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The theory and practice of sampling of particulate material have been the focus of a continued attention on the part of the mining industry worldwide for almost a century now, because of its critical importance at all phases of a mining project, from exploration (bulk sampling, assaying) to production (grade control) and marketing (quality control). A powerful, complete theory was put together by P. Oy (1968 to i982), but its applications to gold ores have met with limited success to date, due to the inadequacy of its sampling variance and minimum sample mass determination formulae to the challenging case of gold ores. This paper presents a research project completed at Inco Exploration and Technical Services Inc., using a geostatistical approach, to shed light on the results of P.Gy's theory, and adapt them to the case of particulate gold ores. After a reminder of Gy's formulae and a simplified demonstration of the generic form of Gy's and more traditional formulae, an empirical approach, based on sampling experiments from four gold deposits, shows how Gy's formulae must be corrected to adjust to the case of gold ores. The generality of the results formula is then substantiated with a geostatistical analysis of the problem. A parameterized formula results, and indications are given on how to fit its parameters to experimental data. Application guidelines and charts to fit any case and budget are presented, along with a detailed practical example.

Jan 1, 1991

By H. Zniber El Mouhabbis

Ore dilution or stope overbreak increases the cost of production and ultimately reduces the profitability of a mining operation. The adverse economic impact of ore dilution is due to the added costs associated with the mucking, haulage, crushing, hoisting, milling, and treatment of waste rock or low grade ore in the hangingwall and footwall. One factor that could have a significant influence on stope overbreak is the undercutting of the stope into the walls. Undercutting stope walls is often necessary, in particular in narrow vein mines. Moreover, equipment size has a role in drift sizing where stope undercutting may become a must in order to fit the mobile equipment while respecting legal clearances.

May 1, 2009

By Peter A. Witt

Highland Valley Copper is a conventional truck-shovel mining and flotation concentrator operation located some 75 kilometres southwest of Kamloops, B.C. (Figure 1). The ore is located within a porphyry copper deposit and is mined at a rate of 135,000 metric tonnes per day from two separate open pits. The nearby mill produces concentrates containing 375 million pounds of copper and 5 million pounds of molybdenum metal yearly. In the Valley pit, head grades are roughly 0.40% Cu at a 0.23% Cu equivalent cut-off. The Valley pit strip ratio currently sits at 0.55:1 with waste quantities made up of separate rock and unconsolidated overburden units. From the mill, concentrates are trucked to nearby Ashcroft and then hauled by rail to ocean port facilities in Vancouver. The mine employs 950 persons and is planned to operate until 2009. The operation is a partnership between TeckCominco with a 63.9% interest, BHP Billiton holding 33.6%, and others 2.5%. Management is shared by the two principal owners and directed by a management committee.

May 1, 2003

By C. Jay Hodgson, Adrienne C. L. Larocque

The Mobrun VMS deposit in northwestern Quebec does not fit the classic Noranda type model, but exhibits characteristics more consistent with a Mattabi-type deposit. The orebodies are underlain by fragmental volcanic rocks indicative of explosive volcanism in shallow water. Felsic volcanic rocks are more abundant than mafic rocks. Alteration assemblages comprise chlorite, carbonate, sericite, and quartz. Broad semi-conformable alteration occurs high in the footwall to the 1100 lens. Mobrun occurs higher in the stratigraphic sequence than most of the other deposits in the Noranda camp. In this part of the stratigraphy, zones of pervasive carbonatization may be more useful guides for exploration than zones of silicification.

