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By Pierre O'Dowd

"The southern Abitibi Belt is the major source of gold in Quebec. Although the northern part of the belt is quickly being developed, the area under study will not be downgraded in the near future because it still contains more than 75% of the known reserves in the belt.Gold is found in a variety of geological environments which are part of a small number of distinct structural patterns. For the exploration geologist, outlining such patterns might be easier and more productive than defining, when it is possible, the exact genetic nature of the gold mineralization.Exploring in the right place has always been one of the main concerns of explorationists. This paper briefly describes the various favourable geological settings, and outlines areas, types of structure and lithologies of high potential for gold exploration in the southern Abitibi Belt.IntroductionThe southern part of the Abitibi Belt is a profilic mining area for precious and base metals (Lulin 1990). Although mineralizing processes and controls of polymetallic deposits are well known and understood (Lydon 1984), almost every gold occurrence is a unique case and numerous models have been proposed (Roberts 1987; Hodgson and McGeehan 1982; Hodgson 1990).This study does not define new genetic models for the formation of gold deposits, neither is it meant to elaborate on the existing ones. A genetic approach, is by definition, theoretical and, to some extent, universal. It helps geologists define types of targets and exploration tools but does not outline or quantify target areas. To tentatively summarize the complexity of a genetic approach, a brief discussion on the various genetic models in vogue in Abitibi and the problems related to their use in exploration can be found at the end of this paper."

Jan 1, 1991

By J. C. O'Rourke

"Gibraltar Mines Ltd. has been operating its open pit copper mine for fifteen years. In that time, over 140 million tons of waste and low-grade material has been placed on the mine waste dumps. Laboratory and in situ testing indicated that enough leachable copper was recoverable from the dumps to justify a leaching operation which would include the construction of a solvent extraction-electro winning plant. The plant, which was commissioned in October 1986, was designed to process 3000 USGPM of dilute copper solution to produce 5000 tons per year of high-quality cathode copper.This paper deals with the start-up and initial year of operation of this plant. The three areas: leaching, solvent extraction, and electrowinning are presented with emphasis on the innovations which were made to help overcome the problems unique to the Gibraltar operation.The Leaching CircuitThe mine waste dumps are leached to produce a pregnant copper solution as feed to the SX-EW plant. To achieve this the dumps are sprinkled with a mildly acidic solution which percolates through the dumps. In travelling through the dumps the solution contacts the surfaces of the liberated copper minerals and dissolves the soluble copper into the form of copper sulphate. The leaching reaction is achieved by a combination of bacterial action to oxidize the sulphides and solubilization of the copper into the acid solution.For the bacterial action to occur the dumps must be near the ideal pH of 1.9 to 2.1 and the necessary elements of the reaction must be present. Because a bacterial reaction is involved there is a lag time between the time when the conditions for bacterial growth are achieved and when the bacteria are in their active growth phase. Laboratory leaching tests with Gibraltar ore indicated that bacterial action takes two to three months to reach its maximum rate (Fig. I).The expected lag time for the planned rate of bacterial leaching prompted Gibraltar to start irrigating the dumps in late June 1986 for a plant start in early fall. The lag time proved to be longer than expected and the dumps took until November to start producing copper at the expected rate (Fig. 2)."

Jan 1, 1988

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 Michael Jerome Wallace

"This paper refers to the wet open circuit ball mil ls in operation at Sidbec-Normines'5 pelletizing plant at Port-Cartier, Quebec, which commenced operations in late November· 1977. Prior to detailing our experiences in the grinding area at Port-Cartier, a general description of the total Sidbec-Normines project is outlined.HI STORYSidbec-Normines was incorporated in 1976 to develop the Fire-Lake - Port-Cartier iron ore mining, concentrating and pelletizing project. The primary objective of Sidbec-Normines is to produce 6mi11 ion tons per year of oxide pellets, of which 3 mi 11 ion tons per year are of a low silica grade suitable for direct reduction purposes. The participants in the project are Sidbec, the Quebec Government steel producer, British Steel Corporation (lnt'l) and the Quebec Cartier Mining Company. Equity interest is held in the company by the participants in the proportions of 50.1%, 41.6:7% and 8.23%, respectively."

