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By Ward E. Kilby, Bruce I. Mcclymont

"Data from the South Fork Pit Area of the Bullmoose deposit in northeastern British Columbia are used to demonstrate the usefulness of microcomputers to the on-site geologist. The technique utilizes a series of digital surf aces in the form of a network of points to define various geologic parameters such as seam positions, topography and coal quality characteristics. Manual digitizing, kriging and combinations of these techniques are used to calculate these digital surfaces, the best method depending upon the type and quality of the data in question. Digital surfaces, each modelling a specific feature when combined and manipulated with the computer form a powerful tool, the deposit model. The resultant model, containing all the surfaces, is used to examine seam and inter-seam thicknesses, deposit volumetrics, seam quality characteristics and the adequacy of sampling.IntroductionCoal deposit modelling using computer-based techniques has become routine. However, a geologist working in the field, without access to mainframe or minicomputer, ha s considerable difficulty implementing these procedures. However, microcomputers, which are portable and inexpensive can handle exploration data, perform various geological analyses, and produce deposit models virtually at the same time as the data is being collected. For the on-site geologist this means that quick access to the results of such modelling could point out inadequacies in deposit definition and trends in coal quality deserving investigation."

Jan 1, 1985

By J. M. D. Wilson, W. Petruk, C. Coté

"The behaviour of minerals in the Mount Wright spiral circuit was investigated to find methods of improving iron oxide mineral liberation and recovery. The feeds, concentrates and tailings from rougher, cleaner and recleaner spirals in one mill line were sampled before and after the sizing screen openings were reduced from 2 mm to 1.6 mm. Sized fractions of each sample were analyzed with the MP-SEM-IPS image analysis system at CANMET to determine mineral quantities and liberation characteristics of the iron oxide minerals. Mass balance techniques were used to calculate stream recoveries of iron, silica, total iron oxide minerals, hema¬tite, goethite, quartz, ferromagnesian silicates, and liberated and locked total iron oxide minerals for each sieve fraction.The reduction of the sizing screen openings increased the ap¬parent liberation of the total iron oxide minerals in the screen undersize from 89% to 94% and raised the over-all iron recovery from 84.4% to 85.3%. Apparent liberation of iron oxide minerals in the cleaner sand also increased from 70% to 85%, but the – 75 µm fraction re¬mained the only fraction containing an acceptable amount of liber¬ated iron oxide minerals in each case. Due to limitations of spiral and grinding efficiency it is recommended that the cleaner sand be screened at 300 m, the +300 m fraction be reground, and the – 300 p.m fraction be rerouted to the rougher spirals.The amount of fine liberated iron oxide minerals (– 106 m) lost in the tailings was 46 TPH in mill line 3. This value represents 41 wt % of the free fine iron oxide minerals, or 7 wt % of the total iron oxide minerals in the spiral feed. Further work is needed to investigate the possibility of recovering this material by other methods.Middling iron particles (010 50%) and presumably free gangue in the size range –300+ 75 m tend to be recovered by spirals by entrapment."

Jan 1, 1990

By Francis Breton, Lei Pan, Kun Liu

"Effects of Fe, Cu and Mg minor alloying elements on creep properties of 8xxx aluminum conductor alloys were investigated by conducting compressive creep tests at 100 ºC under a constant load of 69 MPa. The microstructural evolution was studied using an optical microscope and electron backscatter diffraction technique. The creep properties were characterized by the primary creep strain in the first stage and the minimum creep rate in the second stage of creep deformation. Results showed that additions of all three minor alloying elements (Fe, Cu and Mg) demonstrated a beneficial effect on the creep resistance but individual alloying elements exhibited different impacts on creep properties. The additions of Fe and Mg showed a strong impact on both the primary creep strain and minimum creep rate, while the Cu addition displayed a favorable effect on the primary creep strain but no impact on the minimum creep rate. Microstructure observation revealed that with increasing Fe contents, the volume fraction of fine intermetallic dispersoids gradually increased while the subgrain size and misorientation angle became smaller, providing strong barriers to the dislocation and grain boundary movements. The addition of Cu moderately reduced the subgrain size and misorientation angle and in combination of solid solution strengthening it promoted the retardation of the creep deformation during the primary stage.INTRODUCTIONThe 8xxx aluminum conductor alloys offer significant advantages such as low density, high conductivity to weight ratio and low price over copper conductors and become an attractive candidate in electrical engineering and building wiring (McQueen, Chia, & Starke, 1986; Birol, 2009; Engler, Laptyeva, & Wang, 2013). The mass resistivity of aluminum conductors is one-half of copper conductors. Thus, only one-half the weight of aluminum is required for equivalent capacities. Therefore, this unique property coupled with a relatively low price of aluminum offers significant economic advantage to develop aluminum conductor alloys for electrical applications in the electrical wire industry."

