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By J. E. Dutrizac, J. B. Sunstrum

"Lead was among the first metals known to man, and there are indications that lead was produced by the reductive smelting of galena in the mountains surrounding the Black Sea as early as 3000 B. C .(1). The early metallurgists soon discovered that silver could be recovered from lead bullion by cupellation techniques, and that the litharge product of the cupellation process could be reduced to give silver-free lead metal. By the Graeco-Roman era, lead-silver ores were being extensively exploited for both their silver and lead contents. Although silver was always of prime concern, a number of industrial uses existed for lead that included piping and sheeting. Subsequently, lead metallurgical practice developed steadily in different parts of the world, and was a relatively mature technology by the time of the European colonization of North America. In the United States, lead was produced on a small scale as early as 1720, and by 1800, the extensive Pb-Zn deposits of Missouri-Kansas and the upper Mississippi Valley were being developed(2). Many of the early American lead ores were low in silver, and seemed to have been worked primarily for their lead contents. Major uses of lead at the time included bullets, pipe and sheeting. During the period 1860-1890, the silver-rich galena ores of the western United States were exploited, and a number of lead smelters were operated. Although all the smelters produced lead metal, the recuperation of the associated silver and gold values was generally of prime concern. In Canada, lead deposits are abundant, and galenarich outcrops frequently attracted the attention of the early explorers and prospectors. In fact, the first record of a Canadian ore deposit seems to have been the observation of an argentiferous galena vein near Lake Temiscamingue in 17 41(3). The presence of silver-bearing lead deposits naturally attracted the attention of the mining community, and many outcrops were worked throughout the country. Because of the problems associated with transporting lead concentrates to existing American or overseas smelters, attention soon shifted to the establishment of local lead smelting facilities. The purpose of this article is to examine those early smelting efforts and to try to identify the first ""successful"" Canadian lead smelter."

Jan 1, 1999

By R. G. L. McCready, R. O. van Everdingen, M. Kalin, R. van Everdingen

"This paper is based on data generated between October, 1986 and October, 1989, during the implementation of an Ecological Engineering close-out for a zinc mine in Northern Ontario, where 690000 tonnes of tailings contain 41.1% pyrite and 4.1% pyrrhotite. The objective was to determine rates of sulphide oxidation and iron precipitation in the tailings. Those rates are relevant for the assessment of the long-term effectiveness of biological systems employed in the removal of heavy metals from contaminated drainage issuing from tailings.Data were collected on the topography and stratigraphy of the tailings basin, subsurface temperatures and groundwater levels, and chemical analyses were made of water samples collected from a number of piezometers. The thickness of unsaturated tailings varied widely between individual piezometer locations; annual maxima ranged from 0.67 m to 3.55 m, and annual minima from 0 to 2.68 m. Sulphate concentrations ranged from 36 mg/L to 20180 mg/L; iron concentrations from 0.36 mg/L to 9938 mg/L; and zinc concentrations from 0.33 mg/L to 1123 mg/L.The geochemistry of the piezometer samples suggests that some of the initial Fe content, resulting from metal sulphide oxidation, is subsequently precipitated in the tailings pile and, possibly, in the underlying unconsolidated sediments, probably in the form of iron-hydroxide and Na-jarosite. Other metals such as Zn may be co-precipitated with Fe. The discharge of a portion of these metals to the environment would be delayed by this process. Some of the initial sulphate may also be lost from solution, probably by precipitation of gypsum; approximate saturation degrees for gypsum ranged from -3.27 to +0.47. The precipitated minerals will somewhat decrease the porosity and permeability of the tailings.Sulphide-oxidation rates, calculated using the analytical results and the estimated amount of annual recharge at the tailings site, average 0.0057 mol/kg/yr. Calculated iron-precipitation rates average 0.000386 mol/kg/yr. It is concluded that the characteristics of seepage from mining wastes may change with time, reflecting the solubilization of precipitated secondary minerals."

