper decreases with falling temperature from about 4 per cent. at 525" C. to about 1 per cent. at 300" C. and is apparently still diminishing at lower temperatures. In 1921 there was published a brief account of the results obtained in the National Physical Laboratory, Teddington, England.2 This laboratory, however, reported that 5 per cent. of copper is soluble in aluminum at 540" C., and that this amount decreases to about 3 per cent. at 20" C. In the same year Ohtani and Hemmi3 reported a solubility of 4.8 per cent. copper at 520" C., 2.6 per cent. at 460" C. and 1.5 per cent. at 420" C.

Jan 1, 1926

By C. O. Aksoy, H. Kose, K. Heasley, E. Yalçin

"In Turkey, the growing demand for lignite has resulted in increased lignite production and an emphasis on improving underground recovery. This has resulted in a gradual increase in the application of the retreat longwall mining method and also in the extraction of the safety pillars protecting permanent underground structures (drifts, shafts) in mines prior to their abandonment.Extracting these safety pillars presents a special hazard to mine employees and the stability of the mine. Often subjected to high stresses, the behaviour of these pillars is not always predictable. Special care must be taken to ensure that infrastructure does not collapse prematurely, and that working conditions remain safe. Otherwise the extraction must be halted, and the remaining coal lost.This investigation reports on a study of the stability of safety pillars during the final extraction phase of the Hustas mine in Soma, Turkey.” The model construction was based on mine plans, and material properties were derived from physical properties testing. Observed subsidence effects from other mining events were used to calibrate the overburden lamination thickness.The model results compared well with surface physical disturbance occurring as a result of mining, as well as with conventional longwall stress redistribution theory and underground observations. The study has shown that properly calibrated models can be useful in assessing mine safety during remnant pillar extraction. The LAMODEL results demonstrate the convenience of this easy to use software."

Jan 1, 2004

By K. H. Heino

Two 200 T/da Lurgi Turbulent Layer fluid-bed roasters have been in operation at Canadian Electrolytic Zinc Limited since 1966. Zinc concentrate containing 52-54% Zn, 9-11% Fe,31-33% S is treated to produce calcine containing 60-62% Zn,10-12% Fe, .l-.3% S as sulphide and .8-1.2% S as sulphate. Each roaster train comprises a waste heat boiler, two cyclones in parallel and a two-field electrostatic precipitator. Distribution of calcine collected is 10-15 percent from bed overflow, 55-65 percent from boilers, 20-25 percent from cyclones and 2-6 percent from precipitators. A Monsanto designed contact plant produces acid at a rated capacity of 365 T/da (100% H2304 basis) from roaster gases diluted to 7.0-7.5% S02. The process steps are gas cleaning and cooling by weak acid scrubbing, acid mist removal in electrostatic precipitators, gas drying in 93.5-96.5% H2304, conversion of sulphur dioxide to sulphur trioxide in a four-pass converter containing vanadium pentoxide catalyst, and absorption of the sulphur trioxide in 98.1-98.3% H2S04. Product acid is 93.2% H2S04. Conversion efficiency is 97-98 percent and overall sulphur recovery from concentrate is 87-90 percent. The adjacent fertilizer plant has a capacity of 135 T/da of P205. The products are 54 percent phosphoric acid, run-of-pile triple superphosphate, and diammonium phosphate. The plant is of Ugine-Kuhlman design and consists of an air-cooled multiple tank reactor, a tilting pan filter, and a forced circultation vacuum evaporator. Triple superphosphate is made on a continuous belt and partially cured in a rotary dryer. Diammonium phosphate is produced by ammoniating phosphoric acid in a single stage reactor. The slurry of ammonium phosphate is granulated in a blunger and dried.

Jan 1, 1970

By G. T. Adel, J. D. Miller, C. L. Lin

Washability analysis determines the maximum separation efficiency that can be achieved in the operation of coal preparation plants. Unfortunately, there are no on-line techniques for determining this most fundamental information. It now appears possible to determine coal washability on-line through the use of x-ray computed tomography (CT) analysis. This paper presents the results for the development of a CT-based, on-line coal washability analyzer. The development of appropriate hardware and software necessary to adapt a CT analyzer for use in determining coal washability is discussed. Testing of the analyzer using well-characterized coal samples is presented.

Jan 1, 2000