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By Simon D. Handelsman

?Artisanalmining ?Application of human energy (e.g. pick & shovel) directly to produce minerals ?majority no technical skills and lack adequate working tools ?Small-scale mining ?Application of modern technology for limited scale mineral production Significance of ArtisanalMining ?15 million people in artisanaland small scale gold mining ?55 developing countries ?Use of mercury damages health of 100 million people

May 1, 2006

By H. Agnerian

"A survey of 925 mineral property transactions worldwide was carried out between 1987 and 2004. These mineral properties with reported Mineral Resources include 457 gold properties, 322 base metal properties, and five platinum group element (PGE) properties. Based on constant Canadian dollars, the results of the survey indicate that:• In general, unit metal values in the ground (e.g. $/oz Au or $/lb Cu) have kept pace with the changes in the respective metal prices during the period of the survey.• A general decrease in unit gold values in the ground has occurred from about $57/oz Au in 1987 to about $11/oz Au in 2002 for Canadian mineral properties, with the exception of a peak of $58/oz Au in 1996.• A similar trend for mineral properties has occurred in the United States, but with lower unit values, in the range from $1/oz Au to $10/oz Au during the last three years of the survey. • Unit metal values in the ground for gold properties in Canada were higher than for those in the rest of the world.• Unit metal values in the ground for copper properties in Canada were in the mid-range compared with other countries in the rest of the world; the United States and Australia were higher and Latin America, Asia, Africa, and Europe were lower than in Canada.• Unit metal values in the ground for zinc properties in Canada were also in the mid-range compared with other countries in the rest of the world; Latin America and Europe were higher and the United States, Australia, and Asia were lower than in Canada.• Unit metal values for nickel properties in Canada were highest compared with other countries in the rest of the world, followed by the United States, Australia, Africa, and Asia.• Gold values in the ground in Canada have ranged from 6% to 14% (with an average of approximately 10%) of the gold price.• The unit values of base metals in the ground have been in the order of 2% of the quoted metal price, during the period of the survey."

Jan 1, 2006

By The Staff

Introduction For many years the necessity of mechanization of all the underground services in the collieries of the Dominion Steel and Coal Corporation in the Sydney, Pictou, and Cumberland coalfields has been foremost in the minds of those concerned with the engineering and operating policies of the Corporation, and a gradual policy of the mechanization of these services has been undertaken over the years, until nearly all the underground services, except the actual loading of the coal at the face, have been mechanized. Early in 1928, experiments in mechanical loading were conducted which finally culminated in the installation of mechanical loading units in our No. 20 colliery in 1938. This attempt at mechanization of the final underground services proved unsuccessful because the workmen were not ready to accept complete mechanization, and the fruition of many years of study was delayed until the year 1947. During the ensuing years, the engineers and operators of the Corporation had studied the possibility of adapting to conditions prevailing in Nova Scotia mines the many types of loading equipment in use in the United States and Europe. The varied physical conditions which exist in the fields operated by the Corporation rendered it extremely difficult to apply the equipment and technique successfully employed in the United States, and more recently in Europe, and special techniques have had to be devised to meet the need of each individual colliery. Prevailing physical conditions in the Nova Scotia fields find no parallel for mechanization with those of any other part of the American continent.

Jan 1, 1949

By Göran Bäckblom

The Swedish mining companies LKAB and Boliden, the equipment supplier Atlas Copco, the general contractor NCC and the specialist company Reflex joined forces in 2005 to advance the field of rock technology for the purpose of mining and underground construction. The creation of the company Rock Tech Centre is part of the strategy developed by the Swedish mining industry to strengthen co-operation with industry, academia and society. This paper outlines the current R&D strategies developed for the industry in general, discusses what focus areas are selected and provides some details on the current RTC project portfolio, especially faster and better drifting.

May 1, 2006

By Emilio Giuseppe Sarno Severi

Mining and Petroleum Similar Industries ?Asset Intensive ?Exposed to market volatility ?Highly Risky Operations ?Scarcity of Experienced and Trained resources ?Operating in diverse countries with different cultures ?Increasing Environmental and Safety Regulations

Nov 1, 2011

By René Lefebvre

Numerical simulations of the physicochemical processes related to acid rock drainage (ARD) in the Sugar Shack South rock pile of the Questa Mine were performed with the TOUGH AMD simulator. The objectives were to 1) identify the main processes responsible for observed present-day conditions, 2) provide an estimate of gas phase humidity transfer, and 3) give a preliminary indication of the potential impact of closure involving the installation of an alluvial soil cover. Modeling results indicated that the mine rock geometry (thin layers placed along steep mountain slopes) favors the onset of strong thermal air convection. Simulations also indicate that resloping the piles would not significantly reduce convection and oxidation. For simulation purposes, a potential cover with the permeability range of a local alluvium was investigated and found to be of questionable effectiveness since its permeability can exceed the value of 3x10-13 m2 required to prevent thermal air convection and reduce ARD.