Jan 1, 1992

By W. A. Jones

"The Hollinger gold mine is in the Porcupine area, Cochrane district, in northern Ontario. The adjacent town of Timmins, which has a population of 25,856, lies 490 miles by railway north of the city of Toronto.The property was staked in the late summer of 1909. The original four claims were staked by Benjamin Hollinger, after whom the mine was named. Development commenced in the autumn of 1909 and mill construction in the winter of 1910. A catastrophic forest fire swept the district in 19ll, destroying the surface plant, so that the original 300-ton mill was not completed until June 1912. From that date up to December 31, 1945, 44,280,750 tons of ore have yielded 13,636,796 oz. gold and 2,613,354 oz. silver. The total value of the production has been $363,638,315.0REBODIES AND MINERALIZATIONThe orebodies consist of quartz veins, stringer lodes, and scockworks, with associated pyritized wall-rocks. The width of individual orebodies, as exposed by mining operations, ranges from 1 to 70 feet and averaged approximately 10 feet in 1942. The vein system forms a complex pattern in a favourable zone, which in places is 6,000 feet long and 4,000 feet wide. The deepest level being developed is 5,150 feet below the surface. Few veins are persistent for long distances, although some are continuous for more than 1,000 feet longitudinally and vertically.The host rocks consist of a closely folded series of Precambrian lava flows of Keewatin age. These are schistose and hydrothermally altered, but are believed to have originally been of intermediate to basic composition. The schistosity strikes N.70°E. and dips to the south at 75° to 80°. The structure of the formations is complex, but in the principal ore-bearing zone the general dip of the flows is to the south at a slightly flatter attitude than the schistosity."

Jan 1, 1954

By J. A. Haskin

The Engineering Department at Flin Flon has the following responsibilities: 1. Underground Surveying. 2. Open-Pit Surveying and Supervision of Blasting. 3. Mine Geology and Examination of Prospects. 4. Design of Plant Additions and Improvements. 5. Miscellaneous Field W ork, Reports, etc. The entire surface plant, as well as the underground workings, are located by co-ordinates based on a triangulation system covering an area approximately one mile long by one-half mile wide. The origin of this system and one end of the base-line is the southwest corner of Township 67, Range 29, west of the Principal Meridian. The bearing of the base-line was determined from the township boundary, which runs true north, so that all other bearings are almost astronomical. This is a great convenience when our surveys must be tied to mineral claims, rights-of-way, and land subdivisions. As the base-line extends across Flin Flon lake, its measurement across smooth ice was simple and accurate. The length, slightly over 1,600 feet, has been adjusted for the temperature and tension of the standard tapes used and checked sufficiently to ensure a maximum error of 1 : 15,000. Each angle was repeated ten times, but as no special refinements were used, errors of as much as four or five seconds are possible. The elevation of the origin of the triangulation system was approximately 1,050 feet above sea level. The use of this true elevation in our surveys would have resulted in negative elevations underground as soon as the mine reached a depth greater than 1,050 feet. To avoid this complication we therefore added 4,000 feet to this elevation. The collar of the Main shaft, for example, which is actually 1,072 feet above sea level, appears in our records as at elevation 5,072. Underground surveying, which follows standard practice, is done by two mine surveyors who share three helpers. The meridian is taken underground from two wires hung in either of the two vertical shafts.

