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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 Claude Hubert, Alex C. Brown, Michel Richard, Rejean Sirois

"The Pierre Beauchemin gold deposit, situated 15 km NW of Rouyn-Noranda, produces an average of 1050 tonnes of ore a day. Reserves are 1.2 Mt grading 5.6 g/t Au. The deposit, located in a shear zone within the Flavrianpluton, in the Blake River Group, is hosted by four major rock types: meta-trondhjemite, meta-tonalite, hybrid rocks and meta-diorite (the prefu; ""meta-"" is omitted thereafter). The trondhjemite, tonalite, and metamorphic hybrid rocks are spatially distributed as a series of sub-horizontal tabular sills that form a genetically related group. Early fractures that permitted either the emplacement of pre-diorite felsic dikes or the formation of a visible alteration caused by early fluid circulation might have been developed in response to initial strain created during regional deformation. These fractures were precursors to the emplacement of diorite and the development of major shear planes. Sub-horizontal north-trending en échelon diorite lenses are observed at the mine. Together, these diorite lenses constitute larger scale diorite zones which are observed at the scale of the pluton. The diorite zones are interpreted as evidence of an episode of sub-vertical extension sustained by the layered trondhjemite, tonalite and hybrid system. The anisotropy created by the less competent sub-horizontal en échelon diorite was an important factor. In shear zone development and subsequent mineralization. The tonalite and diorite accommodated further defomation and developed the main N030°/40° ESE shears along with the spatially associated N060°/50°-60° SSE, and the N000°/10°-40° E shear zones. The N000° shears are interpreted as refracted or deflected equivalents of the N030° shears. The combined shears compose the Eldrich shear system which trends approximately N030°/40° ESE and has had a general reverse movement with a minor sinistral component.Gold mineralization occurs in all four major rock types as en échelon sigmoidal shaped lenses associated with the major N030° deformation system. The distribution of ore lenses can be divided into two categories: The first category occupies sub-horizontal diorite zones. The lenses trend N-S, dip 100 40° E, and are associated with shear zones, sub-horizontal fractures or tonalite-diorite contact planes. The second category of lenses is found in felsic granitic rocks; the lenses trend either N0300!40°ESE or N060°/50°-60° SSE. Local hydraulic breccia zones are observed near ore zones occupying felsic granitic rocks. These breccias may have resulted from high fluid pressures brought upon by confinement as a result of impermeable zones in an anisotropic tonalite-diorite system."

Jan 1, 1990

By Alfred Guenkel, Kyle Marte, K. Nikolaisen

"The INCO SO2/air process was patented by INCO in 1985 (Borbely, Devuyst, Ettel, Mosoiu & Schitka, 1985). Well over 80 mining operations worldwide are using this process to oxidize cyanides to cyanates in the tailings from gold, silver and copper mining operations. Though it is possible to supply sulphur dioxide (SO2) in either gas or liquid form, the majority of these SO2/air operations use sulphite-based salts as the ‘SO2’ equivalent reducing agent. This is probably due to the lower capital cost, safety concerns with, and availability of liquid SO2. Sodium metabisulphite (SMBS) is commonly employed, and its cost is an order of magnitude higher than that of sulphur. The increased capital expenditures (CAPEX) for a sulphur-burning SO2 gas plant could have a pay-back of less than two years. In addition, a SO2 gas plant offers potential for significant heat recovery from the sulphur burning process. In terms of safety, the SO2 gas plant only requires the storage of relatively benign sulphur and the site inventory of SO2 will be limited to the plant gas inventory. Furthermore, a SO2 gas plant will not trigger the US Clean Air Act Section 112(r) accident prevention regulations, as may be the case for liquid SO2.Most existing mining operations using ‘SO2’ are gold operations, but in recent years there has been an increased interest in larger sulphur dioxide gas plants for hydrometallurgical facilities that employ leaching of cobalt/manganese minerals or precipitation of manganese/iron (e.g. Tenke Fungurume and Boleo).This paper discusses the different processes which require SO2 as a reducing agent and compares the capital and operational costs of a SO2 gas plant to conventional process alternatives using sulphite-based salts or liquid SO2.The heart of the SO2 gas plant is the sulphur furnace. The most economical way to produce SO2 is by combustion of sulphur in air producing up to 19.5 vol% SO2. Examples of large and small sulphur furnaces are discussed in terms of design, operation, and impacts on operational costs.The SO2 gas produced in a sulphur-burning plant has to be contacted with the liquor to be reduced. The choice of gas-liquid contacting device is important, and two contacting devices are discussed: 1) Gas sparging into the bottom of leach tanks (high, nearly constant, back pressure due to liquid head) and 2) absorption columns (low gas-side pressure drop)."

