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By Charles S. Ney

"THE WRITER'S acquaintance with Kootenay King mine is limited to a brief geological excursion in the vicinity. This paper has been compiled largely from maps of J. Sullivan, mine geologist in 1952, kindly made available to the writer by Transcontinental Resources; a 1942 private report by W. L. Brown; and personal communications with J. Sullivan. Remarks on regional geology, and the comparison with Sullivan mine, are specifically the writer's.The Kootenay King mine is situated in the westernmost range of the southern Canadian Rockies at an elevation of 7,300 feet. From the Sullivan Mine at Kimberley, B. C., it lies 19 miles due east. Six miles of this distance are represented by the pronounced topographical and structural gap of the Rocky Mountain Trench. Mineral claims were locat-ed in the area prior to 1898. The deposit was first seriously explored in the late 1920s, and a small orebody was then outlined and developed. Mining and milling were undertaken in 1951.52. The orebody, amounting to about 14,000 tons, was mined out, and no significant new ore was revealed.The orebody is a replacement of dolomitic argillite by fine-grained galena, sphalerite, and pyrite. The sulphides are finely laminated in the direction of bedding, following out minor folds faithfully. The grade of ore estimated from diamond drill holes (W. L. Brown, 1942) was 11 ·I per cent zinc, 11 ·I per cent lead, and 3·8 ounces per ton of silver. The mineralized zone extends about 250 feet horizontally, 160 feet vertically, with a width of six to twenty feet.General GeologyThe country rocks are sediments of the Aldridge formation of late Precambrian age. For several thousand feet stratigraphically above and below the mine, the formation consists mainly of thinly laminated argillite and siltstone. Several horizons contain appreciable carbonate. About 2,800 feet stratigraphically below the mine, predominantly dolomitic beds mark the top of the underlying Fort Steele formation. The medium-bedded fine quartzites, which characterize about 10,000 feet of the Middle Aldridge in the Purcell range, have a restricted development amounting to only 2,000 feet in the Rocky Mountains. This member is about 2,200 feet above the mine horizon."

Jan 1, 1949

By J. E. Gill, A. R. Byers

"The property of Sladen Malartic Mines, Limited, is in the northern half of Fourniere township, Abitibi county, Quebec. The group is large, but the part so far found to be of economic interest is a strip 3,200 feet long between the Canadian Malartic property to the west and the East Malartic and National Malartic properties to the east ( Fig. 1).The original claims were staked in 1923 and 1924 by St. B. Sladen and H. S. Kennedy because gold was known to occur on claims nearby. In 1928, detailed mapping of the Malartic Gold Mines property (now Canadian Malartic) showed the presence of an east-west gold-bearing zone traceable almost to the Sladen boundary. During the winter of 1929, a hand-driven drill was used to secure short sections of core through casing pipe sunk to bed-rock. This proved the existence of the main Sladen ore zone, but it was not until the price of gold was increased, in 1934, that serious work was attempted.Underground work was started in 1935 and milling in January, 1938. To the end of 1945, 168,882 oz. of gold were recovered from 1,641,974 tons of ore, an average of 0.103 oz. a ton.GEOLOGYThe bed-rock underlying' the Sladen Malartic property consists mainly of Precambrian greywacke and volcanics that appear to form a conformable series with a general northwest strike and vertical to steep northerly dip. Except where disturbed locally by drag folds or faults, tops face north. Intrusive rocks that range in composition from diorite to granite cut the sediments and volcanics. The basic types, generally quartz diorite and diorite, form dykes or sills in the volcanic rocks. The more acidic types, mainly quartz syenite and related porphyries, intrude the volcanics and the sediments in the form of dykes, sills, and larger, stock-like bodies.The orebodies are parts of slightly shattered or brecciated zones, mainly in greywacke, but partly in syenite, diorite, and volcanics, that have been more or less replaced by fine-grained quartz, pyrite, and minor amounts of sphalerite, galena, sylvanite, petzite, and gold."