Jan 1, 1935

By R. W. Boyle, W. Dyck

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

Jan 1, 1980

By Terry J. Briggs, Ian J. Kelso

"Abstract- The Century mine in northwest Queensland,Australia, is a major sediment-hosted zinclead- silver orebody, located within the western Mount Isa Basin. Stratiform, carbonaceous shalehosted mineralization is developed within Proterozoic shale and siltstone sequences of the Lawn Hill Formation. CRA Exploration Pty Ltd (CRAE) discovered the Century deposit in 1990 and open pit mine development commenced in 1998 under the ownership of Pasminco Century Mine Limited (PCML). The deposit contains an identified mineral resource of 105 Mt @ 12.1% Zn + 1.7% Pb + 46 g/t Ag (Broadbent and Waltho, 1998) which is currently being mined at a rate of 5 Mt/y.In December 1998, an exothermic reaction was observed between explosives spilt onto shale drill cuttings from a blasthole. This reaction is referred to as ‘reactive ground’ and has the potential to cause premature detonation of explosives in a charged blasthole with catastrophic consequences. The incident led to a subsequent geological and geochemical assessment of the different rock types at the Century deposit and their potential reactivity with ammonium nitrate-based (AN) explosive products. This assessment was conducted in an attempt to characterize zones safe for blasting with uninhibited AN-based explosive products. A direct result of this study is a set of procedures outlining the use of different AN-based explosives in areas defined as potentially reactive or non-reactive for the different lithologies present. Black shales contained in both the waste and mineralized sequences contain the sulfide mineral pyrite (FeS2), which is thermodynamically unstable with AN contained in commonly used explosives. The potential reactivity of partially weathered pyrite can ultimately result in the decomposition of an AN-based explosive. It was determined that only pyritic black shales showed any potential to react with AN, irrespective of their position in the stratigraphic sequence. Other lithologies present at Century, including Proterozoic sandstone and Cambrian limestone, are classified as non-reactive with AN. This is a geological risk management situation present at the Century mine."

Jan 1, 2001

By Morris Wong

"An antimony-bearing gold ore was studied in a cyanide slurry containing 7 -20 ppm dissolved oxygen. 1 g/L NaCN at pH 9.5 and 48 hours retention time is required to extract 94 % of the gold from the 90 % minus 200 mesh ore by conventional cyanidation at a 7 ppm O2. As oxygen concentration was increased to 20 ppm, the conditions required to extract the same amount of gold were: 0.25 g/L NaCN, 16 hours retention time and 80% minus 200 mesh. Compared to conventional cyanidation, oxygen-enriched leaching was shown to reduce retention time by three times and cyanide consumption by four times. The kinetics of gold dissolution related to the dissolved oxygen concentration in slurry was expressed by following equation:XAu = 100 * (Kt + I0) / (1 + Kt + I0)where XAu is the percentage gold extraction; t is the retention time; K and Io are the slope and the intercept of the plot of XAu I (100 -XAu) vs. t respectively. The values of Kand Io depend upon the concentrations of cyanide and oxygen, the fineness of grind and the origin of ore sample etc.. A preliminary cost analysis was also performed to justify the use of oxygenating slurries.IntroductionGeneral:It was proposed almost 149 years ago (Elsner 1846) that oxygen is required for the dissolution of gold in cyanide solution. However, little attention were paid to oxygen application in the gold milling industry during the following one hundred years. Due to the constraints of gold price, reagent costs, refractory ores, and environmental issues in the past decade, the oxygen enhancement in the cyanidation process was introduced to the gold milling industry for increasing operation efficiency, reducing production costs, and meeting the requirement of environmental standard.The advantages of oxygen-enriched cyanidation have been reviewed by several investigators (Hague 1992, Jara et al 1992, Henderson 1989, Loroesch et al 1988, Elmore 1987, Cathro 1963). Common to all conclusions are the following advantages."

Jan 1, 1995

By Claudio Rocha Rosales

"Peñoles is a mining group with integrated operations for the smelting and refining of non-ferrous metals, and it is also a chemicals producer. Peñoles is the world’s foremost producer of refined silver, metallic bismuth and sodium sulphate. It is also the leading Latin American producer of refined gold, lead and zinc. The Met-Mex Peñoles Zinc Plant commenced operations in 1973 with an annual production capacity of 105,000 t/y zinc. Since then, production capacity has increased, and at present, it stands at 240,000 t/y zinc. In this paper, the leaching process and particularly the operation of the jarosite circuit is reviewed, specifying process improvements to treat high iron calcines.INTRODUCTIONTwenty years ago the Electrolytic Zinc Plant in Peñoles was treating high quality zinc concentrates with low concentrations of iron (~7.5 wt.%) and silica (1.8 wt.%). Nowadays the quality of zinc concentrates is poor with a typical iron content of around 11 wt.% and silica at 2.0 wt.%. This has created numerous processing problems in our roast-leach-purification- electrolysis sections. In order to meet customer requirements it is necessary to find processing solutions to assure the operational continuity of the plant. These processing solutions are the focal point of this paper.REVIEW OF THE LEACHING PROCESSThe purpose of the leaching process is to recover the metals, mainly zinc, through a series of extraction stages with sulphuric acid solutions, which progressively increases the zinc dissolution. After the series of acidification stages, the solution, which is rich in iron, is sent to the jarosite circuit to separate zinc from iron (Garcia & Valdez, 1996). The first step of the leaching process is the neutral leach circuit which produces a zinc sulphate solution that is sent to the purification section for the removal of impurities such as cadmium, cobalt and copper. Then this solution is sent to the tank house to feed the electrolytic cells."

