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By Payal Bakshi, Sangeeta Tiwari, Mohini Saxena, Anusha Sharma

"Recycling and reuse is perhaps the most favorable and beneficial way to solve the problem of industrial waste disposal. However, the use of industrial wastes like fly ash, tailings etc is in practice for brick manufacturing, land filling, embankments filling etc. These existing practices do not add enough value to the wastes. Moreover, some wastes may have very good potential to be incorporated in valuable products like paints. This study reveals about the value addition of industrial wastes like fly ash, copper tailing and blue dust as an extender in paints. Comparative properties and characteristics of these three wastes for application in paints are discussed. Fly ash, copper tailing and blue dust were characterized for physical, chemical and mineralogical properties and used in different paint formulation based on their potential to the particular application. Fly ash has been found very useful in enamel, primers and flat paints while copper tailing was found suitable in oil based paint formulations. The performance of fly ash and copper tailing paints were evaluated by comparing with those of conventional whiting extender. Blue dust was compared with conventional iron oxide for use in priming coats. The extender properties of these new material and paints prepared out of blue dust, fly ash, copper tailing, were found better than the existing related materials.IntroductionIndustrial development and environmental reforms should run parallel to benefit each other. This is the best policy, which must be adopted by industries for the sustainability of the current modern life. The high degree of industrialization and urbanization adds scads of waste resulting problem of pollution and challenges of their safe disposal. The nature of waste depends on the process that produces waste as its by-product. Therefore, policies/plans are prepared by keeping the view of possible potential hazards. Among the different wastes some are generated in the industries, which have great importance in the country’s economy, like fly ash waste from thermal power industry and mining waste. In India the total contribution of thermal power in the total electricity generation is more than 70% and it produces about 123 million ton of the fly ash annually [1]. On the other hand mining plus extraction of ferrous and nonferrous metals also produce waste in million tonnes of categories [2]. Since the generation of this waste is very high they cover large land area for their disposal and dumping [3]. The storage of this waste in the dumping sites for long time duration may cause problem of land, surface water, ground water and fertile soil contamination [4]. As far as Indian fly ash is concerned it does not contain toxic element over the limits of hazardous waste [5]. However, in general most of the metals deposits are found in the form of pyrite based ores. Mining wastes contain toxic elements e.g. heavy metals in varying amount, which get mobilized when pyrite get oxidized to sulphate producing sulphuric acid [6,7]. These ores contain very less amount of targeted metals in themselves thus they require extensive preprocesses for concentration of the metals [8]."

Jan 1, 2008

By T Martin, M Martin

Mining of sulfide-based ore deposits typically creates a potential for release sulfate and acid to groundwater, and many operating mines are thus predicted to require long-term post-closure treatment of groundwater. Toward the goal of selecting an optimal treatment, this paper compares costs of four demonstrated technologies for long-term treatment of sulfate-contaminated groundwater: in situ biological reduction (ISR), in situ permeable reactive barrier (PRB) composed of compost and zero-valent iron (ZVI), and two ex situ æpump-and-treatÆ technologies ù ex situ biological reduction, and ex situ nanofiltration with lime precipitation. Costs for each technology are estimated for treatment of a model aquifer system containing sulfate-contaminated groundwater, with the goal of treating 1 m3/min of groundwater from 1500 mg/L sulfate to a target level of 500 mg/L. (Metal removal is typically achieved by these technologies, and thus excluded from this analysis.) Costs are estimated for each technology as a function of treatment duration (ten to 60 years) and, for PRB and ISR, required replacement frequency (ten to 60 years). Estimated costs include capital and operation and maintenance, using a seven per cent net discount rate. Results indicate that ex situ technologies (biological reduction and nanofiltration with lime precipitation) are the most expensive. In situ technologies may be much less expensive, with PRBs having by far the lowest potential cost, followed by ISR as the next least expensive. However, actual costs for the in situ technologies are very sensitive to poorly constrained parameters, particularly the ZVI content of the PRB, and the effective life of PRB and ISR systems. This large cost uncertainty supports delaying remedial actions where viable, and suggests that feasibility analyses are generally warranted for specific applications to focus on these sources of uncertainty.

