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By G. Beckstead, L. F. Sawatsky, R. Nutbrown, D. Long, G. Gould

"This paper presents a case study for optimizing water management at Gregg River Coal Mine located near Hinton, Alberta. The optimization study was derived based on a review and assessment of existing management practices and plans for future operation and final reclamation. A proactive water management strategy was developed to assist mine staff in adopting a rational, environmentally sound, and economical approach to water management. A plan was developed to provide for safe operation, self-sustaining operation, low maintenance, minimum negative environmental impacts, and anticipation of future needs for existing and future operations. A plan for final reclamation drainage was developed including establishment of effective, sustainable, and maintenance-free water management systems. The study demonstrated the importance and benefits of a proactive approach for integrated mine water management. This approach is recommended for other similar mine operations because it will improve environmental compliance and reduce costs of water management. IntroductionGregg River Resources Ltd. (GRRL) operates an open pit coal mine on the eastern slopes of the Rocky Mountains, located about 45 km south of Hinton, Alberta. GRRL opened the mine in 1982 and expects to continue operations after Year 2000. GRRL has progressively reclaimed the disturbed area and expects to complete mine-land reclamation shortly after mining in the area is complete. The existing disturbed area and reclaimed area covers approximately 9.5 km2 and will expand to about 10.5 km2 upon mine closure.Like other coal mines in the mountainous areas of Alberta and British Columbia, the GRRL coal mine is subject to highly variable temperatures and moisture conditions. Average monthly precipitation varies from a low of 12.6 mm in January to a high of 113.0 mm in August, based on the precipitation record at the mine site since 1982. Despite the relatively high rainfall during the spring and summer months, excess water problems during rainfall events are quickly replaced by water deficit conditions during periodic droughts. Surface runoff varies from near zero in the winter to a mean annual monthly maximum of 78 mm during the month of June. The mean annual surface water yield is 345 mm."

Jan 1, 1997

By Werner Unland, Nick Holl

"The paper presents the engineering design and experience gained so far with a dewatering system for an open-pit uranium mine planned at Key Lake in Northern Saskatchewan.The results of hydrogeological investigations preceding engineering exhibit an extensive unconfined overburden aquifer hydraulically connected with the underlying Athabasca Formation and basement rocks, both of relatively low hydraulic conductivity. The overburden aquifer is embedded in a bedrock trough, with the deepest depression between the planned pits. The hydrogeological data were used in a numerical dewatering model simulating groundwater flow atdifferent stages of the pit development.Based on the model results, the engineering design had to provide for different pumping rates, varying between 1.0 and 0.21 ms/sec. This problem was solved by using a double line pressure system. The design concept for the complete peripheral discharge system and the well design used for 26dewatering wells are discussed.Experience with reverse-circulation techniques applied for the construction of the dewatering wells is presented. Theoretical and practical considerations which led to the constructionof wells of different diameters are discussed.The design of the in-pit dewatering system provides for seepage ditches, horizontal drains, pit-bottom sumps and inpit dewatering wells where required."

Jan 1, 1980

By Tapio Leskinen, Jouko Koskinen, Timo Nlltti, Lappalainen Seppo, Pentti Vanninen

"The Courier 300 on-stream analysis system manufactured by Outokumpu has now been in operation at five of the company's concentration plants long enough to give a reliable basis for closer examination of the benefits achieved in control of these processes. The first system was installed at the Pyhiisalmi mine at the end of 1967. Since November 1970, two analyzers have been in operation at Pyhiisalmi feeding the assay data to a process computer which controls the whole plant and does the optimization of the process. In October 1970 the Courier 300 analysis system was installed at S. T. O. LESKINEN J. KOSKINEN the Keretti concentrator of the Outokumpu mine, followed soon by installations at the Vihanti Mine in May 1971, at the new Vuonos mine in November J 971 and at t he Kotalahti mine in December 1971. Since installation of the Courier analysis systems, considerable improvements have been achieved in the control and the economy of the plants. Immediate knowledge of the changes in the process and the possibility for rapid action has resulted in evener production and higher capacity and recovery. The recovery has been increased most significantly at the Vihanti concentrator, which operates on a complex ore. Consumption of chemicals has been reduced at all concentrators. Due to the fact that process control and partly also bilance analysis are not achieved by conventional methods any more, it has been possible to reduce the personnel of the laboratories drastically. Two of the plants, viz. the Pyhiisalmi and the Vuonos concentrators, are computercontrolled. The Vuonos concentrator started operation in January 1972 under computer control from the very beginning. The Pyhiisalmi concentrator has been computer-controlled since 1970. The Courier 300 analysis system was described in detail at the AIME Centennial Annual Meeting (1971). In this paper, less attention is paid to the analysis system. Only the sampling system of the Courier 300 is described in more detail. In the development of sampling equipment, effective use has been made of the company's wide knowledge of the problems involved and the test field offered by its concentrators. The sampling equipment includes secondary samplers and six basically different types of primary samplers, with which most sampling situations can be managed."