Jan 1, 1993

By W. W. Muir

"The Baker Mine used a 100-ton per day conventional cyanide leach mill located in northern British Columbia. A production decision was made in early 1980 and the mill was on fine in April 1981.Ore reserves had been estimated at 100,000 Ions averaging 0.9 oz gold per ton and 18 oz silver per ton . Mineralization consists of electrum (Au/Ag) and acanthite (Ag2S).This paper will review plant design and focus on some problems associated with start-up and early operation with the intention of assisting other mine operators by indicating potential problem areas.IntroductionThis paper will discuss operating strategies employed at Baker which improved over-all mill performance. Also included are some ideas which could be steps toward a more efficient operation in a similar setting.Actual dore production by mine life was 37,606 oz gold and 742,117 oz silver.Figure 1 illustrates the relative stability attained over the mine life. Metal recoveries for gold and silver were 92% and 85% respectively.Safety at Baker was of prime importance because of its remote location. Baker received the Mine Safety Association ""Small Mines Award "" in two of its three operating years. At shutdown in late 1983 the mill had operated for over 850 days without a lost-time accident."

Jan 1, 1985

By Marc Arpin

La Bolivie possède tout pour réussir. Son vaste territoire de plus d'un million de km2 et qui occupe une position stratégique au coeur du continent sud-américain recèle des ressources naturelles (forêts, minéraux, gaz naturel, etc) variées et abondantes. Elle a une frontière avec pas moins de 5 pays dont le géant que constitue le Brésil ainsi qu'avec l'Argentine, l'autre puissance régionale. Pourtant, malgré tous ces atouts, la Bolivie demeure l'un des pays les plus pauvres du continent. Ce paradoxe n'est nulle part mieux reflété que dans le secteur minier. Celui-ci fait face à une situation de crise en dépit d'une tradition minière de plus de quatre siècles et d'un potentiel géologique indiscutable ayant conduit à la découverte et à l'exploitation de fabuleux gisements d'argent et d'étain. Les habitants des districts miniers vivent dans des conditions de pauvreté extrême et les indicateurs de développement humain de ces régions sont parmi les plus bas en Amérique du Sud. La Bolivie réussit difficilement à attirer les compagnies minières alors que, chez les voisins, et particulièrement en Argentine, qui n'a pas de tradition minière, les investissements miniers majeurs affluent. Dans cette conférence, nous chercherons à expliquer les causes de cette situation et proposerons des solutions possibles.

May 1, 2001

By W. Dan Hausel

"Hundreds of diamonds have been found in the United States. The majority of the discoveries occurred during the 19th century as a by-product of placer gold mining. Most of these early diamond finds have no known source, and in many cases are situated in areas unfavorable for diamond exploration based on current models that require kimberlite or lamproite to be emplaced in a stable Archean craton or Proterozoic craton margin.Diamonds have also been recovered in situ from lamproite at Murfreesburo, Arkansas, from kimberlite in the Lake Ellen region of Michigan, and in several kimberlites in the Colorado–Wyoming State Line district. Unverifed reports also suggest that diamonds may have been recovered from other related intrusions at a number of localities in the United States.IntroductionTo date, diamonds have been recovered in situ from kimberlite in the Colorado–Wyoming State Line district and from a kimberlite in Michigan. Additionally, diamonds have been recovered from lamproites in Arkansas. Diamonds have also been reported from kimberlite in Kansas, and from peridotite in New York and Maryland. However, these latter reports have not been verified."

Jan 1, 1995

By Paul Ransom

Sullivan Production and Present Value ? Mined 150 million tonnes ? 6.0% Pb, 5.5% Zn, 2 opt Ag, + Sn, Cd, In ? May 2006 Value + US$43 Billion ? 50% of metal from first 37% (55MT) of tons ? True original size ~ 110-120 Mt @ 17%Pb+Zn ? 30MT unconsolidated fill + HW cave

May 1, 2006

The New Zealand opportunity NZ?s only significant phosphate deposit 1 million tonnes imported annually - mostly from Morocco US $180/t + $75/t freight

