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By M. Amsden

"Since the computer program is far from complete this must be only a brief progress report on the computerization of flotation metallurgy at the Ecstall concentrator at Timmins.Ecstall is currently treating about 10,000 tons per day of two different ore types. One is a low sulphide copper-zinc ore, the other is a high sulphide lead-zinc-silver ore. The concentrator is divided into three individual grinding and flotation circuits. In effect there are three mills under one roof. Normally there are two circuits on copper-zinc ore and oneone lead-zinc-silver ore. At this point in time one of the copper-zinc circuits is under computerized metallurgy control.The first step in this control is the use of an ARL on-stream x-ray analyzer, coupled with a Honeywell 21 computer which types out assays and recoveries for each individual circuit. Our approach to a computer program was to keep it simple. The program was written to coincide as closely as possible to what an operator would do in the performance of his duties based on the in-formation he was getting from the x-ray analyser read-out. Initially, computer control was put on just the copper circuit of the ""A"" Division, controlling the collector addition at three points in the circuit; primary, secondary, and scavengers. The control was based on recoveries and the grade of the copper concentrate. Targets like 95% recovery and 24% copper concentrate grades that it was felt the circuit could achieve were set to start with. It quickly became apparent that this alone was not enough because if on a certain ore the best recovery that could be made was only 94.9% the computer kept calling for more reagents until the circuit was completely out of control. On another ore it might have been quite possible to get 96% recovery but when the set point of 95% was reached the computer stopped trying for any more. To overcome this problem a very simple optimum seeking program was devised by one of the Ecstall staff and it has worked very satisfactorily since being put into use in July 1970.In the copper float only the two collectors R-208 and the xanthate R-317 are under computer control. The addition of frother and depressant is left to the operator. The flotation cells have automatic level controls which can be adjusted by the operator. pH is controlled automatically through a lime loop-sensor arrangement."

Jan 1, 1971

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 R. H. McCorkell

Radon emanation rates. J. and compositions of materials from the Canadian Certified Reference Materials Project (CCRMP) indicate that the fraction of the Rn produced that escapes, E, is usually 1% to 8% and the 238U decay chain is nearly in secular equilibrium. For some Chinese ore standards, E values are higher than f or the CCRMP materials. Standard ore tailings have 2/oPb concentrations higher than expected from their 226Ra concentrations and Rn escape rates (12 - 34%) . Biological materials have E ˜100%, their ash E ˜0%. Pelletizing ore powders at 20 000 lb/in2 reduces J by only 20 - 40% but only 2% of Rn escapes from pressed Li2B4O7 pellets and 0.01 - 0.04% from fused Na2B4O7 or Li2B4O7 discs. Large, difficult-to-explain random variations in J are observed in some situations but several highly reproducible ways of measuring emanation rates have been found.

Jan 1, 1986

By Tony Lipiec, Ian Orford

"A key part of plant design is the selection of the design tonnage which is the hourly tonnage target that the equipment selection is based on. Its selection is critical for ensuring that the nominal tonnage which is critical to the financial performance of the project will be met. With deposits being developed at lower grades, higher tonnages and a higher degree of hardness variability, it is essential to ensure that the design process selects the comminution equipment which will be able to deal with the impact of this ore quality variation.Ore hardness and its variance are key factors to be taken into consideration for the selection of the design tonnage. It is well accepted that for the nominal or name plate tonnage to be consistently achieved, it is necessary to base the design tonnage on a hardness value greater than the average ore hardness of the deposit to be mined. The question is how much greater should this design value be to ensure that the nominal tonnage is met to the satisfaction of the project financial model. Standard industry practice is to set the design tonnage by:1. Select an ore hardness percentile that is greater than the average ore hardness to be the basis of the nominal tonnage. Typically either the 75th or 80th percentile value from the data set will be selected.2. Selection of a design factor to multiply by for determining the design tonnage, i.e. the design tonnage = tonnage (based on the average ore hardness) multiplied by design factor.This paper discusses the selection of the design ore hardness through the use of modelling or the use of a design factor. It further shows the impact of ore hardness variance on the capacity of the process and its ability to meet the nominal tonnage. Insight is provided as to what ore hardness percentile value should be selected balancing the shortfalls of tonnage for ores harder than the design value versus the above target production achieved for ores softer than the design ore hardness."

