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By Yves C. Lizotte, Malcolm J. Scobie

Blast fragmentation is affected by the physical and mechanical characteristics of the rockmass as determined by geology as well as the explosive energy, the partition of this energy, the actual amount of energy transmitted to the rockmass, the blast geometry and the detonation sequence. This paper reviews studies of the influence of geology on blast fragmentation, particularly structure, i.e. the in situ or inherent fragmentation arising from rock discontinuities. Prior structural studies have mainly been limited to comparing equivalent diameters of the in situ blocks with the size distributions of the blasted material. No criteria exist to integrate structure into blast design. This paper reviews on-going studies of the blasted fragment shape and size observed in different rockmasses aimed at analyzing geological control over blasting. Digital photoanalysis was applied {o measure size and shape distributions, to correlate results with block size distributions and shape parameters of the inherent fragmentation. An integrated and comprehensive fragmentation assessment procedure which accounts for geological control is proposed. The potential of this research is to provide a greater understanding of the true contribution of explosives to fragmentation, the effects of fragment shape on system productivity and the reduction of blast-induced damage.

Jan 1, 1994

By P. R. A. Andrews

"Areas of bubble contact were delineated for cassiterite in the presence of chloride, fluoride, sulphate, phosphate, oxalate and citrate ions using dodecylamine hydrochloride collector. The limits of pH, and the minimum collector concentration were established using the bubble pick-up technique.Bubble contact did not occur below the p.z.c. (point of zero charge), estimated to be 4.7, in the presence of chloride ion, and fluoride ion showed only a weak response in the same pH range. These anions are considered to act as indifferent electrolytes. Activation at pH values below the p.z.c. was obtained with sulphate, phosphate, oxalate and citrate ions; the greatest response was by oxalate and phosphate ions extending pick-up to between pH 2.0 and 2.5.Activation appeared to be strongest when anionic protonated species were in maximum concentration. Strong activation also corresponded to the formation of very stable metal ion species, thought to be formed on the surface of cassiterite, between exposed stannic ions and modifying anions; dodecylammonium ions being absorbed on the surface so modified."

Jan 1, 1991

By Wenbin Zhang, George W. Poling

"Prior or simultaneous addition of ammonium sulphate can, not only enhance the activating effect of hydrosulphide ions on xanthate flotation of malachite but it can also eliminate possible detrimental or depressing effects of excess residual hydrosulphide . Use of relatively inexpensive ammonium sulphate appears to reduce the consumption of the much more expensive sulphidizing agent for optimal flotation recovery INTRODUCTIONOxidized copper minerals float poorly with conventional thiol-type collectors. This is believed to be due to the nonadherent character of the collector coating, which is readily re- moved by particle abrasion. Rosas and Polingr reported emulsion flotation seemed capable of resisting abrasion during the flotation of non-sulphide copper ores. It appears that the most common flotation technique for oxidized copper minerals comprises sulphidization with Na .S and subsequent treatment by sulphydryl collector. The principles of this method are available . Malghan recently reviewed the studies on the role of sodium sulphide in the flotation of oxidized copper, lead and zinc minerals.It is well known that in the sulphidized flotation of malachite, the activation process depends strongly on the concentration of the sulphidizing agent, i.e., insufficient sulphide gives poor recoveries because of inadequate sulphidization; excessive sulphide also causes poor recoveries due to the depressive effect of the residual hydro-sulphide ions in solution. A major difficulty in sulphidization, especially with a varying feed nature, is to establish and maintain the optimum sulphide concentration in solution. In plant practice, vigorous control of sodium sulphide is required and stepwise addition of the reagent is recommended to improve the efficiency of the process. Soto and Laskowskim have found that malachite, depressed by an excess of sulphide ions, can restore its floatability if air or oxygen is introduced to the pulp. They attributed this to the oxidation of the HS - ions on the surface of the mineral. Jones and Woodcockuvi showed that a sulphide ion selective electrode could be used to optimize the conditions (e.g . amount of hydrosulphide, sulphidization time, flotation time) more readily and precisely than existing methods in the flotation of oxidized copper ores."