Jan 1, 2016

By Ferrin Prince

New Gold?s New Afton mine is an 11,000 tonne per day block caving mine, located 8 km west of the city of Kamloops, British Columbia. The reserve is estimated at 52.5Mt with grades of 0.65 g/t Au and 0.93% Cu. Block caving requires a considerable amount of infrastructure development before production can begin. Each excavation had its own challenges, often further complicated by interactions with the caving stresses and fault structures. The main areas of interest were: the production crusher complex, the surface to underground infrastructure, and the footprint interaction with the faults. The ore zone at New Afton is bounded by two very weak fault zones, with large fault affected areas and associated splays. The resulting rock mass rating (RMR) is 35-50, with some areas significantly weaker due to clay infill. These conditions created a unique and decidedly difficult environment in which to develop New Afton?s infrastructure. Despite these difficulties, the mine began production on schedule in July 2012. The main production crusher for the mine is a gyratory crusher that allows for direct truck dumping, with a smaller jaw crusher in series to facilitate the initial development and production. The crusher chamber was redesigned several times to minimize the chamber volume, resulting in a 40% decrease in volume without losing functionality. After drilling and geotechnical testing, the location of the crusher was changed to eliminate the threat of fault affected ground in the area; the entire complex moved a short time before mining began on the chamber. The mining techniques used enabled tight control of the crusher pocket excavation, while also allowing for the timely completion of the chamber. The locations of the surface to underground infrastructure were tightly constrained by geology, surface infrastructure, and an existing open pit. The VR5 raise was drilled and abandoned part way through the reaming process when the drill string became stuck in fault gouge material. It was relocated and successfully completed. The VR7 raise hit a region of poor rock strength and deviated more than 11 m. Poor ground conditions caused a 130m section of the raise walls to fail. An innovative rehabilitation program was undertaken to re-establish the raise for ventilation. A surface to underground slick line was also installed to ease the material movement underground. A layout of the Extraction level was changed from herring bone to straight through drawpoints that resulted in a net reduction of 1.3 km of development. The Apex level was added to aid in the breakage over the major apex pillar and allow for geological mapping. A portion of the caving footprint is located within the footwall fault. The low strength, high clay content, and small joint spacing of the material resulted in blast-hole cut-off and drift deformation. Drift rehabilitation was a key factor in the successful development of the cave footprint. Innovative drilling and loading techniques were used to keep the holes open and to impede movement along the joint structures.

Aug 1, 2013

By D. M. Cruden

"When two discontinuities form a tetrahedral wedge in a rock slope, a factor of safety for the wedge can be simply calculated (if motion is driven only by gravity and resisted only by friction) from plots of the normals to discontinuities on an overlay of a polar, equal-area spherical projection to which the great circles of a similar equatorial projection have been added.IntroductionThe stability of slopes in rock masses is often controlled by the discontinuities that exist in the masses. In particular, the geometry of the discontinuities constrains the ways in which the mass may fail. Displaying the orientation of the discontinuities on a spherical projection has been found to be a useful way of identifying possible failure modes. The most concise display shows the orientation of the normals (or poles) to the discontinuities that compose the fabric of the rock mass. Computer processing of large samples of discontinuity orientations is necessary if the sampling biases identified by Terzaghi (1965) are to be routinely corrected. Fortunately, displays of the poles to discontinuities (or pi diagrams) are relatively easily produced on the line printer of a computer or by a computer plotter (Cruden , 1978; Ramsden and Cruden, 1979). Unfortunately, further analysis of the stability of a rock mass has tended to use traces of typical discontinuities on a spherical projection rather than pi diagrams; examples may be seen in Hoek and Bray's book (1981)."