Jan 1, 2014

By J. P. Aiello

"INTRODUCTIONWhile this paper deals specifically with potential coal use at the Cold Lake Project of Esso Resources Canada Limited, a majority of the technical aspects of coal conversion could apply to any oil field and industrial boilers or stearn generators. In 1962, Imperial Oil Limited's production arm (the forerunner of Esso Resources), began examining methods of extracting bitumen from a 50 m thick layer of oil sand buried 450 m below the surface in the Cold Lake area. Almost two decades of research followed to develop an in situ oil sands recovery technology that made commercial production feasible.A $13 billion mega project that would have produced 22 000 m3/d (141 000 barrels/day) of synthetic crude oil (160 000 bpd of bitumen) was shelved in 1981, and ultimately replaced by a phased development strategy, involving smaller production units, to bring blended bitumen production on stream as markets and prices permit.The first two phases came into production in 1985. There are currently a total of six phases in production with four more in the construction stage. Each phase has 130to 140production wells (initially) and a central plant with water treatment, crude treatment and stearn generating facilities capable of producing 150 m3/d (9500 bpd of raw bitumen). The stearn generation section for each phase consists of three 180millionBtulhour oil fieldsteam generators.In its natural state, the bitumen is too thick to be recovered using conventional oil recovery techniques. At the Cold Lake Project, bitumen is recovered by a method called cyclic stearn stimulation. Water is pumped to the central plant where it is treated and heated to produce high pressure, high temperature steam. During the steam cycle, this steam is transported in insulated pipelines to the wells for injection into the oil sand formation. This cycle lasts from 30 to 50 days, heating the formation and causing the bitumen to flow. Following an initial flowback period, a surface based artificial lift system pumps the heated liquid bitumen, mixed with water and gas, out of the well and back to the central plant. The water and gas are separated from the bitumen, with the gas being used to provide some of the fuel for the stearn generators and the water being treated and recycled. Approximately 5% of the water cannot be recycled and is disposed of by injection into deep formations . The bitumen is blended with a diluent and shipped by pipeline to market."

Jan 1, 1989

By S. E. Malouf, R. Hinse

"The Mineral possibilities of the Chibougamau area were first recognized by Richardson in 1870 and Low in 1885, both of the Geological Survey of Canada. The first record of interest by the Quebec Bureau of Mines is in 1904, when J. Obalski reported on the discovery of copper and asbestos by Peter McKenzie in the previous year. Reports of 1906 and 1907 covered the McKenzie Gold Mine find and other deposits. In 1908 Prof. Dulieux described the quartz vein on Merrill Island and the mineralized zone on the near-by mainland. In 1910 the Chibougamau Railway Commission, though commenting favourably on the possibilities of the area, concluded that a railway was not then warranted.In spite of this, prospecting in the district continued. The main Campbell Chibougamau deposit, the Blake showing on Merrill Island in Dore Lake, was found in 1920, and other discoveries were made soon after. Between 1927 the work of Mawdsley, Retty, and Norman tor the Geological Survey and the Quebec Bureau of Mines stimulated further interest. In 1934 Consolidated Chibougamau Goldfields Ltd., a subsidiary of Consolidated Mining and Smelting Co. of Canada Ltd., acquired the Merrill Island and Cedar Bay holdings of Chibougamau Prospectors Ltd. They embarked on an aggressive campaign, directed essentially at the gold quartz veins at Cedar Bay. Exploration there included a 522-foot shaft, 4,732 feet of drifts and crosscuts, and about 34,000 feet of surface and underground diamond drilling. Thirteen holes were also drilled on the replacement sulphide lens at Merrill Island. Operations were suspended in 1936, as the occurrence was too low grade for development at the existing metal prices and under difficult operating conditions.The completion of the Quebec Bureau of Mines highway from St. Felicien to Chibougamau in 1950 was followed by a project of detailed geological mapping by the Bureau. It also encouraged Campbell Chibou-gamau Mines to acquire the assets of Consolidated Chibougamau Gold-fields and commence a program of surface exploration. Between August 1950 and April 1951, 41,183 feet of diamond drilling was done on the Cedar Bay and Merrill Island holdings. The Blake showing off the west shore of Merrill Island developed at depth into a major orebody. A lease was obtained on adjacent ground from Merrill Island Corporation, and underground work commenced in November 1951. Development at Merrill Island to June 1st, 1955, included a four-compartment shaft to 1,225 feet, with 8 levels at 150-foot intervals, 5,317 feet of raises, 16,709 feet of drifts and crosscuts, 52,514 feet of surface and 96,593 feet of underground diamond drilling."

Jan 1, 1949

By Pierre St-Julien, Réjean Hebert, M. Racine, Roch Gaudreau, Pierre Lacoste, Michel Rocheleau

"Les unités volcano-sédimentaires de la ceinture de l'Abitibi à l’est de Val-d' Or ont été regroupées en quatre domaines litho tectoniques. Du nord vers le sud, nous reconnaissons les domaines d'Assup, de Garden lsland de Val-d' Or, de Trivio et de Villebon. Deux de ces domaines (Garden Island et Trivio) correspondent à des couloirs de déformation, d’envergure régionale et d’orientation approximative E-W. Les trois autres domaines sont caractérises par des volcanites defond océanique de type MORB (tholéiites et komatiites) qui sont en position inferieure; ces volcanites sont surmontées par des sédiments marins de type turbidite (domaine de Ville bon) ou évoluent dans le temps, soit vers un volcanisme bimodal d'affinités tholéiitique et calco-alcaline ( domaine d'Assup), soit vers un volcanisme d'arc insulaire (domaine de Val-d'Or). Il n'existe aucune certitude de corrélation litho stratigraphique entre les différents domaines et l'on doit considérer la possibilité que ceux-ci représentent l'accolage tectonique de terranes de nature et d'origine différente.Les principales caractéristiques structurales de la région sont regroupées en quatre étapes de déformation dont trois sont archéennes et la quatrième attribuée à l'orogenèse grenvillienne. Au moins cinq facies métamorphiques ont été reconnus: I) un facies régional faible propre à la province du Supérieur, 2) un facies de cornéennes a amphibole en périphérie du pluton de Pershing, 3) un facies hydrothermal dans la zone tectonique de Cadillac, 4) un facies régional moyen dans la zone de bordure de la province de Grenville, et 5) un facies régional élevé a l'intérieur de la zone structurale du front de Grenville.Le front de Grenville est défini dans la région par l’apparition subite des ortho gneiss au faciès métamorphique élevé et par la disparition du grain tectonique E-W. Le contact brusque entre les deux provinces pourrait correspondre à une zone de chevauchement de la province de Grenville sur la province du Supérieur. La région étudiée est l’hôte d’un gite majeur de sulfures massifs et de plusieurs gites et indices aurifères; elle présente des cibles favorables pour l’exploration minière dans le contexte tectonométa-morphique de l'influence grenvillienne. L'intérêt de la Formation de Val-d' Or pour les métaux de base et le potentiel aurifère des couloirs de déformation d'orientation approximative E-W, sont bien connus. Les structures grenvilliennes NE-SW, qui se développent vers l'est, représentent aussi des cibles intéressantes, en particulier dans les secteurs ou elles intersectent des cisaillements E-W kénoriens."