Jan 1, 1989

By Gerhard Heinrich

"The composition and the possible utilization of a slag sample from a nickel smelter operation in the Sudbury district were investigated. The sample containing 0.136% Ni, 0.10% Co and 0.29% Cu could be characterized as a fayalit e slag with a sulphide bearing glass matrix. While the fayalite contains most of the Co, Ni is enriched within both sulphides and fayalite. Cu is only present in higher amounts in sulphides or metallic droplets. This, as well as the small grain sizes of the mineral phases, inhibits significant enrichment by mechanical methods. However preconcentration is possible by means of magnetic separation. An effective hydro metallurgical approach proved to be oxygen pressure leaching with sulphuric acid. Accordingly, the method proposed for processing the investigated slags is to combine the normal utilization as coarse construction material with pressure leaching of the fines and the ground upgraded magnetic fraction. Besides the recovery of base metals, the leaching process yields fine-grained residues, which, after addition of CaD, will set to a cement-like material, or can be used as a pozzolanic cement colour. INTRODUCTIONDuring the pyro metallurgical processing of nickel ores, the valuable metals nickel, cobalt and copper are enriched within the sulphide matte, similar to the smelting of copper ores. The resulting waste materials are slags, which may still contain valuable metal contents of up to 0.5% or above . These contents are close to the cut-off grades of some nickel-cop per-cobalt deposits. The methods that are used to recover copper from slags comprise in general further pyro metallurgical treatment, often after enrichment by flotation. Another area of application for such slags is their use as construction material. However, this might cause environmental problems if nickel containing dusts are produced.Based on one example, the characerization and a possible way of complete utilization of nickel smelter slag s as construction materials in combination with partial base metal recovery are described in this paper. The investigated slag sample was obtained from a dump site in the Sudbury district, Ontario. Detailed test series for process optimization were not possible during this study due to the limited available sample size."

Jan 1, 1989

By G. J. Desroches

"In order to improve zinc plant productivity, areas of low efficiency were identified. As a result, solution purification was improved, additional dust precipitator capacity was made available and a new zinc casting plant was installed. The effects of these and other changes are discussed and the expenditure and benefits reviewed. HistoryThe operation of the Flin Flon metallurgical complex including an electrolytic zinc plant commenced in October 1930. The design of the zinc plant was based on that of the Commonwealth Mining and Smelting Company at Trail, British Columbia, the Anaconda Great Falls Montana Plant and the Electrolytic Zinc Company plant at Risdon, Tasmania.The plant began with a design capacity of 60 tons per day and currently produces 255 tons per day. Since plant start-up, approximately 3.5 million ton s of zinc metal have been produced, most of it exceeding special high-grade specifications.IntroductionThe facility incorporates the basic process steps of roasting, leaching, purification, electrowinning and casting. Figure I details the flowsheet. During the late 1970s, management recognized that in order for the zinc plant to continue operating, it was necessary to introduce cost-effective practices.Study groups, consisting of design engineers and operations personnel, were formed to investigate the details of plant revisions.The objectives were to improve productivity by reducing manual labour, reducing energy usage, increasing zinc recovery, increasing metal output and reducing costs.This paper details the changes made during the period 1979to 1986 and shows how these changes have affected the productivity and economics of the zinc plant operation."

Jan 1, 1988

By William E. Stanley

The paper examines the problems that arise in undertaking socio-economic impact studies to meet the requirements of environmental-impact legislation governing new mine development or expansion. An approach to undertaking socioeconomic impact studies is outlined, responses to specific problems are discussed and ways in which the mining company can use these studies to meet its general objectives are explored. The paper concludes by examining how a socio-economic study, as required by legislation, can be used to assist the mining company in achieving some general objectives it will have in developing a mine.

Jan 1, 1979

By P. R. A. Andrews, R. K. Collings

"SUMMARYIn cooperation with the Ontario Ministry of Natural Resources, selected deposits of Potsdam sandstone in the Ottawa-Kingston area were core-drilled and evaluated by CANMET as potential glass and foundry sand. Analyses of cores ranged from 84.6 to 98.5% Si02, 0.09 to 0.65% Fe203' and 0.07 to 2.10% A12°3' The cores were largely composed of white to grey, relatively tough sandstone. Impurities included iron as pyrite, iron oxides and staining on the quartz grains; carbonates as calcite; and aluminum oxide as feldspar. Beneficiation included size reduction to grain size, about 600 um, by jaw, roll and cone crushers, and grinding in a muller-mixer. Removal of fines was achieved by attrition scrubbing and washing. Further removal of impurities was achieved by reverse flotation of pyrite, iron oxide, feldspar, and carbonate minerals employing both anionic and cationic collectors. Flotation resulted in a substantial improvement in the quality of most samples; analyses of quartz concentrates showed 98.5 to 99.7% Sial> 0.022 to 0.156% Fe203' and 0.07 to 0.62% A1203. Further reduction of the iron content by wet magnetic separation and acid leaching also was investigated. On the basis of the samples submitted, sands/ones from six of the seven deposits studied have potential as sources of glass sand. Foundry sand tests were inconclusive because of the small quantities available for study."