Jan 1, 2016

By Marc Lavigne

"The Lac Knife graphite deposit was estimated, as part of a feasibility study, using a geostatistical ore reserve estimation method: ordinary kriging. Based on the geology and its interpretation, the results of a variogram study, and the statistical parameters of the grade distribution of the core samples and the composites through the mineralization, the deposit was divided into three separate zones (west, centre and east), and each of these zones was then analyzed and kriged independently.The geological ore reserve estimation results of ordinary kriging are compared to the results of a conventional method (sections) used previously. The two methods gave almost similar reserve figures. The conventional method provided a +2% estimation for tonnage and + 10% for grade because it ignored the spatial relation between sections and samples, and ordinary kriging gave a more accurate longitudinal definition of the deposit.Finally, the paper concludes with the advantages of proceeding with ordinary kriging, namely: higher flexibility (change of cut-off grade), better deposit resolution and enhanced selectivity for the design of the open pit.IntroductionWithin the context of a complete feasibility study for the Société d' Exploration Minière Mazarin Inc., specialists of the Mines and Geology Division of Roche Ltd., Consulting Group, defined a modelling method appropriate to the graphite deposit of the Lac Knife property near Fennont. Many patterns were considered. Based on the geology of this deposit, the planned mining method, the statistical parameters of grade distribution and the variogram study, the process of reserve estimation retained was to use the ordinary kriging in a selective way. It was required because the geometry of the deposit is such that it was unthinkable to apply ordinary kriging without restrictions.This paper gives a detailed explanation of the observations and interrogations which led to the choice of this ore reserve estimation method, and the calculation procedure itself."

Jan 1, 1991

By E. I. Vargha-Butler

The determination of surface tensions and contact angles of coal particles is of great interest for coal beneficiation processes such as froth flotation and oil agglomeration and for the inclusion of coal particles into oil droplets as a means of enriching the fuel. Methods for contact-angle measurements on small irregularly shaped solid particles have not been available in the past. Therefore, we recently developed a novel technique to determine contact angles and surf ace tensions of coal particles. Two types of bituminous and one type of subbituminous coal particles have been tested. The technique used is called the freezing-front technique and it considers the behaviour of the test particles (in the range of 20-250 pm) at the solidification front of carefully selected, well-defined solidifying melts (matrix materials) such as thymol, biphenyl, naphthalene, salol and 2-phenyl phenol.The results show that the technique works well, providing quantitative and detailed information on the wettability of coal particles. The sensitivity of the method is so high that a difference in contact angle of approximately 4 degrees between the smaller and the larger particles is clearly registered.

Dec 1, 1981

By Stuart McTavish

"The Brunswick Mining concentrator processes 10,000 D.S. T. per day of a complex sulphide ore. Four separate concentrates- zinc, lead, copper and bulk-are produced by selective flotation. Several major changes have been made to the Brunswick flotation circuit to improve the process and control. This paper will describe the present flotation practices at the Brunswick Concentrator. Special emphasis is given to the conversion of two of the primary circuits to an SO2-aeration flowsheet, and to the addition of a new secondary zinc flotation circuit that produces a zinc concentrate from the plant tailings. IntroductionBrunswick Mining and Smelting Corporation operates a zinclead- copper and silver mine and concentrator near Bathurst, New Brunswick. Since the start-up of operations in 1964, the flotation flowsheet has undergone major modifications to process the increased tonnages and to produce a greater number of concentrates. The flowsheet for the concentrator as it was in 1971 has been described by Neumann and Schnarru) and by Ersklneco. Since then, in 1971, the No.6 concentrator was converted from bulk flotation to selective flotation to produce a copper, lead and zinc concentrate. A separate zinc upgrading circuit was installed in 1972 and a bulk flotation or dezincing circuit, which reduced the zinc in the lead concentrate, was started in 1973. In 1976, the secondary zinc circuit was initiated.In addition to these new circuits, changes have been made to the existing circuits to simplify their operation."