Jan 1, 1973

By C. M. Diaz, P. J. Mackey

Since its inception, in 1987, the Copper-Cobre series of conferences has established itself as the prime world forum for discussing advances in copper process metallurgy, in particular copper sulfide smelting, copper recovery by leaching and electrowinning and copper electrorefining. All copper producing countries have actively participated in the pyrometallurgy symposia held in the framework of these conferences. In 1999 and 2003, the close to 100 pyrometallurgy papers that were presented had to be accommodated in two volumes. In the present paper, the authors firstly provide a brief historical sketch of the Copper-Cobre conferences and then focus on providing a retrospective look at the dramatic changes that have taken place in copper smelting in the last few decades, primarily as discussed in the Copper-Cobre and precursor conferences pyrometallurgy proceedings. Topics include: the development of autogenous, increasingly more intensive, environmentally sound and energy efficient smelting processes; significant improvements in Peirce-Smith converting practice and at the same time the commercialization of alternative continuous converting routes; the consequent expansion and modernization of smelters; the increasing compliance with strict government environmental regulations in most copper producing regions of the world; the realignment of the copper smelting industry throughout the world; and trends in copper smelting R&D. A look at the future closes these reflections on the contribution of the Copper-Cobre conferences to the advancement of copper smelting technology.

Jan 1, 2007

By R. K. Collings

The annual production of mineral and mineral-based wastes in Canada is conservatively estimated to be 400 million tons. Over 90 per cent of this amount is waste rock or mill tailings from the mining and mineral processing industry. The remainder are slags and ashes from the metallurgical industry, and sludges from the chemical industry. Because of impurity content, remote location, or both, most wastes have limited usefulness and, indeed, are a liability in that they must be transported to waste disposal sites and contained therein. Certain mineral wastes do, however, have potential value. Some contain small percentages of metals that may be economically recoverable with price increases; others consist of mineral material that may be used for such purposes as road base, construction aggregate, railroad ballast, and in the manufacture of bricks or blocks. Increased interest has been shown in mineral wastes in recent years. This paper reviews the types of mineral wastes available in Canada, briefly describes a number that have potential value, and notes possible uses and related research by CANMET and other organizations.

Jan 1, 1978

By Richard W. Hutchinson, Robert W. Hodder

"Pyritic massive sulphide deposits lie conformably above submarine volcanic rocks, and beneath siliceous and manganiferous iron formations. The massive bodies are often zoned, with a barren pyritic top grading downward into pyrite with sphalerite, and occasionally galena, and a chalcopyrite-rich base. They are frequently brecciated, like the lavas beneath them, which contain pyrite and chalcopyrite as coatings on joints and fractures and as dispersed grains. The lavas also exhibit propylitic, argillic and silicic alteration. These sulphide deposits appear to have formed on sea bottom from fluids escaping through the lavas during periods of exhalative and sometimes explosive volcanism. The process may be comparable to the present deposition of base metals from hot brines in the Red Sea. Some porphyry copper deposits resemble the altered, sulphide-impregnated lavas which lie beneath bodies of pyritic massive sulphide. In porphyry copper deposits, pyrite and chalcopyrite occur as veinlets and dispersed grains in intensely fractured tuffs, flows, sub-volcanic intrusions or other host rocks, characterized by propylitic, argillic and silicic alteration. Metal zoning is present; sphalerite and galena often occur in manganiferous veins peripheral to a shell of pyritic veinlets which encloses the mineable zone of chalcopyrite-bearing veinlets. This sulphide distribution appears to have formed from hydrothermal fluids escaping, either passively or violently, from solidifying magma. The fluids may be comparable to hot brines encountered by drilling in the Salton Sea area of California.There are also important differences between massive sulphide and porphyry copper deposits. The former usually occur in differentiated, mafic-to-felsic subaqueous volcanic rocks, whereas porphyry copper deposits are often associated with shallow plutons and subaerial volcanic rocks of intermediate to felsic composition. Calcic-sodic metasomatism accompanies the massive sulphides, whereas potassic metasomatism characterizes the porphyry cop.per deposits. Further, ages of massive sulphides range from Archean (Keewatin) to Tertiary, but porphyry coppers are Triassic to Tertiary in age. These similarities and differences suggest that massive sulphide and porphyry copper deposits are related tectonically and metallogenetically. The former were generated during early stages of crustal evolution, particularly in Archean time, and were repeatedly derived during early stages of younger, individual tectonic cycles. In contrast, the porphyry coppers developed during the later stages of crustal evolution - later in terms of both absolute age and relative age within younger individual tectonic cycles. They appear in younger rocks generally, and specifically in rocks involved in events of the middle and late tectonic stages of orogenic belts. The massive sulphide deposits are of early-magmatic eugeosynclinal affiliation, whereas the porphyry deposits formed in late-magmatic, predominantly subvolcanic settings at the edge of continental plates, where contemporaneous exhalative volcanism was mainly a subaerial phenomenon. Transitional types of deposits exist and are found on continental margins with flanking eugeosynclinal sequences."

