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By Richard Klimpel

"Over the last 15 years, there has been a number of plant scale sulfide mineral flotation tests organized by this author specifically aimed at understanding how flotation rate varies with reagent type and dosage. With regard to the Interaction of collector dosage, particle size, and frother type, the eight specific trends listed in the paper (Illustrated with data from two specific case studies) are qualitatively the same for almost all plants studied in some depth. The Interaction of flotation rate K and equilibrium (Infinite time) recovery R in the plant has been denoted as the R/K trade-off. The effect of collector dosage on the rate of flotation of specific particle size intervals is particularly interesting and once understood, is quite useful In understanding how industrial flotation plants respond to changing reagent type and dosage. The location and magnitude of the R/K trade-off can be different for various ores and various plant configurations, but the RIK trade-off is almost always present. The purpose of this paper is to clearly describe and illustrate the nature of this very Important flotation system Interaction so that operating engineers have a starting reference point for beginning successful plant Improvement studiesIntroductionIt is well recognized by industrial froth flotation plant operators that improving existing sulfide circuit performance is often difficult due to the self-compensating systems nature of the froth flotation system, e.g. Klimpel (1988, 1989, 1995). Extensive plant testing with existing and new reagents by this author throughout the world has clearly shown that one of the more important self-compensating mechanisms is the Interaction between collector dosage, frother type, and particle size. The term self-compensation in flotation simply means that as the plant operator consistently changes one parameter in the same direction, the nature of the froth flotation system Is such that the circuit will eventually begin to resist this controlled parameter change by automatic (uncontrolled) changes In other parameters and the overall circuit performance will go through a maximum or minimum (hence causing the operator to stop the parameter change)."

Jan 1, 1997

By Pierre Lacoste-Bouchet, Ian Orford, Mike Cooper

"The Brunswick Mine ore requires fine grinding to achieve liberation of minerals. With 25mm slugs as the grinding media in the ball mill, there was a limitation of a P80 of 25 microns on the CuPb regrind circuit. Results from pilot plant testing showed that the separation efficiency could be increased through fine grinding with a P80 of 15 microns.Options considered were the capital expenditure for new fine grinding equipment such as a vertimill or the use of the present ball mill equipment with smaller media. The Wheelabrator Allevard Enterprise (WAE) millpebs product was chosen because it was a smaller media with a distribution of steel pellets between 5 to 12 mm.With millpebs the grinding product was reduced in size from a P80 of 25 microns to a P80 of less than 19 microns. The power draw to achieve the finer grind was 13% lower while the media consumption was 57% higher.The millpebs grinding media was also tested on the zinc regrind mill. The purpose of the trial was to understand what the media consumption and power draw would be with the charge lowered to achieve an equivalent grind as obtained with the slugs. The media consumption was approximately the same and the power draw was reduced by 30%.INTRODUCTIONThe Brunswick Mine ore is a complex Pb-Zn-Cu-Ag sulphide ore with the grain size of the minerals varying from submicron blebs to particles up to several mm in size. The ore is ground progressively finer in stages. The first step of grinding occurs in the SAG mill circuit. The product sizing 80% passing 95 microns is fed to the CuPb prefloat stage (Cormier et al 2002). The tails from the prefloat is then further ground to a P80 of 35 microns for the CuPb rougher flotation. The final liberation for the CuPb circuit takes place in the CuPb regrind circuit. The regrinding mill further liberates the cleaner scavenger concentrate, second cleaner tails and third cleaner tails. The reasoning for the progressive grinding is to achieve fine grinding without having galena over ground due to its high specific gravity that preferentially sends it to the cyclone underflow."

Jan 1, 2006

By Oliver Hall

This paper sketches the mining at Noranda as it would be sketched to a visiting engineer. It does not follow quite the usual form, but is presented somewhat along the lines of a conversation. TO give the mining at Noranda its proper perspective, it is necessary to sketch the whole picture and outline: 1. History 2. Metallurgical Problems 3. Exploration 4. Mining Methods 5. Mining Details 6. Filling 7. Pillar Extraction 8. Costs HISTORY Noranda is a relatively young enterprise, with much of the personal character of its founders woven into it. The name Noranda is an abbreviation of Northern Canada and the name has been a happy one for an enterprise based on northern resources, tenacity and courage. The discovery credit goes to Ed. H. Horne. This intrepid prospector returned again and again to his sulphide discoveries on the shores of Osisko lake and, in 1922, after ten years of prospecting, succeeded in interesting S. C. Thomson and H. W. Chadbourne in his discoveries. The early exploration by L. K. Fletcher, from 1922 to 1925, for Thomson and Chadbourne, indicated promising, but somewhat perplexing, ore conditions. Mr. E. Hibbert and Mr. H. L. Roscoe rook on the task of preparing the mine for production and building a smelter in 1926, and this difficult task in this isolated area made steady progress and the smelter was blown-in in December, 1927.