Jan 1, 1986

By E. Contestabile, T. R. Craig

"This report describes the development and effective use of shaped charges as part of the Elliot Lake Mine Roof Stability Project to remotely remove the heads from roof bolts. With a slight modification, these same shaped charges are being recommended f or use to the Brunswick Mining and Smelting Corporation Limited (Mining Division), to cut steel cable bolts.IntroductionThe Canadian Explosives Research Laboratory (CERL) often receives requests to cooperate in solving problems encountered in private industry or other sectors of the government. CERL is equipped to design and test explosive and pyrotechnic devices and the policy is to treat each case individually.The first case reported evolved from a request from a sister laboratory at Elliot Lake. The intent was to remotely remove the head s of roof bolts as part of a mine roof stability project. The device designed for this use was 1000,70 effective in the laboratory and 80% effective in the mine.The second case reported results from a request from a private mining company knowledgeable of the Elliot Lake Project. Their requirement was to remotely cut cable bolts. This request was first treated as presenting a new problem. The solution, however, proved to be very similar to the former."

Jan 1, 1985

By W. G. Brown

"Review of prototype testing of two types of three-phase, argon stabilized reactor/furnace indicates specialized advantages for both, as compared with existing metallurgical furnace technology. They are: the direct-arc, plasma-torch furnace (Knipp); and the indirect, plasma extended-arc reactor (PEAR), developed at the University of Toronto.IntroductionA unique type of three-phase AC, plasma extended-arc reactor, (PEAR), shown in Figure la with gravity feed, has been undergoing development at the University of Toronto, during the past fifteen years: argon gas, introduced axially through drilled-holes in ordinary graphite electrodes, increases the length of indirect arcing and improves electrical stability while decreasing electrode consumption as compared with the performance of AC electric arc furnaces (EAF 's). Meanwhile and independently, a three-phase, transferred arc plasma furnace (Fig. 1b) has been developed by the Krupp Research Institute in Germany, using argon-stabilized, water- cooled plasma-torches with thoriated tungsten electrodes.Accordingly, the two types of furnace are similar as to both three phase and stabilized operation, but otherwise they differ in several respects, ranging from indirect versus direct-arcing to principal intended application and current state of practical development. Between them they raise questions: (I) of their present or fore seeable economics in comparison with existing processes they are intended to replace , (see Heinke and Neuschutz, 1986, for a review of the merit s of AC and DC plasma-torch furnaces); and (2) of the state of development and accessibility of their plasma theory-in-common."

Jan 1, 1991

By Jacques H. Nantel, James F. Archibald

"Frozen tailings material may be required, in the future, to perform the support functions of underground backfill in mines located in zones of permafrost, notably in the high Arctic. To date, no knowledge of physical support parameters of such materials has been obtained in Canada due to the limited mining activity in permafrost regions. It is therefore judged essential that the important physical and mechanical properties relating to frozen tailings material be assessed.Innovative mining methods will be required to cope with the harsh environmental conditions that exist at mining sites in the Canadian Arctic. A preliminary laboratory study, in the absence of field trials at limited Arctic mining sites, has been completed in which the physical and mechanical properties of several types of representative mill tailings mixtures have been investigated. A variety of tests, to determine frozen slurry and pelletized tailings strength properties, moduli of compression, optimum water contents, deformation characteristics and the like, have been conducted.The results of this study will be used to establish guidelines for the suitability of tailings use in such northern mining environments. IntroductionFrozen tailings material has never been used to backfill underground openings in Canadian mines. At present only one Canadian mine, the Polaris Mine situated on Little Cornwallis Island, is investigating the use of frozen backfill. However, the Canadian North covers a large territory and potentially contains a large number of mineral deposits. The understanding of concepts related to mining activities in sub-zero frozen ground is therefore important.The implications of using frozen tailings as underground backfill material are far reaching both from the point of view of mine stability and disposal of tailings in the arctic regions.The strength of frozen materials (soil, backfill, etc.) may take on different meaning depending on the engineering situation. Both the concepts of rupture and excessive deformation need to be considered.The functions usually required to be performed by the backfill material in underground mines may be summarized asfollows:a) working floors; such as in cut-and-fill stopesb) stope wall support; for cut-and-fill and delayed fill stope~c) self-support when backfill is exposed on one face; during pillar recovery phases.For fill functions such as working floors, the short-term response of the frozen fill is the dominant concern. For the other fill functions, the long-term deformation response of the frozen fill, or creep, is of interest."