Jan 1, 1972

By S. Nakamura, M. Takaoka, Y. Mizuno, T. Kusakabe

"Mercury is produced as a byproduct from gold mines and, historically, has been transferred to recyclers and used for various industrial processes and products. However, under the Minamata Convention on Mercury, adopted in 2013,the intentional use of mercury in processes or products will be minimized in the world. Surplus elemental mercury should be treated as a hazardous waste and consequently, should be converted to a more stable chemical form for long-term storage or permanent disposal. A stabilization method has been developed to react mercury and sulfur using a planetary ball mill. The planetary ball mill can produce high-energy conditions and enable samples to react within a short period of time. A series of tests has been conducted to confirm that mercury sulfide can be produced using a bench scale machine with 2400 mL pot. Under a suitable experimental condition with 25 mm ball diameter, 30%of ball filling rate and 1.05 of molar ratio of S/Hg, a powdered mercury sulfide product was successfully obtained in 60 minutes. The leach ability and atmospheric emission from the product was extremely low. This stabilization method would contribute to an environmentally sound mercury management strategy for the gold industry.INTRODUCTIONDue to its unique physical and chemical properties, mercury has numerous applications, including in industry and medicine. However, mercury and its compounds are highly toxic. For example, in Japan, methyl mercury in waste water caused Minamata disease in the 1950s, which affected large numbers of people (Minamata Disease Municipal Museum, 2001). In recent years, environmental pollution caused by emissions and releases of mercury into the atmosphere, waterbody and land by various emission sources such as coal fire plants, artisanal and small-scale gold mining (ASGM), non-ferrous metal industries, waste incinerations etc. has also become a significant problem. To protect human health and the environment from anthropogenic emissions and releases of mercury and its compounds, the Minamata Convention on Mercury was adopted in 2013(UNEP, 2013)Because ground gold ore typically contains mercury at 0 to 100 ppm, mercury is often produced as a byproduct from gold mines (UNEP, 2016). In gold mining operations, mercury can be released from the ore by roasting or by leaching in cyanide solutions and recovery onto activated carbon. According to Yanget al. (2016), 85.1%of mercury in a concentrate was vaporized during a roasting process. Both processes finally discharge elemental mercury as a byproduct. In the USA, more than 100 ton of mercury was discharged from large-scale and industrial gold mining (Miller, 2004)."

