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By H. D. B. Wilson, D. T. Anderson

"Canadian sulphide deposits generally have characteristic compositions. Metal associations in ore deposits and in igneous rocks are similar.Nickel: sulphur ratios, copper-nickel ratios, and nickel :cobalt ratios in copper-nickel sulphide ores are characteristic for most orebodies. Nickel, as sulphide, is highest in ores associated with ultrabasic rocks, whereas copper is highest in ores associated with gabbros. Cobalt is relatively constant. The metal ratios in copper-nickel ores are due, first, to the composition of the related igneous rock or magma; second, to the partition ratios of the metals into the sulphides when separating from the silicate rock or magma; and third, to the conditions at the time of deposition. Each factor is discussed and the evidence indicates the Frood orebody is a magmatic segregation in place and Falconbridge is an injected sulphide melt. Many Canadian orebodies containing copper, zinc, or lead also have relatively constant compositions. •Copper and zinc are concentrated in gabbro and diorite, and copper-zinc ores in northwestern Quebec are spatially related to gabbros and diorites. The same three factors con-trolling metal ratios in. copper-nickel deposits should control the metal ratios in copper-zinc deposits if the metals come from the gabbro-diorite intrusions.The ratios of base metals in common sedimentary rocks are not similar to the ratios in orebodies. Base metal elements are mobilized during high-grade metamorphism, but much fundamental research is needed before the relation between metamorphism and ore deposits can be established on a sound basis. The distribution of elements in sulphide orebodies suggests that cop-per, zinc, and lead are deposited simultaneously .in relatively constant proportions rather than successively in a paragenetic sequence. Zoning may be due to different ratios of metals being deposited under changing temperature, pressure, and chemical environment, or may be the result of other processes."

Jan 1, 1959

By Y. Shu, A. R. Laplante

"A laboratory Knelson Concentrator (LKC), 7.5 cm in diameter, was used to study the performance of a similar unit at plant scale, 76 cm in diameter, This paper reports on sampling and assaying procedures and problems, and discusses the performance of the plant unit.The sampling procedure was based on the extraction of large samples, 5-15 kg, and their complete processing with a LKC, to characterize size-by-size performance. The procedure yielded reliable estimates of overall gold assays, after blatant assaying errors were corrected. The plant centrifugal unit concentrate was sampled in situ, which yielded very erratic grades, with a large overall negative bias. Due to the very heterogeneous nature of the concentrate, which can be linked to flow patterns in the riffles of the unit, it is recommended that for future testwork, the concentrate be recovered separately, blended and then sampled.The plant centrifugal separator, when operated at a feed rate of 25 t/h, yielded total and free gold recoveries of 41 % and 62%, respectively. This is much lower than previously reported data, and suggest that at 25 t/h, the unit is overloaded. However, recovery did not decrease with time over a loading cycle of 2 hours, which suggests that present loading cycles, generally 1 hour, can be significantly increased. INTRODUCTIONThe study of gold gravity circuits is plagued with the statistical problem of determining the behaviour of particles occurring with a very low frequency. This problem becomes particularly acute with increasing particle size. For example, considering the average weight of a gold flake in the 840-1200 µm (20/28 mesh) to be about 5 mg (Banisi et al, 1991), a reasonably high grade feed of 10 git would require, for a relative error (standard deviation) of 10% on grade, 100 particles, that is 0.5 g of gold, or 50 kg of 840-1200 µm material (a Poisson distribution is assumed). Gy's formula, for the same data, yields a weight of 33 kg, with a shape factor of 0.1 and a size distribution factor of 0.5. Both results are comparable; to obtain the suggested weight of material, a stream containing 5% weight in the 840-1200 µm would require 1 tonne samples (of unsized material). The problem rapidly disappears with decreasing particle size. Still with Banisi et al's data, a gold flake in the 300-420 µm range averages 0.5 mg in weight; the same 10% error on grade corresponds to a weight of 5 kg. Gy's formula for the same size distribution and particle shape factors as above predicts a required mass of 1.4 kg. This would correspond to an unsized sample weight of 10 kg with 14% weight in the 300-420 µm, for example."