Jan 1, 1954

By F. Parada Luna, F. Parada Torres, J. Salvo, A. Reghezza

In the pyrometallurgical processes of copper concentrates, arsenic is preferentially concentrated in the dusts carried by the gases as a result of their volatility and subsequent condensation in the gas cleaning train. In this work results obtained by alkaline leaching of the dusts from Ministro Hales roasting plant, belonging to Codelco, are presented. For the Ministro Hales plant the alkaline leaching would be an alternative to the recirculation of dusts to the roasting furnace. The experimental study considered NaOH and NaOH plus NaHS leaching of two samples of dusts, with 4.24 and 2.55% arsenic respectively. The results show that it is feasible to extract 90% of the arsenic and 50% of the antimony at room temperature using only NaOH, and when the temperature increases to 95ºC, the extractions of As and Sb increase. Copper remains in the solid residue with less than 1% dissolution, and the NaOH net consumption was approximately 140 kg/t of dust. In leaching tests with NaOH + NaHS, two types of methods were tested, baking plus leaching with water and direct leaching with a NaOH+NaHS solution. With the second method arsenic extraction increased to 95% and copper dissolution was less than 0.01% at 85ºC, NaOH 100 g/L, NaHS 90 g/L, 1.5 h and L/S ratio of 5/1 (L/kg). The bismuth remains stable in the solid in all cases. INTRODUCTION The presence of arsenic in copper concentrates has increased with time and the environmental regulations are becoming increasingly more stringent. In concentrate treatment pyrometallurgical plants, most of the arsenic is volatilized condensing in the gas train and is collected along with the entrained solids to form the flue dusts. Under these conditions, in smelters of copper concentrates with high arsenic content, the dusts are being treated in hydrometallurgical plants to extract and stabilize the arsenic, which also means that most of the copper contained in the dust is co-extracted with the arsenic. This affects the recovery of copper in the smelter and an alternative such as the selective alkaline leaching of arsenic, would allow the copper from flue dust to be recovered.

Jan 1, 2019

By R. E. Whitehead

"Abstract - Biotite from aplite that appears to have crystallized simultaneously with disseminated Cu-Mo mineralization at MacLeod Lake in northern Quebec yields K-Ar dates in the range 2032 ± 44 Ma to 2097 ± 47 Ma. The mean value, 2065 Ma, is some 400 Ma younger than the 2485 ± 64 Ma date obtained from a hornblende concentrate separated from a granodiorite sill that is spatially, but apparently not genetically, associated with the disseminated Cu-Mo deposit. The mean biotite date of 2065 Ma at MacLeod Lake, a K-Ar date of 2055 Ma reported from an area 50 km east of MacLeod Lake and published dates of about 2000 Ma for galenas from the Chibougamau Cu-Au ores, suggest a widespread thermal and mineralization event, in that part of the Superior Province east of James Bay, that differs from the common late Archean mineralization that is related to volcanism and metamorphic-intrusive activity.Sommaire - Les datations au K-Ar sur des biotites qui auraient cristallise en même temps que la minéralisation disséminée en Cu et Mo dans une aplite mise en place au Lac MacLeod, au nord du Québec, s'échelonnent entre 2032 ± 44 Ma et 2097 ± 47 Ma. La datation moyenne, soit 2065 Ma, est d'environ 400 Ma plus récente que la datation obtenue pour un concentre de hornblende extrait d'un filon-couche de granodiorite qui est spatialement mais ne serait pas génétiquement associée à la minéralisation disséminée en Cu et Mo. La datation moyenne de la minéralisation au Lac MacLeod, soit 2065 Ma, la dataion au K-Ar à 2055 Ma citée pour une région à 50 km à l'est du Lac MacLeod, de même que les datations d'environ 2000 Ma établies pour des galènes qui se trouvent dans le minerai aurifère et cuprifère à Chibougamau, suggèrent un phénomène thermique répandu qui aurait produit la minéralisation dans cette région à l'est de la Baie James, dans la province du lac Supérieur. Ce phénomène serait diffèrent de celui qui aurait produit la minéralisation répandue d’âge archéen tardif qui elle serait associée à l’activité volcanique et à l‘activité intrusive accompagnée de métamorphisme."

Jan 1, 1992

By R. E. Bou Serhal

Many miners are exposed to dangerously high levels of noise on a daily basis. Over the past 15 years, a continually increasing number of occupational hearing loss has been reported from the mining community in the United States, of which, more than 95% is attributed to prolonged noise exposure. Although the noise exposure levels may differ between coal miners and metal and non-metal miners, in the absence of noise control at the source, the solution is the same: use of personal hearing protection devices (HPD). While protecting the miners? hearing it is also essential to no longer hinder their ability to communicate. With access to an advanced HPD that is customized to the miners? ears we are able to combine these two requirements. Using an intra-aural instantly custom molded HPD miners are protected from high levels of noise. The HPD is equipped with wireless capabilities, and contains both a speaker and an In-Ear Microphone (IEM). Therefore, the miners? speech may be captured from inside the ear and transmitted to the remote listener. This IEM signal is relatively noise-free since it is isolated from the background noise. The IEM speech signal, however, is ?boomy? and is missing some high frequency content, making fricative consonants hard to understand. Nonetheless, the IEM speech signal is correlated with the natural speech signal and may be manipulated through statistical techniques to more closely resemble natural speech. By improving the intelligibility and quality of the IEM signal, numerous applications may be enabled. One use of the enhanced IEM signal will be for radio communication. Using wireless radio communication in a noisy mining environment is sometimes the only practical and affordable solution to allow communication between miners equipped with personal hearing protection devices. Traditionally, one of the weaknesses of such wireless radio communication lies in the lack of designating receivers: whether they are the intended receiver or not, all those carrying a radio receiver are subjected to the broadcasted signal. The current work will detail a new concept of a "radio-acoustical virtual environment" where the radio signal will only be received by miners within a given spatial range, such range depending on the user?s vocal effort as well as the ambient and perceived background noise levels.