Jan 1, 2017

By T. A. Rickard

FROM AN OLD MANUSCRIPT THE text for this contribution to the history of the Patio process is a page taken from a Spanish manuscript of Bartholomé Arzaj Sanches y Uela, now in the possession of the library of Brown University (see Figure 1). I owe the photographic print to the courtesy of the librarian. The date, as will be seen, is 1577. The subject is the treatment of silver ores at Potosi, in old Peru, but in modern Bolivia. The group of drawings (see Figure 1) is entitled "Arrangement of the machines in which the ores of silver are ground on the bank of Potosi, the one of two batteries and the other of one". First, top-left, we have the sketch of a "machine of one battery" operated by a water-wheel. From this mill the water runs in part along an aqueduct supported by pillars to the two-battery mill and in part down the hillside, where men are at work with "tubs, which are called cochas, in which are washed the charges of ore and amalgam". Along the top of the page are shown the "various offices in which are stored the copper, the salt, the ucha [llama dung, used for fuel], the lime, and other materials for the re-duction of the ores". To the right of these are "apartments and warehouses in which is stored the ore brought down from the mountain"?the cerro de Potosi. Through the arched portal come the "sheep [llamas] on which is brought down the ore". To this is added a reference to "the second gate, which closes the whole works". On the level above the aqueduct are seen six men carrying skin-bags that contain water. This is being sprinkled on the ore. On the extreme right are six more of these water-carriers. Next comes the two-battery mill. The mortero, or mortar, is marked on the right. Another description says, "Stone, on which the stamps crush the ore, which is called a mortar". To the left of these are indicated "screens of wire in which the ore is sifted after it has been crushed". Farther to the left, near the ore-washers, is a "shed, in which is placed the metallurgist to test the ore of the boxes". Along the left side of the page are sundry buildings, one of which is labelled a "shed in which are placed the cones of silver". Near it, lower, to the right, is the "furnace, in which are roasted the sulphides to facilitate treatment". Farther to the right is the "place where the Indians mix the charges of sulphide ore, which is called a buitron". Alongside is a "buitron in which the charges are prepared for reduction". Inside is a container, to which are attached the words "measure, called topo" On the extreme right is "another buitron in which the Indians mix the ores with their feet, as well as the heaps for the preparation of the reduction". Near the bottom, on the extreme left, is the "chapel where the holy sacri¬fice of the mass is celebrated". Under the first buitron is the "place where they dry the moist ores, called pãpear".

Jan 1, 1935

By J. Jerome MacDonald, Stephen B. Flewelling

"In the base metals industry of today, operations required to continuously improve productivity and reduce unit production costs. Consistent with these goals is the improvement of the muck haulage cycle from open stopes with the No. 12 Mine.Existing haulage from open stopes occurs with LHD units experience at Brunswick indicates that LHD units have a higher unit cost per tonne mucked than loader-truck combinations when the haulage exceeds approximately 150 m.The reintroduction of open stoping at Brunswick made existing loader-truck combinations unsuitable. It was recognized that application of recent continuous loading technology could potentially result in significant unit cost savings and productivity increases.A review of the technology resulted in the testing of the Voest Alpine AL60 loader in the open stopes located on the 1000m Le vel. Testing began in January 1988 and finished in late November 1988. The test will provide experience as well as quantified production and maintenenace parameters f or the purpose of quantifying production unit costs and system productivity, ultimately resulting in a decision to further utilize or not utilize the loader as a primarymucking unit at Brunswick .INTRODUCTIONBrunswick Mining and Smelting Corporation Ltd. produces approximately 10500 tonnes per day from the No. 12Mine located 35 km southwest of Bathurst, New Brunswick Mining is the largest (single mine) producer of zinc in the world having mined over 52 million tonnes since 1964. Many years of production still remain with proven and probable reserves in excess of 90 million tonnes. Reserves grade at an average of 3.65% Pb. 8.99070 Zn, 99g/mt Ag and 0.32% Cu . The operation employs 1446 people with approximately 800underground including operations and maintenance staff.In the No. 12 Mine two major mining methods are presently utilized . The primary method is Mechanized Cut and Fill (MCF) accounting for approximately 6500 tonnes per day from four major areas. The remaining 4000 tonnes per day is produced from open stopes with the majority of 2800 tonnes per day being produced by the 1000 level.The 1000Level contains 30 million tonnes segmented into open stope blocks 30 m high, 15m and 30 m on strike , with a transverse width of 10m to 50 m (Fig. I). Stopes are organized on a primary secondary stope arrangement along strike. Primary stopes are undercut and overcut full size with the back fully cablebolted. All stope and lateral development is driven 4.0 m high. All stopes to date are flat bottom (Fig. 2). Primary stopes are filled with cemented rock- fill , secondaries with conventional rockfill, Haulage from the stopes is primarily with 8-yard LHD units ; mucking is approximately 50% conventional and 50% remote. Haul distance to date has averaged approximately 150m with current stope haul distances of approximately 300m. Haulage is presently to a small jaw crusher situated on the 1000 Level. From August 1989on, haulage will be to a set of ore passes feeding to a large gyratory crusher situated on the 1200 Level. Start up of the gyratory crusher corresponds to an increase in tonnage from 1000 Level to 4100 tonnes per day."