May 1, 2013

By Eirik Sundvor, Hany Doss, Peter Vogt, Kathleen Crane

"Abstract - In 1989-1990, the SeaMARC II side-looking sonar (11 to 12 kHz) and swath bathymetric system imaged more than 80 000 km2 of the seafloor in the Norwegian-Greenland Sea and southern Arctic Ocean. One of our main goals was to investigate the highly oblique (115°) junction between the Mohns and Knipovich ridges, and to determine, using new morphotectonic data, whether or not this junction is stable or in a state of readjustment. These findings are presented in this paper.Along the northern Mohns Ridge, we observed a highly segmented rift valley undisturbed by transform faults. In contrast to the elevated Reykjanes Ridge, the Mohns Ridge is dominated by deep en echelon fault-bounded troughs, which are bordered by high rift valley volcanoes. These segment boundaries are interpreted to be sites of recent to present volcanic activity, based on their high acoustic reflectivity. The northwestern flank of the Mohns Ridge is marked by shallow quasi-linear ridges parallel to the present orientation of the Mohns Ridge. The orientation and geographic position of these ridges suggest that they were formed at the Mohns Ridge. Only in the northernmost cases are some of these off-axial ridges non-linear. The degree of the non-linearities increases slightly with proximity to the Mohns-Knipovich bend. In addition, because no transform fault was detected between the Mohns and Knipovich ridges, we conclude that the Knipovich Ridge must be obliquely spreading to the NW-SE."

Jan 1, 1999

By C. B. Webber

"Sherritt Gordon's experience with rubber mill liners began in 1964 at the Lynn Lake Farley Mine and continued at the Fox Mine in 1960, and the Ruttan Mine in 1973. This paper describes the history of the performance of rubber mill liners at the three mine operations.LYNN LAKE - FARLEY MINEThe Lynn Lake Farley Mine, which ceased production in 1976 due to exhausted ore reserves, was a 3500 T.P.D. copper-nickel operation; consisting of two parallel rod mills followed by three stages of ball mill grinding. A circuit modification in 1964 necessitated one of the tertiary ball mills to be used as a pyrrhotite regrind mill. This 8' x 5' Hardinge conical mill drawing 225 HP was initially installed with steel liners and lifters. After several months of operation a routine inspection revealed severe channeling and washing behind the liners which was threatening to erode the ball mill shell. It was thought that the highly acidic condition in the mill, due to grinding 100 º6 pyrrhotite, was accelerating the wear of the steel liners. At this time it was decided to reline the mill with rubber liners to see if the wear problem could be relieved, and in December 1964 a set of Skega rubber liners were installed."

Jan 1, 1981

By Pascal Turcotte

_ Introduction _ LaRonde _ Controlling Instability in Haulage Drifts _ Failure mechanism in drifts walls _ Failure mechanism in drifts back _ Field Behaviour of Support Elements _ The ?Hybrid Bolt? _ Conclusions

Feb 1, 2008

By T. A. Postolski, A. J. Sinclair

"Abstract - Where block (selective mining unit) grade distributions can be determined, the effect of average random errors of block grade estimates on metal recovery can be evaluated quantitatively. That is, for a given estimation error and cutoff grade, it is possible to calculate the quantity of metal that is lost as a result of misclassifying ore blocks as waste, as well as the dilution that ensues from misclassifying waste blocks as ore. Example calculations using a computer program GAINLOSS and realistic block grade distribution parameters for both a porphyry-type deposit and an epithermal gold deposit illustrate some fundamental relations that are important in quality control of estimation and in reconciliations concerning a comparison of ore estimates with production.Where the cutoff grade is on the lower tail of the grade distribution, metal arising from dilution can be much less than metal lost through misclassifying ore as waste. Hence, average grade of milled material could possibly be higher than expected (estimated) and tonnes milled will be smaller than estimated.Where the cutoff grade is on the higher tail of the grade distribution, tonnes arising from dilution will be greater than tonnes lost by misclassifying ore as waste. Thus, tonnage mined will be greater than estimated and the average grade milled will be less than estimated.The software GAINLOSS permits rapid comparison of the tonnes and average grade for all blocks selected as ore for a variety of levels of block estimation errors. These comparisons can be incorporated in a financial analysis to evaluate whether or not it is worthwhile to improve a high, average block estimation error to some lower value."