Aug 1, 2013

By Peter Heine

Diavik Diamond Mines Inc. (DDMI) has treated systems development as an integral part of the construction project when it built Canada?s second diamond mine. Fully aware that the proper collection and processing of operational data is needed to drive business decisions during production and planning, key Performance Indicators (KPIs) were defined to effectively manage operations when the mine went into production. The Production Information Management System (PIMS) allows KPIs to be measured, reported, and transferred to other systems in a reliable and convenient manner. PIMS provides a mechanism to collect real time production data from the various control systems around the site and acts as a tool to ?drill-down? and establish the important cause-effect relationships that impact the value of a KPI. Thus PIMS provides a data collection, storage, management, and an environment in which cause-effect relationships can be analysed and statistically verified. This understanding of the real cause-effect relationships, close to the source of the data and in the context of the operation, can be used to improve both operations and maintenance performance. The PIMS includes the PI Database implementation, its associated external interfaces with the various Control Systems, Cleaning Packaging Weighing System, Environmental System, Enterprise Asset Management (EAM) System, and Diamond Tracking System (DTS), as well as laboratory data collection, custom applications (downtime tracking, data reconciliation) and user clients. The interface with the EAM system provides material flow quantities and equipment availability on which the plant performance reports are based. The transfer of equipment usage to the EAM system is fed into the maintenance notifications and work orders functionality in the EAM. The environmental system interface provides the mechanism for capturing water consumption data required to generate the various regulatory reports. The interface to the cleaning-packaging-weighing system provides the mechanism for capturing batch data for transfer to the DTS for diamond product tracking at the production splitting facility in Yellowknife. The PIMS lab application provides a means of getting sample data into the system for critical metallurgical metrics, and the data reconciliation application provides a means of manually verifying the data before it is published or passed onto other systems.

May 1, 2003

By David S. Robertson

"IntroductionCanada has been one of the world's leading producers of uranium since atomic energy first became of practical importance in 1942. Indeed, prior to this time, between 1933 and 1940, Canada produced significant amounts of pitchblende, as a source of radium, from the Eldorado Mine at Great Bear Lake*.Uranium was first reported in Canada, however, in the middle of the 19th Century in the vicinity of Theano Point on Lake Superior north of Sault Ste. Marie. In the early 1900s, uranium-bearing minerals were reported in Quebec north of the St. Lawrence River and in pegmatites in Quebec and eastern Ontario. In 1930, G.A. LaBine and E.C. St. Paul visited an occurrence of cobalt reported by the Geological Survey of Canada in 1900, and brought the Eldorado Mine into production in 1933, as noted above, producing, in the main, silver and pitchblende.In 1942, developments in atomic energy were such that the Government of Canada, at the behest of Great Britain and the United States, requested that Eldorado, which had been closed in 1940, be re-opened to produce uranium for the Manhattan Project. In September, 1943, all discoveries of radioactive minerals were reserved to the Crown and their staking by private interests was banned. In January, 1944, the government nationalized Eldorado and formed a Crown corporation, the present Eldorado Nuclear Limited, to take over the assets. That year marked the formal beginning of uranium exploration in Canada.The Period 1944 to 1948In 1944, the Crown Corporation and the Geological Survey of Canada began exploring for uranium deposits, largely in the areas of Great Bear Lake and Lake Athabasca. In that year, A. W. Joliffe of the Survey visited properties on the north shore of Lake Athabasca near Beaverlodge Lake and recommended that Eldorado acquire claims in the area. In 1946, P. St. Louis and E. Laurum, prospectors working for Eldorado, found the pitchblende veinlets that led to the Ace Mine. Shaft sinking began in 1951, and production, through a carbonate leach plant, began in 1953.Continuing drilling on the Eldorado claims led to the Verna ore bodies, some 6,000 feet northeast of Ace, on which shaft sinking started in 1954."

Jan 1, 1986

By S. V. Burr

"IntroductionOne of the more puzzling geological enigmas, in the writer's opinion, is how rock can move sideways with respect to other rock. In the writer's study of regional northwest-southeast folding in North America (Burr, 1980), he may have found the answer to horizontal offset.Horizontal OffsetSideways movement or horizontal offset does not mean that the movement was, or is, horizontal. As indicated below, the actual movement may be at a high angle from the horizontal. However, it is the horizontal VECTOR, or offset, which is measured on surface by geologists. Vertical movements (thrust, normal, reverse) can be understood, as they involve blocks of rock , generally separated on four to six sides, which have been pushed up or subsided. Vertical faulting across dipping strata can produce APPARENT horizontal offsets. However, this thesis deals only with ACTUAL offsets.The ProblemFigure 1 is a plan view of a simple hypothetical horizontal offset. Three vertical dykes are intersected by a fault lineament. Two dykes exhibit no movement; the middle dyke has been offset a distance (a). Vertical faulting would have produced no horizontal offset. How can a horizontal distance (a) disappear or appear? Rock expands or contracts in infinitesimal amounts.SolutionThe map (Burr, 1980) shows the writer's interpretation of regional northwest-trending folds in Ontario. The writer made two observations about horizontal offsets during his thirty year study of these folds:1) All major horizontal fault lineaments strike at a high angle to the northwest fold axes. Many of the northwest-trending lineaments are fault s, but their movement is essentially vertical.2) Close to many of the known horizontal offsets (and several lineaments), there are areas where plunge reversals are so frequent and close together it is impossible to draw in all the fold axes at a scale of 1 inch = 4 miles, and, in the highly faulted Espanola area southwest of Sudbury, Ontario, all the indicated folds cannot be easily separated at a scale of 1 inch = 1 mile. The writer has observed folds less than 30 feet apart next to the strong Puskitamica-Chibougamau fault in northern Quebec."