Jan 1, 1990

By R. Kumar, A. K. Verma

"In this paper, an anisotropic shear behavior of rock joints is investigated under constant normal load condition (CNL). The natural rock joint roughness is transferred to artificial rock samples using room temperature vulcanizing (RTV) silicon rubber. Artificial cylindrical rock joint samples (90 mm diameter and 50 mm height) are made by mixture of cement, sand, and water in the ratio of 1:1.5:0.45 (by weight). Direct shear tests are conducted on these samples in twelve directions at 30º apart at four different normal stresses (250kPa, 500kPa, 1000kPa, and 1500kPa). Thus, 48 samples are tested in laboratory. Test results shows that the peak shear strength (?) of rock joints is anisotropic in nature due to change in morphological features in particular shearing direction. Joint roughness coefficient (JRC) is determined in each shearing direction by trial and error method using Barton’s model (1973). Further, cohesion ( c) and angle of friction ( ? )of joint surface in each direction is determined. Finally, a relationship between angle of friction ( ? ) and joint roughness coefficient (JRC) is developed. Experimental shear strength and Barton’s shear strength is compared and it is found that effect of anisotropic decreases with increasing normal stress. INTRODUCTIONRock mass is generally composed of discontinuity which refers collectively to joints, faults, bedding plane and folds etc. ISRM (1978) defined joint as a line of break of geological origin along which there is no visible displacement. Joint is three dimensional discontinuity which composed of two matched/mismatched surfaces. The presence of joints in rock mass plays an important role to define its overall shear strength, deformability behavior, in-situ stresses and hydro-geological properties. The shear strength of rock joints is important in the design of near surface/ deep geotechnical works (mining excavation, dam foundation, power plants, underground caverns, and slopes).In last 40 years, many constitutive models are developed to predict shear strength of rock joints under CNL conditions (Patton 1966; Landanyi and Archambault 1969; Barton 1973; Goodman 1976; Barton and Choubey 1977; Plesha 1987; Desai ad Fishman 1991; Amadei and Saeb 1992; Maksimovic 1992; Wibowo et al. 1993; Huang et al. 1993; Amadei et al. 1998; Indraratna 2000; Homand et al. 2001; Wang et. al. 2003; Samadhiya et. al. 2008; Asadollahi and Tonon 2010; Ghazvinian et al. 2012). These constitutive models have limitations to predict shear strength of rock joints in one direction. Till date, among all constitutive models, Barton (1973) model is widely used in practice due its simplicity. It includes parameters such as joint roughness and joint compressive strength (JCS) to quantify joint surfaces."

Jan 1, 2015

By S. Onuki, T. Ando, M. Tejima, R. Enomoto, Y. Okada

"Al–Si–Cu–Mg die casting alloys are widely used in the automotive industry, and recent advancements in high pressure die casting technology enable heat treating products to achieve desirable properties. However, CO2emissions associated with die-casting production lines should be reduced to address the global warming issue. In this study, the relation between the microstructures and thermal stability of precipitates formed during isothermal aging at 423, 485, and 558 K of Al-11.0 mass%Si-1.8 mass%Cu-0.2 mass% Mg (JIS ADC12) was investigated using Vickers hardness tests and transmission electron microscopy (TEM). The hardness tests and TEM results showed that Si compounds and Al–Cu compounds enhanced the mechanical properties of the alloy and that the predominant compound affecting the mechanical properties differed between specimens treated at 423 K and those treated at 558 K. Moreover, additional aging at 423–458 K resulted in a material with a Vickers hardness higher than that of the isothermally single-aged material. Differential scanning calorimetry, differential elemental analysis, and selected-area diffraction analysis using TEM suggested that the combination of Al–Cu compounds precipitated at low aging temperatures and Si compounds precipitated during preliminary aging increased the mechanical strength, resulting in a maximum hardness of 130 HV.INTRODUCTIONAl–Si–Cu–Mg die casting alloys have well-balanced mechanical properties, machinability, and cast ability. These alloys are therefore widely used in the automotive industry because of their superior characteristics, accounting for>90% of the aluminum used in die casting production. As a substitute for iron-based materials, these alloys have contributed to the development of lightweight vehicles.Dong et al. (2014) reported that the heat treatment of die casting materials is effective for changing their dimensions under high temperatures as a result of the precipitation of solid-solution elements. Recent advancements in high pressure die casting technology have led to the application of heat treatments of Al–Si–Cu–Mg alloy products to reduce fuel consumption and satisfy customer needs. Although attempts to reduce CO2emissions have focused on fuel consumption in cars and fuel efficiency regulations have become increasingly strict in recent years, CO2discharged from production lines should also be reduced without adversely affecting the properties of the product. Heat treatments for artificial aging after solution treatment (T6 treatment) and after casting (T5 treatment) are CO2discharge sources, which together with a holding furnace account for >80 % of the CO2discharged during the casting process."