Jan 1, 1979

By Philip Thwaites

"Automatic flotation control in varying degrees of sophistication is practiced at a number of concentrators throughout the world. Practically, the variable conditions in the flotation circuit and the difficulty in maintaining accurate instrumentation information all add to the complexity of automatic control.Further development at the Kidd Creek concentrator has resulted in a 3075 mtpd copper flotation circuit capable of running twenty-four hours a day under the direct control of an optimizing supervisory computer.Three main streams, continuously monitored by an XRF analyzer, supply circuit information to the optimizing computer. Circuit efficiency calculations (from smelter contract terms) and target setpoints allow the computer to determine the economic optimum direction to move the circuit based on the grade/recovery relationship. Circuit manipulation is accomplished by precise pumping of collector and frother reagents together with accurate control of flotation bank pulp levels. The former uses computer controlled diaphragm pumps and the latter is achieved by control loops utilizing a float type level transmitter, variable position dart plugs and a process computer working to setpoints determined by the optimizing computer.The development of a fully automated flotation circuit is by no means complete, but increased knowledge and major hardware improvements have allowed its progression to a higher level of sophistication.IntroductionThe Kidd Creek concentrator is located twenty-six kilometers east of Timmins, Ontario, within a metallurgical complex including a 109,000 mtpy zinc plant and a 59,000 mtpy copper smelter and refinery. The concentrator treats two ore types, a copper-zinc ore and a copper-silver-Iead-zinc ore, at a combined rate of 4,500,000 mtpy. Flotation of the crushed and ground ore yields copper and zinc concentrates from the copper-zinc ore and copper, lead and zinc concentrates from the copper-silver-lead-zinc ore."

Jan 1, 1983

By D. L. Mckay

"Texasgulf's 25-foot-diameter Kidd Creek No. 2 Shaft is situated next to the company's No .1 Shaft, so that both shafts can share common surface and underground facilities. Shaft construction, covering a span of five years from collaring to commissioning , started by freezing the 50 feet of grey varved clay covering the bedrock. After first completing the concrete headframe, sinking to the 5/00-foot horizon utilized two different shaft construction methods. The top 2800 feet was excavated by slashing a 6-foot-diameter borehole and hoisting the broken material up the 3OO0-foot-deep No.1 Shaft; the remaining 2300 feet was excavated by conventional benching methods. Prior to installing the 5100 tons of galvanized shaft steel on the completion of the sinking operation, the sinker was able to install all thirteen armoured electrical cables in one pass with his modified galloway stage. All loading and lip pockets were installed along with the steel guides and bun tons during the equipping operation.IntroductionOn November 21,1978, Texasgulf completed the No.2 Shaft at Kidd Creek Mine. This shaft is the largest shaft (volumewise) in North America and is unique in many respects . The purpose of this paper is to outline some of the unique features of this shaft as well as to record how this facility was constructed.Soon after sinking No.1 Shaft, Texasgulf started a program of exploratory diamond drilling from the 2800 level in No. 1 Mine. The drilling outlined sufficient ore below that level to justify a second mine. Feasibility studies demonstrated that production could be increased by one-third exclusively from this deep ore . The study called for a 25-foot-diameter shaft to the 5100-foot horizon to service a mine producing 5,000 tons of ore per working day.The site chosen for No.2 Shaft and head frame is 350 feet southwest of No.1 Shaft. This, along with No.1 Shaft, will be away from the influence of ground subsidence when mining at depth. The collar and head frame foundations are located on a flat rock ledge some 70 feet below surface. This flat ledge was found by test holing through the overburden by diamond drilling. It is fortunate that this location best satisfied the surface layout and logistics for both surface and underground ore and waste handling."