Jan 1, 2003

DURING a period extending over about the last two and a half years, the performances of various types of rock drills in use at the mining department of the Zinc Corporation have been recorded as closely as possible. The main objective was the collection of sufficient data regarding internal water-feed hammer drills, when used for the same class of work as the older type reciprocating drill, to enable a decision to be formed on the following points:(1) Whether internal water-feed, hammer drills could efficiently perform the work previously carried out by reciprocating drills without excessive repair cost to the former.(2) If so, was it good policy commercially to eliminate the reciprocating type completely, writing these off together with the accompanying spares, accessories, and solid drill steel, and replacing with internal water-feed hammer drills, with lighter fittings, and the requisite amount of hollow drill steel.(3) What were the best types and makes of hammer drill for different classes of mining work(4) What was the economic aspect of the introduction of the principle of internal water feed to every type of machine throughout the mine.Prior to the strike of workers in May, 1919, Ingersoll-Rand E47 31-in. reciprocating drills had been used for stoping and for the major portion of driving, cross-cutting, and similar work. The accompanying columns and fittings were all 4t in. The drill steel used was solid octagon. Up to 1916 these machines were also used for winzing, after which time jackhamers were more successfully employed. These latter were of the usual types, varying in weight from 40 to' 46 Ib. Ingersoll BCR 430, Holman No. 2 spool valve, Sullivan DP33,. Waugh Clipper 55. These machines were all dry head; the drill steel was hollow hexagon, shanked.

Jan 1, 1923

By Mark W. Booth

THIS paper deals with the subject of inbye power as the mechanical . engineer sees it, and invites discussion, particularly from the electrical engineer. At the outset, it will be necessary to clear the ground a little by some definite statements. In this question of compressed air versus electricity for the transmission of power and driving of machinery in underground workings, we must see that there are definite fields that compressed air on the one hand, and electricity on the other, can fill to much greater advantage than the other, and in which there can be little or no direct competition, at least for the present. It is under conditions where these fields overlap, and as to where the line must be drawn between the two, that the respective merits of each must be considered. Steam was the first means used to transmit power from the surface of a mine to drive underground machinery, and so long as distances were short and means possible to ensure fairly efficient use of the steam, and where the fire hazard was not in evidence, it gave fairly efficient service. I have in mind an instance during my own connection with the mines-a duplex pump at the shaft bottom at Princess colliery, driven by steam, and which used the steam compound and condensing; and, but for the mechanical difficulties, it was a not inefficient proposition. The steam was taken direct from the surface boiler-plant through a line down the shaft, and used directly at the pump, being in effect a straight conversion from coal to steam to work done. It was the mechanical arrangement, and failure of the pump in that direction, that led to its removal and replacement by something better in the way of an electrically driven turbine pump. Given modern mechanical arrangements, such as a steam-turbine drive on a high-speed turbine pump, and all other parts made in accordance with best practice in the art of steam generation and use, a steam-driven unit fully as economical would have been obtained. But the central electric supply station, and all that that means or promises (not always the same), turned the scale against steam.

Jan 1, 1936

Factors influencing Cost, systems available for transporting minerals, particularly on land, and methods used for comparing costs are reviewed. A case study presents comparative costs graphically for coal haulage, assumed in Queensland, by 31-6" gauge rail, belt conveyor, 120 ton road train, slurry pipeline, and transmission of electrical energy by transmission line. The effect on comparative results of alternative methods of including initial capital and capital replacement costs is considered and approximate break-even points between systems, bearing in mind the assumptions taken, are discussed.