Jan 1, 2018

By Kazutoshi Haga, Batnasan Altansukh, Yasushi Takasaki, Mohd Yazid Bin Abdul Razak, Atsushi Shibayama

"In this study, the volatilization behaviour of arsenic (As) and antimony (Sb) from arsenicantimony bearing minerals such as tennantite (Cu12As4S13)/tetrahedrite (Cu12Sb4S13) in copper ore was investigated using a roasting process. Specifically, the thermal decomposition of tennantite and the arsenic volatilization were investigated under nitrogen and air atmosphere. The roasting experiments were performed at temperatures ranging from 500 to 1,200oC and different retention times using a laboratory electric furnace. The results indicated that over 95% volatilization was achieved at 700oC. On the other hand, when the temperature increased to 1,200oC, volatilization rate of Sb reached to 90%. Therefore, volatilization of Sb requires high temperature compared to As. It suggests that As and Sb can be separated using temperature difference. Selective removal of As is possible at 700oC, subsequently Sb can be recovered as a volatilized product by increasing the temperature to 1,200oC. Meanwhile, volatilization of Cu and Fe was not observed on these conditions. The content of As and Sb in the residues obtained from roasting at 700 and 1,200oC was less than 0.3 and 0.75 mass%, respectively. The Cu content in solid residue at 1,200oC increased to 34 mass%. The clean copper concentrate with lower amounts of arsenic and antimony was obtained from the process. INTRODUCTIONRecently, arsenic (As) and antimony (Sb) contents in copper (Cu) concentrates are increasing year by year. Arsenic and antimony are present with Cu minerals such as enargite (Cu3AsS4), tennantite (Cu12As4S13) and tetrahedrite (Cu12Sb4S13) that associate with copper sulfides [chalcopyrite (CuFeS2) and bornite (Cu5FeS4)]. They are normally concentrated by flotation and treated in copper smelters. However, treatment of copper concentrates containing high amounts of arsenic and antimony by a smelting process is difficult due to environmental implications of toxic arsenic and antimony. As demand for copper continues to grow, the need to source copper from several types of resources including those that contain high levels of As/Sb becomes necessary. Therefore, a novel processes for the treatment of As/Sb bearing Cu concentrates are required. Reviews on copper resources such as enargite, tennantite and tetrahedrite containing high amounts of arsenic and antimony are covered widely in literatures by many researchers (Padilla et al., 1998; Welham, 2001; Vinals et al., 2003; Guo & Yen, 2005; Balaž & Achimovicova, 2006a; Filippou et al., 2007; Mihajlovic et al., 2007; Lattanzi et al., 2008)."

Jan 1, 2016

By D. Campbell Mackenzie

THIS paper is submitted primarily to put on record the results of a number of experiments carried out at the Wingdam mine of Consolidated Gold Alluvials of B.C., Ltd., with the object of establishing a reliable method of concentration of 'wash' and a maximum recovery of gold from the concentrate. It will be generally admitted that, during the last two decades, a tremendous amount of research work has been carried out, both by private companies and Government departments, with a view to improving the technique of the recovery of gold .from quartz veins or lodes. So thorough have been these investigations that milling practice at a modern quartz gold mine is almost an exact science, carried out with meticulous accuracy. In placer mining, on the other hand, very little departure from old traditional methods can be noted, with the exception perhaps of the more modern gold dredges. This may be due to the fact that the concentration of placer gravels, up to a certain point, is comparatively easy. Only a simple equipment of sluice box and riffies seems necessary where the gold is coarse and the gravel free from clay. In the early days of placer mining, when claims were rich and fairly 5hallow, nobody worried very much about fine gold; and, if clay was present, an enlarged dump box was constructed in which the gravel was subjected to a rather desultory puddling. That large quantities of gold are lost in normal placer practice is generally admitted. Some 26 years ago, I had inspectorial charge of many bucket dredges on the Ovens river, in Australia. Occasionally it was necessary for a dredge to work through its own tailing to reach a new section, and almost without exception enough gold was recovered from these tailings to defray the total cost of moving the dredge. Since those days, the technique of gold recovery in dredging has been considerably improved and it is still receiving close attention. Outside of dredge mining, most placer operations, deep or otherwise, are as they were-notorious gold losers, paying little attention to the recovery of fine gold and taking an absurd amount of time over each clean-up. In the Cariboo district of British Columbia, the alluvial gold is usual y associated with varying quantities of black sand (magnetic) and grey sand (non-magnetic); and it can be taken as axiomatic that, if you save all the black sand and grey sand, you save practically all the gold. It is on this hypothesis that we have carried out our research work and brought our recovery system up to its present highly satisfactory standard.