Jan 1, 1981

By E. Kouame, M. D. Everell, D. Hodouin, T. Kasongo

"To properly operate a mineral beneficiation plant it is often necessary to have systematic evaluations of the performances of a piece of machinery, of one of the processes or of the plant as a whole. The state of a process can be described from a collection of data: flow-rate values, particle-size distributions and chemical compositions of the various streams of the circuit studied.However, the obtained raw data are generally redundant and inconsistent from a material balance viewpoint, due to natural disturbances, sampling errors, unreliable instrument readouts and laboratory analysis inaccuracies.BILMAT is a computer program which upgrades mineral processing data, adjusting them in a least-squares sense in order that they may verify material conservation constraints. As a consequence, the program is able to calculate the best values of unmeasured flow rates (such as. circulating load ratios). BILMAT is a very general program in the sense that it can accept various data from any type of mineral processing circuit, and its usefulness is demonstrated here for three processing circuits of a pyrochlore concentration plant: grinding, desliming and flotation.IntroductionComputation of the material balance around any metallurgical process, that is the determination of stream flow rates and composition. is a basic step in its evaluation. In spite of the simplicity of the physical laws of mass conservation. and of the very fundamental need to know the flow of components around a process. it appears, particularly in mineral processing. that material balancing is quite often a difficult task. The reasons for this involve the problems associated with collecting measurements in mineral dressing plants. They are generally poorly instrumented and invariably data must be collected through hand sampling. Due to the heterogeneous nature of the streams, segregation occurs and samples are often not adequately representative, a problem made worse by the lack of circuit sampling facilities, which are often completely ignored by plant designers. Furthermore, the nature of the measurements to be performed (for instance particle size distributions in the fine range) frequently gives inaccurate results."

Jan 1, 1981

By J. W. Reynolds

"The use of approved fire-resistant hydraulic fluids in equipment operating underground was made mandatory in British Columbia by amendment to the Provincial Mines Regulation Act effective January 1, 1975. This paper covers the action taken to achieve this, in the largest underground metal mine in British Columbia. Some evaluation of costs and effects based on experience to date is included.BackgroundAn amendment to the Mines Regulation Act of the Province of British Columbia was passed in July 1974 as follow s: "" Rule 111 (b) Hydraulic Fluids. After the first day of January 1975, all hydraulic fluids used in equipment operating underground shall be of a fire-resistant type accept able to the chief inspector.""Unquestionably, its intent was justified as a safety improvement measure. The list of fires in mines involving machinery where hydraulic fluid was a factor in the fires is long and sobering. In British Columbia, in a 22-month period from 1972 to 1974, there were six fires in metal mines resulting in extensive equipment damage and injury to four workmen. Nevertheless, no provision was made for research or development of systems modifications compatible with conditions encountered in B.C. and, as late as June 1977, the Mining Association of British Columbia continued to argue that fire suppression systems should be considered sufficient until a universal suitable hydraulic fluid could be developed. The sole concession was to issue permits for continued use of mineral oil in equipment bought prior to 1975 in order to allow producing operations to continue. The burden of developing the necessary technology was laid on the industry.The Sullivan Mine at Kimberley, the largest and longest established underground mine in the province, and Cominco's most important source of ore s for its smelter at Trail, B.C., was at the outset of a conversion from conventional to trackless mechanized mining methods in pillar recovery operation s. Introduction of fire-resistant hydraulic fluid s had to be done in a period of change, in many instances on equipment for which no data existed, and called for research which had to be done while retraining the work force.The introductory phase is complete. All equipment is on FRF, the requirements of Rule 111(b) are being satisfied, and a number of research and development programs continue."

Jan 1, 1980

By O. Neves, C. G. González Beca, M. C. Sol Pereira, H. Sousa Vale, R. A. Guedes de Carvalho, M. N. Sampaio

"As sulphide ores are usually concentrated by flotation , the material is very finely ground <37um), wet and impregnated with the chemicals used in flotation. These conditions are favourable to the development ofchemical reactions which are largely influenced by atmospheric conditions (temperature, humidity, wind, etc.) as well as by the geometry of the piles.Physical and chemical properties of a chalcopyrite concentrate were determined by laboratory tests. The knowledge of these properties was considered necessary in order to establish a correct design of the storage and handling operations of the sulphide ore concentrates obtained by flotation.The properties studied in this work were: transportable moisture limit, flowability, effects of vibration, hygroscopicity, ""caking"", chemical alterations by humidity, oxidation and self-heating (auto-ignition, dry and wet oxidation, self-heating by humidification and its prevention).INTRODUCTIONIn order that the handling and the transport of a powdered material may be carried out correctly, it is necessary to know certain physico-chemical characteristics. Transportable moisture limit (TML), apparent density (both aerated and compacted), compressibility, angle of repose, angle of spatula, cohesion and angle of slide(l-4) are characteristics (among the flowability parameters) worth knowing.Moisture content of the powdered material is fundamental, because if this is too low heavy losses will occur when the product&apos;s particle size is very small, if it is too high, separations between solid and liquid phases may occur. In the latter case, the centre of gravity of the system may change suddenly and cause accidents, mainly if transport is by sea or by river."