Jan 1, 2016

By R. V. Typliski, G. J. Labarre

"Hudson Bay Mining and Smelting Co., Limited, Flin Flon, Manitoba, has developed a metallurgical wastewater treatment system employing two-stage lime neutralization, sludge recirculation and conventional mechanical thickening.Laboratory investigations, initiated in 1970, led to the construction of a pilot-scale minewater treatment plant, at Schist Lake Mine, in early 1971. This plant, treating 600 cubic metres /day of wastewater containing 30-60 mg/1 zinc. demonstrated metal extraction efficiencies of 99% under normal operating conditions.On the basis of this pilot work, a second pilot-scale treatment system was established in 1975 at the Flin Flon metallurgical plant to treat zinc refinery process waste water. Through numerous improvements, including automatic con¬trol equipment, the system has evolved to a full-scale operation capable of treating 800 cubic metres/day of wastewater containing 4-11 gm/1 zinc. Treated effluent contains 0.2-20 mg/1 zinc. The hydroxide sludge produced returns to the metallurgical process for the recovery of metals while clarified effluent discharges to waste.With the system, Hudson Bay Mining has achieved its original goal: the reduction of environmental contamination, while realizing a return from metal recovery.IntroductionWith the goal of reducing environmental contamination, Hudson Bay Mining and Smelting began laboratory testwork, during the late 1960's, into the methods and technology for neutralizing and removing heavy metals from acid mine drainage. Based on lime treatment, early tests showed much promise and culminated in the construction of a pilot-scale test plant at Schist Lake Mine in early 1971.This pilot system utilized technology developed in laboratory studies, and related to work then underway by state authorities in Pennsylvania coal mines. The basic operation consisted of a lime slurry tank, baffle reactor and a free-settling basin. Through three summers of operation, the Schist plant, capable of treating 600 cubic metres/day of minewater, produced a heavy-metal extraction of 99% under normal conditions. Although successful in proving the treatment chemistry, the plant displayed deficiencies in sludge densification and handling."

Jan 1, 1980

By Robert B. Parsons

"The mining industry is adjusting its sights to take a shot at the rising level of non-profit taxes which threaten to handicap the industry 's competitive advantage. In the case of this particular expedition, the industry evidentally has a strong ally - the country's mines ministers. It appears that an initiative is under way to assess the impact of non-profit taxes on mining in Canada as compared to other mining nations.IntroductionSeventy-five years ago, Canadians were unceremoniously introduced to something called income tax. Since then, income tax has been identified as the source of countless evils ranging from the common cold (attributed to stress) to the demise of the economy. Some years later, the provinces jumped on the bandwagon with their own income taxes, and later on with mining taxes. These profit-based taxes had some appeal to elected officials and to a segment of the populace. The onus of the taxes varied directly with the ability to pay (in theory, at least). Previously, the principal sources of government revenues had been property taxes and customs duties, which were viewed as possibly being somewhat regressive in nature.The mining industry has always had concerns about different aspects of income and mining taxes - the tax rate is too high or the depreciation rate is too low, and so on. Recently, though, the industry has turned its sights on the expanding array of taxes which are not based on profit, including payroll taxes, capital taxes, water taxes, sales taxes, fuel taxes, property taxes, environmental levies and user fees. The industry is seriously concerned that the rising level of non-pro fit-based taxes is not only hampering the industry 's ability to reinvest in Canada, but is also placing Canada at a disadvantage in terms of trying to attract new investment capital."