Jan 1, 1982

By F. L. Pellet, S. Maiolino

"The determination of rock bolt contribution to reinforced joint shear strength is still a challenge, especially for assessing reinforced rock mass stability. For this purpose, several experimental studies have been performed during the last decades, mostly on small scale rock joint and bolt specimens. However, to accurately assess the rock bolt contribution, scale effect has to be accounted for. For this reason full scale lab tests are required as an alternative to in situ tests for which boundary conditions are more difficult to control.In this paper the results of an experimental program carried out with a large shear box, dedicated to test bolts in full scale conditions is presented. Different rock bolt diameters and different rock bolts inclination with respect to the shear plane were tested Experimental results were satisfactorily compared with an analytical model which is able to compute the contribution of the rock bolt to the shear strength of the joint as well the associated displacement of the joint. These conclusions allows us to validate the model in order to help the designer of reinforced rock engineering structures, such as slope or underground openings, to precise the optimum installation of rock bolts. INTRODUCTIONFully grouted rock bolts are commonly used for rock masses reinforcement to ensure the stability of rock slopes and underground cavities. For this purpose, the determination of rock bolt contribution to reinforced joint shear strength is of the utmost importance. During the last decades, several experimental studies have been performed to investigate the behavior of bolted rock joint. For obvious reasons, almost all the experiments were achieved on small scale bolted rock joint specimens (Bjurström, 1974; Azuar, 1977; Dight, 1983; Egger & Fernandes, 1983; Schubert, 1984; Spang & Egger, 1990; Egger & Zabuski, 1991). These studies made it possible to assess the influence of different parameters such as the bolt’s diameter, the number of bolts, the angle between the bolt and the joint, the nature of the rock, the friction and the dilatation angle of the joint. However, to accurately assess the rock bolt contribution, scale effect has to be accounted for (Vallier et al., 2010). For this reason full scale lab tests have been performed as an alternative to in situ test for which boundary conditions are more difficult to control."

Jan 1, 2015

By J. Ouellet

The authors participated in a study for a new mine shaft in Saskatchewan?s Athabasca Basin where ground freezing is desired to reduce water inflows. Constraints were imposed on the design in terms of permissible water leakage (less than 10 m3 / hour) inside the shaft. Since the shaft goes through highly permeable formations, water control measures had to be incorporated into the design. Ground freezing was selected as the most reliable and proven method for the sinking and a hydrostatic liner design was implemented. While exploring various technological options available for providing a fully hydrostatic liner, the authors came up with a new design concept that offers a more cost effective solution while meeting all the design constraints. The new liner concept is a composite design using concrete and a sprayed-on membrane system. A proof of concept program comprising laboratory testing and field testing was successfully implemented. This paper presents the innovative liner design, as well as the adapted sinking methodology, developed for this type of project. The new concept offers a new cost effective high performance hydrostatic mine shaft liner comparing favorably with traditional methods. RESUME: Les auteurs ont réalisé une étude pour un nouveau puits de mine dans le Bassin de l?Athabasca en Saskatchewan. Des limites sévères sont imposées sur le design en terme de volume d?infiltration d?eau acceptable (moins de 10 m3 / heure). Comme ce puits de mine traverse des formations très perméables la congélation des terrains a été sélectionnée, une méthode éprouvée et fiable. Un design hydrostatique a été choisi pour le revêtement du puits. Dans le processus de design, les différentes technologies disponibles ont été étudiées. Les auteurs ont développé un nouveau concept de revêtement composite offrant un rapport performance/prix plus avantageux que les technologies conventionnelles. Le nouveau design proposé est composé de béton et d?une membrane projetée. Un programme expérimental a été réalisé pour démontrer la faisabilité du concept. Cet article présente le nouveau concept ainsi que la méthodologie de fonçage de puits adaptée. Ce concept innovateur offre un revêtement hydrostatique performant à un cout avantageux par rapport aux techniques de construction utilisées jusqu'à maintenant.

