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By S. D. Graham

"The extractive metallurgy of gold is largely controlled by mineralogical factors such as the presence of refractory gold, particle size, gangue mineral associations, presence of preg-robbers, oxygen and cyanide consumers, and locking characteristics in base metal sulphides. Light microscopy and automated mineralogy techniques are regularly used to characterise the effect of these variables on gold recovery. However, despite their widespread application there are a range of limitations when grades are low. Typically, the analysis of low-grade samples requires the preparation and analysis of numerous polished blocks. This is often costly and time consuming, with complexities regarding analysing statistically representative particle quantities and problems overcoming stereological bias. This is compounded by the “nugget” effect and high tenor nature of gold mineralisation. X-ray Microscopy techniques have the potential to overcome many of these problems, in particular the statistical representivity of analysis, removing stereological bias and effectively locating phases that are present in ultra-trace proportions. In this paper we outline the potential benefits of employing X-ray Microscopy (XRM) in characterising gold losses from a typical Cu-Au porphyry project, which were illusive in 2D based analysis. The talk will also outline additional application areas of XRM within the extractive value chain."

Jan 1, 2017

By William Petruk, J. R. Schnarr

The behaviour of minerals during flotation of a fine-grained pyrite-rich base metal ore was investigated by studying the feeds. concentrate and tailings from every flotation cell in the mill of Brunswick Mining and Smelting Corp. Ltd. Bathurst. New Brunswick. The samples were analyzed for Zn, Pb, Cu. Fe, Ag. Sb, Sn, Bi, In. Hg, Mo and As. Mass balances were computed and recoveries of all metals from each cell. from each circuit and from the mill were calculated. The percentages of sphalerite. galena and chalcopyrite occurring as free grains in each product were determined with an image analyser, as were size analyses for free mineral grains and middling particles, and compared with screen analyses. The results show that the behaviour of minerals is defined much better by the recovery of free mineral grains than by the recovery of elements. The recovery of free sphalerite in the zinc circuit is 96 wt %, whereas Zn recovery from the same circuit is 88 wt %. the difference being due to unliberated sphalerite. The recovery of free sphalerite from the mill feed is 88 wt % and Zn 78 wt %. The recovery of free galenafrom the mill feed is 88 wt % and Pb 64 wt %. The recovery of free chalcopyrite is 75 wt % and Cu 45 wt %. Size analyses show that more sphalerite and chalcopyrite can be liberated by regrinding. but the galena is so fine grained that regrinding would not necessarily improve liberation. Silver occurs mainly in tetrahedrite and galena; the main loss in the tailings is due to large (+ 26 um) and small (-1 um} free and unliberated tetrahedrite grains.

Jan 1, 1981

By L. E. Djingheuzian

Introduction For many years the immense deposits of bituminous sands of northern Alberta have been attracting the attention of our governments, engineers and .scientists ( 6). Undoubtedly these bituminous sands represent one of the greatest potential resources of liquid fuels in the world. However, their commercial development has been retarded by the complexities of the technological problem and by the economic factors involved in the processing of the sands. The technological problem has been attacked from several angles ( 2, 3, 1, 11), and cold water separation of bitumen is one of them. To the ore dressing engineer, the cold water process is of utmost importance since, by the use of standard ore dressing machinery, it accomplishes most efficiently and economically the separation of bitumen from bituminous sands. The significance of using ore dressing equipment was especially enhanced by the fact that the scientific principles which underlie the design of standard machines were demonstrated in the cold water separation process with an extraordinary clarity. This greatly facilitated the analysing of limitations and advantages of different machines during-the pilot plant investigations and helped to complete the investigations practically in the allotted time.

Jan 1, 1952

By Fred G. Hewett

The Cassiar orebody is located in a sill-like serpentinite body intrusive into the Devonian-Mississippian sedimentary rocks of northern British Columbia. Subsequent intrusion of the Cassiar batholith formed the McDame synclinorium and produced chrysotile asbestos. Present information indicates an orebody of approximately 16,277,000 tons amenable to open-pit mining, but continued diamond drilling will be needed to accurately define total tonnage and grade. A program has been developed at the mine to provide the required information on a daily basis, and work is continuing on more effective dollar-value prediction methods.