Jan 1, 1990

By David C. Mulder

As a result of the dramatic growth of Edmonton, Alberta during the past decade and the rapidly increasing price of natural gas, Edmonton Power has received provincial government approval to construct and operate an 8O0-megawatt- capacity coal-fired thermal generating station at Genesee, about 50 km southwest of Edmonton. The power plant will consist of two 400-megawatt steam-driven thermal generators, with provision in the future for two or more additional units. The coal will be hauled by truck from the nearby three-million-tonnes- per-year strip mine, with economically recoverable reserves exceeding two hundred million tonnes. Edmonton Power controls a major part of the coal leases in the area, and is working with other lease holders in developing the most efficient mining plan. A natural depression was selected as the site for the 600-hectare cooling pond to provide water for the power plant. Edmonton Power is working closely with the people in the area to ensure that their rights and concerns are communicated and recognized by all parties.

Jan 1, 1981

By C. Beaudry

Outline ? General Geology of Frieda River Project Area ? Exploration History ? Geology of Porphyry Deposits ? Porphyry Resource ? Update of Nena Deposit and High-Sulfidation Mineralization ? Nena Resource

May 1, 2006

By O. E. Humeniuk, A. I. A. Salama, M. W. Mikhail

"Performance testing of prepartion plants is referred to as plant audit. It evaluates plant performance under normal operating conditions and is aimed at improving yield through knowledge of actual process performance and interactions between different processes in a plant.This paper describes the program to evaluate the performance of several Canadian preparation plants. The performance test requires the planning of a sampling campaign that would include all circuits in the plant. Samples are analyzed to determine washing characteristics (screen and float-sink), ash and per cent solids of feed and products of separation units in the plant. Performance characteristics are established based on analysis and material balance. Specific recommendations can then be made to improve recovery, and identify and explain potential causes of apparent deficiencies.The objectives of the program are to reduce the cost of production by improving recovery of clean coal and, in tum, to reduce disposed tailings, establish baseline information on coal preparation plants performance in Canada, and define R&D work needed to optimize plant operation."

Jan 1, 1993

By G. W. Poling, J. R. Hoffert

"A common practice in preventing gravity-fed tailing pipeline from plugging due to sanding is to add lime at relatively high dosage rates. The lime affects the rheology of the slurry ""vehicle"" which transports the coarser particles in a heterogeneous flow-regime. This paper describes a study of the effect of lime on the critical deposition velocities and friction head losses of Highmont Mining Company's mill tailing. Testwork using an extrusion viscometer was related to the surface chemistry changes brought about by the high lime additions. Lime additions which typically increased the tailing pH from 8.4 to 11.5 reduced the critical deposition velocities from approximately 3.2 to 1.7 m/sec.IntroductionCommon practice, at least in base metal concentrators in British Columbia, is to use high lime additions in tailing slurries when the pipeline becomes partially plugged with sands. Some concentrators, such as the Highmont Mining Co. Ltd. plant, add lime on a continuous basis during low tonnage periods and during startup/shutdown transition s. Although such high lime additions often solve temporary plugging problems, the effect of lime on slurry rheology has not been well understood.Slurry pipelines are designed to transport solids in suspension with a minimum of wear and without particle deposition. These two objectives are in opposition because the velocities required to maintain solid suspension can sometimes cause high head losses in the pipeline, as well as wear problems. Thus, the optimum velocity for a long distance slurry pipeline is in the range that provides just enough turbulence to keep coarse particles in suspension."