Jan 1, 1980

By Kiran B. Chaudhuri, Michael Koch, Jorg Lema Patino

"Stricter pollution regulations together with other factors have stimulated interest in the modernization of the traditional sinter - blast-furnace lead operation. In this context, direct smelting methods have again come into prominence. Consequently, the Kivcet Lead Smelting Process, which belongs to this group, has attracted the attention of the lead producers of the world. The Kivcet Process offers the advantage that lead and zinc can be simultaneously recovered in a single furnace unit. KHD Humboldt Wedag A G, Koln, Federal Republic of Germany, has signed a contract with the Sociedad Complejo Metalurgico de Karachipampa of Bolivia for the delivery of a turn-key Kivcet plant, which is expected to be the first plant of its kind in commercial operation. This paper describes the design and expected operating data of the plant."

Jan 1, 1980

By Ben Cottrell

In this paper technical drilling is referred to as the process of collecting engineering and metallurgical parameters for subsequent mine design considerations. Throughout the development of a mineral exploration property there are various opportunities to ?piggy-back? technical drilling onto the routine exploration drilling process, helping to reduce future costs and delays associated with re-drilling and missing information. This article is written from the engineers and metallurgists point of view, to outline some best practices for exploration borehole planning and technical data collection requirements.

Aug 1, 2013

By R. G. Gillies, C. A. Shook, W. H. W. Husband

"The current state of knowledge concerning pipeline transport of coal in water slurry form is summarized for the four categories: coal-water fuels, conventional fine coal slurries, coarse coals and stabilized mixtures of coarse and fine particles. Surface chemistry is of dominant importance for coal-water fuels and considerable progress in understanding the role of surfactants has been achieved in recent years. Little generalized research on the flow of conventional fine coal slurries has been done in the past 17years, although innumerable conceptual designs and feasibility studies have been undertaken. Coarse coal slurries are fairly well understood but the model which explains their flow requires more data to be established fully. Stabilized flow mixtures should probably obey the same model as coarse coal slurries but experimental results for stabilized mixtures have not yet been incorporated in a comprehensive theory.IntroductionPipeline transportation of coal in a water vehicle has been employed for many years in coal washeries. It has been used less frequently in hydraulic mining and thermal power station fuel supply systems. The suitability of coal for pipeline transport arises from its moderate specific gravity relative to the carrier fluid and the fact that throughputs are often large and steady over long periods of time."

Jan 1, 1990

By Dale L. McKinnon, Jean-Pierre Farant

"Dust collectors are used to reduce airborne dust levels generated during crushing at a mining operation and their adequate maintenance is required to ensure effective control of health and safety hazards associated with dusty conditions. Although the pitot traverse technique is commonly employed to balance these dust collectors, the use of an alternative-the equilibrium concentration tracer gas method-is proposed in this paper. This method, which uses sulphur hexafluoride as a tracer gas, can be used to balance airflows in dust collectors and to determine hood collection efficiency accurately and effectively. This technique also reduces some of the errors inherent to pitot tube measurements. IntroductionIndustrial dust collectors are used to reduce dust genera ted by processes like cru shin g, and these must be periodically checked and balanced to ensure continuing effective control of health and safety hazards associated with dusty conditions.To date, the pitot tub e traverse technique has been commonly used to balance air flows in dust collectors. However, this technique is both time -consuming and cumbersome. The equilibrium concentration tracer gas method, described in this paper, is proposed as a viable alternative as it is faster, more reliable, and amenable to automation. The use of tracer gases for such task s is not new. This approach has been used in a variety of mining situations to determine volumetric flow rates in underground openings, leak age through mine stoppings, the effectiveness of face ventilation, pollution migration patterns, mine air recirculation rates, and rock porosity."

Jan 1, 1986

By M. P. Filion, K. Ferguson, L. L. Sirois

Acidic drainage is the largest single environmental problem facing the Canadian mining industry today. Technologies to prevent acidic drainage from occurring in wasterock piles and tailings sites, and on the walls of open pits and underground mines, need to be developed and demonstrated. There are two groups in Canada which have accepted this challenge: the national Mine Environment Neutral Drainage (MEND) program and the British Columbia Acid Mine Drainage (BCAMD) Task Force. This paper summarizes the activities of these two organizations.