Jan 1, 1978

By M. Tiedtke, T. Crary, P. MacPhail, M. Ounpuu

"The SGS carbon-in-pulp (CIP) / carbon-in-leach (CIL) model was used to evaluate and optimize the carbon circuit at Alamos Gold’s Young-Davidson Mine. Plants samples (pulp and carbon) were tested in a series of small scale batch experiments to determine rates of gold leaching and adsorption on carbon from a combined flotation concentrate/tailing product. The testwork data, along with existing plant data were then used to model the performance of the existing CIL plant and optimize the circuit to improve performance and overall recovery. The primary purpose of the modelling exercise was to determine the most cost effective way to lower solution losses at the plant. The modelling was also used to learn how the various parameters in the CIL circuit affect solution losses and the gold inventory tied up on the carbon. This resulted in recommendations for minor changes in the CIL flowsheet, which were successfully implemented in the plant. The approach, results and plant changes are presented.INTRODUCTION This paper presents the details from testwork conducted on samples from Alamos Gold’s Young- Davidson gold mine which is located near the town of Matachewan, approximately 60 km west of Kirkland Lake in Northern Ontario, within the southwestern part of the Abitibi Greenstone Belt. The objective of the test program (completed by SGS Canada in 2013) was to investigate and model the combined flotation concentrate / tailing carbon-in-leach (CIL) circuit and suggest ways to improve gold recovery. The metallurgical test program included mineralogy, rheology, and carbon modelling testwork. Plant information was provided during the test program and used in the optimization process. Upon completion of the testwork, plant changes were recommended and implemented at site (October 2014). Plant results from 2014 to 2017 confirmed the modelling results."

Jan 1, 2018

IDS: The Company IDS Ingegneria Dei Sistemi, founded in 1980, is a world leader as a high technology solution provider in selected niche defense and civil market sectors ? Private and Independent ? IDS?s Quality System is certified ISO 9001 Ed. 2000 ? IDS has the highest Military Clearance ? Continuously measure (24x7) mine wall movements with sub-millimetric accuracy at long distance (up to 4 Km) ? Provide reliable early warning for progressive slope movements potentially leading to mining wall failures ?Map the long term evolution of slow moving slopes to get a better knowledge of the rock mass strength

Aug 1, 2013

By A. P. Yovanovic, P. Bonadia, M. N. Moraes, C. F. Matos

"The magnesite flotation operation of Magnesita Refractories, in Brumado – Brazil, has four parallel ball mills working in closed circuit with hydrocyclones to grind the ore for the reverse flotation of silicates in mechanical cells. The modification proposed here opens the circuit for three of the ball mills so that their products are collected together to feed a high frequency screen. The oversize of the screen feeds the fourth ball mill in closed circuit for secondary comminution. The product reports to dewatering and flotation. The undersize undergoes desliming and subsequent flotation operation. According to pilot tests performed, the modification on the magnesite flotation operation will promote a 20% increase in the productivity, 8.5% increase in mass recovery and a 26% reduction in water consumption. The gains are due to the fact that the silicates in the feed are concentrated in the smaller particles (<100µm) with over grinding in the existing closed comminution circuit, making it impossible to take advantage of this natural selectivity. The combination of open circuit, more selective milling, and the high frequency screen, ensures high efficiency particle size classification. This renders the process more effective, both to the screen undersize with its high contents of silicates and fines, and to the screen oversize, which is almost without fines and has a higher magnesite grade. This situation allows for the setting of specific parameters for the flotation of both the primary milling undersize and the secondary milling product.INTRODUCTIONMagnesita Refractories, with its operations site in Brumado, is the largest producer of high purity dead burned magnesia (DBM) in Latin America. The magnesite ore is extracted from an open pit mine, named Pomba, after which it is crushed and handpicked. Due to the presence of disseminated silicate the ore has to be ground and concentrated by means of reverse flotation, before feeding the high temperature kilns for sintering and conversion to DBM.The magnesite flotation unit has an impact crusher to reduce the size of the feeding ore, which is transported to the mills by conveyor belts. The milling operation is composed of four ball mills in closed circuit with hydrocyclones. The closed circuit generates a high circulating load with excess fines, which is detrimental to the selectivity of the desliming and flotation operations, reducing both mass andmetallurgical recovery.In the current scenario, the capacity of the flotation plant is insufficient to meet the demand of the DBM kilns, forcing the operation to work with high purity ores in order to attain the highest possible mass recovery in the concentration process. This is not an ideal situation as it requires the mining site to work with higher stripping ratios.The objective of this work was to improve the magnesite flotation unity to increase its productivity. After a detailed analysis, the milling was identified as the limiting process, which made the authors to focus on this operation."