Jan 1, 1937

By J. Ling, L. Racine, P. Cousin, A. R. Laplante

"The problem of gold recovery by gravity from high density gangues is first discussed from a fundamental perspective. Three potential difficulties are presented: poor liberation and fine gold particle size, concentrate bed erosion by coarse dense particles, and slow percolation of gold particle through the high density gangue bed. Two case studies are presented to illustrate the impact of these three problems on circuit performance and the feasibility of gravity recovery. It is concluded that careful testing can determine the scope of the first problem and feed preparation can limit the extent of the second. The third problem is clearly the most serious and results in lower gravity recoverable gold recoveries than would be achieved with the same recovery effort on a low density gangue application.IntroductionSince the commercialisation of the Knelson Concentrator in 1980, gold gravity recovery has experienced a surge in interest, in Canada and the United States first, then Australia, Zimbabwe and Russia, and finally worldwide. Table I details number of units sold in various countries, which corresponds approximately to their importance as gold producers 1. Exceptions. such as the low rank of South Africa or the importance of Zimbabwe, despite the much higher gold throughput of the former, are generally the result of differences in the effectiveness of the local marketing strategies, as well as a reflection of the ""philosophy of recovery"" prevalent in the country.The impact of the Knelson has not been felt equally in all plants. Typically, a highly successful application falls into the following categories: (a) alluvial or hard rock recovery for which gravity is the sole recovery method, (b) hard-rock free gold milling with ores which have a very high gravity-recoverable gold (GRG) content, (c) low grade sulphide copper-gold ores which include flotation ahead of cyanidation and (d) free-milling gold ores ahead of Other applications can be successful, but are not as common. The ""explosion"" of centrifuge concentration, of which the Knelson is the most successful but by no means the sole model, has created a certain euphoria which can be summarized as ""gravity recovery will improve the profitability of gold milling, no matter what."" This presentation is about applications that are not so successful, or even downright unsuccessful. The rationale is not to down-play gold gravity recovery, but to set it in its rightful position: metallurgically, it is not appropriate all of the times; it can become inappropriate if poorly designed; and it can be metallurgically possible but economically unjustifiable. In this paper, this will be illustrated and explained rather than demonstrated, with the use of both industrial plant surveys and laboratory test work. Alternate gravity approaches will be suggested."

Jan 1, 1997

By Karl Marquis

La Mine Mouska exploite depuis deux ans certains blocs de réserve à l?aide de la méthode long trou. Même si, à priori, la méthode long trou semble rentable, le contrôle de la dilution ainsi que le respect d?un échéancier serré s?avèrent primordiaux afin que cette méthode d?exploitation puisse être justifiable économiquement. La présentation suivante fera la comparaison économique entre cette méthode et celle de chambre-magasin.

Oct 1, 2001

By Michel Grenier, Robert Scott, Mahendra Gangal, Andrew Dasys

This research project which took place at Brunswick Mining and Smelting's Brunswick mine, was undertaken to assess the ability of new diesel engine technology to meet current and anticipated underground worker exposure limits to diesel exhaust contaminants. Concentrations of CO, CO2 NO, NO2 SO2 and respirable combustible dust (RCD) were measured in a remote mine area where a haulage truck was operating under simulated duty cycles. The 26 ton haulage truck was equipped with a 300 hp, electronically controlled engine. Fresh air was supplied to the area as per the MSHA certification document for this particular engine. Results showed that while the vehicle operated in a normal and realistic duty cycle, the concentrations of contaminants remained well below the ACGIH guidelines for the threshold limit values (TLV)for exposure. While operating at 630/0 of maximum horsepower (26 tons of ore in the dump), the time-weighted average exhaust gases concentrations (CO, CO2 NO, NO2 and SO2) never exceeded 150/0 of their respective TL Vs. ReD concentrations in this instance were measured at 0.59 mg/mJ or 39.30/0 of the exposure limit suggested by the Canadian ad hoc Diesel Committee. The Air Quality Index was 1.55 or 520/0 of the maximum suggested value for exposure.