Jan 1, 1989

By W. Hrynewich

"The Central Canada Potash facility, a division of Noranda Metal Industries Limited, is located 3 kilometres south of the junction of Highways 16 and 2, about 75 kilometres south-east of Saskatoon, Saskatchewan. The site occupies 26~ hectares of land and includes well-equipped machine and electrical _shops, a 325,000-tonne muriate storage and reclaim warehouse, a stores warehouse, processing plant (mill), two mine shafts (one for production, the other for servicing and personnel transport), as well as a two track product loading complex. Access to the site is by road and railway (CN and CP Rail). Suppliers of operating and maintenance use both means of transportation, whereas muriate shipments are entirely by railway.Currently five continuous mining machines - one a Goodman 526 borer and four Marietta 911-A units produce some 10,500 tonnes of ore per operating day. A smaller Voest-Alpine Af.1-100 boom type miner is used for floor cutting which was previously done by a production unit. In order to maintain safe, competent ground conditions only 35 percent of the total ore is extracted. At this rate mine reserves are deemed adequate for about 100 years. The plant has operated since August 1969 and has a current practical annual capacity of 1,270,000 tonnes of muriate. As of June 30, 1981, some 11.9 million tonnes of muriate had been produced from 31.4 million tonnes of ore. Of the total muriate, 6.1 million tonnes was granular, 4.8 million was coarse, and 1.0 million was standard. No standard product has been made since August 1978."

Jan 1, 1982

By R. P. Nicholson

On September 3rd, 1947, the Acadia Coal Company put into operation a central coal preparation plant. This plant was designed to process all coals mined by this Company in the Pictou County coal field. The problem, briefly, was to make a three-way separation of the raw coal fed into the plant. It was desired to remove all stone, metal, and wood from the marketable coal and to produce a splint coal from the raw material which would be suitable for making power. This splint coal product is belted directly to the power plant which is adjacent to the wash plant. To attain this object it was therefore necessary, first, to wash the raw coal, removing all splint, stone, metal, and wood. The stone, metal, wood, and highest-ash splint had then to be removed by a second washing in order to obtain the lower-ash splint for use in the power plant. The plant is on the east side of the Allan colliery, in the town of Stellarton (see accompanying photograph). It is of steel frame construction with wood siding, covered on the exterior with corrugated galvanized iron. Considerable concrete had to be poured for the foundation piers as the plant site is largely 'fill' ground. This necessitated the construction of extremely heavy footings containing a large amount of reinforcing steel. In all, 884 cu. yds. of concrete were poured for the various pits, tanks, and foundations; its average test was about 3,500 lb. per square inch. At the time the plant was constructed, the steel shortage in Canada and the United States was acute and, in order to obtain steel for the structure which is four stories high- an unused part of the Allan bankhead was torn down. Enough steel was obtained from this demolition to fabricate the new plant.

Jan 1, 1948

By J. H. Crocket, J. Andy Fyon

"All gold deposits within the Tisdale Group are spatially associated with carbonate alteration zones; however, not all carbonate alteration zones are host to gold mineralization.Within the komatiitic and magnesium tholeiitic volcanic rocks at the base of the Tisdale Group, discordant and stratabound carbonate alteration zones occur throughout. The alteration zones consist of a ferroan dolomite - quartz inner assemblage which is enveloped by calcite-quartz and an outer chlorite-rich subzone. Factors which control the distribution of carbonate alteration include joint systems, flow contacts, synvolcanic faults and bulk composition of the host rock.Two environments of carbonate alteration can be recognized: 1) The subsea-floor alteration zones developed within the volcanic rock succession where hydrothermal fluids did not discharge directly onto the sea floor. Insignificant gold production has come from this alteration environment. 2) Seafloor zones, which represent points of hydrothermal fluid discharge into the hydrosphere, are spatially associated with felsic volcanic rock complexes and are host to syngenetic, auriferous cherty dolomite.Carbonatized, magnesium tholeiitic basalts, which are host to the auriferous, cherty dolomite, are enriched in antimony (>0.35 ppm), boron (>30 ppm), gold (>5 ppb) and lithium (> 35 ppm), but are depleted in copper ( < 70 ppm), with respect to carbonatized basalts not associated with gold. Orerelated, altered komatiitic flows can be screened using gold (>5 ppb), lead(> 10 ppm), arsenic(> 70 ppm) and antimony (>1 ppm).The habit, petrographic characteristics and certain geochemical characteristics of the altered volcanic rocks suggest that the hydrothermal fluid was of sea-water origin."