Jan 1, 2016

By B. Simonian

"The Mount Milligan mine, located in central British Columbia, began operation in September 2013 with a design nameplate capacity of 62,500 dry metric tonnes per day, using a conventional SABC grinding circuit followed by bulk sulphide flotation. Over the last few years, many changes have been implemented in order to debottleneck and optimize the comminution circuit to meet and exceed the nameplate design capacity as well as improve the metal recovery of the flotation circuit. This paper discusses changes, grouped into mechanical modifications, projects, and operational changes. The implementation of these changes as well as ongoing challenges and potential solutions will also be discussed.INTRODUCTION The Mount Milligan mine began production in the fall of 2013. It is a conventional truck-shovel open pit gold and copper mine and concentrator with a 62,500 tpd design capacity copper-gold flotation processing plant. The average head grades for the life of mine are 0.18% Cu and 0.39 g/t Au. At full capacity, the mine produces an average 400 dry tonnes per day of copper-gold concentrate containing approximately 21.5% Cu and 50-60 g/t Au. Current mine life is 21 years. This paper will discuss a selection of the operational, engineering, and administrative changes that have been made at the site since the publication of An Update on Reaching Nameplate Throughput and Continuous Improvement at Mount Milligan (Yue, 2017). PLANT FLOWSHEET CONFIGURATION Run-of-mine ore is crushed through a 60-110 gyratory crusher to a top size of seven inches. Up to 100% of the gyratory crusher product is diverted to the secondary crushing circuit by a moveable gate. The material diverted to the secondary crushing circuit is fed to a vibrating double-deck screen. The screen oversize is placed on a surge pile and reclaimed by four vibrating feeders to feed two 1250 kW standard head cone crushers in open circuit. The screen undersize is combined with the crusher discharge and returns to the conveyor feeding the coarse ore stockpile along with primary crusher product that by-passes the diverter gate."

Jan 1, 2019

By R. Bashir, J. F. Hatley

ABSTRACT Northern Saskatchewan boasts some of the world’s largest known high-grade uranium deposits. The successful mining of these deposits cannot be accomplished without overcoming technical challenges, however, which include high grade of the uranium ore, specialized mining methods to deal with groundwater at high pressures, and poor ground conditions. This paper discusses the various mechanisms of inflows at three of Cameco Corporation’s sites in northern Saskatchewan. The risk of inflows is quantified in terms of unique challenges due to hydrogeological conditions, rock mass integrity, and uncertainty in geological conditions. Strategies in the case of an inflow are also briefly described.

Jan 1, 2013

By K. K. Mamyrbayeva, G. Z. Guseinova, V. A. Luganov

This paper presents the results of theoretical and technological studies on leaching ore of Bozshakol deposit. The major oxidized copper mineral of Bozshakol ore is atacamite, and the content of malachite, azurite, cuprite, copper sulfide minerals is insignificant. The chemical composition of the test sample, with mass, %: 0.9 Cu; 55.00 SiO2; 22.40 AI2O3; 9.00 Fe2O3; 2.24 CaO; 3.56 MgO. Technological studies on leaching the ore were carried out with sulfuric acid 5, 15, 25 and 50 g/dm3 in one and multistage (up to 4 stages) countercurrents. Size of ore 1 mm, the ratio S : L 1: 4. It was found that with increasing of acid concentration from 5 to 50 at one-mode leaching the copper extraction reaches 55 %. At multistage leaching the copper recovery reaches 73 %, a productive solution containing g/dm3: Cu - 4; Fe – 1.2; Al – 1.1 at a pH of 2.8. When leaching the cinders obtained by ore roasting in an atmosphere with limited value of oxygen, with increasing of the roasting temperature from 650 up to 750 °? the extraction of copper from the Bozshakol ores reduced from 31.5% down to 24.4%. The rate of filtration is markedly increased compared with the crude ore. When leaching the cinders obtained by roasting in a nitrogen atmosphere a marked increase up to 87.2-88 % of copper extraction from Bozshakol ores is observed. As an extractant for the extraction of copper from the pregnant solution used LIX984N extractant in kerosene. On the basis of the calculations it was built extraction isotherm and modeled processes of copper recovery with the circuit 2E + 1P, allowing up to 93 and 96% extraction. The through extraction of copper from the ore in the electrolyte was 67 %. Technology is recommended for semi test.