Jan 1, 1937

By I. Tinto

"IntroductionEnvironmental pollution problems, caused by gases or effluents, by high noise levels or from air borne dust, are placing increasing demands of the operators, from statutory authorities, and from the general public, to adopt more stringent dust control measures. This requires that a successful plant dust control program be an integral part of any plant operation. While a successful environmental program is a key factor in maintaining productivity and good labour relations, anticipation of future environmental regulations during the plant design phase can avoid the high cost of plant upgrading, or dealing with pulmonary conditions in the work force in subsequent years.Particulate removal, conventionally called dust collection, involves the removal of coarse- and fine-particulates from the airstream. Although generally limited to that required for contro!Hng plant generated air emissions, particulate control equipment also has a wide application for air cleaning and filtration inside the plant, as well as for recovery of usable products.In the late 1950s and early 1960s the public became more concerned with the problems of air pollution. This increasing concern was stirred by an awareness of the extremes to which water pollution had progressed in many of the rivers and lakes, by increasing perception of air pollution as detected by visible smog and harsh chemical odors, by the alarming rate of increase of smog plagued cities and communities, and by specific incidents where ill health could be traced to specific episodes of air pollution. Particle size is perhaps the most important single factor affecting particulate control. Large particles (over 5 microns) constitute the major mass fraction of most sources of emission and economical control equipment dqes exist for their efficient control. However, while these larger particles frequently represent the source of soiling and unsightly smoke plumes, they are not a major health problem as they tend to settle relatively quickly, are more easily scrubbed from the atmosphere by rain, and are not able to penetrate the respiratory system. By far the most important form of particulate air pollution is the fine particulate, (below 3 microns) which is harmful to the pulmonary system. The adverse effects caused by this particulate matter on human health are for the most part related to injury to the surfaces of the respiratory system. Particulate material in the respiratory tract may produce injury itself, or it may act in conjunction with gases, altering their sites or their modes of action. A combination of particulates and gases may produce an effect that is greater than the sum of the effects caused by either individually (i.e. synergistic effect)."

Jan 1, 1989

By F. D. Ball, D. E. Gemmell

"Coal mining in New Brunswick dates from the year 1639, when, as today, coal was produced from the area surrounding Grand Lake near the geographic centre of the province. Current production by N.B. Coal Ltd. is approximately 453,000 tonnes per year and comes from a single, thin Pennsylvanian age seam. The history of development and understanding of the geological setting of New Brunswick coals is discussed, beginning with the work of A. Gesner in 1839. The lithostratigraphic sub-divisions of the Pennsylvanian strata proposed by various workers are summarized and the present hypothesis that only biostratigraphic subdivision is possible is detailed. The Pictou Group, which includes the New Brunswick coals, consists of repetitive sequences of sediments consistent with flu vial and upper delta plain sedimentation. The development of coal mining in the province has progressed from horse-powered stream-bank stripping operations prior to the late 1800s through to the present use of open stripmining techniques utilizing eight walking draglines of various sizes. During the intervening years the introduction and eventual phasing out of underground mining occurred with the near termination of the coal mining industry in the province prior to 1973. The coal seam worked by N.B. Coal Ltd. averages 46 cm in thickness, has never been found to be more than 76 cm thick and has both high ash (12.4 -31.1%) and sulphur (4.9-7.8%) content. The wide variety of coal beneficiation studies and tests completed on New Brunswick coals have lead to the conclusion that a washed New Brunswick coal product could be expected with 10 to 12% ash, 5 to 6% sulphur, and 30,225 to 31,388 1/gm. Recoverable coal reserves, including those in the area licenced to N.B. Coal Ltd. and those licenced, although not yet in production, to both the St. Lawrence Cement Co. Ltd. and Provincial Holdings Ltd. at Lake Stream and Beersvil/e, total 60 million tonnes. At present, coal in New Brunswick is removed from prospecting and staking and all existing and future coal exploration and development will be administered by agreement under the Mining Act. Based on the results of a recently completed drilling project undertaken by the Department of Natural Resources, seventeen coal exploration blocks which carry minimum work commitments will be offered to the New Brunswick Electric Power Commission on a sixty-day first right ofrefusal basis. Those blocks not taken up by the Power Commission will be offered under the same commitments to the private sector."

Jan 1, 1985

By F. K. McKean

Abstract A discussion of the difficulties involved in the production of a clean molybdenite concentrate introduces the paper. There is a brief description of the various methods for producing molybdenite that will fit marketing specifications, and it is shown how these methods are nor suited to the milling of ores containing small quantities of bismuth and copper. A process that was developed in the ore testing laboratories of the Bureau of Mines at Ottawa for the La Corne mill of the Molybdenite Corporation of Canada is described in derail and some results from pilot plant operation of this milling scheme are tabulated. A description of the La Corne mill, where the writer supervised the introduction of the process on a commercial scale, adds to the interest of the paper. In conclusion, there is a review of the conditions necessary for proper operation of the flowsheet. An attempt is made to show how the process might be satisfactory in other problems similar to that at the La Corne mill. Introduction The production of a molybdenite concentrate, free from impurities and high in grade, has been a difficult task when milling some types of ore. Whereas ordinary flotation techniques will produce acceptable concentrates of the baser sulphides, in many cases they will not serve in the treatment of molybdenite, for this mineral must be of optimum grade and purity to bring the best prices, or to find a market at all . Faced with the necessity of meeting stringent specifications, the mill operator has been forced to employ expensive processing methods, and to be satisfied with low recoveries, in order to make grade with his product. The most objectionable and common impurities found with molybdenite are: (a) gangue minerals and iron sulphides, which lower the grade; and (b) copper, bismuth, or antimony, which are deleterious in the final uses of the molybdenite products. It is difficult, and sometimes impossible, to market a concentrate with over 0.5 per cent of any of the latter impurities, or with a grade much below 90 per cent MoS2. Various methods, which have been highly successful in specific applications, have been developed to meet the problem of producing a suitable concentrate. All of them, however, have certain defects precluding their use when treating ores of a nature different from the one for which they ? were designed. The writer has developed a process for molybdenite ores having an objectionable impurity content. A description of the procedure involved, and of the La Corne mill of the Molybdenite Corporation of Canada, where the process has been adopted, should be of interest to flotation operators.