Jun 1, 2012

By H. H. Wober, P. J. Morgan

"The proposed classification of ore reserves based on geostatistical and economic parameters is designed to be used as a communication tool to identify for investors, bankers, and financial analysts the degree of risk associated with a specific class of ore reserves. The class designation identifies the degree of accuracy of the estimates for tonnage and grade of annual production volumes of ore. It thus enables decision makers to quantify the risks and uncertainties inherent to the ore reserve in achieving over-all economic goals and annual financial objectives.IntroductionDuring the evaluation, in 1980, of the Creede silver deposit in Colorado, U.S.A., as a potential bulk mining project, the authors were faced with the problem of conveying to senior management the risks and uncertainties that are implied by using common or conventional classification systems to categorize ""ore reserves"". The estimated numbers for tonnages and grade of the projected ore reserves represented the only entry for future revenue from the project. It was, therefore, necessary to convey to management the level of risk associated with those numbers when implementing further stages of development, and to convey what impact significant differences would have on the profitability of the project. The desire to quantify these risks led to a revised definition of ""ore"" and to a re-examination of the classification system of ""ore reserves"". Traditional methods for estimating tonnage and grade of reserves do not include a quantification of the errors of these estimates. The advent of geostatistical methods made this quantification possible. The proper application of geostatistical methods, with adequate geological input and a reliable set of sample data, is essential to its validity.Others have proposed a classification of ore reserves based on geostatistical parameters. Diehl and David (1982) recognized ""that concepts of proven, probable and possible have to be linked to the production rate or time horizon"". Royle (1977) had also addressed this problem. Forsyth (1987) pointed out that bankers are concerned with ""the probability that the variability will fall within a certain degree of accuracy"" especially for grade estimates, and would like to see grades expressed together with the error of the estimate at the 950Jo confidence level."

Jan 1, 1993

By J. B. Evans

"THE OPPORTUNITY to participate in a CIM seminarconcerning the direction of mining engineer trainingis perhaps fortuitous as it comes at a time when theengineering schools which have invited inspection byCanadian Accreditation Board of the Canadian Councilof Professional Engineers are examining theirrespective curricula.Currently there exists considerable soul-searching in various engineering schools across Canada where questions such as the following are being posed.- Have we achieved the correct balance between subjects related to technical methodology and subjects devoted to conceptual planning and design?- Should we include ""applied humanities"" in our program? (Whatever ""applied humanities"" might mean!)- Should courses covering pollution control, ecology, economics, technical writing, professional conduct and ethics be offered as compulsory subjects, as electives, or not at all?- Is it a better scheme to have students complete an Arts or Science degree before entering engineering?- Do we place too much emphasis on pure mathematics and sciences with the result that insufficient time is available for instruction in the engineering applications of these subjects?- Should English expression and/ or literature be included in the engineering programs and, if so, should these subjects be presented by the English department or by English lecturers captive to the engineering school? - Is the present program directed primarily at research activities rather than preparing engineers for their future roles in industry? - Should we teach our own versions of mathematics and physics? - Do we have sufficient engineers on the faculty strength or are we overloaded with Ph.D.'s from the pure sciences? - How flexible should the program be, what should the compulsory vs elective course content be, and what freedom of choice, if any, should be given with respect to elective courses? - To what extent should students contribute to faculty decisions?- How far should we pursue computer science subjects and do we need to make our students proficient in computer programming? - How far should operations research and mathematical models be carried in the undergraduate program when the students have limited experience in the real engineering world? - Realizing that an engineer never completes his education, should the undergraduate program cover ""bare bones"" engineering and should instruction in particular disciplines be left to graduate schools and / or industry?"

Jan 1, 1972

By Euler De Souza, Kyle Penner

"Booster fans play an important role in assisting main fans by boosting the air pressure of the ventilation air passing through it. The correct placement of underground booster fans in a mine is a critical decision since it not only influences the mine’s operational capacity in terms of costs and efficiency, but also affects safety and environmental conditions. There are many booster fan installations around mines that are inappropriate, in some cases resulting in system failure (fan stall) or in a major negative impact on mine production. Inadequate booster fan placement has resulted, in some cases, in extensive air leakage which has compromised the economics of the operation and, in other cases, in recirculation of contaminants which has seriously compromised safety. This paper provides guidelines for proper design and siting of booster fans underground, to not only achieve an economic ventilation system but to create safe work conditions and efficient atmospheric conditions. Procedures for the planning, installation, operation and maintenance of booster fans are also presented. A case study illustrating the design and siting of a major booster fan installation in a potash mine, supported by computational fluid dynamics analysis, is presented.INTRODUCTIONIn large underground operations with complex ventilation networks offering high resistances, the never ending increased required quantities of air can almost only be supplied through use of high pressure main fan systems. These fans may, in effect, induce significant leakage losses in the system, having a negative impact on safety performance. In such cases, booster fans can be designed to decrease the main fan pressures and reduce the leakage flows, improve system efficiency and reduce costs. In well designed systems, booster fans can thus be used to redistribute the airflow and assist the main surface fans.When the pressure supplied by the main fans is insufficient to meet the ventilation circuit pressure requirements, booster fans are placed in strategic locations where pressure boosting is essential. Booster fans are typically sited in sections of the circuit offering high ventilation resistance and requiring this form of ventilating air motivation. Booster fans are thus required when the mine expands, and so the resistance of the ventilation network, and the primary fans no longer can supply the required pressure and flow."