Jan 1, 1973

By M. L. Requa

IN the fall of 1910, the Nevada Petroleum Co., operating in the Coalinga field in California, determined to drill a number of wells with rotary tools, in order to prove conclusively the relative value of the rotary as compared with the standard rig. At that time, the rotary was but little known in California and its proposed introduction met with considerable criticism on the part of the operators and quiet opposition among the standard drilling crews. There were few men available who were competent to handle a rotary outfit, and these few had had but little experience in California fields. Machines as then built were much inferior to those now used, being lighter in construction and of a poorer quality of material. The shells encountered at depths between 1,700 and 2,000 ft. seemed to be too hard to permit of successful drilling, and at these depths the rig was changed over to standard tools and by them completed. Because of the heavy depreciation, the time lost in converting from rotary to standard, and the comparatively small profit, it was concluded that unless in the future there was some material improvement in rotary rigs, nothing would be gained by drilling with rotary tools upon this property. Little or no drilling has been done upon the property since that time, but in the meantime a large number of men have been trained in the use of the rotary; in fact, many standard-tool men have abandoned standard drilling and, starting in as roustabouts, have become thoroughly competent rotary operators. The improvements in the machinery have been such as to remove many of the objections and the rotary drilling of to-day is in every way superior to that of 1910. In contrast, a well drilled recently by the Kern Trading & Oil Co., in the east side Coalinga field may be cited. The first 2,500 ft. was drilled in 24 days and it is confidently expected that the water string will be set with a rotary at 3,400 ft.-a record obviously far beyond that made in 1910 by the Nevada Petroleum Co.

Jan 2, 1915

This work describes preliminary laboratory testwork to study the relative leaching response to cyanidation and thiourea leaching of an oxidised low grade gold ore. The potential of an alternative lixiviant like thiourea to commercial processing of mine dumps is discussed by specific reference to selected case histories. It would appear that on the basis of the available published information, there could well be a case advanced for the acid thiourea leaching of sulphide concentrate tailings being subjected

Jan 1, 1994

By Milka Madahana, John Ekoru

Particulate matter is one of the leading pollutants in the mines and in the communities in the areas near or surrounding the mines. The availability of large Clinical and environmental data sets obtained from the mines has brought about the need to use better approaches to analyze the collected data. Prediction and control of particulate matter concentration in the mining environment is essential in minimizing air pollution and protecting the respiratory health of mine workers. The main objective of this research work is to analyze datasets obtained from a mining area using machine learning techniques which will then be followed by the real-time estimation of dust particulate in the mines. Datasets collected over 8 years in an hourly sampling rate from an open cast mine is cleaned and analyzed for trends and patterns. Artificial Neural Networks (ANNs), Support Vector Machines and decision trees techniques are applied to analyze and predict the particulate concentrate. The estimation of particulate concentration using Artificial Intelligence will significantly contribute toward the development of quantitative tools to be used in correlating of cause and effect of dust concentration in mines. The results obtained in this work will form a sub-unit of a respiratory health monitoring system for mine workers. The system is essential in the collection, analysis, and tracking of mine workers’ respiratory health, thus it forms part of an early intervention system for mine workers. From the existing theory, we expect the Artificial Neural Network will perform much better than previously used classical models.

Jun 25, 2023

By Neil C. Hario

The purpose of this paper is to explore the various designs and design alternates available for large fine crushing plants, and to discuss objectives and criteria upon which these designs are founded. Common practice for size reduction of run-of-mine or run-of-pit ore in preparation for subsequent treatment utilizes a three-stage system. First stage, or primary breaking is generally carried out with jaw or gyratory crushers. The usual output from this stage provides a feed for fine crushing in the size range of 6 to 10 inches. This discussion is limited to the process of fine crushing, and we will further limit it to what is commonly called secondary and tertiary crushing. We define a large crushing plant as one with a minimum capacity of 2,000 tons per hour. Generally, these plants utilize cone type crushers to reduce ore for rod, ball, or pebble mill feed.