Jan 1, 1947

By J. D. Bateman

DURING the summer of 1942 several chromite deposits were discovered in the Lac du Bonnet district about eighty miles northeast of Winnipeg. The deposits are confined to the Bird River complex, a folded composite sill of basic and ultrabasic igneous rocks of early Precambrian age that lies to the north of Bird river (1). The outcrop of the sill represents the trace of an easterly plunging anticline, the south limb extending from the northeast bay of Lac du Bonnet eastward to Bird lake, and the other limb, 6 to 12 miles north, lying between Maskwa river and Euclid lake. The south limb of the sill is 20 miles long, and the northern counterpart 15 miles; but its continuity is interrupted by faulting and by the intrusion of younger granite, so that the total length of outcrop of both limbs does not likely exceed 25 miles. The sill is intrusive into early Precambrian greenstones and associated sedimentary rocks. The chromite discoveries have been made at several localities on both limbs of the sill, and in each case at the same geological horizon. The sill comprises an upper layer of hornblende gabbro and a lower layer of serpentinized peridotite, and altogether is from 500 to 3,500 feet thick. The chromite zone is within and near the top of the peridotite layer, and thus about midway of the sill, both of the limbs of which stand nearly vertical. This geological relationship was established early in the exploration of the initial discoveries and served as the basis for successful predictions of later discoveries. The chromite zone probably persists throughout the length of the sill, but its width and chromite content are not everywhere the same. The initial exploration along both limbs of the sill has shown a combined total length of about 12,000 feet of chromite ore varying in width from 6 to 12 feet. In many places the overburden is thick and obscures the continuity of the chromite zone; but as investigation proceeds the known deposits will be extended and new ones will be found. Thus, although a total length of 12,000 feet of ore has been investigated up to the present time, there is an indicated length of approximately 31,000 feet along which the ore is more than 4 to 5 feet wide. This estimate is based on geological considerations and represents a total tonnage in the order of 15,000 tons per vertical foot of depth.

Jan 1, 1943

By M. Reda

Erosion corrosion is acceleration in the rate of corrosion attack in metal due to the relative motion of a corrosive fluid and a metal surface [1]. The increased turbulence caused by pitting on the internal surfaces of a tube can result in rapidly increasing erosion rates and eventually a leak [2]. Engineers and scientists in different disciplines have worked for many years on methods of complete protection from flow induced corrosion or partially reducing corrosion of this type, and they have proposed or applied different countermeasures [2]. The accident at the Surry-2 nuclear power plant (USA) and other similar incidents have once again disturbed the scientific-technical community. This has forced the plant operating personnel and the management to undertake wide-ranging measures in order to discover on time, assess, and monitor such cases. In these extreme cases, protective coating and even metallic bulk material coating (the substrate itself) may suffer considerably from mechanically induced damage [2]. At high velocities, the "leaching out" dissolution reaction of particular, for example, steel elements at the solid-liquid interface become the rate determining step. One practical example for the difference between erosion corrosion and flow induced corrosion is the famous steam eruption from a Japanese [3, 4] nuclear power plant cooling system at the pressurized water reactor (PWR), turbine and in the condenser-steam generator connecting pipe. The rupture occurred in the carbon steel pipe connecting the condenser and steam generator. The damage occurred in the easily eroded/corroded carbon steel at a vulnerable location downstream of an orifice plate, In general, when there is significant reduction in pipe thickness through erosion/corrosion, distinctive patterns are visible, but those patterns were not observed in this case, Thus one can speculate that another phenomenon is possibly controlling, Fluid flow within pipes reduces the pipe thickness leading to an eventual rupture, The thickness reduction mechanism is erosion/corrosion, One example is "cavitations corrosion", Erosion and corrosion occur at the same time, and it is difficult to determine which one is dominant, Therefore, the term, "erosion/corrosion," is used [5]. In the atomic energy field, the terms "Flow Accelerated Corrosion" and "Flow Induced Corrosion" are also used. It is the purpose here to show that during flow induced corrosion a different mechanism may possibly be operating and in some cases it can be different from the well known erosion corrosion. Another example is fretting failures in early Canadian nuclear steam generators which was thought to be due to excessive vibration caused by high.