Jan 1, 1992

By Earl Sweet

This overview of economic and financial market prospects examines the current and future state of the global and Canadian economies with a focus on commodities. Housing prices and global conflicts all have effects on the overall economy and metal prices and mining developments in particular. This presentation outlines where we are today and where the author and BMO see the economy moving over the next few years.

Jan 1, 2016

By Camilo I. Morales

At present, many open pit mines are designed to operate and feed a processing plant with ore characteristics that were defined in the design of the mining business, which are not necessarily the same as those found during the operation. This is the case of Spence Mine, whose design was conceived considering the extraction of primary iron sulphides, however, with the extraction of such minerals, it was necessary to mine and process minerals with secondary enrichment. Then it is necessary to work with a greater selectivity degree in the operation, however the capacities of the mine and operational resources were not designed for this. Thus, the changes in the general conditions of the operation should be considered as a series of geometallurgical restrictions to achieve production schedules. The above facts impacts directly in the short-term planning, whose task is to achieve the goals proposed in productive long-term planning because the company does not have the tools necessary to plan with the degree of selectivity required, so it should be done by hand, discretizating and scheduling the production polygons. Naturally, this manual process in the short term planning deviate sharply from the areas in operation, opening new fronts and benches that result in a significant increase in costs and a loss of the value of the mineral resource to miss the exploitation sequence. The research presented in this paper is related to the development of a mathematical programming model to support the short-term planning, so as to integrate the challenge of selective mining with geometallurgy restrictions, to provide the planner a guide to construction of the polygon of exploitation, guiding the fines production to achieve the goals of the proposed production plan.

Oct 1, 2009

By D. E. Pickett, J. A. Roland Faucher

"INTRODUCTIONMineral processing is undergoing growth and change at a rapid rate. The modern flotation process was discovered only in the mid-twenties, and uranium leaching in the mid-fifties. The growth of individual unit equipment size is also a recent development and is perhaps most dramatically illustrated by the increase in size and connected horsepower in grinding mills. It is also apparent in much of the equipment used in mining operations.These technical and mechanical developments have influenced the design of plants but not to the expected degree.Many metallurgists believed until recently that grinding an ore in the absence of steel results in a loEs of selectivity in the flotation. However results reported by Thornton (1), Fahlstrom (2), Canmet (3 & 4) have shown on the contrary better selectivity and improved recovery at a lower operating cost. According to the Aitik (2) plant results, the net difference in cost plus tailing value in favour of the two-stage autogeneous grinding circuit is in the order of 40 cents per ton of ore milled. Even a greater difference was indicated by the results obtained by Canmet on a sample supplied by Mattagami Lake Mines Ltd.Yet how many plants are taking advantage of these findings and how many plants are being converted. How much more improvement of the results would have been obtained if these tests had incorporated optimization of classification efficiency? By nature the mining industry and the mineral processors are conservative, and tend to emphasize the possible penalties for an unsuccessful innovation out of proportion to the potential rewards even for a significant improvement."