Jan 1, 1985

By Rod Maclean, George Stewart

"This paper covers the planning, development, and mining of an underground pillar complex using a bulk mining method. To achieve the long-term goal of continually improving our operating processes, all cut-and-fill and undercut-and-fill workplaces at Thompson Mine have been evaluated for their conversion potential to the bulk method referred to as VerticalBlock Mining (V.B.M.).Since December 1979, conversion has occurred to the extent where 80% of the ore hoisted is mined using Vertical Block Mining (V.B.M.).Productivity and profitability have been substantially improved by converting to this bulk method, especially in pillar areas.IntroductionThompson Mine is a 6000 ton per day underground nickel mine located 3 miles outside the City of Thompson, 400 air miles north of Winnipeg . The mine has two shafts, T-l and T-3 (Fig. 1). The T-l shaft is 4427 feet deep with the deepest mining taking place on 4000 level. The T-3 shaft, which is the area of this presentation, extends to a depth of 2607 ft with the deepest operating level at 2400 ft. The two shafts, some 10 000 ft apart, are connected underground on numerous levels. The ore mined at T-3 is trammed on 2400 level to the T-I main ore pass and hoisted through the T-l shaft. Eight hundred and fifty thousand cubic feet of fresh air per minute ventilates the mine.The 1400 level 310 pillar complex is located in the T-3 mine area. There are six rib pillars, two crown pillars and one main pillar (boundary pillar between cut-and-fill stopes) within the complex (Fig. 2). The majority of the ore in this area has been mined using mechanized cut-and-fill techniques .Undercut-and-fill mining was the method originally planned to recover these pillars. It is slow, labour-intensive, and requires extensive use of timber during the mining cycle. The extraction drifts in the area are heavily timbered due to their proximity to the ore and the age of the facilities. Increasing drift maintenance costs and the successful mining of two main pillars with V.8.M. techniques provided the stimulus to review the mining plans for this area."

Jan 1, 1988

By Garth A. Maki, Douglas M. Smith

"The largest Canadian Lamella thickener installation to date, with 13,600 square feet of effective clarification area, is being successfully operated at the Rocanville Division of the Potash Corporation of Saskatchewan. This paper describes the operational principles and design features of the Lamella thickener, the testwork leading to its specification f or the Rocanville phase two expansion project and a review of the performance of the full- size unit which was ultimately installed.IntroductionThickening equipment operating in the Saskatchewan potash industry is generally situated within the main refinery building or a separate insulated an d he a ted structure due to the harsh winter climate and the need to closely control plant brine temperatures to en sure good flotation circuit performance Because these thickeners process large volumes of slurries, conventional basin-type designs tend to be large, ranging up to 46 metres in diameter or more. As part of the general effort to minimize capital costs, the potash industry has begun investigating the use of alternate high capacity, compact thickener designs. In this regard, PCS Rocanville undertook a program in 1978 to evaluate a LamellaTM thickener and one other type of ‘High Rate' thickener as alternatives to a conventional design for the phase two expansion project.This paper describes the operational principles and design features o f the Lamella thickener (manufactured by Axel Johnson lnc.), the testwork leading to the specification of a Lamella for the Rocanville expansion project and are view of the performance of the full size unit (1486 m2 of the effective settling area) which was ultimately installed and is now successfully operating in the Rocanville plant."