Jan 1, 1998

By J. W. Scott, G. G. R. Falkenberg

"The Syncrude project is an integrated complex designed to mine, extract and upgrade bitumen from the Athabasca tar sands to produce 129,000 barrels per day of synthetic crude oil. This paper provides a brief general description of the location, history and geology of the tar sands and the overall organization and structure of the project. The major emphasis and the bulk of the detail presented deals with the extraction plant; a material handling and mineral processing installation designed to treat 312,000 tons per day of tar sand feed.History of the Tar SandsThe Athabasca tar sand deposit was recorded in 1788 by Peter Pond, however the first serious geological mapping and exploration did not begin until 1915. Little interest was taken in the deposits until the Second World War, when the governments of Canada and Alberta conducted an extensive exploration program. In addition a number of pilot plants were sponsored during the 1940 -1950 decade. The hot water separation process was successfully demonstrated on the pilot plant scale during 1948 -49, and was established commercially in 1967 in the G.C.O.S. plant. The Syncrude plant, due to come on stream in April 1978, also employs the hot water separation process."

Jan 1, 1978

By T. Yalcin

"This paper describes a process that can provide high rates of dewatering and that has the potential of bringing significant savings in tailing disposal costs. It involves a three-stage, closed-circuit operation employing a hydrocyclone, a settling cone and a thickener. The mill tailing, with adjusted pulp density, is first fed to the hydrocyclone. The underflow from the hydrocyclone goes to the settling cone for further dewatering. The overflows from the hydrocyclone and the settling cone are sent to a thickener which, with the addition of a flocculant, gives a clear liquid effluent and a dilute underflow which is recycled back to the hydrocyclone. The underflow of the settling cone is a material with as high as 78% pulp density and is suitable for direct disposal or for use as mine backfill.INTRODUCTIONDewatering is an essential process for the disposal of mill tailings and forms an additional cost item in milling operations. Cost-cutting measures are therefore desirable in this area as in any other stage of milling. One way of achieving this is through increased rate of dewatering which would require smaller dewatering units with correspondingly lower capital and operating costs."

Jan 1, 1987

By Pierre Bernard

Contents of the presentation ? Overview of Brunswick 12 Mine ? Brief description of the mine geology ? Summary of the current Mineral Reserves and Additional Mineral Resources ? Review of the ongoing mining remnants evaluation process, particularly emphasizing the integrated work between the mine geologists, the mine engineers and the metallurgists ? Challenges and opportunities to extend the mine life

Apr 1, 2005

By Jean Robitaille

La découverte du gisement du puits Penna a nécessité une révision complète de tous les circuits du concentrateur Laronde pour permettre, premièrement, l?adaptation du procédé à la nouvelle minéralogie et deuxièmement, l'accroissement du tonnage de 1800* à 4500 t/j. Aux circuits existants de broyage, de flottation du cuivre, de cyanuration et du charbon en pulpe, les circuits suivants furent greffés : broyage semi-autogène, flottation du zinc, manutention des concentrés (épaississage, filtration, entreposage et chargement automatisé), décantation à contrecourant et précipitation au zinc, destruction des cyanures et remblai en pâte. Cet article se veut une revue des différentes étapes du projet d'agrandissement du concentrateur qui se sont effectuées par phases progressives de 1997 à 2000. De plus, une description des nouvelles technologies sélectionnées et implantées est présentée.