Jan 1, 1979

By Glenn Clarke, Ben Holloway, Barry Lumsden

"BHP-Billiton’s Cannington mine in Northwest Queensland produces a silver-rich lead concentrate and a zinc concentrate. The plant has the capacity to treat 3.1 mta and for the year to June 2006 mean feed grade was 10.3% Pb, 461 g/t Ag and 3.7% Zn. Recovery losses of lead and silver are primarily in the fine fractions. In 2001, the plant was modified when it was determined that splitting the lead feed into two size fractions, and floating the two fractions separately increased fine lead and silver recovery. The fine lead fraction (called the split float) is 80% <16µm and about a quarter of the lead and silver is in the finest fractions. In 2006 to further improve lead and silver recovery a statistical plant trial of magnetic aggregation conditioning of the split float feed was undertaken. While pure galena is diamagnetic the Cannington galena contains iron within the galena mineral resulting in the galena having a paramagnetic, magnetic susceptibility. The trial demonstrated to greater than 98% statistical significance an increase in lead recovery, and to 99% statistical significance an increase in silver recovery, and to greater than 97% significance an increase in the lead in concentrate in the split float circuit. These economic recovery benefits with magnetic aggregation of the split float feed, has led to the incorporation of this technology in the plant.INTRODUCTIONThe BHP-Billiton owned Cannington silver, lead and zinc mine is located in the Australian state of Queensland about 800 kilometres west of the coastal city of Townsville. This is a pastoral area characterised by large cattle farms, low population density and very little infrastructure. The area has two seasons, a wet, hot summer and a dry, mild winter. Because of its isolation and lack of infrastructure the mine has always operated on a fly in fly out basis.Geological drilling defined the resource as 43.8 million tonnes grading 11.6% lead, 4.4% zinc and 538 ppm silver (Walters and Bailey, 1998). At times during its history Cannington has held the crown as the single largest lead and silver mine in the world (Torrisi and Smith 2003). Another unique characteristic of the operation’s history is the rapid progress from discovery to operation. The first drill to intersect the ore body was in 1990 and the mine and plant was commissioned in 1997."

Jan 1, 2008

By Shahé Shnorhokian

Remnant pillars evolve towards the latter stages of mining in an orebody and act as repositories of mining-induced stress. It is therefore crucial to plan ahead as to their dimensions and location with respect to the local geology. A case study of a Canadian base metal mine was examined for the planned mining sequence scenario to identify the stress regimes acting on a future remnant pillar occurring between levels 4800 and 5100, near a brittle dyke. A mine-wide model was constructed in FLAC3D and input parameters included results from field measurements of the direction and magnitude of the principal stresses, and the geomechanical properties of the different formations. Prior to the study, the model was calibrated based on a 6-year record of mining-induced seismicity and rockburst events, focusing on large event clusters. The mining sequence followed at the mine was then used in the model to study the planned future extraction program around the remnant pillar. It is shown that the distance of the remnant pillar from the dyke and another orebody plays an influential role on the stress levels developed in the pillar.

Jun 1, 2012

By B. Meddings, O. J. I. Evans

"This paper is mainly concerned with the present and probable future role of hydrometallurgy in the production of base metals.The economic and technological bases of the hydro-metallurgical plants built and operative in the mid-fifties are compared to those in production in the mid-sixties. The trends observed are extrapolated and some estimates made concerning the probable size and type of hydrometallurgical plants which will be operative in the mid-seventies.The impact of emerging socio-economic factors such as pollution control and the market demand for specialized metal products which favour hydrometallurgical recovery processes are also discussed."

Jan 1, 1971

By E. Lampard

Mines must manage variability from within their operations as well as those imposed upon them by external forces. This paper provides a framework to manage both internal and external variability through the review and adjustment of the mine plan where applicable to minimize lost production and maintain operational safety. KEYWORDS Mine plan, mine schedule, variability, safety, lost production, float days, update, adjustment, operations, Maintenance

May 1, 2011

By H. A. Schnell

"The Phase I Mill at Cluff Lake in northern Saskatchewan was designed to accommodate a very specific high grade uranium ore. As this high grade ore was depleted, construction was undertaken on a new mill circuit (Phase 2) to process a more conventional lower grade ore deposit.The Phase 2 Mill expanded the ore throughput to 800 tonnes per day. The new process employs crushing, grinding, leaching, counter—current decantation, solvent extraction and magnesia yellowcake precipitation and drying to produce a yellowcake concentrate.This paper discusses the process decisions required for this expansion and outlines the new mill circuit. Included is a review of the start up in August, 1984.1.,INTRODUCTION The Cluff Lake Uranium deposit is located 720 kilometers north of Saskatoon in northern Saskatchewan. A high grade deposit was discovered in 1970 with the first production taking place in 1980 after lengthy hearings. The original mill (or Phase 1 Mill) was specifically designed to accommodate this very high grade ore for a very limited plant life.Pilot plant work was carried out in 1981 on the other lower grade ores found in the immediate Cluff Lake area. Feasibility and process selection studies were then undertaken as the Phase 1 mill was nearing its end of life. These studies were completed in March 1983 when approval was given to construct the new Phase 2 mill."

