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By C. A. Lorenson

THE Marmoraton mine is located a mile southeast of Marmora, a village in Hastings county, Ontario, 130 miles east from Toronto and 143 miles west of Ottawa. Highway No. 14 from Belleville intersects Highway No. 7 at Marmora and the branch line of the Canadian National railway between Trenton and Bancroft passes through the property just east of the open pit. HISTORY AND GEOLOGY The Marmora orebody appears as the whorls of a thumb print on the aeromagnetic survey map released by the Ontario Department of Min.es early in 1950, the survey of the district having been made in 1948 and 1949, under the joint auspices of the Department of Mines and the Geological Survey of Canada. The 'anomaly' appeared to have interesting possibilities, so Bethlehem representatives carried out a ground survey and began drilling the property in June, 1950. A total of 40 holes, aggregating 37,000 feet, were drilled to delineate the ore-body. The Marmora orebody is a re-placement of Precambrian limesto.ne with zones or bands of magnetite interlaid with metamorpl1ic rocks and cut by later intrusives. The ore and associated Precambrian rocks are overlain by 100 to 150 feet of flat-lying Palaeozoic limestone. It is a low-grade deposit averaging 33 per cent iron and requires concentration to raise it to 65 per cent iron grade.

Jan 1, 1958

By William C. Boggis

EKATI? Diamond mine is located 300km northwest of Yellowknife and 200 km south of the Arctic Circle in the Northwest Territories of Canada. The site is a remote tundra area with continuous permafrost and is in the Canadian Shield. EKATI? is the first diamond mine in Canada and is operated by BHP Diamonds Inc., which is a part of BHP Minerals. BHP Diamonds Inc. is an incorporated company based in Vancouver, BC. The mine is a joint venture of BHP Diamonds Inc. (51%), Dia Met Minerals Ltd. (29%), Charles E. Fipke (10%) and Dr. Stewart L. Blusson (10%). The mining area lies within the 10,960 ha land leases held by the joint venture. The total claim block covers 344,000 ha. To date over 136 kimberlite structures have been identified within the claim. Five of these provide the basis for the present mine plan with three others identified for future inclusion. At full capacity the mine will produce 4.5 million carats per year, approximately 4 % of world production and employ approximately 600 people, with over 33% being Aboriginal from the Northwest Territories.

May 1, 2001

By Wayne F. Luinstra, Frédérique Vincent, André R. Laplante

"A test to determine the amount of gravity recoverable gold (GRG) in an ore is described. Typical GRG results are given; possible uses are discussed. A case study of how the technology was used at Hemlo Gold Mines is presented.IntroductionUsing gravity to supplement either flotation or cyanidation is a well established practice in the gold industry. It differs from most recovery circuits in that most of the gold recovered by gravity would be recovered in any case by the circuit downstream, be it flotation or cyanidation. The economic justification of gravity is therefore based on small margins (for example. a net smelter return of gravity gold of 99%, as opposed to 94% by flotation). It was easily demonstrated when either flotation or cyanidation were relatively inefficient processes, and labour costs low (as gravity can be labour intensive). Over the past thirty years, the advent of better flotation machines (flash, column, high capacity), more effective collectors, and better control systems has increased flotation's metallurgical performance, thereby decreasing the incentive for gravity recovery. Cyanidation technology has undergone similar changes, with the advent of activated carbon, oxygen and lead nitrate addition, and improved impeller design.Today, gravity could remain an attractive option only inasmuch as it can be implemented with very low capital and operating costs. This has resulted in a relative shift away from gold gravity recovery (except for alluvial deposits), in the seventies and eighties. For example, as of the early nineties, gold gravity recovery has disappeared from the typical South African flow sheet. The advent of the Knelson Concentrator, at the beginning of the eighties, foreshadowed a resurgence of gravity recovery, as gold's grinding and classification behaviour (Banisi, Laplante and Marois, 1991) makes it possible to achieve adequate gold recoveries with very simple, Knelson based, gravity circuits (Laplante et al., 1994).Consider for example the Hemlo recovery circuit, made of a single 61 cm x 122 cm (2' x 4') screen, an automated 76 cm (30"") Knelson feeding a Gemeni table, to produce a 70-80% Au concentrate which accounts for approximately one fourth of the gold production (about 3 million grams per year). The low capital and operating costs make it possible to justify the gravity circuit on the basis of savings in other operating costs (e.g. fewer carbon stripping and reactivation cycles). Any improvements in recovery, which would at Hemlo be difficult to measure, would be additional benefits. Such a circuit is rapidly becoming a standard in Canadian and Australian operations.The justification for installing a gravity circuit must be first based on an estimate of how much gold it will recover (irrespective of what the potential benefits are). This is a function of the nature of the gravity circuit to be installed; however, Knelson circuits, by their very simplicity and efficiency, limit the options. This leaves the amenability of the ore to gravity recovery as the single most important parameter. Various approaches can be used to characterize this response (Woodcock, 1993). The present work sought a route which would fulfil the following criteria:"