Jan 1, 1994

By J. A. Finch, J. Vergouw

"Flotation selectivity becomes progressively more difficult to achieve with decreasing particle size. One of the possible problems is agglomeration, i.e. particles clustering together. One mechanism derives from the surface charge all particles develop in water. Agglomeration driven by surface charge is known as coagulation. One form is coagulation between particles of two different minerals of opposite charge. Another is coagulation between particles of the same mineral when conditions are such that the charge approaches zero.The paper describes the method of measuring charge and a new automated settling technique for measuring agglomeration. The phenomena are illustrated with galena and pyrite. Methods of assessing the importance of agglomeration in the plant environment are still to be developed.IntroductionFine particles can agglomerate (i.e. adhere to each other) under certain conditions. This is counter-productive for any physical separation process, including flotation. One cause of agglomeration derives from the surface charge that all particles develop in water. Agglomeration due to surface charge effects is referred to as coagulation. There are two forms of coagulation. If the surface charge is brought to near zero then the force of repulsion is lost and particles coagulate. This usually applies to particles of one mineral type and the technical tenn is homocoagulation. On the other hand, if two particles have opposite charge there is an attractive force. Since the particles in this case are usually different minerals this is known as heterocoagulation.Both types of coagulation can cause minerals to be misplaced. In the case of heterocoagulation this is fairly obvious. One example that may be familiar is slime coating (fines of one mineral coating larger particles of another) But even homocoagulation can be a problem an agglomerate of the same mineral particles can entrap particles of another mineral."

Jan 1, 1997

By Ronald L. Bruening

IBC has developed a cost effective MRT process for the selective extraction and recovery of gold, copper, and a mixture of other base metals from mine leach cyanide solutions. This is a two-stage system in which the gold is first recovered selectively as a high purity salt. The copper is then recovered separately as a sulfate salt. The other base metals are then collected as a cyano anion concentrate. Three different SuperLig® materials are used. The first stage can be used alone on gold leach solutions where copper and other base metals are not a problem. The two or three stage system is used when the base metals build up in the leach solution, resulting in increased costs. The cyanide or cyanide/cyano anion concentrate can be recycled back to the leach solution until the build-up of other salts, etc. becomes too high, or this too can be controlled with a percentage bleed stream.

Oct 1, 2003

By Anthony A. Zoobkoff

Gotcha! ? What Did I Just Agree To? ? You can?t repeat or use anything you heard. ? Did you agree before you knew what you were going to be given, told or shown? ? If you knew what you were to be given, told or shown, would you have still agreed?

May 1, 2013

By T. Mulpeter, G. Deschênes

Sadiola Gold Mine is an open pit operation located in Mali and owned by Anglogold Ashanti Ltd, Iamgold Corporation and the State of Mali. A conventional cyanidation process (cyanidation/CIP) treats 15,000 tpd (of a mixture of oxide and sulphide ores) at an average grade of 3 g/t Au. In 2002, the increasing tonnage of sulphides caused an increase in production costs and a decrease in gold extraction (from 95% to 70-76%). Because of the short retention time of the leach circuit (18 hours) and the refractory nature of the sulphides, gold leaching kinetics are critical. A laboratory investigation was initiated to evaluate actual plant practice and identify a means of improving the gold leaching rate. The study was conducted on a "soft" sulphide gold sample containing 4.96 g/t Au, 0.4 g/t Ag, 2.5% pyrite, 1.2% arsenopyrite, and 0.2% pyrrhotite with a P80 of 74microns. Pre-leaching significantly influenced the gold leaching rate. A short pre-leach (2 hours) performed at pH 8, a DO of 8 ppm, and with 300 g/t lead nitrate added at the start or the end, produced the preferred conditions. Pre-leaching for more than 2 hours was detrimental to gold extraction. The cyanide concentration had to be maintained at 450 ppm NaCN to reach a maximum gold extraction within the retention time of the circuit. A maximum extraction of 81.3% Au was obtained in 20 hours (leach residue at 0.97 g/t Au). A high level of dissolved oxygen in the pre-leach was not of any benefit. Extension of gold leaching beyond 20 hours indicated that all leachable gold had been extracted by that time. Plant modifications were made to allow leaching of oxides and sulphides in separate circuits. Treatment of oxide ores is now performed at a higher rate. Sulphides are processed at a lower throughput, with 300 g/t lead nitrate, to allow a retention time of 22 hours. Combined throughput remains the same, and total gold recovery has increased from 76% to 80% for an additional gold production of 16,000 ounces per year.