Jan 1, 1970

By Z Li, Z Yan, M Yang

Recycled cement paste (RCP) is a kind of CaO-rich recourse, which is separated and recycled from waste concrete. Due to its chemical composition, RCP could partially replace lime as a flux and be loaded into electric arc furnace (EAF) steelmaking process. Rapid RCP assimilation and liquid slag formation have significant effects on EAF steelmaking process. In this study, the dissolution behaviours of RCP and lime at temperatures of 1400°C to 1500°C were investigated using static dissolution experiment. The dissolution rate, interfacial microstructure, as well as dissolution mechanisms were discussed. It was found that the dissolution rate of RCP was much higher than that of lime. RCP could be completely dissolved into slag in 60 secs at 1500°C. The dissolution rate of lime was in range of 3.910-4 g/(cm2ꞏs) to 2.610-3 g/(cm2ꞏs), while that for RCP was in range of 1.510-3 g/(cm2ꞏs) to 4.910-3 g/(cm2ꞏs). As for lime dissolution, a dense layer of dicalcium silicate (C2S) generated at the dissolution interface, and the diffusion of Ca2+ in C2S layer was the limited step for lime dissolution. Some of calcium ferrite could be observed at the interface between lime and C2S layer. As for the RCP dissolution, there is no obvious product layer of C2S formed at the dissolution interface, Fe2+ continuously diffused into RCP layer during the dissolution, and small amount of FeOꞏMgO solid solution generated and gathered at the dissolution interface. The dissolution rate of RCP in EAF slag was quite higher than that of lime, partially add RCP as flux in EAF could accelerate the slag formation process.

Aug 21, 2024

By H. Sevim, M. K. Mohanty, Y. P. Chugh

The economics of “Conventional” and “Advanced Fine Coal Cleaning” (AFCC) circuits for refuse pond reclamation as well as active mine/preparation plant applications have been evaluated. The benefits of innovative tailings management schemes on the profitability of mining enterprises have also been evaluated as a part of this study. Analyses have been conducted for scenarios involving generation of two different product qualities of 5.81%, a quality that was obtained during AFCC circuit operation at SIU, and a more typical 8% product ash. The inability of the conventional circuits to achieve profitability while producing the lower ash content product is revealed from this approach. The spiral-only circuit is unable to produce a 5.81% ash product. A different version of the same circuit involving spirals and column flotation produced the desired product quality but could not maintain profitability. The AFCC circuits on the other hand, achieve net present values greater than $6 million, with return on investment in excess of 50%, in the water-only as well as dense medium mode, with and without tailings management. Addition of tailings management system to any of the circuits is found to enhance the project economics. The comparison of water-only and dense-medium applications of the AFCC circuits indicate a better suitability of the dense-medium circuit for producing premium quality products. Application of AFCC circuits in active mines offer better economics compared to applications in refuse pond recovery operations. The superior project economics are a result of reduced overhead cost and use of existing material handling facilities. Sensitivity analysis reveals the significance of capacity utilization on project economics.

Jan 1, 2002

By S. R. Gilbert, C. O. Bounds, R. R. Ice

"Many ores are upgraded to produce auriferous sulphide concentrates (principally pyrite) which are not amenable to direct cyanidation. Often these refractory gold-bearing concentrates contain undesirable elements such as arsenic, antimony, etc., which present environmental and/or processing problems. The pretreatment techniques most often considered for processing these concentrates prior to cyanidation include roasting, pressure oxidation, chemical oxidation, pre-aeration, and recently, bacterial oxidation. In this paper, the technical performance and economics of bacterial pre-oxidation vs the classical roasting approach for treating a specific auriferous pyrite concentrate are compared. The results provide flow sheets, preliminary operating and capital cost estimates, and rates of return for the two processes and demonstrate the potential advantages of bacterial pre-oxidation. IntroductionThe increased attention that gold has received in the last ten years has prompted mining companies to look for lower grade orebodies. Many of these low-grade ores do not respond well to conventional direct cyanidation recovery techniques. As free-milling ores decline, gold is sought in refractory environments which, ultimately, drive up the treatment costs.There are a number of reasons for a gold ore to be termed refractory, most of which can be identified by thorough mineralogical examination. The mineralogical occurrence of gold can, to a great extent, impact the method by which the ore is treated and, consequently, the capital and operating costs. Gold in a refractory ore can often occur as discrete particles locked in a host of silicate or carbonate gangue, or associated with sulphides or sulphosalts. Depending on its particle size, gold can be liberated and recovered by mechanical size reduction such as regrinding. This approach is most effective for gold in the 10 - 50 micron range. Gold can also be associated with clay, iron hydroxides, or other minerals which tend to coat the gold surface and inhibit the penetration of the cyanide solution, resulting in poor plant recovery. Because of their resistance to cyanide dissolution, gold tellurides and other refractory gold minerals are also responsible for losses to the tailings. For these reasons, it is paramount that the mineralogical occurrence of the gold be thoroughly and accurately determined before selecting a final process flowsheet."