Nov 1, 2002

By R. H. Watson

A paper read by J. A. H. Church before the Northern Alberta Branch last February described a method of working thick inclined coal-seams from the dip upwards, with the view to preventing fire risk by using water for the stowage of waste workings, and extinguishing a conflagration should such break out, was favorably received; but there was some doubt as to the cost of such a method, as it involves a retreating system, which might prove too expensive for practical adoption. Let us consider a typical leasehold such as we are familiar with in the Rocky Mountains. The leasehold is crossed at one end by a mountain river, which is cut by the outcrop of a seam of coal, 30 feet in thickness, pitching as shown on the drawing herewith, which also shows the general line of the outcrop in elevation through the leasehold, which is two and a half miles long and contains 1,500 acres. The average lift is about 600 feet above river level and drilling has proved that the coalseam can be tapped at a depth of 1,500 feet, 2,000 feet down the pitch from the outcrop in the river, all as shown in the illustration, which is typical enough of our mountain conditions to serve the purposes of comparison of methods. We take up the leasehold and invest $30,000 to pay for drilling and prospecting; $30,000 more for the price of the lease to locator or holder, and say $100,000 to pay for a railway spur (it might cost twice this sum, but we are only comparing two methods in the same location) and "now we are ready to consider equipment and the cost of development.

Jan 1, 1924

By Arnaud Faucher, Claude Bazin, Marilène Renaud, Mohamed Mai Manga

"Grinding circuits are operated in the presence of disturbances due to ore changes and equipment malfunctions. One of the challenges faced by the operators is the detection of a disturbance, the identification of its origin and of the action needed to minimize its impact on the process. Hydrocyclones overloading, worn or partly plugged apexes of the units of a nest, surging and internal liner detachment may cause a loss of classification capacity and quality. Recovery losses in downstream leaching or flotation circuits justify the effort in developing techniques for the detection of hydrocyclone malfunction. Two approaches for the detection of hydrocyclone fault are presented in this paper. The first proposed method uses the conventional sensors available in a grinding circuit allowing real time detection of hydrocyclone abnormal operation. The second technique is based on the use of vibration sensors installed on the hydrocyclone of a nest. This last technique enables the classification of the hydrocyclone malfunction as surging, overloading and plugged apex. The two techniques can be combined to increase the robustness of the problem detection.INTRODUCTIONHydrocyclones are used for many purposes ranging from water separation, fines removal and obviously for the preparation of the feed for subsequent concentration circuits in a mineral processing plant. The strategic role of the hydrocyclones for the preparation of the feed to the concentration circuit is often underestimated because of their simplicity of operation and robustness (Napier-Munn et al., 1996; McIvor, 1988). It remains that despite their simplicity, hydrocyclones have been studied extensively from 1970 to 1990 with the objective of developing mathematical models of their operation to be incorporated into grinding circuit simulators (Napier-Munn et al., 1996; Lynch, 1977; Flintoff et al., 1987). Such a development effort is already an acknowledgement of the importance of these units in a mineral processing plant. Several authors (Del Villar et al., 1996; Plitt and Kawatra, 1979; Casali et al., 1998) proposed soft sensors to estimate the classification efficiency of hydrocyclones using sensor measurements. However, few of the proposed sensors are used in the industry. More recently several papers were published to discuss the application of 3-D modelling of the hydrocyclones (Narasimha et al., 2005; Cullivan et al., 2004) and sensors (Bond et al., 1999; Neesse et al., 2003; Hou et al., 1998; Olson and Waterman, 2004; Williams et al., 1995; Willams et al., 1997) to help in the detection of hydrocyclone malfunction."

Jan 1, 2010

By Christopher F. Gleeson, Roger D. Thomas

"Abstract - Wawa has long been a center of mineral exploration activity and diamonds have been reported in the area since the 1930s. However, diamond exploration did not begin in the area in earnest until 1991. In 1993, Sandor Surmacz and Marcelle Hauseux of Saminex began a prospecting program in the area which culminated in their discovery of the "Sandor" diamond occurrence in an outcrop on the east side of the Trans-Canada Highway. They subsequently optioned the property to Spider Resources Inc. who obtained the necessary exploration permits from Algoma Central Corporation and undertook an exploration program. The diamonds at the "Sandor" diamond occurrence are hosted in the matrix of a spessartite dike composed of actinolite, biotite, and albite. Of the 64 similar dikes that have been sampled and analyzed to date, eight have been found to contain a total of 231 diamonds. Most of the diamonds are of high quality, although they are small.The dikes are non-magnetic and do not have any other geophysical characteristic that can be used to differentiate them from the adjacent country rock. All of the dikes discovered to date have been found by prospecting or by geological mapping by the Ontario Geological Survey. The dikes do not contain pyrope garnet or chrome diopside, two of the commonly used kimberlite indicator minerals. Instead, low Mg, high Cr, Zn-rich chromite, and ilmenites of variable composition, some of which lie within the field of kimberlitic Mg-ilmenite compositions, characterize the diamondiferous dikes. A regional-scale till sampling program for heavy minerals, and geochemical analysis of the <0.177 mm fraction, was used to further define the areas of occurrence of the spessartite dikes and to indicate additional areas for diamond prospecting. Heavy minerals in till that were most useful include: actinolite, chromite, and ilmenite. Geochemical analyses of the dikes show them to contain elevated concentrations of Ni and Cr and to be enriched in Ba, Co, V, Ca, Fe, and Mg relative to the surrounding rocks. Elevated concentrations of Ni and Cr in till occur overlying and down-ice of known lamprophyre dikes."