Jan 1, 1986

By C. John Sullivan

Is there a relationship among the Sudbury nickel ores, the Cobalt deposits, and the Huronian, Athabascan and North Australian uranium provinces? All are within intracratonic basins, with ore near basal unconformities. There are similarities in the geochemical "fingerprints" pointing to basic igneous rock affiliations. Some show a time relationship to basic igneous activity and to great rift systems. Many basic dykes show radioactivity at their edges. A spatial relationship between the uranium-cobalt-nickel ore at Midwest Lake, Athabasca, and basic igneous rock is reported. These remarks are not meant to imply, necessarily, a direct genetic connection between the basic dykes, etc., and the ore. In some cases, the relationship may be similar to that noted between basic and ultrabasic intrusives and volcanic massive sulphide ore deposits. Crustal thinning in these areas, associated with graben, appears to have favoured volcanic activity, as well as basic intrusives. Probably, the potential of the Huronian should be re thought . Athabascan-type ore may be present . New data on mineralization in Turner Township, Ontario, suggest this. The mineral possibilities of other basins should be reconsidered. It is to be remembered that the Athabascan and Kombolgie basins were neglected for many years. Intracratonic sedimentary-igneous basins contain many of the greatest mineral fields of the world.

Jan 1, 1979

By Bruce K. McKnight

"The new H-W orebody was evaluated by considering its production at three technically feasible rates. Capital and operating costs were estimated f or each rate and production sequences were established using mining cutoff limits generated by an ore value algorithm. The viability of the project was then examined by conducting economic and financial analyses of each of the three production cases under various world economic scenarios. These analyses employing standard profitability measures demonstrated all three cases would be economically viable; the production case which maximized the Net Present Value of the project was the one selected for development and production.IntroductionWestmin Resources Limited and its predecessor company, Western Mines Ltd., has been continuously carrying out mining operations near Myra Falls in Central Vancouver Island since December 1966. The mining operation, whose current production rate is approximately 830 tonnes / day (t/d), based on ore derived from the Myra and Lynx Mines, produces copper, lead and zinc concentrates which contain significant quantities of gold and silver in addition to the base metals.In the late 1970s, with Lynx and Myra ore reserves slipping to unacceptably low levels, a major increase in mine-site exploration was initiated, with outside geological staff brought in to complement the efforts of the mine's geological staff. By the end of 1979, this program had resulted in the discovery of a significant new orebody on a deeper horizon some 300 metres below the one which hosts the Lynx and Myra deposits. Throughout the period 1980-1982, major drilling was carried out to evaluate the new H-W deposit and by the end of 1982, drill-indicated mineral inventory for the deposit compared with Myra, Lynx and Price ore reserves was as shown in Table I. During this same three-year period, a major development and evaluation program was underway comprising environmental monitoring and control programs, shaft sinking test mining, bulk sampling and mill testing, rock mechanics studies and preparation of the Stage I and II reports for the multidisciplinary Environmental Land Use Committee (ELUC).As a result of the preliminary engineering work and exploration drilling, it became rapidly apparent that the H-W deposit could support a considerably higher production rate than the current one. Accordingly, engineering studies including scheduling of development and production as well as capital and operating cost estimates, were focused on three possible production rates: 1350, 1800 and 2700 t/d. Our task was to select the optimum one of these three rates."