Jan 1, 2005

By Jim Wickens, Andy Cole, John Pekrul

"The roaster facility at Barrick Goldstrike Mine treats double-refractory sulfidic/carbonaceous gold ore. Prior to oxidation in the roasters, the ore is ground in two parallel dry grinding circuits utilizing 19’ Ø x 75’ double rotator mills. The design plant throughput was 12,000 dstpd. However, process improvements have significantly increased throughput rates by 25-50%. The current daily throughput record is 20,078 dstpd. Grinding circuit improvements contributing to this success include feed size reduction, ball size optimization, automatic ball charging optimization, and operating strategy changes. In addition to the changes already made, this paper discusses future optimization initiatives including on-stream particle size analysis and “real-time” ore hardness determination techniques.INTRODUCTIONBarrick Goldstrike Mines is located 29 miles Northeast of Carlin, Nevada, USA. Through 2001, the property has produced over 23 million ounces of gold making it one of the premier gold operations in the world.Prior to Barrick’s ownership, the property began producing gold from a small heap leach operation in 1975. The small-scale heap leach production continued through 1986, and in January 1987, Barrick acquired the property. Increased exploration activity by Barrick significantly increased gold reserves. In August 1988, a 4,500 dstpd mill facility was commissioned to treat high grade oxide ore. Also, heap leach operations had significantly expanded. In 1990, the mill was expanded to 6,000 dstpd, and the first autoclave was commissioned to treat refractory sulfide ore. By the end of 1996, the mill and pressure oxidation facility had expanded to 16,000 dstpd and heap leaching had ceased as oxide reserves were exhausted. In 2000, a roaster facility was commissioned to treat double-refractory sulfidic/carbonaceous ore."

Jan 1, 2003

By Lyn Jones, Mike Bratty, Rick Lawrence, David Kratochvil

"The treatment of water is required at many mining operations to remove metals and prevent the discharge of contaminated water to the environment. In other cases, water treatment is used to provide better quality water for recycle to the processing plant, or to process bleed streams in hydrometallurgical operations in order to prevent deterioration in metallurgical performance or process equipment. BioteQ Environmental Technologies Inc. has developed and commercialized the BioSulphide® Process which can provide an alternative for these treatment needs and which can also be used for profitable metal recovery, particularly in cases where conventional metal winning techniques are not economical. The BioSulphide® Process is a high rate anaerobic biotechnology for on-site production of H2S from sulphur species, although for smaller flows and metal loadings, the use of chemical sulphide reagent might be more cost effective. Both new projects and existing commercial operations, used for removal and/or recovery of copper, nickel, and zinc, and which range in sulphide generating capacity from 50 kg/day to 3.7 tonnes/day, are discussed. Particular focus is placed on the results of operations at the Raglan Mine in Northern Quebec where dissolved nickel is removed from contaminated water at a high efficiency through the precipitation of nickel sulphide. Plant effluent is discharged directly to the receiving environment and the high-grade nickel sulphide product is combined with the nickel flotation concentrate produced at the mine. The plant at Raglan will replace the existing conventional water treatment plant and will eliminate the need for on-site storage of treatment plant sludge.INTRODUCTIONThe BioSulphide® Process, which produces low cost H2S for use in selective metal precipitation, offers a low capital cost alternative which can operate efficiently and cost-effectively within a wide range of flows and solution grades. Since the technology can be used to recover metals from solutions with low flows and metal grades, the technology is also applicable to environmental applications for water treatment, with the sale of recovered metals providing an offset to treatment costs. For environmental control, the technology has a distinct advantage in being able to meet strict effluent discharge criteria due to the very low solubility of metal sulphides. To date, BioteQ Environmental Technologies has designed and constructed three plants for metal sulphide precipitation, with another three projects in the advanced development stage."

Jan 1, 2006

Cobalt mostly occurs as a minor, but valuable, constituent in most nickel and a few copper ores, however it is not generally found alone in commercially viable deposits. Its concentration and extraction are therefore intimately linked to the recovery of the primary metals and the major task for the metallurgist is to extract cobalt of commercial grade from a much larger mass of either nickel or copper and sometimes both. With a particular focus on Vale Inco?s operations, this paper provides a brief overview of the mineralogy of the main deposits being currently exploited that contain cobalt as a co or by-product. The concentration and then extraction to metal are reviewed and some advances released into the public domain are discussed.