Jan 1, 2016

The major factors limiting the establishing of vegetation on the shale and sandstone wastes at Minto are high exchange acidity and low pH (shale pH 4-5, sandstone pH < 4). In the first revegetation program (1967-74), 3.8 million trees were planted without grading on 4,000 acres (red, jack and scots pine and black locust). White dutch and sweet clover, birds-foot trefoil and red top were established in selected areas, and ponds were stocked with brook trout. In the second phase of reclamation, beginning in 1979, a comprehensive plan will be initiated which will allow for forest production, wildlife development and recreational activities.

Jan 1, 1977

By G. B. Harris

It has long been appreciated that chloride chemistry has a number of advantages over the more traditional sulphate route for both extracting and separating metals. Modern materials of construction, coupled with the development of a more cost-effective and efficient hydrochloric acid regeneration technology mean that chloride-based processes are now more competitive and attractive. Such processing routes are able to recover more of the contained metal values in a feed, including iron in a benign and/or marketable form, and particularly the so-called rare and rare-earth elements which are increasing in demand in our electronic age. These factors, therefore, add appreciably both to the overall economics of a project, but also offer a much more sustainable approach to our dwindling natural resources. Additionally, chloride-based flowsheets can be much more environmentally-friendly, offering practical alternatives to two of the biggest headaches faced by the industry, namely cyanide usage and the generation of red muds. This paper presents a brief theoretical background, focusing on the latest developments of ferrous iron oxidation and the low-temperature, selective hydrochloric acid regeneration process, which is the key unit operation for any chloride flowsheet, and discusses in general the advantages of the chloride approach. Several recently-developed flowsheets, which have been tested at the miniplant level, are presented for gold, base metals, titaniferous magnetites and lead-silver ores wherein the advantages of considering a chloride approach are highlighted.

Jan 1, 2014

By John Convey

"IRON ORE was first smelted in Canada about the year 1736 at Les Forges, Quebec -a most appropriate name for such a historic locale. During these early years of the 18th, 19th, and first quarter of this century, about six million tons of ore were mined and employed in the manufacture of items ranging all the way from kettles, stoves, and cannon for the early settlers to weapons of war for the allied armies of World War I.But this symposium on iron ore is not concerned with this early period. Instead, it is concerned with the iron ore industry of today -an industry which will soon be-come one of the two most valuable mineral industries in Canada.The subject in its broad geographic and economic aspects will be introduced by Mr. Buck; Dr. Harrison will describe the geology and origin of the principal iron ore deposits; Dr. Downes will go into beneficiation as it applies to iron ores; and Mr. Gertsman will discuss primary processing. The symposium will not deal with the steel end of the industry, because that phase itself could well form the subject of a symposium."

Jan 1, 1955

By Pritii Tam Wai Yin, Tony Leong, Neetu Bansal, James Vaughan, Amiel Boullemant

Mercury is a toxic element that is associated with many industrial activities including the Bayer process. Mercury is known to enter the Bayer process with the bauxite feed material and has the potential to accumulate as the process solution is recycled. This paper explores the speciation of mercury in Bayer process streams. The process parameters considered are the oxidising-reducing potential, basicity and temperature. The results of a refinery circuit survey are presented and compared with the thermodynamic predictions. Electrochemical measurements on relevant synthetic systems are also included.