Jan 1, 1989

By A. D. Hunter, J. M. JR. Moore

"The Aldermac deposit and associated prospects occur in the Noranda Archean volcanic centre of the Abitibi greenstone belt. Over two million tons of are grading 1.65% copper, with minor gold and silver, were mined between 1933 and 1943. The deposit has features in common with the Vauze and Millenbach mines in the same district.Recent detailed mapping, geochemistry and petrography combined with data from former diamond drilling and mining operations permit a re-interpretation of the Aldermac area. The orebodies and spatially related sulphidic zones are con eluded to be volcanogenic.The host rocks are a chemically bimodal suite of mafic and rhyolitic flows, volcaniclastics and related subvolcanic intrusions. Massive sulphide lenses are spatially related to a quartz-feldspar- phyric rhyolite dome and siliceous, volcanic-exhalative sedimentary rocks, The succession is cut by gabbro and syenite, related to later stages of volcanism, and by post Kenoran diabase dykes. Regional metamorphism attained the low greenschist facies, but the rocks lack penetrative deformation.Hydrothermal alteration is primarily an intense silicification of mafic volcanic rock s enveloping, or lying below, sulphidic zones. Chlorite and sericite alteration zones, including dalmatianite, are more locally developed and have a closer spatial relation to mineralization, as do epidote, calcite and calcic garnet.The structure is complex and cannot be explained by isoclinals folding; rather it reflects concurrent faulting, and extrusive and intrusive activity. Fault blocks were rotated and exhibit opposed facing; the original selling was possibly a submarine graben.Sulphidic zones occur within mafic and felsic volcanic rocks and at the interface between the two; there is thus potential for discovery throughout the known stratigraphic succession. Exploration conducted according to a volcanogenic model is advisable, particularly in the light of recent discoveries in similar rocks of the Amulet area."

Jan 1, 1983

By T. G. Carter

"This paper traces the introduction and use of fire-resistant (FR) fluids in hydraulic systems at underground mines in British Columbia. Reasons are given for the decision to make compulsory the use of such fluids in underground mines after January 1, 1975 by provincial government legislation. Examples are quoted of fires caused, or supported by, mineral oils which have resulted in loss of life and property.The basic principles of hydraulic systems are reviewed in order to present the effects of changing a system operating on mineral oils to one which must operate on FR fluids. Approximations of the cost of system conversions are given for certain machines and a brief survey is made of the various pieces of underground equipment which contain hydraulic systems operating on FR fluids.IntroductionIn 1948, Admiral Ben Moreel , who was then president of Jones and Laughlin Steel Corporation in the United States , wrote:""Although safe and healthful working conditions can be justified on a cold dollars-and- cents basis, I prefer to justify them on the basic principle that it is the right thing to do. ""(1)This paper will outline the introduction and use of fire resistant fluids at mine operations in the Province of British Columbia. That it was the right thing to do in enhancing the safety and well-being of every person employed at the mines was the prime consideration. That it may be justified on a cold dollars-and-cents basis is left for the reader to judge. The prime purpose of this paper is not so much to trace the efforts made by the individual mines to successfully introduce and use fire-resistant (FR) fluids in hydraulic systems, but to present an over-all picture of what occurred, in a collective sense, at the majority of mines which have ended up with a working FR fluid system. The problems attendant to the introduction of an FR fluid will be recounted in a manner which, it is hoped, will prove useful to those contemplating system conversions. Because the predominant fluid now in use is an invert, or water-in-oil. emulsion, this will be the FR fluid considered.Fluid Loss from SystemsThe quantities of various fluids required for use in hydraulic systems are extremely high. This statement in itself may appear to be redundant in that the capacities of the various systems should demand that a fixed amount of fluid is required for the initial filling of a system, with a further amount of fluid scheduled for use during subsequent fluid changes, i.e. at the recommended change intervals . If only this were the case!"