Jan 1, 1988

By F. Milton Lewis, James L. Coburn

In this paper the economics for a conventional comminution system, consisting of three stage crushing and one stage ball mill grinding, are compared to a semi-autogenous/ball mill comminution system. Capital and operating costs for a copper concentrator in Arizona at two levels of production, 20 and 40 thousand short tons of ore per day, are estimated. In order to minimize the investment and operating costs, large size equipment is selected for all four cases. These cost data are evaluated and conclusions are presented.

Jan 1, 1982

By Orville R. Lyons

THE relative performance of coal washing equipment, or the effectiveness with which any type or make of equipment removes impurities from coal, has been most difficult to evaluate in the past. The most widely used yardstick is the Frazer and Yancey efficiency formula developed in 1922,1 but Yancey in a later article states that "washers treating coals of different density composition or operating at different densities of separation cannot be compared directly on the basis of this criterion."' Prior to and since 1922, a variety of other methods has been used for comparison purposes, including the distribution curve, the error area, and the "ecart probable" or probable error. Yancey and Geer in discussing these methods conclude, "Performance can be evaluated in a number of different ways, with the choice of the proper method to use being dictated by the objectives of the investigation and the data available."' It is true that performance can be evaluated in a variety of ways, but if the equipment is to be evaluated on an effectiveness basis, there should be only one universal comparison method. Varying methods have been used because one universal comparison method has not been found or developed. In the article previously quoted, Yancey and Geer state in clear terms the primary concept for a universal comparison method: "One of the simplest, and certainly one of the most obvious evaluations of washery performance is the quantity of sink material in the washed coal and the float material in the refuse. If the washery products are tested at the density at which the washing unit is operated, the sink in the washed coal and the float in the refuse represent material that has been misplaced." The quantity of misplaced material was used as a criterion of washery performance by Lincoln in 1913,8 by the United States Bureau of Mines in 1938,4 by Hancock in 1947,3 and by the national French research agency Cerchar in recent years.2 In 1950 Anderson' proposed the use of this criterion as an efficiency value to replace the Frazer and Yancey formula. However, none of the above-mentioned investigators used the misplaced material concept in a manner that would provide universal coal-cleaning equipment comparisons. The Correlation Theory The ideal coal cleaning process would treat all sizes and would make a perfect separation at any given specific gravity. All material lower in density than the desired value would report in the coal product and all material higher in density would report in the refuse product. Unfortunately, no known cleaning process achieves this goal and there seems little likelihood that any process yet to be invented will do more than approach it. When coal is treated in volume under operating conditions, it is impossible to avoid mechanical entrapment, fluctuations in throughput and effective gravity of separation, and the creation of turbulent currents, even when a true heavy-liquid bath is used and the feed is closely sized and contains little intermediate gravity material. This being so, it is possible to appreciate the difficulties inherent in trying to obtain a perfect separation when treating a wide range of sizes and a feed containing high percentages of intermediate material, using turbulent currents to help create the effective separation gravity, under operating conditions which normally tend to be on the overload side. When coal is separated from refuse in any coal cleaning equipment, some refuse always reports to the coal and some coal to the refuse; the writer therefore assumed that there should be a relationship between the total amount of misplaced material produced by any given piece of equipment and the difficulty of separation as represented by the percentage of near gravity material in the feed. With small amounts of near gravity or ±0.1 material in the feed there should be less misplacement of material than would occur with large amounts of near