May 1, 2009

By Tad S. Golosinski

The paper describes Online Master of Engineering Program at the University of Missouri-Rolla and discusses the experience gained during the first three years of its offering. This innovative educational program is designed primarily for mature students with an engineering or science background who want to acquire or upgrade their knowledge of mining engineering. Students can specialize in several different areas of mining, such as surface mining, underground mining, mine environment and the like. All courses constituting the program are taken online, through the Internet, without the need to attend the UMR campus in person.

May 1, 2004

"The Detour Lake Mine is located in northeastern, Ontario, approximately 300 kilometres northest of Timmins and 185 kilometres by road northeast of Cochrane. The mill processes a free milling gold ore using both gravity circuit and leaching with a state-of-the-art carbon-in-pulp circuit. The grinding circuit is a conventional SABC circuit with a tonnage of 55,000 mtpd at 95 microns with the addition of pre-crushing upstream of the SAG mill.The present paper will describe the methodology used to successfully prove that gold of the Detour Lake Mine ore is mostly concentrated into the fine size fraction of the blasted rocks. It will describe the various options possible to create value from this ore property.IntroductionThe purpose of this test work was to determine the grade of the pebbles and consider if there was value into discarding the pebbles and replacing the material by fresh feed.1. Pebble samplingThe pebbles were sampled and assayed for grade every day for 3 months. After 3 months of sampling, the results showed that the pebbles only carried 28% of the fresh feed grade. This phenomenon can possibly be explained by the fact the gold is located into area of fracture in the rocks.January-March 2014 – Pebble assay average• Line #1 = 0.222 g. Au/t• Line#2 = 0.275 g. Au/t• Head grade = 0.88 g. Au/t2. Crushing and grinding circuit surveysWith the above information, a grinding circuit survey was planned. The purpose of the survey was to build a JKSIMMET model of the grinding circuit for mill optimization but also to understand the deportment of gold into various size fractions that may explain why there is so little gold in the pebbles. All streams and size fractions was analysed for particle size analysis and for gold.Since Detour has dual grinding lines, both circuits were sample at the end of May. Two supplemental survey of the secondary crushers were also done and a survey of the gyratory crusher."

Jan 1, 2016

By D. du Plessis, R. Stobbs

The Canadian Clean Power Coalition (CCPC) was created in 2001 to protect and enhance Canada’s vast coal and other carbon-based resource wealth, and to ensure that environmental public policy decisions recognize that these resources are a Canadian asset, not an environmental liability. To achieve its mandate, CCPC has undertaken to demonstrate technologies for use in new coal-fired power plants to control air emissions including carbon dioxide (CO2), nitrogen oxides (NOx), sulphur dioxides (SOx), mercury (Hg), and particulates (PM). The first phase of the demonstration project assessed three existing technologies for coal-fired power plants that reduce all emissions, including CO2, to levels equivalent to or below that of gas-fired power plants. The technologies considered in this phase included integrated gasification combined cycle (IGCC), ‘oxyfuel’ combustion, and amine scrubbing of pulverized coal. These technologies were evaluated with three coal types: Eastern bituminous, western sub-bituminous, and lignite coals. This paper presents the studies, conclusions concerning the capital and operating costs of these technologies, the cost of electricity produced (with CO2 removed), and the cost per tonne of CO2 avoided. It also discusses the potential for using captured CO2 for enhanced oil recovery, and possible technology enhancements for improving the competitiveness of gasification of low-rank western Canadian coals. The second ongoing phase of CCPC’s work includes the design and construction of a commercial-scale plant to demonstrate the feasibility of integrated gas combined cycle (IGCC) and/or polygeneration technologies and the potential for using the CO2 generated for enhanced recovery of conventional oil.

Jan 1, 2006

By N. M. Soonawala

Detailed regional geophysical and geochemical surveys, ground surveys and drilling were undertaken to follow-up the federal-provincial Uranium Reconnaissance Program (URP) over the 27,OOO-km2 Kasmere-Munroe area in northwestern Manitoba. An analysis of the URP data showed a fair correlation between the regional airborne radiometric results and lake-bottom geochemical results, both of which also correlated with the regional geology. More detailed surveys defined anomalies with greater precision within the broad zones outlined by the URP. The ground follow-up indicated that URP anomalies can be caused by varying degrees of uranium concentration in the rocks. Mineralization of interest was found in leucogranites at several locations within the survey area. Anomalous uranium concentrations were also found in coarse pegmatites and Hudsonian granitic intrusions. It can be concluded that large-scale regional surveys can be followed up in relatively short periods of time to yield meaningful results if a properly planned exploration sequence is employed, each stage of which successively narrows the area of interest.