Jan 1, 1987

By G. Macfarlane

"In 1985, a crushing plant, located in the feed preparation area of Kidd Creek's copper smelter, was modified. The plant, which originally had treated only copper-bearing materials for the smeller, at a rate of 10 tonnes per hour, was expanded to 40 tonnes per hour capacity and modified to treat silica ores to meet the smelter’s flux requirement and also gold ores from the company's Hoyle Pond property. In the storage area adjacent to the crushing plant a 300 tonne per day gold mill was constructed.This paper outlines the construction and first year of operation of the crushing plant and gold mill with emphasis placed on the capital and operating advantages to be gained by utilizing an existing metallurgical infrastructure.IntroductionThe Kidd Creek Mines Ltd. copper-zinc-silver orebody is located 24 km north of Timmins, in Northern Ontario . Mining of this massive sulphide de posit commenced in 1966 and, in the ensuing years, a concentrator and zinc plant were constructed at the metallurgical complex, approximately 17 km by rail from the mine site . At this same site, in 1981, a copper smelter and refinery were commissioned. The smelter which utilizes the Mitsubishi smelting process for the continuous production of blister copper from copper concentrate, has been comprehensively described in previous papers.At the time of the smelter construction, a small (10 tonnes per hour) crushing plant was installed in the smelter's feed preparation area. This plant was used to treat the copper-bearing revert material produced by the smelter, and, apart from periods when slags and matte were produced through furnace drainage at the end of each operating campaign, it was generally under-utilized. This fact, together with the requirement to devise a more cost effective method of providing silica flux for the smelting operation, was the basic rationale for modifying and expanding and crushing plan t. At the same time, development at Kidd's Hoyle Pond orebody had progressed to such a stage that a gold mill of 300 tonnes per day was warranted."

Jan 1, 1988

By P. W. Udo Graefe

"An interactive computer model of a dragline, windrow and spoil pile has been developed for modelling oil sands mining operations. The model is to be used in conj unction with models of other oil sands mining equipment, such as bucket-wheel excavators and reclaimers, conveyors, hoppers, bins, trucks and shovels, to study mining operations on a minute-by-minute basis. From the orebody description and mining selectivity, layers of ore and reject are determined and mined by the dragline. Reject is dumped on the pit floor; ore is either dumped on a windrow for later removal by a bucket-wheel reclaimer or directly dumped into a hopper feeding a conveyor. The model allows for the presence of pads as a dragline base, key cuts and a number of interruptions to the mining operation such as shift changes, boom checks, engineering services, repairs, delays, preventative maintenance, etc. A CRT display keeps the model operator informed about equipment status and performance. The model operator can direct the course of the simulation by issuing commands to the model at any time during the simulation.IntroductionThis paper describes an interactive computer model of a dragline, windrow and spoil pile. The model was designed by the author based on information and specifications provided by Canstar Oil Sands Ltd. The model design was implemented and tested on Petro-Canda's IBM computer system by DMR and Associates Ltd. of Calgary, Alberta.The model was designed to form three components of an oil sands mining operations model, described in reference I. Figure 1 shows the interaction of those components with the remainder of the oil sands mining model.An ""Ore/Waste Body"" data base provides information to the dragline model about the location and composition of or e and waste in the mine. The dragline model is used to simulate mining this ore/waste body, to build up spoil piles for center reject (waste) and to dump ore onto windrows or hoppers (situated over conveyors) to deliver the ore to the extraction plant. The windrow model component provides a data base which will be used by a model of a bucket-wheel reclaimer which simulates the loading of ore from the windrow on to a conveyor via a bridge."

Jan 1, 1982

By A. J. Keen

"Cominco's Polaris Mine is an underground lead/zinc operation located on Little Cornwallis Island in the Canadian High Arctic. Polaris is the world's most northerly base metal mine, situated at about latitude 75°N and longitude 97°W (Fig. 1). The closest community is Resolute Bay about 100 km to the southeast.During the construction and development phase Cominco adopted a unique scheme to save time and money. Many facilities including concentrator, powerhouse, shops and warehouse were built on a barge in southern Canada. The barge was towed 4800 km to the site in August 1981. Production at Polaris began in November 1981. The current mining rate is 1 million tonnes per year. The mine employs 250 people with about 190 on site at any time.GeologyThe deposit is a Mississippi Valley type orebody and lies near the top of the Ordovician Thumb Mountain Formation. Ore minerals consist of sphalerite and galena with the gangue being predominantly dolomite with calcite and marcasite. The orebody is situated completely within permafrost, consequently the voids frequently found in the ore are filled with ice. The orebody strikes north-south, dips approximately twenty degrees to the east and takes up an area about 400 m wide by 700 m long. The upper portion of the orebody (60 m to 150 m below surface) lies in an area called the Panhandle Zone. The deeper portion of the orebody known as the Keel Zone extends to 300 m below surface (Fig. 2). The deposit is strongly zoned with a high grade zinc/lead and iron-rich tabular core overlain and underlain by vein, stringer and breccia mineralization. The ore reserves at January 1991 were 14.6 million tonnes at 3.6% Pb, 13.7% Zn, 3.2% Fe, after producing 8 million tonnes at 3.9% Pb and 14.2% Zn since November 1981"