Jan 1, 2014

By M. Kumral, E. Dilhuydy, B. Ozdemir

"In-pit crushing and conveyor (IPCC) systems have become increasingly popular in the mining industry as an alternative to truck-and-shovel systems. Deciding whether to haul waste rock by truck or IPCC represents an important gap to be investigated in semi-mobile IPCC systems. A cost analysis is presented to evaluate the feasibility of crushing and conveying waste material in IPCC systems, along with sensitivity and risk assessments through Monte Carlo simulation. Although the capital expenditure is higher, results showed waste crushing will be feasible in deep pits and over long mine life and haul distances.RÉSUMÉ La popularité des systèmes de concassage et de transport dans la fosse minière (IPCC, de l’anglais In-pit crushing and conveyor) n’a cessé de croître dans l’industrie minière ; ils constituent en effet une option intéressante de remplacement des systèmes par camions et pelles mécaniques. Le choix entre le transport des stériles par camions ou par IPCC est encore difficile à faire et devra faire l’objet d’une étude approfondie au niveau des systèmes d’IPCC semi-mobiles. Une analyse des coûts est présentée pour évaluer la faisabilité du concassage et du transport des stériles par les systèmes d’IPCC, de même que des évaluations de la sensibilité et des risques par le biais de la simulation de Monte Carlo. Si les dépenses en immobilisations s’avèrent être plus élevées, les résultats ont cependant montré que le concassage des stériles sera envisageable dans des fosses profondes ainsi que sur une longue durée de vie et de longues distances de transport. INTRODUCTIONThe mining industry is currently facing many challenges such as low commodity prices, increasing environmental concerns and regulations, lower grades, deeper mines, and greater waste-rock quantities. As a result, the profitability of mining operations has decreased, with many mines stopping or suspending operations, and many new projects being postponed. Mining companies have therefore focused on improvements and innovations to increase the cost effectiveness of operations, for example, increasing production capacity to take advantage of the economies of scale (Crowson, 2003) and decreasing the unit operation cost. Changes in production capacity force mining companies to evaluate alternative material handling systems.During the past few decades and within material handling systems, significant improvements have been made to truck performance and capacity. Consequently, operating expenditures (OPEX) associated with fuel and maintenance have risen swiftly. In a mining operation, truck hauling systems now represent 40–50% of the capital cost and approximately 45% of the OPEX (Tutton & Streck, 2009). Additionally, material transportation costs are expected to increase due to longer haul distances arising from deeper and larger mines (Fleet et al., 2015). Finally, large fluctuations in fuel prices result in high risks associated with truck OPEX. According to the United States Energy Information Administration (2015), the cost of crude oil—the largest component of the retail price of diesel fuel—increased from US$24.31/barrel in 1986 to US$46.92/barrel in 2016, peaking at US$151.72/barrel in 2008 (Macrotrends, 2016)."

Jan 1, 2017

By V. J. Southey

HOLLINGER mining operations have required development by 24 vertical shafts having a total length of 28,160 feet. Seventeen of these are sunk from surface and seven are interior shafts. Four surface shafts and three interior shafts are now active and, excepting the recently completed new main ore-hoisting shaft, the remainder are either abandoned, or are used for passes or for ventilation. The new shaft bottoms 60 feet below the 3,050 level, but there is much work to be done before it is placed in service. The term &apos;shaft practice&apos; is intended to include shaft construction and a general treatment o typical equipment. It is a large subject for an organization the size of Hollinger and this paper is intended only to present the salient features. The writer wishes to thank the management for permission to publish this information and to acknowledge the advice and criticism o A. G. Irving, chief engineer, and the willing co-operation of other members of the Hollinger staff.

Jan 1, 1939

By The Mill Staff

Introduction The Hollinger cyanidation mill commenced operating in June, 1912, and 49,779,978 tons of solids had been discarded by the end of 1950. The present tailing disposal plant has been in operation since 1924 and holds 39,762,250 tons. The tailing disposal plant is on the property of Kayorum Gold Mines, Limited, about a mile and a quarter south of the mill. The area available for tailing covers 640 acres and is roughly l ¼ miles long by ¾ mile wide. Tailing was first deposited on the McKay veteran claims, l mile long by 1/2 mile wide, which were later included in the Kayorum property. This site sloped from northeast to southwest, the slope amounting to 106 feet between the northeast corner and the low land to the southwest. The low land was wet and a considerable area was covered with muskeg up to 17 feet deep. It was drained by a brook running across the property in a southwesterly direction. Later, when more ground was needed, the low-lying mining claims adjacent to the south and west boundaries were acquired to allow for the extension of the plant. This entailed some extra difficulties and expense that would not have been encountered had the whole of the present site been acquired with the original purchase. At the present time, the top of the tailing deposit stands at two elevations, but eventually the tailing deposited on the extension to the original plant will be raised to the same height as the older deposit. Previous to 1924, all of the tailing was deposited on Low land and in lakes near the mil1. Between 1924 and 1926 an increasing tonnage was sent to the McKay claims until, in 1926, all of it was being sent there. In 1936, the area of the deposit was extended and shortly afterwards all of the tailing was deposited on the new site.