Jan 1, 1996

By G. P. Holmes

"Bell Copper is situated (Latitude 55°0''N, Longitude 126°14W) on Newman Peninsula on the east side of Babine Lake at an altitude of 2,450 feet above sea level. The mine site is approximately 170 miles north west of Prince George and 40 miles east of Smithers. Access to the property is by way of 30 miles of paved road to the Village of Granisle on the west shore of Babine Lake and then a further 9 miles of unpaved road to Bell Copper's west dock. The two mile lake crossing is made using a tug and barge. A compressed air bubbler system maintains an ice free channel for the barge during the winter months.Temperatures range from summer maximums of around 32 °C to winter minimums in the region of minus 40°C. Long periods in winter may be spent with temperatures in the range of minus 15""C to minus 28°C. Freezing temperatures usually start in October and last for six months. The lake normally freezes over in mid December and the ice breaks up in mid May. Snow accounts for approximately two t11irds of the annual precipitation of twenty-one inches."

Jan 1, 1976

By Mark Stephenson, Paul Henharen

"The objective in the mineral processing of primary ores is to liberate and effectively recover each individual economically viable diamond with minimal reduction in its potential value.There are three distinctive phases in diamond processing:Feed PreparationWhere the diamonds are liberated and waste material finer than the economically selected cut-off size is rejected. In this section we will discuss the various techniques utilised worldwide to effect feed preparation.Primary ProcessingWhere the diamonds are recovered into a high grade concentrate suitable for the more expensive and selective final recovery process. In this section we will discuss the development of primary processing from jigs and pans to the more current techniques utilised worldwide.Final RecoveryWhere there is selective rejection of waste particles until only diamonds remain. The diamonds may then be acidised prior to packing and marketing.Within these distinct phases a variety of unit processes have been utilised and this paper will discuss their relative merits, limitations and suitability in respect to the various stages of diamond project development from prospecting through to production.Basic capital and operating costs are compared for the autogenous milling route developed in the CIS and the crushing and scrubbing route utilised in the west."

Jan 1, 1994

By Paul J. Beck

The impact of mining development on the environment can be minimized through comp careful construction and accurate monitoring. Key to meeting these objectives is t information at all stages of design, development, operation and decommissioning. This paper will describe cost-effective means of gathering and managing reliable fiel groundwater flow and surface water regime within and surrounding the mine site, as localized contaminant plumes associated with mine operations, and monitoring ground quality. Recent research has led to the development of inexpensive, easy to install syst sampling/monitoring from different zones within a single borehole. Advances in elect developed small diameter probes and pumps for measuring water levels and taking wat continuous monitoring stations in remote areas can be colldeacttaeldo gaguetrosm aatnidc adlolwyn luosaidnegd from a central facility using telemetry systems powered by direct, battery or solar radio frequency, landline or cellular telephones, and satellite. Two case studies will illustrate the use of datalolgagregres sacnadl et eblaesmient rwyi dien mao nitoring sc and the use of multilevel samplers for managing a contaminant plume.