Jan 1, 1992

By Scott K. Palm, Neil D. Pearson, Jr. Read

"Uncertainty is an important feature of many investments, but it is especially important for mineral properties. Mine properties often entail large investments and fixed commitments. Managers in the mining industry often view price uncertainty as undesirable, due to the possibility that the price of the mined commodity will fall to a level that prevents the firm from recovering its investment and fixed costs. This paper identifies pitfalls in conventional valuation methods and outlines a new approach to valuation based on option pricing theory and dynamic programming. In the process, it identifies generic strategies for organizing mine production in ways that allow the owner to benefit from the uncertainty in mineral prices. IntroductionRisk is pervasive in the mining industry. The operation and maintenance of a mine may involve high, and unavoidable, fixed costs. A mine owner can only begin to recover these fixed costs if the price of the mined commodity is greater than the incremental cost of production. But the prices of many minerals are extremely volatile. The price volatility of copper has been significantly higher than the price volatility of a typical stock, for example. A mine owner cannot be sure that the price of the mineral product will remain high long enough for the firm to recover its capital and fixed costs.Mine owners are able to minimize their losses during periods when the output price is below the incremental cost of production by temporarily closing operations. During these periods they incur only fixed costs. While shut-downs help minimize losses, the y do not help the firm recover its initial investment. If output prices remain low long enough, the fixed costs and original investment may never be recovered. High fixed costs, therefore, lead to substantial downside risk."

Jan 1, 1986

By S. Srinivasan, R. Coté

"India is undergoing a massive electrification program using local coal to provide fuel for thermal generating plants. The Rajmahal mine in northern Bihar is scheduled to produce coal at the rate of 10 million tonnes per year by the end ofa sixty six month contract with MET-CHEM Canada Inc. for mine development. In order to control maintenance performance and warehouse inventory MET-CHEM is installing some proven computer-controlled programs and will provide extensive training to ensure that the systems are properly used. The programs are Maintenance Planning and Control System (MPCS) and Materials Management System (MMS), both of which were developed and used by Quebec Cartier Mining Company at its iron ore mine in northern Quebec. In addition, a Predictive Maintenance Program (PMP) using vibration analysis will be introduced . This paper discusses the three systems.IntroductionIndia is currently undergoing a massive countrywide industrialization and electrification program. Most of the power will be supplied by coal-fired thermal electric plants.One of the mines being developed as part of the over-all program is the Rajmahal Open Cast Coal project in the Godda District of Bihar, approximately 270 km north , and slightly west of Calcutta.The project is based on an agreement signed in January, 1989 between two government organizations - Canadian Commercial Corporation (Ccq and Coal India Limited (CIL). MET-CHEM Canada Inc. (MET-CHEM), an international consulting engineering organization which has worked in India since 1976, is CCCs Canadian executing agency for the project. METCHEM was the engineer-constructor and mining associate for the development of the 7.5 mtpy Kudremukh Iron Ore Project. This US$430 million project was completed on schedule and under the original budget estimate. The Rajmahal deposit is the responsibility of Eastern Coalfields Limited (ECL), a subsidiary of CIL. The project is to produce coal at the rate of 10.5 million tonnes per year in 1994."