Jan 1, 1983

By P. J. Urso

"This paper discusses the mining problems encountered at the Fording River coal mine to date, and how they have been solved. All of the production now comes from surface mining, but underground hydraulic mining is planned and briefly described. Other facets of the operation, including markets, equipment and personnel, are covered. IntroductionFording Coal Limited was formed in 1968 to mine and market metallurgical coal resources held by the company in the East Kootenay region of southeastern British Columbia. The company, at its Fording River operations, holds approximately 48,000 acres of coal lands, having an estimated 2.0 to 3.0 billion long tons of metallurgical coal in place. (Although not a point for further discuss ion in this paper, the company also holds approximately 1.0 million acres of mineral rights. These consist primarily of coal in Alberta, potash and coal in Saskatchewan, and minor industrial mineral holdings of limestone, salt and bentonite in the three western provinces.)Fording Coal is owned 60% by Canadian Pacific Investments Ltd. and 40% by Cominco Ltd. Inasmuch as Cominco itself is owned approximately 54% by CPI, it is seen that the effective CPI ownership of Fording is about 82%. The company, originally staffed and managed by Cominco, has been independent since January, 1978, with its own board of directors, management, staff and employees.The company's head office is in Calgary, Alberta. The Fording River operations are located approximately 45 miles north of the Crowsnest Pass, near the Alberta/British Columbia boundary, only 95 air miles and 200 road miles from Calgary.A contract was signed in 1969 with the Japanese steel mills to provide 45 million long tons of metallurgical coal over IS years, commencing in 1972 at an annual rate of 3 million long tons per year. Despite some start- up difficulties, the mine and plant facilities achieved their rated capacities in 1975, producing and shipping at the 3-million-Iong-ton level. The coal specifications are shown in Table I."

Jan 1, 1980

By K. B. Chaudhuri, I. D. Webb, F. E. Rachor, F. A. Sauert

"A characteristic of all methods of copper concentrate smelting is that slags with a high copper content are produced at some phase of the operation. Recirculation to the smelting unit, electric furnace reduction or flotation are at present used on an industrial scale to recover the copper from these slags. This paper describes the application of the top-blowing technique, originally developed by KHD Humboldt- Wedag AG for refining of metallic copper, for the treatment of copper-containing slags. By this method, a mixture of a reducing gas and commercial oxygen is blown onto the melt at velocities which can be supersonic from the tip of a water cooled lance. The impingement of the reaction gas mixture on the surface of the slag forms a depression in the molten phase and a deflection of the jet which brings the slag into a rotating motion. This causes high rates of heat and mass transfer. Results of pilot-plant tests for batch and continuous treatment of slags are presented, and the potential applications of top blowing for industrial-scale operation are indicated. In combination with the cyclone as a smelting unit, this process appears to be economic for smelters of very low capacity."

Jan 1, 1980

By G. N. Nnadi, R. J. Mitchell

"The response of cemented backfills subjected to simulated blast loadings is studied using centrifugal modelling techniques. Scaling requirements for centrifugal modelling are developed and the results of the similitude analysis show that an acceleration field is required to capture the true dynamic response of backfill models. Experiments have been conducted at various centrifugal accelerations. The effects of ring blasting were simulated by an impact loading on the exposed fill face. The behaviour of the cemented backfill was monitored with accelerometers as well as dynamic stress gauges.Cemented backfill response to impact loadings was found to depend strongly on simulated gravitational stress level. Results also show that the characteristics of the response changes with distance from the impact. Finally, the results of the models were used to predict prototype peak accelerations, velocities and displacements. The results reported in this study indicate that centrifuge modelling may be a useful tool for dynamic studies of blast damage to backfills. Centrifuge modelling can be carried out at low costs and can be used to design prototype instrumentation systems. IntroductionThe response of cemented mine backfills to blast loadings is of concern to the mining industry because these backfills provide local and regional support during ore pillar extraction. Considerable research has been reported on the static stability (Smith et al., 1983) and on effects of reinforcements on cemented backfills (Mitchell, 1989). However, information regarding dynamic aspects of cemented backfill is limited. O'Hearn and Swan (1989) reported on the response of cemented backfill to blast loading, but these blasts were not very dose to the backfill . A schematic representation of a rock-backfill system is shown in Figure I.The mechanics of transfer of explosive energy to rock and subsequently to the backfill, and the resulting responses, are complex. Achison (1968) listed more than twenty factor s affecting the blast fragmentation of rock and grouped these into the categories of explosive parameters, charge parameters and rock parameters. Although significant advances have been made, during the past two decades, in understanding the mechanics of blasting, the degree of uncertainty associated with predicting the behaviour of backfills subjected to dynamic loadings makes the application of simple design safety factors insufficient. Comprehensive testing programs or prototype instrumentation is required to evaluate the risks associated with blast or earthquake loadings on backfill."