May 1, 2001

By F. W. Wellmer

"This paper discusses the changing nature of reserve/consumption ratios, using data from Canadian operations and focussing particularly on copper.IntroductionStudies in world metal consumption and world metal reserves are done regularly by the U.S. Bureau of Mines, the German Federal Institute of Geosciences and Raw Materials, and other institutions. Formerly, on the basis of these data, the following figures were calculated:• a static lifetime of reserves-reserves at year end divided by the consumption of the previous year;• a semidynamic lifetime of reserves-reserves at year assumed constant growth rate.These figures were severely criticized because they were purely theoretical and did not consider the availability of the raw materials.Recent practice is to calculate reserve/consumption ratios (R/C ratios) for periods(1) or for different years(2,3) and to interpret the trend. For base metals, these R/C ratios are increasing (Table 1). This had led to the conclusion that today the relative availability of base metal reserves has increased as compared to 10 or 15 years ago(4,5) As there are normally numerous deposits for each commodity, the anticipated production rates of deposits are irrelevant to this conclusion as long as the statistical distribution of possible production rates does not change significantly.The above conclusion is only justified if the quality and the nature of the data base remain the same. This is certainly the case for the consumption data. It is also assumed for the quality of the reserve figures, which usually take into account only proven and probable reserves. For the nature of the reserve figures, however, this does not seem to be true.Undoubtedly, the reserves increased in absolute terms between 1966 and 1978. To put the absolute reserve figures into perspective, it is common, as said above, to compare them with the mine production, which increased in the free world between 1966 and 1978; e.g. for copper by 40%, for Pb by 17% and for zinc by 36%. However, the reserves are also influenced by other factors, such as average size of the mines, changes in the grade of the mined ore, price-to-cost relationships and the like.The purpose of this paper is to examine the nature of the reserve figures in more detail. Because the completeness of mining data published in Canada is rather exceptional, Canadian .data serve as an example and it is assumed that the conclusions can be extrapolated to other mining countries or at least to those with free market systems."

Jan 1, 1981

By G. J. Desroches, E. Ozberk, M. J. Collins, B. D. Krysa, I. M. Masters

The zinc refinery of Hudson Bay Mining and Smelting Co., Ltd. went through a major modernization in 1993, when the roaster and the calcine and fume leach circuits were replaced by a new Sherritt zinc pressure leach plant. The HBMS plant, in Flin Ron, Manitoba, represents the first commercial application of the Sherritt two-stage zinc pressure leach process, with /00% of the zinc concentrate feed processed in autoclaves. The process design of the new plant required definition of the deportments of elements such as antimony, arsenic, chlorine, cobalt, fluorine, germanium, magnesium, manganese, selenium, tellurium and thallium. This paper describes the deportment of these elements in the commercial zinc plant during the first nine months of pressure leach operation, and compares these results with operations prior to the introduction of pressure leaching, and with the results of tests carried out in the She"itt laboratory during the process development phase.

Jan 1, 1995

By Steve R. McCutcheon, Cees R. Van Staal, John P. Langton, Leslie R. Fyffe

Abstract -The volcanic-dominated Tetagouche Group in northern New Brunswick contains numerous massive-sulfide deposits and intimately associated metalliferous sediments. The massive sulfides are generally hosted by Llanvirnian-Llandeilian volcaniclastic sedimentary rocks of the Nepisiguit Falls, Patrick Brook, and Boucher Brook formations. These formations are intercalated or closely associated with rhyolitic-dacitic flows and/or domes, and minor continental tholeiites of the Flat Landing Brook Formation. Pyroclastic rocks are relatively scarce in the silicic volcanic complex. Tuffites and epiclastic sandstones, of the Nepisiguit Falls and Patrick Brook formations, respective­ ly, generally occur along the fringes of the silicic volcanic complex. The flows and tuffites are intruded by quartz- and feldspar-phyric rhyolite (porphyry) dykes, sills, and domes. Although the volume of the silicic volcanic complex is large, evidence for a previously proposed collapsed-caldera model is lacking. On the basis of host-rock type and stratigraphic position, the sediment-hosted massive sulfides have been divided into three groups: the Brunswick-, Caribou- and Halfmile Lake-type deposits. Syngenetic feeder pipes and asociated alteration zones are generally not present beneath the Caribou- and the Halfmile Lake-type deposits and, if present at all, are poorly developed underneath the Brunswick-type.

Jan 1, 1992