Jan 1, 1990

By Z. M. Zavodni, C. D. Broadbent

"Major rock slides in open-pit mines are rare, but they can develop in spite of sophisticated slope engineering design practice. Experience has shown that, except under high seismic load conditions, the serious slope instability condition is almost always accompanied by the gradual development of one or more tension cracks behind the slope crest, allowing for time-displacement monitoring. Continuous surface displacement measurements with attendant prompt analyses are usually adequate for the prediction of slope behaviour.This paper describes efforts that have been made to quantify parameters associated with slide movement, and compares the displacement records of several large-scale open-pit porphyry copper mine slope failures. Two principal failure stages are recognized in the typical slope failure leading to total collapse: (a) REGRESSIVE STAGE, during which the failure will re-stabilize if some disturbance external to the rock and structure is removed; and (b) PROGRESSIVE STAGE. during which the failure will gradually displace at an accelerating rate to the point of total collapse unless active control measures are taken. Over-all displacement records in both failure stages are apparently of simple exponential form. with a definite break occurring at the onset-of-failure point.Certain common empirical relationships are suggested for the large-scale slope failures analyzed in this study. Based on plots of displacement rate vs time prior to collapse, a semi quantitative empirical relationship is presented for failure collapse prediction. The comparison of historical slide displacement records furnishes the operating engineer with practical guidelines to estimate the characteristics of future large-scale slope failures."

Jan 1, 1980

By S. H. Lucas

"This paper describes the development and application of single-stage fluidized bed continuous ion exchange to the following processes: removal of base metals from mine water; cyanide removal from gold mill waste; uranium removal from sulphuric acid leach liquor; and uranium removal from mine water. IntroductionThe recovery of metals by fixed bed ion exchange has been practiced for many years. However, the operation has been semi-continuous and can only treat clarified liquors. Substantial savings could be made if the operation were continuous, thus reducing the resin inventory, and if the costly filtration and clarification steps could be eliminated. Work has been in progress over the past twenty-five years on the recovery of metals from unfiltered ore leach slurries, heap leach liquors and mine drainage waters containing a small amount of suspended solids. In the case of dilute waste solutions, with suspended matter and high flow rates, filtration is usually prohibitively expensive.Continuous ion exchange operation using fluidized beds has made it possible to reduce considerably the resin inventory and to allow the treatment of unclarified liquors and slurries. The subject has been reviewed on several occasions (1. 2. 3, 4). Although fluidized beds in multi-stage columns are an improvement over previous resin-in-pulp treatment, they have complicated internal structure and I or involved automatic control systems to move the resin stages.Slater and Prud'homme performed studies on multi-stage fluid beds, and concluded that the performance of single, deep fluidized beds of resin with continuous or semi-continuous flow of both phases would be worth investigating, particularly in the case of very dilute metal solutionsoi, A pilot plant was designed and built at CANMET in a cooperative project with Dr. Slateroi. The use of deep fluidized beds of resin (with a large height/diameter ratio) for treating dilute solutions has the attraction of engineering simplicity. Extreme simplicity of equipment design is considered of paramount importance when treating, for example, mine waters at high flow rates, possibly of the order of several million gallons per day. This paper describes the development of the concept of single-stage fluidized bed continuous ion exchange arid reviews the pilotplant development of processes for the treatment of base metal mine water, cyanide waste water, uranium leach liquor and uranium mine water."

Jan 1, 1980

By P. Farsangi, A. Hara

"Kidd Creek Mines places approximately 10 000 tonnes of consolidated rockfill on a daily basis. Consolidated rockfill is comprised of sized aggregate mixed with different ratio and amounts of binders. Consolidated rockfill, with closely controlled specification, is employed for subsequent pillar recovery and improved passive active/ ground support. With the same amount of cementing agents consolidated rockfill mass yields a higher, compressive strength, modulus of elasticity, cohesion, and angle of friction compared to a typical cemented hydraulic fill mass.The consolidated rockfill at Kidd Creek Mines is designed to sustain not only the gravitational loading of the overlaying fill material, but also to some extent the dynamic loading applied during blasting of the adjacent stopes and/or pillars.The quality control of materials used in Kidd Creek's rockfill system could be effected in three stages of handling: (1) at the surface backfill plant; (2) during transportation; and (3) placement of the fill.The emphasis in this paper is placed on individual stope backfill design and quality control measures at Kidd Creek Mines to place fill in large-scale open stopes which eventually could be exposed on all four walls during pillar recoveries."