Jan 1, 1985

By C. H. Dionne, J. A. Holmes, B. J. Huls

"Falconbridge Limited is presently developing flowsheet alternatives for implementation at the Strathcona Mill. The main thrust for change arises from the need to reject more pyrrhotite from the concentrate. To maintain appropriate levels of sulphur in the smelter concentrate, however, enhanced removal of non-sulphide gangue minerals must be realized. This could best be done in the Non-magnetics flotation circuit.Laboratory investigations were followed by an on-stream pilot test at Strathcona. The application of a column cleaner on the Non-magnetics concentrate resulted in significantly lower levels of fine gangue at equivalent sulphides recovery.This paper discusses the laboratory and pilot testing and results, with a focus on piloting. Also included is a discussion of the Strathcona Mill Flowsheet and the role of the Nonmagnetics flotation circuit.INTRODUCTIONFalconbridge Ltd. is currently developing means to reduce atmospheric sulphur dioxide emissions. One alternative is a reduction in the amount of pyrrhotite in Strathcona Mill concentrate, which reports to the smelter. Pyrrhotite in the Falconbridge ores contains an average of 39 percent sulphur and 0.64 percent nickel. By comparison, pentlandite contains an average of 33 percent sulphur and 34 percent nickel.A consequence of the rejection of pyrrhotite is a decrease in the sulphur assay in the remaining concentrate. This decrease causes downstream difficulties in the roasting process. Therefore, the content of non-sulphide gangue in the mill concentrate must also be reduced in order to maintain sulphur grades adequate for roasting. The resulting concentrate will have lower sulphur dioxide emissions while maintaining nickel production."

Jan 1, 1991

By S. Schwartz, T. Bilney

"Over recent years a number of techniques have been developed as part of the AMIRA1 P9 project to represent mathematically or measure directly sub-processes occurring in flotation cells. These include the determination of:•,?a machine parameter (bubble surface area flux, Sb);•,?a froth parameter (froth recovery, Rf );•,?entrainment and drainage parameters (???and d); and•,?an ore floatability parameter (Pi).These flotation parameters allow the plant engineer to develop a model of the flotation circuit that can be used as a powerful tool to diagnose the performance of that circuit. The recently developed JKSimFloat V6 flotation simulator uses these parameters to predict the flotation circuit performance that would arise from changes in operating conditions and circuit configuration.This new flotation technology was applied to the Placer Kanowna Belle Gold Mine in Western Australia. A series of plant surveys, cell characterisation measurements and laboratory batch tests were conducted to determine the flotation circuit model parameters to be used within the JKSimFloat V6 simulatorA thorough flotation circuit analysis was then performed at Kanowna Belle using JKSimFloat V6. The recommendations that were provided by this circuit analysis have resulted in large improvements in the flotation performance at Kanowna Belle. These include a 1.3% increase in flotation gold recovery and a substantial decrease in operating costs. The analysis is still ongoing, however, the benefits already obtained from applying this technology have resulted in a US$1.3 million a year improvement to the mine."

Jan 1, 2005

By D. Komljenovic

This paper presents the use of queuing theory in truck fleet sizing in open pit mining operations. In the first part, a methodology is developed by which an optimal ratio between the size of shovels and trucks may be determined. The second part gives the fundamentals of the queuing theory and its application in analyzing shovel and trucks. The stochastic character of involved times (loading, truck cycle, etc.) is taken into consideration. Using the above theory, the optimal truck fleet size in actual conditions may be estimated. Minimal operating cost of the working system (shovels-trucks) represents the basic requirement of the analyzed process. The methodology was validated on Caterpillar front-shovel ? truck combinations.