May 1, 2004

By K. V. Konigsmann

"The mill at Mattagami Lake Mines treats 3,850 tons per day. The mill feed is a heavy sulphide ore. It is composed ofPyrite & Pyrrhotite - 40%Marmatite - 17%Chalcopyrite - 2%Magnetite - 10%Talc and other Silicates - 31%The sulphides are intimately associated with each other and often finely disseminated throughout the siliceous gangue. An abundance of middling particles exist in the circuit and a delicate balance has to be maintained between concentrate grades and metal recoveries.The metal content of the mill feed fluctuates between 0.4 to 0.7% for copper and 6 to 13% for zinc.The flow sheet of the concentrator is shown in Figure 1.Eight sample streams have been selected for on-stream analysis, namely1. Copper Concentrate 2. Copper Cln Tailings 3. Copper Tailings 4. Zinc Concentrate5. Zinc Rgh Concentrate6. Zinc Cln Tailings7. Zinc Rgh Tailings8. Flotation TailingsThe mill feed is not being analysed. Only a very coarse sample could be obtained which would be difficult to pump, and the fluctuations in the copper content are too small in absolute terms to have much of a bearing on circuit operations."

Jan 1, 1969

By Neri Roux, Ross Gilders, Yemi Oyediran, Leo Cheung

Quest Rare Minerals Ltd. is developing their Strange Lake B Zone REE deposit located in Northern Quebec. An extensive batch bench scale testing program was launched to investigate open circuit rougher flotation performance under different conditions. The objective of the flotation study was to demonstrate the effectiveness of rejecting gangue minerals to flotation tailings while concentrating rare earth elements as well as yttrium, zirconium, and niobium into a concentrate suitable for further hydrometallurgical processing. Mineralization of interest at Strange Lake occurs within peralkaline granite-hosted pegmatites and minor aplite. The REE and high field strength element (HFSE)-bearing phases are hosted primarily in pegmatites as relatively fine-grained phases or pseudomorphs. The gangue phases comprise of quartz, feldspar, amphibole, and pyroxene. Potential ore minerals comprise predominantly of kainosite, gerenite, gadolinite, zircon, pyrochlore, and gittinsite. These complex mineralogical characteristics increase the difficulty for processing and provide a challenge for choosing the correct reagent scheme in rougher flotation. Process options including reverse flotation to reject the gangue minerals and de-sliming prior to the direct rougher flotation were investigated. Optimization trials were conducted using various reagent regimes and under different flotation conditions. Encouraging results were found using a combination of collector reagents supplied by Cytec Canada Inc.

Jan 1, 2014

By J. Gillies

"Extraction of Gold in Ketone from a suitably prepared sample of sulphide or siliceous ore or concentrate permits a rapid and accurate determination by Atomic Absorption. Mill test products and exploration development samples are acid-treated to effect complete dissolution of silver, copper, zinc, lead, molybdenum and iron present. Interferences in the determination of molybdenum are described. Results by Atomic Absorption are comparable to those afforded by other acceptable methods.INTRODUCTIONDuring recent years much has been written and discussed concerning Atomic Absorption Spectrometry as a technique for determining the concentration of metallic elements. We would, however, recommend a perusal of the paper presented by J. C. Ingles at the annual meeting of Canadian Gold Metallurgists held in Ottawa, January, 1966. The paper was published in the August, 1966, issue of the Canadian Mining Journal, and contains general information on Atomic Absorption.The purpose of our paper is to present certain specific method procedures for the determination of several precious and base metals normally present in Mine, Mill and Smelter products. Of necessity, analytical results should be as accurate as the most suitable available method permits, whether for metallurgical balance, settlement purpose, ore beneficiation programs, or exploration property development. The success of such programs is relative to the accuracy, speed and economics of the analysis. It is obvious, therefore, that in presenting our paper we consider the technique has achieved the desired evaluation."

Jan 1, 1968

By Anton Sizo, Alexey Klyashtorin

The Gunnar uranium mine operated from 1953 to 1961 on the shore of Lake Athabasca, in 1964 the site was abandoned, and the mine pit flooded. In 1981, the pit was found to be stratified with a chemocline at ~70 m by a sharp change in conductivity and TDS. The uranium concentration was ~700 μg/l on the surface, 320 μg/l at 50 m, and 2,900 μg/l at 110 m. The Ra-226 concentrations were 0.1, 0.06, 0.12, and 0.4 Bq/l. By 2017, the chemocline was ~60 m deep. The highest uranium concentration was 1150 μg/l and the lowest concentration (220 μg/l) was attributed to the chemocline. In the deeper column, the uranium content varied around 750 μg/l. The uranium distribution down the pit profile was aligned with DO, ORP, and TDS. Ra-226 averaged to ~0.4 Bq/l, Ra-226 and uranium profiles were inversely related with the respective maximum and minimum attributed to the chemocline. The uranium and radium amount accumulated in the pit by 2017 was 3,420 kg and 2,163 GBq, respectively. The uranium and radium apparent accumulation trend estimates to 64 kg/y and 41 MBq/y, respectively. A stable chemocline and no concentration gradients in the other parts of the pit suggest that the contaminants are unlikely to inflow from the bottom opening or outside the pit, but rather are mobilized from the pit walls. The contaminants are not in a steady state, and concentrations of both uranium and radium are likely to increase slowly in the future if no remediation occurs.