Jan 1, 1982

By K. R. Brown, P. L. Krawchuk

In-situ recovery (ISR) is under increased investigation as a viable methodology for the extraction of minerals in ore bodies that may be uneconomic to mine by traditional methods or for resources situated in locations where underground and surface mining operations are not feasible. Although the concept of ISR is relatively new with respect to hard rock applications, it is not a new technology; it has been used successfully in potash and uranium mining and also in heavy oil extraction for years where it has been proven to be both technically and commercially viable. ISR allows for the extraction of metals and minerals without the need to mine large amounts of ore that need to be processed at surface. The ore stays in-situ and only the valuable metals and minerals are produced to surface in a process that produces very little to no waste rock or tailings. This paper aims to discuss the geological conditions that make ISR amenable to an ore body as well as the application of maturing technologies such as advanced directional drilling and the creation of artificial permeability that will help to access deep ore bodies and enhance or create favorable recovery conditions for ISR in the ore body. INTRODUCTION In-situ recovery (ISR), also referred to as in-situ leaching, is a potentially disruptive technology for the mining industry. ISR is under increased investigation as a viable methodology for the extraction of valuable minerals that may be uneconomic to mine by traditional methods or for deposits situated in locations where underground and surface mining operations are not feasible. Although the concept of ISR is relatively new with respect to hard rock applications, it is not a new technology; it has been used successfully in potash mining, uranium mining and also in heavy oil extraction for years where it has been proven to be both technically and commercially viable. Recently there has been increased interest in deploying ISR in hard rock applications due to a number of factors: the drive for environmental stewardship is stronger than ever, technological advances in drilling, blasting, and leaching are strengthening ISR economics, and a large inventory of ore bodies exists that are not feasible to mine or extract using conventional open pit or underground mining techniques. ISR allows for the extraction of valuable metals and minerals without the need to mine large amounts of ore that need to be processed at surface. The ore stays in-situ and only the valuable metals and minerals are produced to surface in solution via a recovery process that produces very little to no waste rock or tailings. Where the right conditions exist, the ISR process has a much smaller environmental footprint (Hiam-Galvez, 2018) and potentially much lower capital and operating costs than using conventional mining techniques.

Jan 1, 2019

By Richard J. H. Barclay

The Lac Otelnuk iron property is located in the Province of Québec in the Northern Labrador Trough. The Labrador Trough is an extensive geosyncline of Proterozoic rocks, unconformably underlain by Archean gneisses. The Lac Otelnuk iron property hosts two significant magnetite iron-formation taconite deposits that outcrop, are sub-horizontal and are known to extend for at least 25 kilometres along strike. The host sedimentary rock package that includes a cherty iron formation is weakly metamorphosed to greenschist facies.

May 1, 2007

By C. S. R. Rondestvedt

In February 2010, Teck?s newest Concentrator at Mina Carmen de Andacollo (CDA), Andacollo, Chile was commissioned. The Concentrator is a single line configuration, comprising of one 36 ft diameter x 20 ft long SAG mill driven by an 18,000 hp gearless drive. The SAG mill feeds two ball mills each 25 ft in diameter x 40.5 ft long, each driven by a total of 19,000 hp through dual pinion drives. A series of four phases will be reviewed from operational readiness, mechanical completion, commissioning, and ramp-up. Also, now after 6 months of operation, this paper will review the significant steps taken during ramp-up enabling the announcement of commercial production, in October 2010.