Jan 1, 1981

By M. W. Gilmore, J. Ndlovu, T. Sepoloane, Anglo American Platinum, L. R. Nelson, R. N. Varty, L. J. Bryson, B. Peyper, J. P. Hagemann

"As part of Anglo American Platinum’s Rustenburg Base Metals Refinery (RBMR) expansion, a new Nickel Electrowinning Tank House, with a nameplate capacity of 32 400 tons per annum, was constructed. Commissioning of the new electrowinning technology, which involved world first full deposit plating, commenced in 2011.The old nickel electrowinning process at the RBMR used starter sheets which, along with a divided cell configuration, was very labour intensive and exposed operators to high levels of aerosols that contained both nickel and acid. To overcome this problem two key design intents of the new technology are aerosol abatement and minimising labour in the electrowinning cell environment, with a design target to reduce the nickel occupational exposure limit (OEL) to half the legislated limit of 0.1 mg.m-3.Commissioning of the new technologies adopted to achieve mist abatement is described, together with a review of technical and operational challenges relating to the titanium blank edge strips, blank roughness and sandblasting optimisation, blank quality control, electrode spacing, polyconcrete cell dimensional stability and cell top furniture. The impact of these on process performance is presented and the development of fault checking and rectification methods, to maintain high process performance, is explained.INTRODUCTIONThe expansion of Anglo American Platinum’s Rustenburg Base Metals Refinery necessitated the construction of a new Nickel Tank House to achieve two primary objectives, namely increasing the nickel electrowinning capacity from 22 500 tons per annum (tpa) to 32 400 tpa nickel, and a design to reduce the nickel aerosols to 0.05 mg.m-3 as a time weighted average over 8 hours, which is half of the legislated occupational exposure limit (OEL) of 0.1 mg.m-3."

Jan 1, 2016

By D. S. Harries, R. I. L. Guthrie

"The use and development of scientific disciplines involved in heat transfer technology, related in general to steel masses and in particular to batch annealing, is demonstrated.The measurable benefits of cooperation are tabled; on industry&apos;s side, this includes a 10 to 20 per cent decrease in annealing time, and a corresponding higher productivity potential. The returned benefits to the academic engineer include the development and refinement of new computer algorithms related to heat transfer mechanisms.IntroductionThe object of this discussion is not really to describe an improvement in the batch annealing operation, but to use the project as an example of joint work between industry and the academic field. Both parties have derived a number of substantial benefits. These benefits will be listed, of course, but the batch anneal project will be used just as a typical example of cooperation. Why it happened will be explained, what actually happened will be described and finally some thoughts about areas in the field of heat transfer, where cooperation will continue in the future, will be highlighted. The paper will close with a message.Industry OverviewAt the outset, perhaps it should be understood that, in general, the Research Department at Dofasco believes that the universities and their staff are not destined for the problem-solving role in industry. Its opinion is that the university staff is best suited to the gathering and dissemination of knowledge. There are other more professional problem solvers for industry, including the steel industries&apos; own individual investigation groups. Dofasco, therefore, looks to its academics for consultation. It benefits by bouncing ideas off them, and through rubbing shoulders with them in such a way that their knowledge can be tapped, and thus establish more effective scientific routes. In addition, the younger and brighter engineer s can broaden their scope through the intellectual contact with experts. The university professors can, it is believed, benefit through closer contacts with result-orientated research. They can also develop an appreciation of the fiscal control necessary for industrial investigations, without government subsidies, grants, etc., and with massive overheads."

Jan 1, 1982

By J. G. Mieczaniec

"A key element for a successful mining operation is the grinding circuit. It must be reliable, maintainable and must perform to or above design capacities for 95% of the time. Once chosen and installed major modifications are either prohibitive or very expensive. The selection of mills and drive system is probably the single most important decision made within the concentrator complex.Economic studies revealed that the optimum expansion for Lornex was a 68070 increase in plant throughput. This indicated that the new mill would have to be larger than the existing 32-foot mills.The Lornex ore has a wide range of hardness which has a great influence on mill throughput (Table 1). To smooth out the fluctuations, a blending program ha s always been in effect at the mine, coupled with ball charge control for the primary mills. As can be appreciated, an optimum balance is difficult to achieve. A sudden change to so ft ore with a high ball charge puts the mill into jeopardy with a constant speed drive.The blending philosophy would continue after the expansion, but with the addition o f a second primary crusher, it was envisaged to have the new mill process harder ore to maintain production targets. In theory, the new mill would have a proportionally larger ball charge than the existing mills.This scheme makes the new mill vulnerable for processing so ft ore for a fixed speed mill. Thus, variable speed seemed an attractive option.In early 1979 prominent mill manufacturers were invited to submit proposals for the new mill. Fixed and variable speed mills were to be considered with mill size ranging from 32-foot to 36-foot diameter. The proposals were to include the drive, ball charge, % critical speed, couplings, gears, etc. - the complete package. The recommendations were variable with most manufacturers submitting several alternatives. Recommended mill speed varied from 68% to 78% critical. The concept of variable speed was becoming more attractive, Lornex ultimately decided on a twin DC motor drive system for a 34-foot diameter mill.After start-up it was discovered that best throughput occurred at 78% mill critical speed or top speed of the drive motors. The motor speed range was adjusted upward from 186 to 200 rpm. Today the mills operate at 79% critical or a motor speed o f 193.5 rpm.Clearly the selection of the variable speed mill allowed the flexibility to optimize throughput."