Sep 1, 2004

By D. Kong, A. Klyashtorin, L. McGilp, I. Wilson

This work reports on applying iron oxide coated sands media in an experimental permeable reactive barrier to remove uranium species from uranium containing water. A field experiment was conducted at the legacy Gunnar uranium mine & mill site that was abandoned in the 1960s with limited to no decommissioning. The flooded Gunnar mine pit presently contains about 3.2 million m3 of water contaminated by dissolved U (1.2 mg/l), Ra-226 (0.4 Bq/l), and minor concentrations of other contaminants (As, Se, etc.). The water is seeping over the pit rim into Lake Athabasca posing environmental and health concerns. Iron oxide coated sands can be used to immobilize uranium species through an adsorption process. Methods for preparation of hydrous ferric oxide sorbents and their supported forms onto silica sands are described. Fourier transform infrared spectroscopy (FTIR), and powder X-ray diffraction (pXRD) were used to characterize the physical-chemical properties of the sorbent. The monolayer uranium uptake capacity (Qm) was estimated through the Langmuir adsorption isotherm. Adsorption kinetic experiments were performed in a lab environment before moving to a field trial. Based on these lab results and the input on field-scale appropriate geometries, a pilot permeable reactive barrier was fabricated and a field test conducted near the Gunnar pit in June 2019. This pilot test provided technical data and information needed for designing a full-scale permeable barrier based on the iron oxide coated sands, which may be applied for in-situ water treatment at Gunnar and other legacy uranium sites.

Jan 1, 2020

By Bern Klein

Motivation for Improved Energy Efficiency in Mining ? Canada potentially adopting Greenhouse Gas Emissions and Intensity Legislation in 2013 ? BC -Climate Action Plan and Energy Conservation Agreement in 2008 ? BC Energy Plan 2007 -BC Hydro is mandated to save 50% of new load growth ? TSM Initiative -Energy Management, Energy Reporting, Setting Energy Targets, GHG Management BC MOU for Energy Conservation Agreement -MEMPR, MABC, BC Hydro

May 1, 2010

By R. E. Michelutti

"The Sudbury area environment became drastically altered commencing in 1871 with lumbering, repeated fires and early open roasting operations from 1887 to 1929. These combined factors had such a pronounced effect that approximately 24 square miles were totally denuded of any vegetation and 87 square miles were semi-barren to somewhat stressed.The two nickel producing companies in the area have reduced the sulphur dioxide emissions considerably over the years. Falconbridge Limited, alone, reduced sulphur dioxide emissions by 77% from 1971 to 1984.These reductions have resulted in improvements to the air, soil, lake and vegetation quality in the Sudbury area.INTRODUCTIONThe Sudbury area environment became drastically altered commencing in the 1870's with lumbering, repeated fires and early open roasting operations from 1888 to 1929. These combined factors had such a pronounced effect on the Sudbury environment that approximately 24 square miles (64 sq. km.) were totally denuded of any vegetation and 87 square miles (225 sq. km.) were semi-barren.Over the years, the Sudbury area environment has improved considerably. This has been due to emission cutbacks by both Inco Metals and Falconbridge Limited, coupled with industry, government and community reclamation programs. Before the improvements are discussed, a better understanding of the historical degradation is required."

Jan 1, 1986

By Daniel D. M. Adams

Over the last 100 years industry in general has been trying to advance the ability to maintain systems through management processes. Mining has looked to utilize onboard data systems in our mobile mining fleets to compliment these processes. Combining the two concepts build the concept that Rio Tinto refers to as asset health. These programs can generate a cost shift between $20,000 and $100,000 per piece of equipment annually. Even with the documented savings many operations struggle to justify the systems, why?