Jan 1, 1991

By James D. Bliss, David M. Sutphin

"On the basis of differences derived from genetic, descriptive, and grade-tonnage data, graphite deposits are classified here into three deposit types: disseminated flake, amorphous (microcrystalline), or graphite veins. Descriptive models have been constructed for each of these deposit types, and grade-tonnage models are constructed for disseminated flake and amorphous deposit types. Grade and tonnage data are used also to construct grade-tonnage models that assist in predicting the size and grade of undiscovered graphite deposits. The median tonnage and carbon grade of disseminated flake deposits are 240 000 tonnes and 9% carbon and for amorphous deposits, 130 000 tonnes and 40% carbon. The differences in grade between disseminated flake and amorphous deposit types are statistically significant, whereas the differences in amount of contained carbon are not .IntroductionIndustrial minerals, including mineral construction materials such as sand, gravel, and crushed stone, have historically exceeded metals in global production value. Today, industrial minerals are being used in new and sophisticated ways in response to the late 1980s emerging information- and service-based economies' requirements for less metal and more industrial minerals . Natural graphite is an example of one such mine ral that has been known since ancient times, and is now finding new applications. Graphite was first used as an industrial mineral in Bavaria in 1400 A.D. where it was an important component of crucibles. Traditionally, natural graphite has been used in heavy industries. New applications, such as development of carbon-magnesite bricks a decade ago and the current advancement of alumina/graphite refractories, are increasing its graphite use. Other new uses for natural graphite, such as the use of graphite powder as a thermal blanket for the continuous casting o f steel (Northern Miner, 1987) and the addition of expanded graphite to polyurethane foam where it acts as a flame retardant (Industrial Minerals, 1987), are increasing the demand for this mineral. To meet future demand for graphite and other industrial minerals, knowledge of the characteristics of the deposit types in which they are found is important. A further look at the deposit types for graphite and at construction of descriptive and grade-tonnage models is warranted .The graphite industry recognizes three basic types of graphite , distinguished primarily on the basis of grain size: crystalline flake, lump and chip, and amorphous. These naturally occurring forms of graphite are products of metamorphism and are found in schists of various types, gneisses, and marbles. Temperatures of at [east 400°C are thought to be required for graphite formation (Landis, 1971). Synthetic graphite, also known as artificial or manufactured graphite, is a man-made carbon product and offers higher purity and lower crystallinity than natural graphite. Several authors have written about graphite's unique physical and chemical properties and the uses for individual types of graphite (Krauss et 01., 1988; Graffin, 1983; and Harben and Bate s, 1984)."

Jan 1, 1990

By Zygmunt Wegrzyk

Poland is rich in hard and brown coal reserves, which have become the most important, and steadily developing, primary source of energy. With 192.6 mill. t of hard coal production and 40 mill. t of exports in 1978, Poland is fourth and second in the world respectively. The paper discusses the various mining methods adopted in the country's history of mining. Particular emphasis is placed on longwails, their mechanization and equipment assembled in the form of fully mechanized face complexes suited for varied geological conditions. Measures to improve the recovery ratio of coal resources are dealt with. investment efforts undertaken in the coal industry are outlined. An account is given of the ever improving techno-economic results of coal mining. The country also exports a wide range of mining equipment and transfers its well-founded technical know-how to others. There are strong cooperative links with many foreign partners.

Jan 1, 1979

By John De La Vergne, Morri Medd

"This paper describes the sinking of a circular concrete ventilation shaft through 110feet of water-bearing overburden. The brine circulation freezing method was used to stabilize the walls during excavation. In order to take advantage of the cold winter weather, a custom-built heat exchanger (natural radiator) was installed in place of the standard mechanical refrigeration plant.IntroductionIn the late fall of 1979, Les Mines Selbaie awarded a contract to the firm of 1.S. Redpath Limited to sink a ventilation shaft on its property in northwestern Quebec. The shaft was to be sunk through 110 feet of water-bearing overburden and approximately 90 feet of rock to a final depth of 200 feet below surface. The water table was measured at 16 feet below the surface elevation.After examining the soils report and diamond drill core from the shaft area, Redpath came to the conclusion that freezing was the best way to stabilize the overburden for sinking. LocationThe mine site is 420 statute miles due north of Toronto, just east of the Ontario/Quebec border (Fig. 1). Access to the site is provided by a development road built by the Quebec government from the Matagami/lames Bay Highway. History of Ground Freezing In the Mining IndustryIn 1862, shortly after the development of mechanical refrigeration, a mine shaft was sunk through an artificially frozen water-bearing formation in the United Kingdom. The first application of ground freezing in North America was for a Michigan iron ore shaft in 1888. Since that time, the freezing method has been used to sink hundreds of shafts throughout the world. Its best known application was to sink sixteen shafts through the Blairmore Formation in Saskatchewan.The same method has been used in every case. It involves chilling a calcium chloride brine solution and circulating the brine in concentric pipes, spaced between 3 and 6 feet around the shaft perimeter and usually 4 to 5 feet from the limits of the proposed shaft excavation."

Jan 1, 1982