Jan 1, 1985

By C. A. Roloff

Demands for increased grinding production, lower-grade are and limited capital have resulted in ever larger grinding mills, some as large as 36ft (/1 m) in diameter and 12,000 hp (8,950 kw). The result is often f ewer machines per plant, with a greater dependence on each machine. To meet these reliability requirements, manufacturers of mills, drives, castings and other components have had to be innovative and creative in their design. Only by thorough review of past designs and practices and by model studies confirmed in field tests can the manufacturer have full confidence in the product. Because of the great interest in large mills, a short review of larger machines is presented. This discussion covers material feeding, bearings, the rotating element, drive trains and material discharge.

Jan 1, 1979

By Gary Raymond

The Similkameen Division of Newmont Mines Limited operates a low-grade open-pit copper mine in southern British Columbia. The copper mineralization is extremely erratic.As a result, over seven years of operation, ore control methods have evolved from standard flagging based on blast-hole assays, through an in-pit grab sampling system, to the present method of mining to computer-determined ore-waste boundaries based on geostatistics. These improved ore control methods have resulted in a gain of at least 25% on ore tonnage miffed, at the same grade as obtained using earlier techniques. At the same time, geostatistical ore reserve estimates are now within 7% of milled ore tonnage and grade on a yearly basis, whereas the original polygon ore reserves under-estimated mineable ore tonnage by 15% to 45% and over-estimated grade by about 20%. The author discusses why traditional ore estimation methods give poor results in erratically mineralized orebodies and describes the geostatistical approach to the problem.

Jan 1, 1979

By Owsiacki. Leo

Both the 8 and A-C ore zones of the Heath Steele Mine contain extensive areas of brecciated ore. This texturally distinct ore type has been tentatively defined as a fragmental or conglomerate massive sulphide. It consists of subangular and round fragments of pyrite up to 30 em in diameter, smaller quartz clasts and chloritized, mineralized, lithic fragments in a faintly banded matrix of pyrrhotite and chalcopyrite. The chalcopyrite is concentrated in the pressure shadows of the fragments and appears to reflect remobilization and recrystallization during a deformation which postdates the formation of the clasts. A generally conformable relationship with the underlying banded pyrite-sphalerite ore, discordancy with the host banded argillites, spatial distribution and similarity to Kuroko ores suggest that the fragmental ore is of primary origin and that breccia formation is probably due to soft-sediment slumping or sliding initiated by volcanic explosions, as hypothesized for similar massive sulphides in various Japanese mines. It is nevertheless recognized that certain structural features in some clasts may result from deformation preceding the formation of the breccia, and at this stage a tectonic origin cannot be completely eliminated.

Jan 1, 1979

By Marc C. Betournay, John E. Udd

"In this study, the authors investigated the stabilities of slopes cut predominantly into Trenton limestone in a large quarry in the Montreal region. Analyses of the bench and pit walls indicated that small-scale toppling, ravelling, and wedge failures are the most likely modes of instabilities. Because most of the joints dip steeply, however, the conclusion was made that relatively high walls could probably be justified.The field investigations were supplemented with the determinations of the shearing strengths of artificial discontinuities that were cut into samples of the limestone. The properties of a fine-grained limestone were found to be influenced very much by the magnitudes of the loads that were applied, while the properties of a crystalline limestone were found to be much less, or not load-dependent.The study is one of a number that are being conducted in this area.IntroductionExtensive quarrying has been done in the sedimentary strata of the Montreal area for a very long time. By any standards, it can be said that the Island of Montreal is a major mining area. At present, approximately 20 quarries are being operated in, or near, the city of Montreal.In addition to the millions of tons of limestone that are removed each year, a large number of excavations for foundations are cut into the sedimentary rocks.To date, however, very little information has been published concerning the engineering properties of the local geological materials. No mention is made of this aspect in the references which are presently available.Because of this lack of information, the senior author has initiated a study in which the strength properties of most of the rock types, and of the natural discontinuities contained in these, would be investigated. To date, the results of one other investigation have been published and other work is in process. It is hoped that the end result will be a comprehensive knowledge of the important strength properties of all of the rocks in the entire region."