Jan 1, 2019

By A. Boucher, D. Marcotte

"Abstract- The accepted practice for the estimation of thin (2D) vein deposits recommends the use of the grade x thickness service variable (i.e., the accumulation). Grade estimates are obtained indirectly by the estimated accumulation/estimated thickness ratio. This practice stems from the varying support (thickness) problem and the resulting non-additive nature of the grade variable. We compare the actual performance of the direct grade estimation approach used by some practitioners to that of the indirect approach using accumulation. Our simulated and real data indicate that the direct approach is more accurate for point grade estimation where the grade-thickness correlation coefficient is positive (and vice-versa). Moreover, the relative gain of the direct method increases with the (positive) correlation coefficient. This finding contradicts common thinking that the indirect approach should be the preferred method where grade-thickness correlation is strongly positive. Also, for a given positive grade-thickness correlation, the relative gain of the direct method increases with the coefficient of variation of the grade and thickness. Résumé - L’estimation de veines minces (2D) minéralisées est normalement réalisée à l’aide de la variable auxiliaire épaisseur x teneur (i.e., accumulation). Les estimés des teneurs sont alors obtenus indirectement du rapport accumulation estimée/épaisseur estimée. Cette pratique découle du problème du support variable pour la teneur, l’épaisseur de la veine, qui fait que la teneur est alors une variable non-additive. Nous comparons les performances de l’estimation directe de la teneur, utilisée par certains praticiens, à celle de l’approche indirecte. Pour l’estimation ponctuelle des teneurs, nos données simulées et réelles indiquent une meilleure justesse de l’approche directe quand la corrélation teneur-épaisseur est positive (et vice-versa). De plus, le gain relatif de la méthode directe s’accroît avec la valeur de cette corrélation (positive). Ce résultat contredit l’opinion répandue que l’approche indirecte devrait être utilisée en présence de forte corrélation positive de la teneur et de l’épaisseur. Nous montrons également que pour une corrélation positive teneur-épaisseur donnée, le gain de l’approche directe augmente avec les coefficients de variation de l’épaisseur et de la teneur."

Jan 1, 2001

By Réal Marcotte

L'industrie minière représente un apport économique important pour le Québec ; cette production est très diversifiée et répartie dans plusieurs régions. Hormis les forges du St- Maurice au XVIIe siècle, l'industrie minière québécoise commence en 1862 dans la Beauce avec l'exploitation de placers aurifères dans les affluents de la rivière Chaudière; à la même époque, on met en production les mines de cuivre Eustis, Albert et Capelton, près de Sherbrooke1; à partir de 1881, les mines d'amiante et de chromite au voisinage d'Asbestos et de Thetford commencent à voir le jour. Dans les années ?20, l'extraction de métaux précieux et usuels débute en Abitibi- Témiscamingue; durant les années '50, le fer et l'ilménite de la Côte Nord, ainsi que le cuivre de Murdochville et Chibougamau, commencèrent à être mis en valeur. Plus récemment, soit durant les années '80, le sel des Iles-de-la-Madeleine et le graphite des Hautes Laurentides s'ajoutèrent aux produits que nous commercialisons. Finalement, soit en 1998, le Québec devenait un important producteur de nickel avec le démarrage de la mine Raglan, au Nunavik. L'exploitation d'une mine peut nécessiter l'extraction de grandes quantités de roches stériles, particulièrement lorsque la production se fait à partir d'une fosse à ciel ouvert. De plus, étant donné que les minéraux et les métaux mis en marché ne constituent qu'une fraction du minerai extrait, il faut le concentrer : on se retrouve alors avec des proportions plus ou moins importantes de résidus finement broyés. Ces deux types de rejets totalisent bon an mal an environ 100 millions de tonnes. Lorsque cela est possible, ceux-ci sont, soit valorisés comme matériaux de construction ou retournés dans leur lieu d'origine afin de soutenir les murs des excavations. Toutefois, une bonne partie est accumulée en surface dans des aires aménagées spécifiquement à cette fin ; comme le montre la figure 1, la superficie utilisée à cette fin a progressé parallèlement aux booms miniers que nous venons de mentionner pour atteindre actuellement environ 15,600 hectares, dont environ la moitié sont situés sur des aires encore en activité.