Jan 1, 1992

By Chester J. Kuryliw

"The Geology of the Cochenour Willans mine, Red Lake district, was described by J. P. Nowlan in the Jubilee Volume. The complexity and intense alteration of the rocks obscured any clear relationship between ore and structure, and Nowlan recognized that future work might greatly modify the ideas then existing. This has proved to be the case. Subsequent detailed work has established the stratigraphic sequence in the Keewatin formations which are the host rocks of the orebodies, and has shown that the principal orebodies are confined to two adjoining bands of tuff.On this basis, it is now possible to map the complex structures with greatly increased accuracy. The amount of drilling and underground exploration required to find new ore has been much reduced, and indicated ore reserves are now substantially greater than at any time in the past. Daily production in 1955 averaged 187 tons per day at 0.578 ounces of gold per ton.The sequence of Precambrian events in the mine area is summarized in the following table:Intrusion of post-ore dykes.Ore formation.Intrusion of post-faulting, pre-ore dykes.Deformation: folding and shearing; alteration; faulting.Timiskaming sedimentation.Unconformity.Keewatin: volcanics and sediments, host rocks of the orebodies, described in detail in the next section.KeewatinFrom oldest to youngest, the Keewatin rocks in the mine area can be divided into six formations, which have been given local names:“Grey"" LavaA fine-grained dacitic to andesitic lava, amygdaloidal and locally containing well formed pillows. Certain areas are sheared and chloritized to a degree that makes identification difficult. The rocks are carbonatized throughout; the most persistent carbonate veins parallel the upper contact. Thickness more than 1,000 feet. (Minor mineralization, type 1 I -See section on Ore Formation)."

Jan 1, 1949

By Andrew Neale, Len Anderson, Rob Perry

"P.T Freeport Indonesia's new SAG mill commenced operation in February 1995. The reclaim system consists of four apron feeders and a series of three conveyors. The feeders are used in various combinations and relative speeds to blend SAG mill feed. It was recognized during design of the control system that effective automatic control would require dead-time and gain compensation. The successful implementation of the Dahlin algorithm to provide variable dead-time and gain compensation is described and compared to Pl control.RésumeL'opération du nouveau SAG de P.T. Freeport Indonésie a commencé en Février 1995. Le système de réclamation du minerai comprend 4 alimentateurs à tablier et une série de 3 convoyeurs. Les alimentateurs sont utilisés dans une combinaison variée et à des vitesses relatives différentes pour mélanger l'alimentation du SAG. II a été reconnu pendant la conception du système du contrôle que le contrôle automatique demandera une compensation pour le temps de délai et le gain. L'implantation de l'algorithme de Dahlin a été un succès et ii est comparé avec le contrôle Pl.ResumenP.T. Freeport Indonesia inició la operación de un molino semi autogeno (SAG) en febrero de 1995. El mineral es alimentado desde cuatro alimentadores tipo apron feeders y una serie de correas transportadoras. Los alimentadores se utilizan en variadas combinaciones y velocidades para mezclar apropiadamente el mineral que alimenta al molino SAG. Durante el diseño del sistema de control se detectó que el control automático requería de un tiempo muerto y compensaci6n de ganancia. En este documento se describe la implementación del algoritmo de Dahlin, que ha sido exitosa y ha permitido proveer tiempos muertos variables y compensaci6n de ganancia. Ademas se hace una comparaci6n con el Pl de control."