Jan 1, 1983

By A. Himsley, E. Byleveld

"With reference to the chemical constraints underlying the ion exchange process for uranium recovery, a generalized view is presented on the actual and the achievable efficiency of the process. The area considered is restricted to acid leach circuits in combination with nitrate, chloride or sulphuric acid elution.A comparison between actual plant results and the chemical requirements shows that:(a) Unless new and better ion exchange resins are developed, no dramatic change in equipment size and resin inventory is""anticipated. (b) Elution efficiency, especially for sulphuric acid, may be improved considerably. More concentrated eluates may beproduced at lower chemical consumption. A possibility of achieving additional purification solely by ion exchange is discussed.IntroductionThis paper presents a generalized view of the actual and the achievable efficiency of the ion exchange processes for urannium recovery. The study is restricted to acid leach circuits in combination with nitrate, chloride or sulphuric acid elution. The ultimate efficiency attainable with ion exchange is dietacted by the chemical reactions underlying the process. The most sophisticated technology cannot change the diffusion rates in the ion exchange resin nor make the uranium loading exceed its equilibrium valueThe design of an ion exchange plant for uranium recovery involves a considerable number of parameters as a result of the wide variety of pregnant solutions produced from different orebodies. Therefore, a generalized approach of the ion exchange process inevitably neglects some parameters which maylead to significant variations in apparently identical situations.The uranium concentration in the pregnant solution, together with its pH, are the principal variables which determine resin loading and residence rime in the ion exchange system. On the elution side, the type of eluant and its concentration will determine the contact time required and the amount of eluant needed to strip the resin. The temperature, the properties of the ion exchange resin, the total dissolved solids in the feed solution and the competing ions are other variables that affect the chemical performance; these, however, may be considered as a correction of the principal parameters if one is interested only in general trends."

Jan 1, 1980

By G. W. McDonald

"The mineralogical studies done on Hope Brook ore samples indicated that a heap leach gold recovery process may be possible. In the metallurgicalprogram that followed, column tests confirmed this interpretation. A 9300 tonne demonstration heap leach test, in the late fall of 1985, reaffirmed laboratory results while identifying operating conditions that will have to be addressed in a production heap leach plan. Operating plans, gold recoveries and cost areas will be presented.Another important factor in a successful operation is the environment approach.LocationThe Hope Brook gold deposit is located 5 km inland from Cinq Cerf Bay on the south coast of Newfoundland. Port au x Basques is 80 km west and Burgeo 40 km to the east of Hope Brook. As there are no roads in the area, access to the site is by sea or air (Fig. I). The Hope Brook orebody is near the saddle of three watersheds flowing to Cinq Cerf Brook, Boat Hole Brook and Betty&apos;s Pond. Acceptable solutions to several environmental concerns can be obtained by using the relative sensitivities of these watersheds.The terrain at the site is generally rolling to hummocky with numerous small ponds separated with rock outcrops and patches of peat covering. Tree vegetation is sparse except in the sheltered areas around Couteau Bay and Cinq Cerf Brook. GeologyThe Chetwynd deposit is hosted by the Lower to Middle Ordovician Georges Brook Formation of the La Poile Group, a thick sequence of volcanic and sedimentary rocks.The Hope Brook Zone contains the largest concentration of gold found to date at Chetwynd. The zone occurs as an intensely silicified lens 100 m to 550 m long (average 400 m) by up to 60 m in width and forms the eastern portion of a more extensive, 4 km long , alteration zone trending at 053 degrees Az and dipping 70 degrees southeast.The Hope Brook Zone contains massive finely granular quartz, scattered blue quartz eyes, 3070 to 8% pyrite, minor calcopyrite and trace molybdenite. Barren silicification envelopes the gold mineralization. The zone is flanked to the southeast by a 200 metre thick aluminosilicate alteration zone and to the northwest by a 10 metre barren pyritic siliceous fragmental rock . The original lithology may be either a thick felsic pyroclastic sequence or a quartz porphyry sill that has been altered by hydrothermal processes. Silica flooding of the permeable host was accompanied by gold deposition."

Jan 1, 1989