May 1, 2011

By R. E. Jones

"INTRODUCTIONCominco Limited, one of the free world's largest producers of lead and zinc having majority ownership in Pine Point Mines Ltd., operates a mining and concentrating complex on the south shore of Great Slave Lake in the Northwest Territories. The mine was developed in 1963-64 and commenced operation late in 1964 with the direct shipping of high grade ore. The concentrator started production in November 1965 at a design throughput rate of 5,000 T.P.D.In 1968, the Company acquired the mining rights to a large ore body located to the east of the Pine Point claims and subsequently an expansion program at the concentrator raised the throughput rate to 8,000 T.P.D. The expansion included a third grinding circuit and additional flotation capacity. The flotation capacity took the form of a completely separate circuit from the existing flotation plant. In 1972, the zinc flotation circuits were modified to include two additional stages of cleaning. This change concluded the ""Pre-Consolidation Flotation"" circuit at Pine Point.In considering these modernization programs of the flotation section of the concentrator, the philosophy was to utilize the largest available machine having proven and modern technological changes. With the advent of the large capacity machines available in the late 1970's, it was considered a suitable time to revamp the entire flotation circuit at Pine Point in 1979. The Circuit Consolidation Project was developed to meet this requirement."

Jan 1, 1982

By Tom Olsen, Tom Ohrling, Peter Taggart, Chris Martin

"North American Palladium commissioned the new concentrator at the Lac des Iles mine in the summer of 2001. The 15,000 tonnes per day mill replaced the existing 2,400 tonnes per day facility, as part of a major mine expansion.At Lac des Iles, the grain size distribution of the platinum group minerals (PGM) is bi-modal. Roughly two-thirds of the PGM are relatively coarse-grained, and often associated with fast floating copper and nickel sulphides. The remainder is ultra-fine, and locked in silicates, which poses a special challenge to the metallurgist. The previous mill typically recovered just the fast-floating mineralisation, but to achieve target recoveries, the new mill has to recover a substantial amount of the ultra-fine PGM. The metallurgical challenges associated with the recovery of fine-grained PGM in the talcous Lac des Iles ores, are described.The paper describes the approach used in circuit optimisation, and how metallurgical and operational optimisation raised palladium recoveries by more than 6% in six months, by tuning the circuit to recover some of the ultra-fine PGM mineralisation.BACKGROUNDThe Lac des Iles mine, owned and operated by North American Palladium Limited, is located about 115 kilometres north of Thunder Bay, Ontario. The mine originally opened in December 1993, and prior to 2001 included a mill capable of handling 2,400 tonnes per day of ore delivered from the nearby Roby pit.Following an intensive exploration programme conducted in 1999, and a detailed feasibility study, conducted in 2000, North American Palladium initiated a mine expansion project, the objective being to expand milling throughput from 2,400 tonnes per day to 15,000 tonnes per day. This required construction of an entirely new mill, which was commissioned in May 2001."

Jan 1, 2003

By Ralph Harris, Xia Qi, Frank Komsic

"Gold cyanidation kinetics which are often limited in commercial operations by the rate of oxygen dissolution from gas bubbles into the slurry, are discussed. Emerging technologies for improving the rate of oxygen dissolution are compared and the gas efficiency, energy use and operating performance of inverted cones and a device based on a jet diffuser are examined. The effect of solids loading is also reviewed and a relationship between operating parameters and the kilograms oxygen dissolved per kilowatt-hour is presented.Inverted cones are shown to create slurry settling problems at high gas flow rates and surface tension lowering agents are shown not to greatly improve the overall rates of oxygen dissolution.The novel jet diffuser technology operates at a range of gas to liquid ratios that improves its energy use over conventional technology and the performance of a number of designs are discussed.Gas efficiencies as high as 35% were observed with the novel diffusers at 40 cm diffuser immersion and are expected to increase at greater immersions. Energy use of 1.8 kg 02 I kWh were estimated for prototype full size diffusers operating at 100 cm immersion. Energy use of 3.S kg 02 /kWh is conservatively expected at 600 cm immersion.IntroductionGold ExtractionGold that can not be recovered by gravity separation can be recovered by either the Merrill-Crow process or the newer carbon in pulp (CIP) or carbon in leach (CIL) processes 1. Reductions in operating costs and improvements in gold recovery are central mill operating strategies. The present paper examines the potential for improvement that can be achieved in the gold cyanidation step of gold recovery. This paper introduces a number of emerging technologies and provides a basis for comparison of the performance. It shows that the basis for comparison of a particular technology depends on the type of gold extraction circuit employed and the type of reactants employed."

Jan 1, 1996