Jan 1, 1991

By G. A. Kosick, M. Freberg, L. A. Kuehn

"Highly variable and higher than forecast lead miff feed grades were problems experienced at the Polaris Concentrator. Limited flotation capacity and poor operating flexibility prevented the plant operators from overcoming these difficulties, and resulted in grades and recoveries which were lower than necessary.Two column cells were integrated into the lead circuit in an effort to give the operators the tools they required to achieve the best grades and recoveries possible, regardless of the miff feed grade. Installation of the column cells increased lead concentrate grade from 76% to 79.3% Pb, and increased lead recovery from 9/% to 91.3% Pb.This paper reviews the design, development and operation of column flotation cells in the lead circuit at the Polaris Mine.IntroductionThe Polaris Concentrator was constructed in 1980 on a barge in Trois Rivieres, Quebec, and towed to its pre sent location on Little Cornwallis Island, Northwest Territories, about 1450 km from the North Pole. The plant presently treat s about 2800 mtpd of new feed with average feed grades of 3010 Pb, 13% Zn and 5% Fe. Current production averages 36 000 tonnes of lead concentrate and 220 000 tonnes of zinc concentrate annually.The barge design and min e location dictated that the concentrator .be compact with very little unused floor space. As a result, flexibility was limited in terms of expanding the plant.Mill feed grades were highly variable and above forecast and the lead flotation circuit was found to be under capacity. This caused continued overloads in the lead cleaning circuit, resulting in low concentrate grades and recoveries. It was evident that an increase in lead flotation capacity and flexibility was necessary. The existing floor space did not allow for the installation of additional conventional flotation cells, but one alternative was the installation of column cells."

Jan 1, 1988

By Robert J-M Ethier

"Sequential blasting has been introduced at Mines Gaspe to reduce the harmful effects associated with the instantaneously primed blasting method. Three stages of modifications had been tried to improve this method: trials of a rectangular pattern instead of the square one, trials of lengthened hole collars at the perimeter of the blast and trials of increased inter-row delay intervals from 35 to 50 ms. Results showed that improvements to the instantaneously primed blasting method were limited by the probability of increasing the number of bootlegs.Doing these modifications allowed the definition of a general approach to improvements before the sequential blasting method was introduced. This general approach included: planning of the mining sequence, following up on the drilling, loading and tying-in of each blast, compiling and evaluating blast results, comparing results from one bench to the other and moving along with one change at a time. This approach has permitted the successful introduction of the sequential blasting method in two steps: trials of blast by rows of delayed holes and trials of hole-by-hole blasts. The optimum result has been analyzed by high-speed camera and by computer blast simulations, The actual optimum sequential blast is obtained with a CORDLINE priming system using period 15, permitted in a large drilling pattern, with an intersequences delay interval of 230 ms and an inner-sequences delay interval of 50 ms.While these modifications were being introduced, a careful study of the drilling and blasting costs was made, which showed a sharp reduction of drilling cost per tonne. Even though there was an increased use of packaged slurry the blasting cost per tonne also dropped. The reduction in cost per tonne was explained by the fact that the average energy factor was maintained while the blasting pattern was expanded.The sequential blasting method, successfully introduced at Mines Gaspe, permits blasts of better quality, reduces the drilling requirements by more than 30% and helps to achieve improved productivity while minimizing costs."