Jan 1, 1996

"Industry Guide 7 contains the Securities and Exchange Commission’s basic mining disclosure policy. It provides guidance for Registrants on various matters including the definition and disclosure of proven and probable reserves and the definitions of exploration, development and production stages. The definitions and disclosure instructions contained in this Guide apply to all public mining entities and their public disclosures. As NI 43-101 is law in Canada, the SEC allows mineral resources and mineral reserves prepared under CIM Definition Standards for issuers that are reporting in Canada but are also registered with the SEC. The SEC places conditions on the disclosure of these estimates. US shareholders are cautioned that the requirements under NI 43-101 and the CIM Definition Standards differ significantly from those of Industry Guide 7.Note: this Guide contains the Commission's basic mining disclosure policy. The definitions and disclosure instructions contained in this Guide apply to all public mining entities and their public disclosures pursuant to the rules of Regulation S-K. Questions regarding the application of these definitions and rules to specific situations should be asked of the Commission's mining engineers or geologists at either the Office of Engineering, Division of Corporation Finance, 202-272-3252, or the Branch of Geology, Central Regional Office, 303-844-1000.(a) Definitions. The following definitions apply to registrants engaged or to be engaged in significant mining operations:(1) Reserve. That part of a mineral deposit which could be economically and legally extracted or produced at the time of the reserve determination.Note: Reserves are customarily stated in terms of DoreD when dealing with metalliferous minerals; when other materials such as coal, oil shale, tar, sands, limestone, etc. are involved, an appropriate term such as Drecoverable coalD may be substituted.(2) Proven (Measured) Reserves. Reserves for which (a) quantity is computed from dimensions revealed in outcrops, trenches, workings or drill holes, grade and/or quality are computed from the results of detailed sampling and (b) the sites for inspection, sampling and measurement are spaced so closely and the geologic character is so well defined that size, shape, depth and mineral content of reserves are well- established.(3) Probable (Indicated) Reserves. Reserves for which quantity and grade and/or quality are computed from information similar to that used for proven (measured) reserves, but the sites for inspection, sampling, and measurement are farther apart or are otherwise less adequately spaced. The degree of assurance, although lower than that for proven (measured) reserves, is high enough to assume continuity between points of observation.(4)(i) Exploration StageDincludes all issuers engaged in the search for mineral deposit (reserves) which are not in either the development or production stage."

Jan 1, 2003

By O. S. Perry, G. H. Charlewood, K. Griffin, S Ogryzlo, H. Hopkins, G. MacIntosh

"The gold mines of Kirkland Lake are in Teck township, Timiskaming district, Ontario, latitude 48°9'N. and longitude 80°3'W., at an elevation of 1,038 feet above sea-level.There are seven contiguous producing mines in a length of 3;1/2 miles along the general east-west strike of the deposits. The seven mines are, in order from west to east, Macassa, Kirkland Lake, Teck-Hughes, Lake Shore, Wright-Hargreaves, Sylvanite, and Toburn.The first discovery of gold in the zone was made by W. H. Wright in 1911 on the east claim of the Wright-Hargreaves property, and the following year the Tough brothers and H. Oakes found gold in what is now the No. 1 vein_ at the Toburn. Production began at the Toburn in 1915, followed by Teck-Hughes in 1917, Lake Shore in 1918, Kirkland Lake Gold in 1919, Wright-Hargreaves in 1921, Sylvanite in 1927, and Macassa in 1933.To the end of 1946, 31,938,350 tons of ore had been milled at the Kirkland Lake mines, from which 16,132,700 oz. gold, having a value of $498,517,600, were recovered. The bullion varies from mine to mine between 800 and 900 fine, the average for all the mines being 853. The average fineness of the native gold is 941.The workings of some of the mines are more than 6,000 feet deep and sinking to depths below 7,000 feet is now in progress.GENERAL GEOLOGYThe regional structural geology of the area in which the mines of Kirkland Lake occur is described in a preceding paper in this volume by J. E. Thomson."