Jan 1, 2020

By G. R. Yourt

"Generally speaking the usual approach for controlling dust is three-fold:1. Suppress the formation of dust at the source.2. Confine dust that cannot be suppressed.3. Dilute and remove dust that cannot be confined.This discussion will be concentrated on the second approach, namely, confining and disposing of dusty air.Dust BehaviourBefore delving into actual design principles a brief discussion on behaviour of dust should be helpful. For prevention of silicosis only respirable dust, ranging in size from < 0.5 to 5u needs to be controlled. As particles of this size have insignificant mass and momentum they follow air movements very closely. Consequently if contaminated air around dust sources is confined within enclosures and leakages maintained, strongly indraft dust is automatically controlled and capture velocities are of no concern.Dust SourcesAt metal mining operations ore is prepared for wet concentration by crushing it to mostly <3"" in size. This usually involves a jaw crusher, one or two cone crushers, one or more screens and several conveyors. Dust is actually produced only by the crushing action but it is liberated also when numerous fine particles contained in the ore stream are dispersed by surges and vibration at the discharge of the crushers, screens and headpulleys of conveyors. Furthermore, dust is dis¬persed when the slime adhering to the underside of the return portion of belts becomes dry and is rendered airborne by vibration and by cleaning floors especially where the slime has dropped."

Jan 1, 1971

By Y. Hou, E. Arinaitwe, L. Tong

A series of dense medium separation (DMS) tests were performed on four low-grade copper ores from Mount Polley mine in order to determine their amenability to pre-concentration. The tested ores contained disseminated sulphides, mostly chalcopyrite with low copper grades (0.15% - 0.2%) and relatively high iron grades (4.76% - 5.63%). The samples were screened into different size fractions ranging from 9.5 mm to 50.8 mm and DMS tests were performed on each size fraction at various separation densities. The pre-concentration efficiency at different size fractions and separation densities was evaluated in terms of metal upgrading and distribution. Mineralogical tests were also performed on selected DMS sink and float samples in order to understand the influence of mineralogy on the pre-concentration efficiency. The results showed that significant copper upgrading was achieved for intermediate size fractions only at the highest separation density tested. The copper recoveries generally increased with particle size at all separation densities. However, it was impossible to upgrade the copper content in the ores and achieve acceptable recoveries and mass rejections at the same time. Copper recoveries and grades showed a strong correlation with iron recoveries and grades indicating that it is possible to pre-concentrate copper by recovering magnetite and other iron-containing minerals in the ore. Overall, the results suggest that the optimum separation density and pre-concentration efficiency is a function of particle size, and that the ore mineralization properties limit the performance of DMS as a pre-concentration technology for such low-grade disseminated ores.

Jan 1, 2014

By C. J. Li, J. L. Feng

In the present study we focuses on developing the poroelasticity model that deals with adsorptioninduced permeability change mostly occurred in porous media constituted several phases such as coal. The underlying pressure induced by adsorption is obtained by Virial expansion form, which is substantially founded on statistical physics. The corresponding thermodynamic pressure applying on the pore surface is then used to detail the pore pressure and further to evaluate the macroscopic volumetric strain and the porosity. In the process of physical adsorption, the adsorbed phase is in thermodynamic equilibrium with the gas phase regardless of pore size, and the corresponding chemical potentials are all equal. The relationship between surface strain and volumetric strain is derived under infinitesimal deformation conditions to gain an explicit formulation of pore surface area with respect to volumetric strain and porosity. After the formulation of surface stress modified by the Langmuir isothermal adsorption is derived, the differential micromechanical constitutive equations of porous media are obtained. By supposing that the ratio of current permeability to initial is equal to the cubic ratio of current porosity to initial porosity, the differential equations for permeability change is consequently derived. The numerical simulations by the present model for permeability change of coal induced by adsorption of CO2 and CH4, respectively, are presented under confined stress conditions. The predictions from the model are good in agreement with the experimental data of sorption-induced permeability change of coals in the literature.