Nov 1, 2002

By Dale A. Sketchley

"Abstract - Large differences in gold content between rock fragments or small volumes of unbroken rock, commonly referred to as the nugget effect, are related to sample and subsample weights, crushing and pulverizing sizes, and density, shape and size distribution of gold grains. Gy’s sampling theory, which relates gold grain characteristics to sample weight, fragment size, and error, provides a method of estimating a sampling constant that quantifies deposit characteristics that contribute to the nugget effect.Sampling control tests provide methods for determining the sampling constant. Sampling control charts, comprising nomographs that relate sample weight, error and particle size, use the sampling constant to optimize sample weights and comminution sizes, and to minimize the variability introduced by incorrect or inadequate sampling and sub-sampling procedures. The sampling constant is useful in classifying ores into arbitrary categories according to expected sampling difficulties. Spatial variations in the sampling constant can be used to aid in defining ore types for modeling geological domains, which should have unique sampling protocols. Examples that illustrate the relationship between the sampling constant and expected sampling problems, and the spatial variation of the sampling constant, are given for various deposits. Case histories of sampling programs are presented.Quality control procedures consist of monitoring sample contamination amd accuracy and precision of analytical results. Accuracy, precision and contamination are monitored by including standards, duplicates and blanks into batches of samples being analyzed. Analytical results of standards and blanks are plotted on a batch-and-time basis with control lines at plus or minus two/three standard deviations as confidence limits of the accepted round-robin value. Examples are given that illustrate a laboratory system in statistical control and the various types of problems encountered when interpreting quality control data. These include sample transcription errors, degradation, contamination, sampling errors, detection limit problems, instrumental drift, lack of statistical control, laboratory bias, procedural problems and data tampering.Where bias is negligible, precision, as a function of concentration, can be obtained from duplicates using the Thompson-Howarth precision algorithm, which also allows calculation of the practical detection limit. Examples are given that illustrate variations in precision, with grain size, for different types of duplicates, as well as mineralogical changes and sampling errors."

Jan 1, 1998

"One of the workhorses of the minerals industry is the autogenous grinding (AG) mill and its derivative the semi-autogenous (SAG) mill. These are relatively well understood, but not so Scrubbers. Scrubbers are mechanically of the same appearance/function, so what is the difference between an AG mill and a scrubber? When, or under what conditions should a scrubber be used and when should an AG mill be used? Can the processes within be mathematically modelled?In the past decades, the minerals industry has focused most of its attention on the modelling of unit processes such as cone crushers, HPGR’s, AG/SAG mills and ball mills within conventional processing circuits. Less attention has been placed on the modelling of scrubbers. This paper reviews the difference between scrubber and AG mill modelling, highlighting some of the similarities, with a view to confident comminution circuit designs.Engineering practitioners should be comfortable with these circuit designs and not shy away from considering AG mills for both small and large scale plants. AG mills can treat both harder or softer ore and scrubbers are for only softer ores. This paper differentiates scrubber and AG mill modelling in a similar way that overflow ball mills and grate discharge ball mills are analysed. Smaller scrubbers (600 kW) are suited to small scale soft ore (100-300 t/h) operations. Larger scrubbers up to 2,500 kW have the capability of very high throughput, in excess of 1,750 t/h with soft fine feed ore. AG mills are scalable up to 28 MW and are capable of treating harder ores at rates exceeding 2,000 to 4,500 t/h. Observations are that the specific energy requirements for the mill grinding duty is wide and varied. This is to a large degree the crux of the matter in sizing, as well as, differentiating, scrubbers from AG mills.INTRODUCTIONAG milling is a mature technology that has been used in minerals processing applications for more than 100 years. In the last 50 years, harder metalliferous ores have seen the introduction of steel balls to assist in the grinding/milling action. Tumbling mills used in such a configuration are called semiautogenous mills (SAG). The greater minerals industry is no newcomer to the use of fully autogenous (AG) mills. Perhaps scrubbers are. What is the difference? When should a scrubber be used versus an AG mill? How are they mathematically modelled?"

Jan 1, 2016

By I. Stefanou, J. Sulem

"In the last thirty years, the study of earthquake mechanisms has emphasized the major role of rock friction. Earthquakes appear to be the result of a frictional instability and occur by sudden slippage along a pre-existing fault or a plate interface. Multiphysics couplings play a major role in earthquake nucleation and seismic slip. The effect of shear heating and thermal pressurization may be of prime importance in the control of the stability of shear deformation inside the slipping zone. Other mechanisms in relation with thermal decomposition of minerals affect the pore pressure and the temperature evolution of the system, and can play a significant role in the extreme strain localization observed in field. The actual width of the localized zone is a key parameter for understanding fault weakening mechanisms. On the basis of field observations, laboratory experiments and recent advanced modeling, these phenomena are analyzed and discussed in order to bring some new insights in the understanding of earthquakes nucleation.INTRODUCTIONIn the last thirty years, the study of earthquake mechanisms has emphasized the major role of rock friction. Earthquakes appear to be the result of a frictional instability and occur by sudden slippage along a pre-existing fault or a plate interface (Scholz, 1998; Sykes, 1978). The conceptual model of spring slider system is commonly used to represent the interaction between an elastic crustal block and a fault zone. Stick-slip instability is possible if the softening rate of the shear resistance of the fault zone is high enough as compared to the stiffness of the crustal block. Therefore, weakening mechanisms that can explain how the frictional resistance of a fault decreases with slip and/or slip rate is of prime importance to understand how an earthquake nucleates. In recent years, several drilling programs in active faults (e.g. the Nojima fault in Japan, Boullier, 2011; the Aigion Fault in Greece, Cornet et al. 2004, the Chelungpu fault in Taiwan, Ma et al. 2006, and the San Andreas fault in the U.S.A. Ellsworth et al. 2000) as well as high velocity friction experiments (e.g. Tanikawa et al 2012a; Tanikawa et al. 2012b) have highlighted the physico-chemical processes involved in the nucleation of seismic slip (see for the example the review paper of Niemeijer et al. 2012)."