Jan 1, 2000

By Ben Murphy, Tomi Maksimainen, Antti Rinne, Ishmael Muzinda, Toni Mattsson

"A 500m3 flotation cell was commissioned during 2014 at the Kevitsa Mine in northern Finland working as the first rougher in the copper flotation circuit. Treating around 850tph with an average grade of 0.3% Cu, the TankCell® e500 has run with an average copper recovery of 71% and an average enrichment ratio of 50 during early optimization test work. In addition the cell has proved to be very energy efficient, being operating with a specific energy of around 0.4kW/m3 under production conditions.A review of the installation and operation of the TankCell® e500 at Kevitsa will be presented. Metallurgical performance of the cell during the operating period will be discussed on the basis of on-line analysis and the metallurgical samples collected during early optimization campaigns.IntroductionThe past three decades have seen a rapid increase in the size of mechanical flotation cells. This has been largely driven by demand from the copper industry to process larger volumes of material due to combination of declining head grades and the desire to benefit from the economies of scale (Lynch, Harbort and Nelson 2010). During this period Outotec has been at the forefront of development of large flotation technology with an enviable scale-up track record; 100m3 (1995), 160m3 (1997), 200m3 (2002), 300m3 (2007), 500m3 (2012) to the largest commercially available cell at 630m3 in 2013. This scale-up methodology has been developed, tested and validated (Elgueta et al., 2009; Coleman & Dixon, 2010) by the large installation base of cells in the 300m3 size class operating (over 400). The learning from large cell operation are then reviewed and incorporated back into new designs and scale-up practices (Murphy, Miettinen, and Yanez, 2014)."

Jan 1, 2016

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 J. A. Meech, R. J. Tucker

"A dynamic computer model has been developed to simulate the operation of ore crushing plants. The model was created in a general form to allow imitation of a wide variety of secondary crushing circuits with only minor program modifications. The system can be run on either a microcomputer or a large time-sharing main frame computer as required.The size reduction and separation characteristics of crushers and screens are based on standard manufacturers' design and performance data. These calculations are isolated within various subroutines to facilitate tailoring of the model to emulate any particular plant of interest. Input and output data are presented in a variety of formats. Crusher, screen and conveyor amperages and surge bin levels can be presented in a tabular form along with appropriate alarms arising from improper circuit operation or process interruptions. Graphical representation of the circuit together with strip chart data output are features of the system when operated in a microcomputer environment.The program is currently being used at Queen's University as a teaching aid to demonstrate plant operating characteristics, control strategies and design concepts. The acceptance of, and reaction to the program at this level indicates a good potential for it to provide a valuable tool for operator training."

Jan 1, 1985

By Jim Wickens, Vince Lofftus, Mark Kravetz

"This paper discusses how Bald Mountain Gold Mine (Ely, NV) dramatically reduced mercury production by treating process solutions with a liquid reagent based on organic sulfide chemistry. Cherokee Chemical Co. Inc. (Vernon, CA) markets this reagent as UNR 81 1A.A summary of the bench tests comparing various reagents is given. The ensuing plant trial is discussed and successful mercury removal results are presented. Data is presented to show that early removal of mercury in the process reduces the amount of mercury collected at retort. Data is also presented to show that use of this reagent does not reduce gold recovery. Plant trial treatment costs are discussed. Finally, the potential of this technology to substantially reduce closure costs as compared to traditional mercury removal options is discussed.INTRODUCTIONBald Mountain Mine produces gold. Mercury is a costly by-product of gold production because no revenue is generated from mercury. Traditionally, mercury leaches out with gold and follows the gold throughout the processing circuit until it is eventually condensed and collected at Retort. Minimizing or eliminating mercury production minimizes worker health and safety risks, minimizes environmental liability, and reduces costs. In the long term, minimizing mercury production will make mine closure easier and less costly.Bald Mountain Mine dramatically reduced mercury production by treating process solutions with liquid reagents called UNR 811 and UNR 811A. Several reagents were bench tested before proceeding with a plant trial. The reagent complexes the mercury to form an insoluble organic sulfide precipitate. In independent tests, the precipitate passed acid mobilization tests and is considered stable."