Jan 1, 1952

By H. A. D. Kirsten

The structural competence of mesh- and fibre-reinforced shotcrete was investigated in two series of tests on full-scale panels. It was found that uniformly distributed loading was marginally less demanding than point loading, mesh reinforcement was superior to fibre reinforcement, Dramix fibre and Melt Extract fibre were similar in reinforcing characteristics, and 30 mm to 35 mm was an optimum fibre length. Furthermore, the consistency of the fibre content could not be assured on practical mining sites, and an average fibre content of about 1,5 per cent was the maximum that could be obtained for fibre lengths not exceeding 35 mm. Also, owing to its greater flexibility, shotcrete reinforced with diamond mesh was preferable to fibre-reinforced shotcrete in mining excavations subject to squeezing conditions. In view of the difficulties associated with assuring the consistency of the fibre content in mining applications, it appears that the sophisticated attributes of fibre-reinforced shotcrete can be effectively exploited only in civil-engineering applications in which quality assurance is subject to explicit contractual control. Load-deflection curves, together with measurements of compressive stresses in the surfaces of some of the panels, enabled the ultimate flexural strength of reinforced shotcrete to be rigorously formulated, and rigorous design procedures for shotcrete subject to bending and direct thrust to be developed. The application of the procedures is demonstrated in a practical design case in which the moments and thrusts in a tunnel lining are determined in a numerical analysis of the interaction between the lining, the surrounding rock, and the rock-reinforcing system.

Jan 1, 1992

By M. Roy, Y. Roux, R. McIvor

"La Compagnie Minière Québec Cartier at Port Cartier. Quebec, utilizes six parallel wet grinding open circuit ball mills to grind iron ore concentrate for the pellet making process. In an attempt to improve grinding efficiency, a mix of 605 25 mm (1 in.) and 405 38 mm (1 ½ in.) grinding balls was charged to mill D while the opposite mix was used in all the other mills. Because media costs were higher for mill D, it was necessary to find out if the smaller balls provided a net economic benefit. Therefore, a test program was undertaken to determine the relative efficiency of the charge in mill D compared to the other grinding mills. Mill F was chosen for comparative efficiency measurements.The first problem was to choose a suitable definition of grinding efficiency. In management's view, previous attempts to quantify milling performance using population balance modelling were ineffective because they were unable to relate breakage rate functions to grinding circuit economical in any meaningful way. Work index analysis was also not favoured because of a number of complicating effects, including the fineness of the feed and product streams and open circuit grinding.It was decided to adopt a recent innovative approach which had been successfully applied to closed-circuit ball milling. For ""functional performance analysis of ball milling"" (McIvor et al., 1990; McIvor et al., 1991), a size criterion is first chosen to define ""fines·' and ~'coarse particles"". In this case 45 microns (325 mesh) was chosen to distinguish fines from coarse material. The relative grinding efficiency of each ball mill environment would then be ~e cumulative grinding (disappearance) rate of coarse (plus 45 nucrons) material per kilowatt hour of energy applied to the coarse material in the mill, divided by the grindability (measured in the laboratory) of the coarse material in the mill feed.Because the disappearance rate of coarse particles is also the production rate of fines, the following equation can be written for the efficiency of the mill environment, Effe, as defined above:"

Jan 1, 1994

By D. R. Spink, D. N. Okuhara

Two reagents have been proposed for the commercial extraction of copper from leach liquors, namely LIX64N produced by General Mills Chemicals, Inc. and Kelex 120 produced by Ashland Chemical Company. Each of these reagents consists of an active extractant and an additive. In the LIX system, the additive improves the kinetics of the extraction reaction whereas for the Kelex system, the additive acts primarily as a third phase inhibitor. Equilibrium and kinetic studies have been conducted with both systems. In general, the Kelex system effectively extracts copper at a lower pH than LIX64N; furthermore, the rate of extraction with Kelex is considerably faster than with LIX which indicates that systems employing the Kelex extractants might also be able to employ agitated columns rather, than mixer- settlers which are absolutely necessary for effective extraction with the LIX reagents.

Jan 1, 1973

By George J. Karabin

Over the past few years, the use of fully grouted bolts to support the roof of underground openings has increased dramatically. While conventionally anchored bolts effectively reinforce the strata of over 80 percent of mine workings, the use of resin grouted bolts in areas of severe roof conditions has resulted in a substantial reduction in roof falls and an overall improvement in both safety and production. The Roof Control Group of Pittsburgh Technical Support Center, Mining Enforcement and Safety Administration, has conducted numerous field and laboratory investigations to isolate parameters which contribute to the effectiveness of both conventional and resin bolting systems. Specially developed strain-gamed resin bolts, compression pads and differential sag stations installed during the mining cycle have generated data reflecting the loading characteristics and resulting roof stability associated with each type of support. Additional investigations have produced optimum installation parameters of the two systems including the effectiveness of thrust/torque controls for conventional bolts and various bolt/ hole diameter combinations for fully grouted rods. Theoretical analyses, coupled with the actual field and laboratory findings and numerous installation observations have facilitated basic comparisons of the working mechanisms and failure modes of both resin and conventional systems.