Jan 1, 1954

By G. W. H. Norman

"The Indian Molybdenum deposit occurs in granite near the centre of Indian peninsula, Kewagama lake, in Preissac township, Abitibi county, Quebec, about 40 miles east of Noranda. Quartz veins containing molybdenite were discovered in the vicinity of the deposit in 1901 by J. F. E. Johnston of the Geological Survey of Canada. Innumerable trenches and a few pits were excavated during and prior to the first great war, but there was no production. During the second great war systematic drilling along a favourable zone disclosed by the earlier trenching outlined an orebody of 225,000 tons with a molybdenite content of 0.46 per cent.A 500 on mill was built in 1943, and 91,489 tons of ore underwent treatment, from which 699,568 lb . of molybdenite was recovered. The mill closed down early in 1944 when the wartime need for Canadian molybdenum production eased .GEOLOGYThe general geology in the vicinity of the deposit is shown in Fig. 1. The deposit lies at the north edge of a muscovite-biotite granite mass about 2 miles wide and several miles long. West of Kewagama lake this granite is cut by great swarms of quartz veins . The only other granite in the district known to contain similar swarms of quartz veins, and which similarly has become a source of mo!ybdenite, occurs 20 miles east in La Corne township. The non-molybdenite-bearing granites of the district are either more basic and contain frw dykes except aplite or are distinctly pegmatitic types and contain up to 50 per cent pegmatite dykes with molybdenite only here and there in small amount."

Jan 1, 1954

By M. E. Sirkka

"INTRODUCTIONInco Ltd. has ten operating mines in the Sudbury district. Over the last ten years mining methods have changed at Inco with special emphasis on the introduction of bulk mining replacing conventional mechanized cut-and-f ill and undercut-and-fill methods. At the same time, the technology and size of LH D equipment has also progressed to meet the demands of bulk mining. This paper will concentrate on various aspects of the introduction and operation of electric LHD equipment a t Inco Limited, Ontario Division. BackgroundInco Limited received its first electric 2 cu. yd capacity LHD in July 1975 at Creighton Mine. The unit was tested pulling shrinkage stope muck and then was moved to a mechanized cut and- fill complex on 6600 Level. The electric scooptram received quick acceptance from the operators, due to its quiet operation and low heat generation, in a relatively hot operating environment. Although there are only two of these units operating a t present, this initial favourable experience, both in cost and efficiency, led to the development and purchase of electric scooptrams of the 5 cu. yd size. There are currently eleven 5 cu . yd electric LHDs in operation.The most recent development a t Inco is the move to yet larger electric LHDs in the 8 cu. yd size. The first electric ""Eight"" has now been in operation at Copper Cliff South Mine for over a year. This unit has the distinction of being the first one of its kind in service in North America. At the time of writing, there were ten 8 cu. yd electric ma chines operating in Inco's Ontario Division Mines, with three on order."

Jan 1, 1989

By Jun Shan

The preparation of poly(N-isopropylacrylamide)-monolayer-protected clusters (PNIPAM-MPC) of gold nanoparticles was carried out in a homogeneous phase using three ways, in which three types of PNIPAM ligands were employed. The first type comprised of PNIPAMs with narrow molar mass distributions, synthesized by reversible addition-fragmentation chain transfer (RAFT) polymerization and thus, bearing a dithiobenzoate at the chain end. These polymers were used directly to passivate the gold nanoparticles upon the Schiffrin reaction in a one-pot synthesis. The second type of ligands was derived from the first one through hydrazinolysis, and they therefore contained a thiol end group. The third type of ligands was PNIPAMs obtained through conventional radical polymerization, post-modified to contain thiol end groups. The PNIPAM-MPCs were characterized by high-resolution transmission electron microscopy, and by dynamic light scattering. The one-pot synthesis utilizing the ligands of the first type turned out to be a simple and facile method, compared with the other two ways, with which the size of the gold nanoparticles can be easily manipulated mainly by adjusting the molar ratios of PNIPAM / HAuCl4. PNIPAM is a more efficient ligand to stabilize the gold nanoparticles in water and in organic solvents than alkanethiols. The surface density of PNIPAM chains ranged from 1.8 to 3.6 chain/nm2, which is much lower than that typical for alkanethiols. The thickness of a PNIPAM monolayer bound to the gold core is somewhat larger than the size of the random coil of corresponding free PNIPAM in aqueous solution, which suggests that the conformation of a PNIPAM chain bound to the gold core is to extended.