Jan 1, 1951

By H. R. Bullis

"A 2003 paper discussed the idea that as gold mining companies grow larger it becomes more difficult for them to discover or acquire new gold deposits large enough to replace extracted reserves on an annual basis. Deposits of the size required to replace annual production of very large gold producers (VLGPs) are statistically few. An update in 2005 reviewed VLGPs with respect to production, reserves, and reserve life and observed recent changes in the gold mining industry. Now, more than 12 years later, this paper explores how those VLGPs have fared and how the gold mining industry has changed. KEYWORDS Gold, Gold price, Production, Reserves, Very large gold producer (VLGP) RÉSUMÉ Un article de 2003 examinait le concept selon lequel, à mesure que les sociétés d’exploitation aurifère se développent, il leur devient de plus en plus difficile de découvrir ou d’acquérir de nouveaux gisements aurifères suffisamment grands pour remplacer les réserves qu’elles extraient annuellement. Statistiquement, il existe très peu de gisements de la taille requise pour remplacer la production annuelle des très gros producteurs d’or. Une mise à jour de l’article en 2005 se penchait sur la production, les réserves et la durée de vie des réserves des très gros producteurs d’or ; elle portait sur les récents chan¬gements au sein de l’industrie de l’exploitation aurifère. Aujourd’hui, plus de 12 années plus tard, cet article explore l’évolution et les résultats de ces très gros producteurs d’or, ainsi que les changements qu’a connu l’industrie de l’exploitation aurifère. MOTS CLÉS or, prix de l’or, production, réserves, très gros producteurs d’orINTRODUCTION In 2005, the world’s top five gold producing companies were Barrick (after purchasing Placer Dome; 9.1 million ounces; Barrick Gold Corporation, 2005; Placer Dome, 2005), Newmont (8.2 million ounces; Newmont Mining Corporation, 2005), AngloGold Ashanti (6.2 million ounces; AngloGold Ashanti Limited, 2005), Gold Fields (4.2 million ounces; Gold Fields Limited, 2005), and Harmony Gold (3.0 million ounces; Harmony Gold Mining Company Limited, 2005). During the years leading up to that point, there had been a flurry of activity in the gold sector directed at consolidating production and reserves to create very large gold producers (VLGPs). As defined in Bullis (2005), a VLGP is a mining company whose primary busi¬ness is gold production and that has annual gold production more than 5 million ounces. A stated corporate goal of several of the VLGPs at the time was to become the largest gold producing company in the world."

Jan 1, 2019

By Robert J. Shank

This paper describes the Federal Government assistance available for northern mining development. The paper commences with a discussion of why the Federal Government should give assistance to northern mining development. It then uses a hypothetical case of a company which has a property it wants to develop to list the sectors within the . Federal Government that would have to be consulted in order to obtain federal financial assistance. The paper also outlines the type of information a company must be willing to provide the government in order to qualify for such assistance.

Jan 1, 1978

By C. Jay Hodgson, P. J. MacGeehan

"The gold ore zones of the Campbell and Dickenson mines occur within a km-wide, 2 km-long zone of highly altered and anomalously fissile and deformed volcanic and subvolcanic intrusive rocks, cut by a series of mainly post-deformation mafic and quartz-feldspar porphyry felsic dykes.Three main periods of gold ore-forming hydrothermal activity are recognized. In the first, the mine rocks were extensively altered and gold and arsenic enriched; syn-volcanic auriferous veinlet and low-grade chemical sedimentary ores were deposited in what was a major discharge zone of a submarine hydrothermal system. The chemical sedimentary ores are closely associated spatially and temporally with domal rhyolitic extrusions. Because of its mechanical and chemical properties, this synvolcanic alteration zone became the locus of later deformation, and the site of the generation of auriferous hydrothermal solutions, perhaps by pressure solution processes at depth. These solutions deposited, synchronously with deformation and metamorphism, the second and most economically important type of ore zones in the mine: large foliation-parallel and foliation-oblique banded carbonate- 'chert' Rssure-filling veins.In the third and final mineralizing event, high-grade, pipeshaped zones of siliceous replacement-type mineralization, and quartz-arsenopyrite-gold veins, were superimposed on the previous mineralization types after most of the deformation, but before the emplacement of the late dykes. The source of heat and also possibly the fluids for the third mineralizing period may have been the dyke-forming igneous activity. The textures and mineralogy of the banded carbonate- 'chert' veins suggest a near-surface (<3 km) environment of deposition, but their relationship to the fabric of their host rocks indicates they were emplaced in an environment, presumably at considerable depth (5-10 km?), in which the rocks were undergoing intense penetrative deformation. This anomaly is unresolved."

Jan 1, 1982