Jan 1, 1993

By R. A. Frith

"Belts of Yellowknife Supergroup volcanic and sedimentary strata in the Slave Structural Province are separated by granitoid complexes. The strata and most intrusive rocks are 2.7 and 2.5 Ga old, but rocks around 3.1 Ga have been recognized in some of the complexes and are believed to represent a once-extensive sialic crust. Volcanics were erupted along rifts in this crust and the rifts became sites of crustal foundering and the development of troughs that filled with turbidites of volcanic and granitoid provenance. Present margins of the granitoid complexes may be close to the original margins of deposition of volcanics, with associated mineral deposits, and close to the thickest parts of basins of sedimentation.The Hackett River volcanic belt extends 100 kilometres between Mara River and Back River and is crudely S-shaped, with a northern sector surrounding a core of granitoid basement rocks, a northeasterly facin g central sector and a southerly facing southern sector. Up to 8 kilometres of andesitic volcanics with intercalated felsi c volcanic rocks are preserved within the central sector. The f elsic rocks include thick lenses of volcaniclastics with large, quartz-eye-bearing clasts, but the known metal deposits are not associated with these. Both of the deposit areas, the Hackett River area in the northern domal sector and the 'A itch ' Lake area 40 kilometres to the south in the central sector, are near the tops of relatively thin atypical volcanic sequences above suspected basement rocks. The Yava deposit near 'A itch , Lake overlies a unit of subaerially deposited, welded, andesit ic ash flows and the Hack ett River deposits overlie highly metamorphosed volcaniclastic rocks that include suspected ignimbrites.The deposits are notably argentijerous and lead-rich. They consist of stratigraphically zoned sheet-like bodies of iron sulphides, sphalerite, galena and chalcopyrite underlain by altered rocks containing chalcopyrite stringers. The Yava deposit was formed where hydrothermal fluids exuded through the fractured floor of a flooded caldera basin. The Hackett River deposits may have been formed in a similar setting and it is not inconceivable that all the known deposits in the belt were formed in a single large caldera basin, although there is no independent evidence for the required major fault separation between the Hackett River dome and the 'Aitch' Lake area."

Jan 1, 1980

By G. Bimpilas, G. N. Anastassakis

"Today, one of the most intense problems in mineral processing systems is referred to the treatment of fine (–100 µm), and even smaller size, particles. Fine mineral particles possess specific properties (small mass, high specific surface area, and high surface energy per unit area) that render their separation inefficient with conventional methods (physical, physicochemical) and equipment. As a result, it is not feasible to obtain a concentrate both of high grade and recovery in mineral value and a considerable amount of mineral fine particles is lost to the tailings. As evidence, we could cite numerous examples from industrial practice with considerable loss of mineral values to the tailings because of the fine particles size, e.g. phosphate, tungsten, tin, etc.The objective of the current paper is to investigate the possibility to separate magnetic gangue minerals and rocks (serpentine, ilmenite, and olivine) from magnesite by co-agglomerating all the gangue components and subsequent magnetic separation of the agglomerates.To meet the target, hydrophobic agglomeration tests were initially carried out on single minerals to define the favorable conditions for selective hydrophobic co-agglomeration of the gangue minerals. In this respect, the effect of the following factors was examined: pH, collector dose, conditioning time, and settling time of the agglomerates. Among the combinations examined, the most favorable proved that with amine in the basic pH region. The use of non-polar oil proved beneficial. The preliminary results were confirmed by tests on mineral mixtures. The application of magnetic separation resulted in the removal of gangue minerals and the recovery of magnesite concentrates of commercial grade (magnesite content 95% approximately) and high yield.The current paper contributes to the development of mineral processing, since it successfully copes with a very important problem of mineral processing, such as the efficient separation of fine particles."