Oct 1, 2003

By D. L. Churcher, R. J. Perry

"In recent years much work has been don e to examine alternate underground backfill support systems with the objectives of improving ground support and reducing mining costs. This report deals with the work conducted at Placer Dome Inc.’s Dome Mine to develop a high density backfill system, taking advantage of the unique properties of this fill material, to address a backlog of large open stopes to sustain stable minewide ground conditions for future generations of Dome miners.IntroductionDome Mine began its existence in 1909 with the discovery of visible gold on a dome shaped outcrop located west of Porcupine Lake in northern Ontario. The orebody beneath this outcrop has supported one of Canada’s premier gold mining operations for 80 years. Production during this time of over 46 000 000 tons of ore has resulted in the recovery of II 500 000 ounces of gold and 2 000 000 ounces of silver. Mining continues today at a production rate of 3500 tons per day, with ore coming from both underground and open pit operations.During the life of the mine, different underground mining methods have been used to extract the ore from a complex deposit consisting of single and multiple vein structures.Cut-and-fill mining was introduced in the 1930s using sand fill and later hydraulic fill for support and to provide a working platform. A total of 1 800 000 tons of sand fill and 7 300 000 tons of hydraulic fill have been placed underground in cut-and-fill stopes. Until the late 1960sshrinkage mining was used extensively, particularly in narrow, steeply dipping zones. Bulk mining with small diameter longholes was introduced about this time, to take advantage of low production costs. Longhole blasting has often been used to recover low grade ore from the walls or sills of old shrink stopes, resulting in large open stopes. Backfilling of these open holes has been restricted to smaller stopes and associated with local ground conditions.In the late 1970s it was recognized that the number of large open holes which were being generated presented serious regional and minewide stability problems for future operations. With higher gold prices, economic ore reserves were being tied up in areas adjacent to multi-level open stope complexes. To stabilize these stopes and allow access, backfilling would be required."

Jan 1, 1990

By S. H. Lucas

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

Jan 1, 1980

By Pierre Gaucher

It is possible to selectively mine an orebody by combining the use of the SSW probe with any bulk mining methods as well as narrow vein exploitations. In other words, thanks to the use of the SSW probe, mining operators will be able to reduce and control dilution that is generally associated to every mining operation. Already, quick and substantial savings have been obtained. Falconbridge mines have demonstrated in Sudbury and Raglan operations, that the SSW probe can log nickel, copper and iron sulphides in blast holes. They have raised the grade and reduced dilution by outlining the limits of the ore zones before selectively loading explosives in the blast holes. By lowering the probe in a blast hole, a profile of the ore content is instantly measured and recorded. Logging cored diamond drill holes at the mine have demonstrated that the ore content (nickel, copper, cobalt, PGM) as assayed is proportional to the intensity of the probe?s electromagnetic response (EM), probably because the nickel / pyrrhotite ratio is uniform within a given orebody. The probe also measures the intrinsic conductivity of the sulphides and the percentage of magnetite. All probe readings are displayed and automatically stored at a 10-cm interval within the same reading unit. A 30-meter blast hole takes less than 2 minutes to survey. The whole instrument is light (12 kg) and very simple to use. One operator can carry the SSW system and survey either down or up-holes. The probe works from ?40°C to +50°C. Plotting the measurements of the SSW probe in blast holes (2-3-4-6-10 inches in diameter) before blasting allows to determine the exact shape of the orebody within the waste rock and uneconomic low-grade ore. Thus, only the portion of the holes that contains ore is blasted. The dilution is then much reduced and, by leaving the waste and low-grade ore in place, the mine saves on transportation, blasting, crushing, backfilling, milling, etc. Thousands of dollars are thus saved. Some savings may also be achieved during the development of stopes by guiding development drifts and by replacing diamond drilling by percussion drilling. In some circumstances, when the probe shows that some of the rich ore was not drilled-off, additional blast holes will allow to recover pockets of rich ore that otherwise would have been abandoned in the walls of a stope.

May 1, 2001