Jan 1, 1984

By D. Marotte, M. David

"A new exploration procedure is proposed for volcanogenic massive sulphide deposits of the Abitibi belt. ""Samples"" of acidic volcanic rocks (Si02 > 60%) were compiled from the literature. From the 574 samples collected, 322 were used in each of a number of experiments. Forward inclusion linear discriminant analysis performed on the untransformed geochemical data shows a general depletion in Na20 and CaO along with an enrichment in MgO for a ""proximal"" group of samples relative to a ""distal"" one. The percentage of misclassified samples reaches 20% to 30%, but test sets shows good stability.The exploration procedure was applied to the Normetal deposit, where the authors took 160 samples during the summer of 1977. It quickly leads to the definition of targets for more detailed exploration together with a good recognition of the old known deposit.IntroductionVolcanogenic massive sulphide deposits in the Abitibi belt (Quebec) are mainly found in rhyolites or acidic pyroclastics in the uppermost portion of a calc-alkaline volcanic series (Descarreaux, 1972). The deposits are thought to have formed syngenetically by the action of volcanic hydrothermal exhalations during the waning episode of a felsic volcanic cycle (De Rosen Spence, 1976). Alteration pipes of disseminated sulphides are commonly present. Metal deposition from hot springs, initiating convection cells in a submarine environment, may have been the dominant process in ore formation (De Rosen Spence, 1976)."

Jan 1, 1981

By G. R. Davidge

"Falconbridge Mine has a history of significant microseismic activity dating back to 1955. Since mid-1981, a microseismic monitoring system has been operating in the area around the No .9 Shaft hoist room located 25 m from the are zone at a depth of 1200m. Several mine wide faults dissect this area and have been shown to be microseismic-active due to mining. The microseismic monitor has been interfaced with a small dedicated personal computer and a communications device to perform routine data evaluation and to produce graphical output. The programs are described and selected result s from the monitoring program are presented.IntroductionFalconbridge Mine is located on the southeast rim of the Sudbury Basin. The nickel-copper sulphide ore zone is essentially a tabular orebody striking approximately east-west and dipping at 80 degrees north near surface and rolling to 75 degrees to 80 degree s south by the 1000 level. (Mine level number s represent the approximate depth in feet below ground surface .) This dip is maintained down to the 6000 level. The north wall is comprised of jointed norite and a faulted contact up to 1 m thick exists between the ore zone and the norite. The south wall rock is a greenstone with highly siliceous lenses and often has a faulted contact with the ore zone. No less than five mine wide faults exist, as shown in Fig. I.Mining ha s been carried out from virtually the bedrock contact to depths of 1830 m although the majority of current activity is between 900 and 1500 m below surface. A winze provides access below the 4025 level and the cage hoist room is located on the 3850 level in the north wall 25 m from the contact. Because of concern for the stability of this hoist room, there has been no mining activity within 60 m of the hoist room and by 1980 this block of ore was one of the major blocks of high-grade ore remaining in the mine . Mining of this block was originally scheduled for very late in the decade.Prior to a six-month shutdown in 1982, this mine produced approximately 450,000 tonnes of ore annually, mainly by cut and- fill techniques. Following the shutdown, annual production has been reduced to approximately 200,000 tonnes and production is now mainly by undercut and fill techniques .Mining-induced seismicity has occurred at Falconbridge Mine since at least 1955 and reviews of the microseismic problems for the periods from 1965 - 1970 and from 1971 to 1979 were made by Singh (1971) and Vongpaisal et al. (1980), respectively. The need for instrumentation was addressed by these authors. Following several large events during the spring of 1980, a decision was made to investigate existing microseismic monitors and to apply micro seismic monitoring techniques in Falconbridge Mine. Application of microseismic monitoring to Falconbridge Mine was to concentrate on the No . 9 Shaft hoist room area of the mine, particularly around the 4025 level where a large mechanized cut-and- fill stope was approaching this level. By mid - 1981, the microseismic monitor was fully installed and with in a year , a short-period seismograph and a personal computer had been added to the system."

Jan 1, 1984

By S. Maleksaeedi

On-site evaluation of strength using hardness is common for different materials. The lack of such relation for aluminum alloys has always been felt. In this work, a finite element analysis of plastic deformation and its distribution in Brinell hardness testing is presented. The analysis has been performed for different strength levels of aluminum alloys. By utilizing the Von Mises equivalent strain in hardness, it is tried to derive a correlation between hardness and strength values for the alloys under consideration. The derived correlation is verified using the experimental values available in literature.

Jan 1, 2004

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