Jan 1, 2015

By H. Polkowska-Motrenko

Analytical methods for the determination of uranium and other precious elements in ores, industrial wastes and by-products have been developed. The analytical techniques employed are NAA and ICP-MS. Two procedures of digestion were applied: microwave acid digestion and alkali fusion. Accuracy was checked by using CRMs. Results of analysis of the same samples by the elaborated methods have been presented and compared. The methods will be used for estimating the capacity of uranium production for the Polish nuclear power programme.

Aug 1, 2010

By B. McFadzean, J. Taguta

In order to recover minerals by flotation, the sub-processes of collector adsorption onto the valuable minerals followed by bubble-particle attachment should occur efficiently. This paper investigates both of these sub-processes for various thiol collectors and their mixtures onto base metal sulfide minerals. The aim of the work was to investigate whether there is a correlation between the strength of the collector interaction with the mineral and the subsequent bubble-particle attachment. The collector-mineral affinity was measured experimentally using an isothermal titration calorimeter in order to determine the molar enthalpy of adsorption for each collector-mineral system. The bubble-particle attachment was measured using a microflotation device which essentially determines the hydrophobicity of the mineral. Sodium ethyl xanthate (SEX) and sodium diethyl dithiocarbamate (DTC) and mixtures thereof were used as collectors with either single minerals or binary mixtures of pyrite and chalcopyrite, respectively. The adsorption enthalpies of the collectors were interpreted with reference to the chemical structure of the collector molecule and showed that the greater the positive inductive effect of the collector, which may be qualitatively related to their pKa, the greater their affinity for the mineral surface. In the microflotation studies, recoveries were strongly related to the alkyl chain length but there was no correlation between the enthalpies of adsorption and the flotation recoveries. When mixtures of collectors were used there was clear evidence of synergistic effects and in the case of DTC/xanthate mixtures it is possible that this synergism is related to a DTC promoted oxidation of xanthate to dixanthogen. When mixtures of pyrite and chalcopyrite were used the flotation recovery and enthalpy of adsorption of SEX onto the individual pyrite particles in the mixture increased substantially compared to the case of pyrite alone. Mechanisms are proposed to interpret these important observations.

Jan 1, 2016

By M. B. McClenaghan

"The Ontario Geological Survey completed a regional till sampling project between 1984 and 1988 in the Black River Matheson (BRiM) area of northeastern Ontario. The project was designed to aid mineral exploration in the area by documenting the Quaternary stratigraphy and by providing a database of till geochemistry. This paper summarizes the results from the sonic overburden drilling component of the BRiM project carried out in areas of thick drift (> 5 m). Results from till sampling in areas of thinner drift (<5 m) have been presented in a separate paper. Approximately 1100 till samples were collected from 225 sonic drill holes. For each sample, the non-magnetic heavy mineral (>3.3 S.G.) and fine (-250 mesh) till fractions were analyzed geochemically and visible gold grains were recovered by panning.Gold and other multielement geochemical data were plotted using proportional symbols to highlight anomalous areas and to demonstrate how to evaluate and interpret a very large geochemical dataset. Thresholds for identifying anomalous concentrations of Au in till samples, defined using the 95th percentile, are 1500 ppb and 12 ppb for the heavy mineral and fine fractions, respectively. The heavy mineral fraction of older ti1ls and to a lesser extent, Matheson Till, display a number of significant gold anomalies in the study area, which in a few cases, are associated with base metal anomalies. The vertical position of gold anomalies in Matheson and older ti1ls varies from directly on bedrock to several meters above the bedrock surface and can provide information of relative ice transport distances."

Jan 1, 1994

Overview of Presentation Why Enter Agreements with First Nations Making it Win-Win ?The Principles The Process Making the Process Work Resource Agmts Defined Resource Agreements are legally binding agreements between a resource company and the First Nation(s) whose Aboriginal or Treaty rights may be impacted by the activities of the proposed project

May 1, 2009

By G. Ergin

The interactions taking place between Al-Mg alloys and the bed media during aluminum filtration are not yet well understood. These interactions were studied using the infiltration technique. The reaction interface was analyzed using different techniques such as SEM and TEM. Only spinel formation was observed at the interface regardless of the type of alumina and contact time. This paper discusses the results of the study.