Jan 1, 2010

By M. Romana

"RMR geomechanics classification is being used all over the World as a useful tool for tunnel predesign and construction. Bieniawski (1989) provided guidelines for tunnel excavation and support, according RMR numerical values; guidelines which are now included in almost all tunnelling books, regulations, and Internet portals. But in the last 25 years tunnelling techniques have changed and some of the guidelines have become obsolete.. The conditions of tunnels for the 1989 guidelines application were: horseshoe shape; 10 m width; vertical stress < 25MPa; construction by drilling and blasting. Actually a substantial number of tunnels are mechanically excavated. The tunnel themselves are now bigger: a span of 14 m. is very common for transportation tunnels. The guidelines included the following concepts: excavation method, rock bolts (20 mm diameter, fully grouted), shotcrete, steel sets. There is an increasing gap between the 1989 guidelines and the practice: in many tunnels different to-the-case excavation and support requirements are established. Bieniawski himself has published a paper (Lowson & Bieniawski, 2013) with different guidelines based in Lowson’s practice, with no explanations and/or justification.In this paper a new set of RMR tunnelling guidelines is proposed taking account of the actual industry conditions, and increased safety regulations. The 5 classes’ limit RMR:values (80,60,40,20) are no more useful actually: Values of RMR (70, 50, 30) are used in practice to define quite different tunnelling techniques. Consequently a 10 subclasses system is proposed. Each subclass covers a span of 10 RMR points. As the number of mechanically excavated tunnels has increased a new correction factor is necessary, according the excavation method.The proposed 2014 RMR guidelines are for tunnels 10 to 14 meters wide, the world practice for new roads/railways tunnels. Many tunnels for water/ services are smaller with smaller support needs. Wider than 16 m caverns are out of the RMR system.It is not possible, nor desirable, to propose only a value for support in each subclass. The new guidelines propose an interval of values for certain types of support. More pessimistic values must be chosen for wider tunnels, tunnels with lower RMR values in each class and cases when a higher safety factor is asked for. More optimist ones are chosen for narrower tunnels, and higher RMR values. Some alternatives are almost complementary, others are contradictory. Sometimes several different techniques are proposed Guidelines are not automatic options. Good engineering judgment must be used when choosing between the different alternatives. Guidelines are not mandatory suggestions."

Jan 1, 2015

By D. G. Sinclair

IN surveying the changes that have taken place in the mining industry of Ontario in the past ten years, the striking growth of the gold mining industry first attracts attention. In 1925 there were only twelve regularly producing gold mines in the Province, six of which were located in the Porcupine area, four at Kirkland Lake, and one each at Larder Lake and Boston Creek. In contrast with this, the year that has just closed (1935) has seen production on a steady basis from forty mines, with what may be termed spasmodic production from twelve other properties. This so-called spasmodic production is, however, responsible for the recovery of gold to an estimated value of between $75,000 and $100,000 during the year. Further illustrating the expansion in the gold mining group is the activity in the exploration and development of properties. In 1925, mining operations were conducted at 29 properties which had not reached production, while in 1935 work of this nature was carried on at 87 such properties. It is of interest to note that of the six properties listed as producers in Porcupine ten years ago, five are still producing at rates considerably in excess of the tonnages handled at that time. In fact, the estimated tonnage treated at these mines in 1935 exceeds the 1925 tonnage by approximately half a million tons. In Kirkland Lake, the figures are still more striking. The four mines in production in that camp in 1925 milled 346,584 tons, producing bullion to the value of $5,132,128, while in 1935 these same mines treated approximately 1,643,000 tons, with a recovery of roughly $29,300,000. While Porcupine and Kirkland Lake remain the great gold-producing camps of the Province, the growing importance of the newer areas is readily appreciated when we consider that in 1935 the mills of these areas treated over 1,000,000 tons of ore for a production in excess of $7,500,000. Pro-duction from these newer districts first became of importance in 1930 with the commencement of recovery at the Howey mine, in Red Lake, and the Minto, in Michipicoten. Since 1930, eighteen new gold mills have gone into production in the Red Lake, Little Long Lac-Beardmore, Pickle Crow, Sturgeon Lake, Michipicoten, Sudbury, and Matachewan areas. In respect to silver mining, the period under review has seen a considerable slackening of activity, due to a generally unstable and depreciated price for the product and to the exhaustion of ore reserves in the mines.

Jan 1, 1936

By T. Crivits, I. Bellemans, V. Cnockaert, K. Verbeken

It is well known that slag foaming can significantly complicate copper smelting processes. This is especially true for unpredictable events that generate large amounts of slag foam during a small period of time. These events are here referred to as eruptions and are typical for certain types of liquid bath furnaces such as modern top submerged lance (TSL) furnaces. A few authors already made the link between over- oxidation of the slag phase and slag foaming. In this work, the mechanism is studied in more detail and explained using the smelter chemistry and the behaviour of dynamic foam formation. Finally, the most influential parameters on the foaming behaviour of the process are presented. INTRODUCTION Copper remains one of the most important metals used in our modern society. In 2017, the annual amount of copper produced worldwide was about 23.5 million metric tons. This makes copper the third most produced metal in the world after steel and aluminium. The majority of this copper was produced by copper smelting, about 19 million metric tons, both from primary concentrate and secondary feed material as reported in ICSG’s World Copper Factbook 2018. In the past, copper smelting was mainly performed using Flash smelting, a technology developed before 1949 (Kojo, Jokilaakso & Hanniala, 2000). Nowadays, this type of smelter is being complemented by other smelter types such as ISASMELT and Ausmelt technologies. These are two types op smelter furnaces, based on the concept of a top submerged lance (TSL) furnace. These two technologies were both commercialized around 1990 (Arthur & Hunt, 2005). It is estimated by Wang et al. (2016) that at this time, both flash and TSL smelters contribute equally to the total amount of annually smelted copper. Flash smelting stays the most important technology for extracting copper from primary sources. TSL furnaces, on the other hand, are more suited than Flash smelters for the recovery of copper from secondary materials. This is because of the different ways that the feed material is inserted in both furnaces. In Flash smelters, the ore and fluxes needs to be fine and completely dry in order to be able to blow the mixture inside the burner area of the furnace. The feed material for TSL smelters show less constrains concerning humidity and particle size. The material is dropped from the top inside the molten bath. As a result, a more complex and flexible feed stream can be processed using TSL smelting (Wood, Hoang & Hughes, 2017). This feature will become more important in the future as primary copper sources gets more scarce and the need for a fully sustainable material cycle increases.