Jan 1, 1979

By Dany Pelletier, Yeves Breau, Ahmed Bouajila, Claude Gagnon

"In flotation circuits, direct measurement of the volumetric flow rate delivered by electromagnetic flowmeters is often difficult to obtain particularly due to piping and flow constraints. However, on-line material balance calculation could offer a reliable alternative to flow rate sensing. This paper describes an on-line material balance software package as applied to a copper flotation circuit for mass flow rate estimation mainly based on the output of an x-ray fluorescence onstream analyser. Rule-based procedures for the management of the main calculation modules are designed and integrated to a Comdale expert system. Attempts to diagnose the flotation process from mass flow rate data are also discussed.INTRODUCTIONThe need for improved process diagnosis, control and optimisation has become more evident for the mineral processing industry in the last decade. Therefore, new process variables need to be inferred or measured particularly in highly multivariable processes such as flotation. This process is subject to a wide variety of disturbances mainly generated by frequent and sometimes abrupt changes in physical and chemical characteristics of the ore and due to some irregularities in manual process operating procedures.A flotation circuit scheme is always a complex network counting more than two dozens of slurry streams. Information about those streams (%solids, metal content analysis, and flow rates) are important for process diagnosis, performance evaluation control and optimization (1,3).On-line sensing instruments such as x-ray fluorescence (XRF) on-stream analysers (OSA), gamma density gauges and electromagnetic flowmeters have been developed for the measurement of those variables. Unlike the other variables, stream flow rates could be determined by different ways:"

Jan 1, 2000

By R. Sharon, P. M. Hawley, X. Ochoa

"Development of Barrick Gold’s Pierina gold mine in north-central Peru requires construction of a waste dump for disposal of nonore grade materials stripped from the open pit to expose the ore. The waste dump site is located near the open pit in a U-shaped valley that is underlain by thin colluvial and morainal soils with locally thick peats and glaciolacustrine deposits. Bedrock underlying the site consists of variably weathered volcanics. The site is subject to moderately high seasonal rainfall and significant seismicity. The geotechnical quality of the waste rock is highly variable, ranging from low-quality, degradable, soil-like materials containing appreciable quantities of silt and clay, to high-quality, hard, durable, blocky waste rock.A hybrid dump design concept was developed that requires in-pit segregation of highand low-quality waste, and selective placement of materials in the dump to maximize stability and enhance internal and subsurface drainage. The design includes a rock drain beneath the dump along the axis of the valley and a series of connected finger drains. Extensive peripheral diversions and stormwater control measures to inhibit infiltration of precipitation are also included.This paper discusses the unique challenges presented by the site and waste materials, the design methodology and stability analyses, and planned monitoring systems."

Jan 1, 2003

By A. H. Ransom, M. E. Robb

The Salmita Gold Mine is located 240 air kilometres northea t of Yellowknife on the east shore of Matthews Lake. Salmita is one of several gold showings which are regionally associated with the steep easterly dipping contact zone between mafic, intermediate and fe lsic volcanics to the west and overlying finely laminated argillaceous sediments to the east. The volcanic assemblage forms a narrow 2 m to 7 m wide north to northwesterly trend ing Belt of Yellowknife Supergroup Archaen greenstone rocks. The contact zone exhibits exhalite character and the increased sulphide content provides a di stinct VLF-EM , IP/Resisti vi ty and/or magnetic signature. Gold occurs at or near this contact in three main settings: associated with si liceous vein-like structures; with finely disseminated sulphides within ericitic altered upper fe lsic tuffaceous unit ; or with quartz in narrow shear zones within the overlying sediment . The B-Vein is of the former type with free mi ll ing gold occurring within a highly sil iceous zone having an average widtn of 1.69 m. Alteration is intense within the immediately su rrounding host lithologies and includes the development of si lica, carbonate, sulphide, sericite and chlorite with minor graphite, epidote. tourmaline and scheelite. The mineable zone above the 4th level (-200 m) ha an average strike of 160 m and is divided into three south plunging ore shoots separated by the regional Salmita Shear Zone in the south and by a low-grade structural hiatus in the north. Production from the B-Vein commenced in August 1983 based on a diluted in situ ore reserve of 140,000 t grading 26.6 git Au . As of December 31, 1985, a total of 131 ,537 t grading 27.8 git Au have been mined at a ra.te of 160 t per day and a 96.4 percent recovery, to produce 3,522 ,000 g (113,227 ounces) of gold . Current (January I, 1986) reserves total 45,500 t grading 37 g/t Au down to the 10th level (-500 m) .

Jan 1, 1986