Jan 1, 2004

By E. G. Joe

"The continuing increase in costs for power and fuel are affecting significantly the cost for milling of Canadian ores. In order to ascertain the scope and source of energy consumption in Canadian mills for further energy conservation opportunities, CANMET, in cooperation with the Canadian Mineral Processors carried out a detailed survey on the metallic mineral processing plants. Sixty-seven mills representing almost 300 million tons of annual capacity participated in the survey. The results indicated that in general, copper and iron ore mills, representing the bulk of the tonnage, consumed relatively low levels of energy per ton. Nickel-copper, zinc-copper and lead-zinc ore mills were moderate in energy consumption, while uranium and some copper-lead-zinc mills reported much higher energy consumptions due primarily to the increased heat requirements of the ore pulp. While the plant heating costs varied somewhat with the location in Canada, it was apparent that the lowest heating costs were achieved by those built with proper insulation and energy conservation design. If the cost per million British Thermal Units is taken as $2.00, the mill energy costs generally ranged from 40¢/ton up to $1.20/ton.IntroductionEnergy consumption in milling has always been of significant interest to Canadian mineral processors who traditionally calculate their milling costs down to pennies per ton. Even the measure of an ore's grindability is expressed empirically by Bond in kilowatts hours per ton milled. Some of the mills located in isolated areas were obliged to generate their own electrical energy from fossil or petroleum-based fuel at a cost substantially higher than that obtainable from captive or public hydro-electric installations. These mills practiced energy conservation long before the advent of O.P.E.C. energy cost inflation. One example of this was the use of electric boilers to produce process and heating steam during off-peak electrical load periods."

Jan 1, 1978

By L. V. Bell

"The gold deposits of the Caribou mine are representative of those of the Province of Nova Scotia as a whole. The ore shoots within the deposits, in common with those elsewhere within the Province, are associated usually with interruptions or irregularities in the broader and uniform fold structures. The structural factors localizing ore deposition as applied more specifically to the Caribou mine are demonstrated chiefly by the close relationship of the two principal ore shoots, known as the main orebody and the 'high-grade' vein, with a pitching drag fold developed in the southeast limb of the major northeast-southwest trending anticlinal fold or elongated dome that constitutes the primary structural feature of the Caribou gold district. It is of interest to note that the evidence indicates that this drag fold, in common with other folds subsidiary to the primary fold, formed as a result of movement similar in nature to that which caused the primary fold itself. Presumably, therefore, these folds are coincident in time.The property described here as the Caribou mine is an amalgamation of two adjacent properties formerly known as the Ross and the Nova Scotia Gold, but which since 1934 have been controlled and operated as a unit by The Consolidated Mining and Smelting Company of Canada, Limited. The Caribou mine is situated in the Caribou gold district about 45 miles to the northeast of the city of Halifax and about 6 miles to the north of the Moose River gold district.The history of mining at Caribou dates back to 1867. The first recorded gold production is for 1869. From 1869 to 1933 the gold produced from the mines of the Caribou district aggregated 47,249 oz. from 93,234 tons of ore treated. The grade of the ore on the basis of the gold actually recovered from the district would thus average 72 oz. gold a ton, which, as will be noted from the following, is precisely the grade of the ore treated from the Caribou mine during operations by The Consolidated Mining and Smelting Company. These operations commenced in September, 1934, and milling on a commercial basis was under way during 1938. Since that time production, apart from a short period in 1945 during which efforts were concentrated on development work, was continuous to early in 1947, when the mine closed down owing to exhaustion of the known orebodies. Gold produced from the Caribou mine by The Consolidated Mining and Smelting Company from September, 1934, to the close of 1946 aggregated 43,205 oz. from 90,754 tons of ore milled; the grade of ore treated averaged 0.50 oz. gold a ton. The indicated average grade of O.S oz. gold a ton for the ore of the Caribou district throughout its mining history ranks somewhat higher than the 0.37 oz. a ton average for the ore milled to date from the several gold districts throughout the Province as a whole according to the production statistics of the Provincial Department of Mines."

Jan 1, 1954