Jan 1, 2019

"In the first half of 1990, Cambior earned $1104 million, or $0.37 per share, on revenues of$66.4 million. The company&apos;s strong cash position, attained late in 1989 and in the beginning of1990 through the exercise of warrants, put working capital at $83.8 million as of June 30, 1990. Long-term debt was reduced by over $10 million during the weak gold market, leaving the company with a debt to equity ratio of only 0.03. Production for the six first months of 1990 was 107 720 oz. Au, an increaseof24% over the corresponding period in 1989.Financial Profile/Results and Corporate Financial StrategiesTotal revenue for the 1989 fiscal year was $116,883,000 compared to $109,848,000 for 1988. The 1989 results achieved despite a lower average gold price, resulted from a higher level of commercial gold production and a 115% increase in investment income over 1988. Investment income rose to $3,568,000 in 1989 due to higher levels of cash assets.Cambior&apos;s mine operating costs increased 19% to $68,561,000 in 1989 from $57,472,000 in 1988. This increase resulted from rising operating costs at the Doyon Mine ($250/oz to $274/oz) due to the shift from open pit to underground operations, as well as from the start -up of three new mines with relatively higher costs. These costs have reached a stable level that should be maintained for the next few years.In 1989, net exploration expenses were $6,838,000, down 31070 from 1988, partially due to the Canadian tax reforms introduced since 1987. Depreciation and depletion grew by $1,432,000 in 1989, reflecting a higher volume of commercial production from the Pierre Beauchemin, Lucien C. Beliveau and Chimo mines.Net earnings for 1989 amounted to $17,661,000 or $0.73 per share compared to $17,438,000 or $0.74 per share in 1988.Cambior has an exceptionally liquid and strong balance sheet. As at December 31, 1989, net working capital stood at $85.6 million, with a total assets base of $338.6 million. With a long-term debt (including gold loans) of $21.5 million and shareholders&apos; equity of $277.2 million, Cambior has a debt to equity ratio of 8%. All but about $100,000 of long-term debt is in the form of gold loans.The total market value for Cambior&apos;s stock as at September 10, 1990, stands at $420million, based on recent stock market prices. Cambior&apos;s stock and warrants are traded on the Montreal and Toronto Stock Exchange."

Jan 1, 1990

By Z. Zhang, X. -G. Chen, F. Breton, A. Arici

"The microstructure of Al-10Si-0.3Mg-0.5Mn alloy with 0.27% Zr and 0.19% V addition was quantitatively analysed and compared to the base material. The addition of Zr and V in the AlSi10Mg alloy exhibit a beneficial effect on the Si modification, resulting in the reduction of the size of eutectic Si particles on both as-cast and T6 conditions. The quantitative image analysis results show that the Zr and V addition and T6 treatment can reduce the size of the intermetallic phases. In addition, the presence of Zr and V is favour to the partial dissolution of the intermetallic phases during the T6 treatment. Different heat treatment procedures were tested to optimize treatment parameters. The mechanical properties of the alloys were evaluated using hardness measurements and tensile tests. The results show that the addition of Zr and V can remarkably improve the mechanical properties on as-cast and T6-treated conditions. The correlation between the microstructure and mechanical properties of the alloys is discussed.INTRODUCTION Al-Si-Mg foundry alloys have been widely used in automotive and aerospace industries due to their high strength-to-weight ratio, heat-treatable capability and excellent castability. The alloying elements such as Cu, Mn, Fe are traditionally added into Al-Si-Mg alloys to improve their mechanical properties (Mohamed, Samuel, Samuel & Doty, 2009). In recent years, the attention has been paid to the effect of the transition elements such as Zr, V, Ti, etc, as minor alloying additives, on the properties of aluminum alloys (Elhadari, Patel, Chen, & Kasprzak, 2011; Kasprzak, Amirkhiz & Niewczas, 2014; Shaha, Czerwinski, Kasprzak, Friedman & Chen, 2015; Shaha, Czerwinski, Kasprzak & Chen, 2014; Mahmudi, Sepehrband, Ghasemi & 2006; Meng, Cui, Zhao & Zuo, 2013). Kasprzak et al., (2014) and Shaha et al., (2015) reported that additions of Zr, V, and Ti in Al-Si-Cu-Mg cast alloys can improve the alloy mechanical properties at high temperature by forming stable intermetallics and precipitates on as-cast and heat treatment conditions. On the other hand, the mechanical properties of near-eutectic and eutectic Al–Si cast alloys depend not only on the chemical composition but also, even more importantly, on microstructural features such as the morphology and size of eutectic silicon and intermetallic particles, which are present on both as-cast and heat treatment conditions (Mohamed et al., 2009). However, there is rather limited information on the effect of Zr and V on the microstructure features and mechanical properties of near-eutectic Al-Si-Mg cast alloys."

Jan 1, 2018