Jan 1, 2016

By R. W. Boyle, W. Dyck

"Scientists throughout the world have observed radioactive disequilibrium in the uranium decay series. Radiometric and fluorometric determinations of U in plants from the Goldfields district and in sediments and rocks from the Beaverlodge district exhibit marked radioactive disequilibrium. The eU/U ratio in plants rooted in mineralized ground shifts from greater than unity to less than unity for those rooted in background soils. A similar shift is observed in the transition from rocks to stream sediments and to lake sediments in the Beaverlodge district.Results of a Rn-Ra-U survey of lake bottoms in the Key Lake district show that anomalous hydromorphic U from U mineralization is parallelled by anomalous Rn, even though there is much less Ra in the lake-bottom sediments than one would expect from the amount of U present. IntroductionDisequilibrium in the U decay series has been studied for many years and is observed in almost all surficial environments. Rosholt (1959) measured U, 234Pa, 23O'fh, 226Ra, 222Rn, and 210Pb in a large number of U ores and outlined six types of situations to explain disequilibrium. Davy (1974) outlines the geological significance of radioactive disequilibrium situations with respect to exploration. Levinson and Coetzee (1978) discuss possible mechanisms by which separations of parent and daughter elements are affected and suggest analytical pitfalls that may be encountered in the use of radiometric techniques when disequilibrium has occurred. Generally, it is the oxyphile and easily complexed parent 238U (with its associated isotopes 235 U and 234U) that is leached from rocks and soils at a faster rate than are the daughters, 230Th, 226Ra, 2loPb, and 206Pb. This mobilized U is adsorbed on sediments in streams and lakes, or enters groundwaters, where it is eventually precipitated under reducing conditions, occasionally in oregrade concentrations. This surface leaching and"

Jan 1, 1980

By A. Nielsen

"The Polaris mining operation (Fig. I) is located on Little Cornwallis Island in the Canadian Arctic, an area with a very limited shipping season and a hostile climate. The operation consists of the underground works, barge-mounted concentrator facilities, concentrate storage and shipping facilities, a personnel accommodation complex and an airstrip.The electrical energy for the operation is provided by four 2.3 MW diesel-generators installed on the barge. The power is distributed through a 4.16 KV system connected directly to the generators and to the large motors. All other loads are supplied at 600 V or 120 V.The power plant has been designed as a Total Energy System with maximum energy efficiency and thus designed for maximum heat recovery. The waste heat recovered on a year 'round basis is expected to be sufficient for the process heating and drying as well as the space heating for offices and accommodations. The engine cooling system consists of two glycol-water loops, a "" cold"" loop for charge air and lube oil cooling and a ""warm"" loop for jacket water cooling. Each loop is complete with heat exchangers f or heat recovery and out-door forced air radiators to dispose of excess heat. A water loop in the concentrator area transfers the heat from the jacket water loop and the exhaust gas boiler to the space heating equipment, the process and the heating system for the accommodations. Part of the exhaust gas is diverted to the concentrate dryers.IntroductionThe Polaris Mine is a highly mechanized, trackless mine with access through a decline. Drill equipment is electrically powered to minimize fumes and heat. Dry development and production drilling eliminates the need for an underground water supply system.The concentrator, power generators, work shops and auxiliary facilities are located on a barge which is beached in a prepared berth on the shore of Crozier Strait. Oil storage tanks, concentrate storage and accommodation buildings are built on shore. The concentrate storage building ha s a 12-month capacity and is fed from the concentrator by a conveyor system.The accommodation building houses up to 240 people, includes administration office, dining and recreation facilities. The complex is housed in on e interconnected building with public use areas alternating with accommodation units.At full production annual concentrate shipments will amount to about 220,000 wet tonnes. This will be accommodated by about 8 shipments during the regular shipping period-August 15 to October 15. Re-supply of general cargo and fuel will be carried out annually in about 3 shipments within the same period."

Jan 1, 1984

By C. E. McCoy

This paper will discuss briefly loader shovel requirements and then present a detailed survey of electrical drive systems to meet these requirements. The survey in general will deal with the various electrical provisions for shovel drives, along with the advantages and disadvantages, and in particular will discuss the MG set and static D. C. drive systems. Also included will be some insight into the requirements for optimum electrical power at the shovels.

Jan 1, 1979