Jan 1, 2012

By C. Jay Hodgson, P. J. MacGeehan

"Of the 135 current or past-producing 'gold-only' mines in the Superior Province of the Canadian Shield which have produced more than 100,000 tons of ore, the 33 mines which have produced one million ounces or more of gold have accounted for over 84o/o of the total production of 137 million ounces. The published literature on these 33 deposits was reviewed in order to determine the geological setting of the mineralization on a regional (mining camp) and local (mine) scale. The dominant rock types underlying the mining camps are mafic, variolitic, tholeiitic volcanic rocks with a significant ultramafic (komatiitic) component, and clastic and/ or exhalative sedimentary rocks. Felsic volcanic rocks are generally subordinate. Most mining camps are located at or near the top of the mafic-ultramafic sequence, and generally there is a major sedimentary-volcanic contact present which may be 1) a stratigraphic superposition of volcanic and sedimentary units, 2) a rapid, along-strike gradation from dominantly volcanic to dominantly sedimentary units due to facies change, or to the formation of structurally controlled sedimentary basins during volcanism, or 3) frequently (at 40o/o of the camps) a juxtaposition of sedimentary and volcanic sequences across a major fault zone or 'break', which in many instances may represent a reactivated, syn-volcanic basin-margin fault.In contrast to the regional setting, on a local or mine scale over 90o/o of the 33 mines contain felsic intrusions and/ or extrusions within mine workings. The felsic rocks may be quartzbearing (commonly quartz-feldspar porphyries) or dominantly quartz-free (commonly feldspar porphyry syenites or monzonites), and appear to have been emplaced in two different depth-zone environments: an epizonal intrusive environment, and a volcanic vent environment. Although the distinction between epizonal intrusive and volcanic intrusive-extrusive rocks is uncertain in many deposits, the two are tentatively separated on the basis of the grain size of the rock matrix (medium to coarse grained vs aphanitic to fine grained), homogeneity (multi-phased and/ or compositionally zoned vs compositionally homogeneous), and the shape and relationship to stratigraphy of the igneous body (circular or oval in cross section, and cutting the stratigraphy at a low angle vs lenticular and either stratiform or cutting the stratigraphy at a high angle). In addition, felsic intrusions of the volcanic environment commonly grade into tuffs or breccias of extrusive origin, whereas epizonal intrusions do not.Stratiform ore zones of possible exhalative sedimentary origin appear to be confined to deposits in which volcanic felsic rocks are present, and are absent from deposits in which only epizonal felsic intrusions are present. Pervasive carbonate- rich alteration, commonly in roughly stratabound zones of regional extent, and carbonate-rich chemical sedimentary rocks are present in over three-quarters of the deposits containing quartz-bearing felsic volcanic rocks, but are much less common in deposits with quartz-free felsic volcanic rocks. The host sequences of epizonal felsic intrusions which are quartzfree are more commonly sedimentary than those of quartz-bearing epizonal felsic intrusions.Quartz, carbonate minerals, arsenopyrite, scheelite, tourmaline and Cr-muscovite ('juchsite') are the most common mineralogical associates of gold. Arsenopyrite is essentially restricted to the volcanic environment, occurring in about twothirds of the deposits with quartz-bearing felsic volcanic rocks and less than half of the deposits with quartz-free volcanic rocks. Tourmaline, on the other hand, is almost ubiquitous in deposits associated with epizonal intrusions and also occurs in about two-thirds of the deposits associated with quartz-bearing felsic volcanic rocks, but in only one of the eight deposits associated with quartz-free felsic volcanic rocks. Scheelite is essentially restricted to ores of quartz-rich fel"

Jan 1, 1982

By J. T. Kalala

Particles for heap leaching in conventional circuits are prepared using Cone crushers. Cone crushers are used to crush the ore to the size range between 12-25 mm for heap leaching operations. Leaching from these relatively coarse particles occurs only at the surface and in subsurface regions, which are accessible from the surface by cracks and pores. Since leaching is active in accessible areas, the method of comminution that forms multiple cracks and pores would offer an advantage for heap leach. An investigation was performed in this study to assess if the HPGR forms more cracks and pores than a Cone crusher. X-ray computed tomography (CT) was used to characterize the degree of cracks and pores formed by these two comminution devices that can be used for heap leaching feed preparation. In order to compare the effect of the comminution method on crack network generation, a bulk sample of the zinc ore was crushed by the HPGR at three different pressure settings (45, 95 and 120 bar), the sample ore was also crushed with a Cone crusher for comparison purposes. Products from the Cone crusher and HPGR experiment were then screened into five size fractions ((-25 +19), (-19+16), (-16+9.5), (-9.5+5), (-5+4.75) mm), Subsamples from the ((-25 +19), (-16+9.5), (-5+4.75) mm) size fractions were analysed. The spatial distribution of the crack network of sphalerite particles was characterized using a highresolution industrial X-ray CT system. This paper describes the use of image analysis techniques including image segmentation, which uses a combination of thresholding and other methods to characterize and quantify crack and mineral dissemination in the sphalerite particles after comminution with the HPGR and Cone crusher. The results are validated with those obtained using traditional techniques such as physical gas (with N2) adsorption, mercury intrusion porosimetry, SEM and QEMSCAN.