Jan 1, 1976

In Finland fogging in mine ramps is a common problem. Fog forms when warm and humid exhaust air moves upwards the ramp, where the skin temperature of surrounding wallrock decreases to about 10¦C. Changes in fog thickness and lengths of fog-filled parts of the ramp depend at least on climatic conditions outside, diesel equipment usage in the mine, and the length of working sequence at a time. In this study fogging problems in three different mines, namely Orivesi mine, Louhi mine and PyhSsalmi mine, with different ore types, different layouts, and different depths are presented and compared. All these mines use diesel equipment and the ramp is used as an exhaust air route. In two of these mines the ramp is used as a haulage road and in the other for supply and repairing, checking, and surveying purposes. Measurements of humidity, temperature, air velocity, and dust content were performed in these mines. The measurement results, as well as field observations are presented, compared, and discussed. The results show that fog is observed in areas with high relative humidity values when air temperature and dew point temperature are less than 3¦C apart from each other. Air velocity values were high in many areas. In addition, measured particle concentration values were mainly high in areas, where fog was observed. A short literature review of recent fog formation studies is also presented. It demonstrates the complicacy of the fog formation process. Similarities and differences concerning fog problems in the ramps are found and presented. As already a lot of new information is gained, the continuation of research is considered plausible to find more economical ways of solving and preventing this problem in the future.

Jan 1, 2005

By S. S. Nosseir

INTRODUCTION General Quebec Cartier Mining Co. (QCM) , Canadian subsidiary of U . S. Steel Corporation, operates two iron ore concentrators at the North Shore of Quebec, Canada. In its own operation at Mount-Wright, over 45 million tons per year of crude ore are ground, fully autogenous in six 32' x 11-1/2 Dominion Engineering mills. At Fire Lake/Lac Jeannine operation, property of Sidbec Normines Incorporated (SNI), over 11 million tons per year of crude ore are now ground, fully autogenous, in twelve 18' x 5' Hardinge mills. The Lac Jeannine concentrator was originally constructed by Q.C.M. in 1960 and. was since operated, in beneficiating the Lac Jeannine ore deposit until fully depleted in 1977, at an annual capacity of 18 million tons of crude per year. In 1976, the new company, Sidbec Normines, was incorporated by Sidbec Corporation, British Steel Corporation and Quebec Cartier Mining Co., in proportional holding of approximately 51 to 41 to 8 percent respectively. Since the acquisition of the Fire Lake ore deposit and Jeannine concentrator by the new company, QCM has continued managing the mine and concentrator operaions for S idbec Normines Inc . Flow-Sheet Design Jeannine concentrator comprises twelve identical mill lines grinding the ROM from 8" top size to 14 mesh nominal size range. Driven by 600 H.P. motor at 13.4 RPM corresponding to 73% critical speed, each mill is operated in a closed circuit with two 5' x 12' vibrating screens dressed with 2.5 x 25 mm slot opening cloth. The ground product or screen undersize at an average grade of 33.4% Fe is beneficiated by gravity separation to produce 66.5% Fe concentrate grade at 42.7% weight recovery. At the rougher stage, each mill line con-tains 128 Humphrey 13- 1/2" pitch spirals which produce a final tailing, middling and concentrate products. The rougher concentrate - so produced - is classified by 64 Humphrey pocket sizers into an underflow, which is a final concentrate product and an overflow which is further upgraded by 64 cleaner spirals. The cleaner concentrate together with the pocket sizer 's underflow are dewatered on a 13'[6] horizontal filter. The rougher middling together with the cleaner tailing are recycled in a closed circuit with the spirals. The rougher tailing after dewatering by hydrocyclones is disposed of by a

Jan 1, 1982