Jan 1, 2015

By Mary Kama, Kaul Gena

"Magnetic skarn ore (MSO) is one of the major copper bearing ore extracted by the Ok Tedi Copper Mine in Papua New Guinea. Copper minerals are recovered by flotation while the iron not associated with copper are discarded as tailings. The objective of this investigation was to determine the iron ore reduction kinetics for the Ok Tedi MSO and ascertain if it can be processed to produce sponge iron for a mini steel plant in Papua New Guinea.SEM-EDAX analyses of the Ok Tedi MSO indicated 10.1% C, 30% O, 0.6% Mg, 1.1% Si, 21.1% S, 0.8% Ca and 36.2% Fe. Most of the iron is in sulfide form. Both naturally occurring and roasted sinters of Ok Tedi MSO samples of different particle sizes were reduced by charcoal carbon at three different temperatures and seven different reduction times. Analyses of the reduced products indicated a metallic iron content of more than 65 wt.% on average.Results showed that there was no significant difference in reduction between fluxed and control materials. Only a slight increase in kinetics with reduced particle size, hence the reaction rate constant (K) did not vary much within the temperatures investigated. Reaction kinetics increases with increasing reduction time at 900°C. Therefore, more iron reduction is observed with particles of 106 µm size. In addition, the results also confirmed that the reduction energy used was higher at 800°C and lower at 1,000°C. In summary, iron reduction can be improved but close monitoring of temperature and reduction times are essential to determine the reaction kinetics of the Ok Tedi MSO. INTRODUCTIONMost mines in Papua New Guinea mine large quantities of iron in the form of sulfides, carbonates, oxides and sulfates that are eventually disposed off as tailings. The Ok Tedi magnetic skarn ore (MSO) contains a significant amount of hematite, magnetite, siderite, ilmenite and pyrite that can be a source of iron. The mineralogical compositions of Ok Tedi MSO reveals more iron minerals than copper, yet all iron containing minerals are collected in the tails from the main copper production stream. Therefore, this study is conducted to investigate the possibility of producing direct reduced iron (DRI), or sponge iron, from the MSO. Such sponge iron may serve as the feedstock for electric arc/induction furnace based mini steel plants in Papua New Guinea."

Jan 1, 2016

By R. A. Frith

"Belts of Yellowknife Supergroup volcanic and sedimentary strata in the Slave Structural Province are separated by granitoid complexes. The strata and most intrusive rocks are 2.7 and 2.5 Ga old, but rocks around 3.1 Ga have been recognized in some of the complexes and are believed to represent a once-extensive sialic crust. Volcanics were erupted along rifts in this crust and the rifts became sites of crustal foundering and the development of troughs that filled with turbidites of volcanic and granitoid provenance. Present margins of the granitoid complexes may be close to the original margins of deposition of volcanics, with associated mineral deposits, and close to the thickest parts of basins of sedimentation.The Hackett River volcanic belt extends 100 kilometres between Mara River and Back River and is crudely S-shaped, with a northern sector surrounding a core of granitoid basement rocks, a northeasterly facin g central sector and a southerly facing southern sector. Up to 8 kilometres of andesitic volcanics with intercalated felsi c volcanic rocks are preserved within the central sector. The f elsic rocks include thick lenses of volcaniclastics with large, quartz-eye-bearing clasts, but the known metal deposits are not associated with these. Both of the deposit areas, the Hackett River area in the northern domal sector and the 'A itch ' Lake area 40 kilometres to the south in the central sector, are near the tops of relatively thin atypical volcanic sequences above suspected basement rocks. The Yava deposit near 'A itch , Lake overlies a unit of subaerially deposited, welded, andesit ic ash flows and the Hack ett River deposits overlie highly metamorphosed volcaniclastic rocks that include suspected ignimbrites.The deposits are notably argentijerous and lead-rich. They consist of stratigraphically zoned sheet-like bodies of iron sulphides, sphalerite, galena and chalcopyrite underlain by altered rocks containing chalcopyrite stringers. The Yava deposit was formed where hydrothermal fluids exuded through the fractured floor of a flooded caldera basin. The Hackett River deposits may have been formed in a similar setting and it is not inconceivable that all the known deposits in the belt were formed in a single large caldera basin, although there is no independent evidence for the required major fault separation between the Hackett River dome and the 'Aitch' Lake area."

Jan 1, 1980