Jan 1, 2012

By James J. Holmes

ACCIDENT prevention as we know and understand it today is of comparatively recent origin. As a matter of fact, the Industrial Accident Prevention Association of Ontario, which is the oldest organization of its kind in the Dominion, celebrated its thirtieth anniversary only this year. On the other hand, one could say that the history of accident prevention is almost the history of mankind itself. In a general sense, every forward step that man has made throughout the centuries has been largely due to two things: (1) a desire for increased comfort, and (2) the need for safety; and, always, the need for safety has taken precedence over the desire for comfort. While it is true that hundreds of volumes have been published, dealing with the many aspects of accident prevention work, you cannot go to a library and pick out a book that has all the answers in it. Such a book has never been published-we are still looking for some of the answers. However, a good deal of real progress has been made.

Jan 1, 1945

By A. Lloyd, Montgomery, J. Jorgesen, W. J.

"Coals are ranked according to their degree of metamorphism or progressive change in the natural series from lignite to anthracite. The reader is referred to ASTM Standard D-388, ""Standard Classification of Coals by Rank"" for a detailed description of the system in use throughout North America including Canada.Fixed carbon and calorific value (expressed in Btu/ lb) calculated to the mineral-matter-free basis have been chosen as the main characteristics from which rank is established. Higher rank coals are classified according to fixed carbon content, lower rank coals according to the calorific value on a ""moist"" mineral-matter-free basis. Certain adjacent groups are differentiated according to their agglomerating (caking) properties. ""Moist"" refers to inherent moisture. Referring to Table 1, coals having calorific values greater than 14,000 Btu/lb on a moist mineral-matter-free basis and fixed carbon of 69% or more on a dry mineral-matter-free basis are classified according to fixed carbon. Coals having calorific values of less than 14,000 Btu/lb on a moist mineralmatter-free basis are classified according to calorific value on a moist mineral-matter-free basis. Agglomerating coals having 69 to 86% fixed carbon on the dry mineral-matter-free basis are classified in the low or medium volatile group of the bituminous class according to their volatile matter content. Agglomerating coals having calorific values between 10,500 and 11,500 Btu/lb are classified in the high volatile C bituminous coal group; nonagglomerating coals in the same calorific value range are classified in the sub-bituminous A coal group."

Jan 1, 1985

By Alan Reed

"Afton Mines Ltd. (N.P.L.) operates an 8500-tpd open-pit copper-gold mine near Kamloops, R.C. The pit is currently 480 feet deep and will reach an ultimate depth of 800 feet. The east, south and west walls of the pit are composed of Triassic andesites (Nicola Volcanic Formation) and contemporaneous sub-volcanic intrusions of micro-syenite, microdiorite, diorite and picrite-basalt (Iron Mask Pluton). The north wall of the pit consists of Tertiary sandstones, shales and mudstones (Tranquille formation) and dacitic dykes, sills, flows and tuff s (Kamloops Volcanic Formation).The Afton copper-gold orebody (30 million tons of open-pit ore with a grade of 1.0 per cent copper, 0.016 oz/ton gold and 0.12 oz/ton silver) occurs within the Triassic Iron Mask Pluton adjacent to its faulted contact with Tertiary rocks. The Triassic-Tertiary contact is a complex fault system dipping 65 degrees southward and causing the Triassic rocks to overlie the Tertiary rocks. The detailed trace of this contact is very intricate due to the mutual interference of three principal fault systems.The dominant geological structures in both Triassic and Tertiary rocks are faults striking WNW and dipping 65 degrees southward, i.e. sub-parallel to the orebody (strike faults). Cross faults and oblique faults are abundant and highly significant in some sectors of the pit. The Tertiary sedimentary rocks show an early series of northwesterly plunging folds deformed by a later series of northeasterly striking folds. Major lithological and structural boundaries divide the pit into 8 structural domain s within which the pitwall slope stability is controlled by the local geological structures and lithology.Assay data from production blasthole samples and exploration diamond drillhole samples show a log-normal frequency distribution, with a fairly low relative variance. Variograms computed from production blasthole samples show good continuity with elliptical horizontal zones of influence which are truncated by the major faults that bound the ore zone."

Jan 1, 1983