Jan 1, 2001

By J. Lovering, M. Allan

"INTRODUCTIONThe Afton Mines property is located 180 air miles east-northeast of Vancouver and nine miles west of Kamloops, a city of 60,000 people in south central British Columbia. The deposit is located adjacent to the Trans-Canada Highway at an elevation of 2,231 feet above sea level.Interest in the property dates back to 1898, when a 330 foot shaft was sunk on the Pothook claim immediately south of the existing operations. Since that time many companies and individuals have investigated the possibilities of the area but it was not until 1972 that Afton Mines Ltd. located a significant zone of copper mineralization. The feasibility study was completed in the period 1974-1975, by Teck Corporation Ltd., who control Afton through 50% direct ownership and 65% ownership of Isa Mines, who own a further 23%. Construction began in March, 1976. Ore production started in August, 1977, and the first ore was fed to the mill on December 8. This was the third mine to be brought into operation by Teck since 1975, the others being Newfoundland Zinc and Niobec.The orebody is a porphyry copper deposit which has undergone supergene alteration. It is 1,700 feet long in a west-northwest by east-southeast direction, 300 feet wide and dips at 55° to the south. The mineralization in the supergene zone, which accounts for two-thirds of the known deposit, is rather unusual in that the copper is present primarily as native copper and chalcocite. The grade is one percent with the distribution: 70% native, 25% chalcocite, and 5% chalcopyrite with bornite, covellite, and cuprite."

Jan 1, 1979

By T. R. Raponi

"The Williams Mine, Canada's largest gold mine, processes over 6,000 t/d of ore utilizing SAG/ball milling, leaching and CIP, carbon stripping, electrowinning and refining to produce gold dare bullion.Grinding accounts for over 50% of milling costs with grinding media, power and mill liners making up over 75% of grinding costs. Since 1991, several changes in these areas have been made which have significantly reduced costs, while fineness of grind and mill recovery have both been increased.This paper discusses the changes made in the grinding circuit over the past few years and assesses their impact on operating costs and mill operations.IntroductionThe Williams Mine is located in the Hemlo area of northwestern Ontario, about 40 km east of Marathon. It is operated by the Williams Operation Corporation, which is equally owned by Teck Corp. and Homestake Canada Inc. Canada's largest gold mine, the Williams Mine commenced operation in October 1985 at a rate of 3, 000 t/d. Mill throughputs have typically exceeded design, averaging 3, 397 t/d in 1986, and 3,380 t/d in 1987. In August 1988, production was increased to 6,000 t/d. Throughput averaged 4,212 t/d in 1988. Production increased over the next few years averaging 5,477 t/d in 1989, 6,248 t/d in 1990, 5,987 t/d in 1991, and 6,283 t/d in 1992. With a recovery of over 95.5%, production for the first three quarters in 1993 is averaging 6,408 t/d.GeologyThe main ore body is a steeply dipping conformable zone of microcline-quartz-muscovite rock. It occupies the contact zone between hanging wall metasedimentary rocks and foot wall felsic quartz feldspar porpyritic rocks of possible volcanic or subvolcanic intrusive origin. The ore body is enveloped by muscovite schists which are thickest on the footwall side. The schist is enriched with tourmaline, pyrite, and green mica containing vanadium and barium. The ore body contains pyrite and barite which occurs mostly in laminae and bands to veinlets and breccia fillings. The gold occurs as free (native) grains, 20 to 30 µm in size at grain junctions of the microcline and quartz. Visible gold is rare in the main zone. There are commonly two or more discernable layers of economic value which are separated by weakly gold mineralized material which usually consists of muscovite schist, barite rich zones or sulphide poor-microcline-quartz rock and in the upper and western fringes by biotite rich schist with feldspathic fragmental characteristics. Molybdenite is a common constituent as very fine disseminated grains in the microcline-quartz rock and is most prominent with gold enrichment. Stibnite and realgar are two of the more common accessory sulphides."

Jan 1, 1994

By R. Amelunxen

"The molybdenite associated with copper porphyry ores is generally treated in a flotation circuit consisting of a separation and a cleaning section. In the last few years column flotation cells have replaced the numerous counter-current stages that were previously used in the cleaning section. Despite this, little has been published regarding the number of column cleaning stages required. Therefore a Moly flotation database is assembled with the objective of verifying the existence of a common upgrading profile. This can then be used as a guide in the prediction of the number of stages required to achieve a given concentrate grade.INTRODUCTIONAlthough the first use of Column Flotation was on copper flotation products, the popularity and understanding of their operation was not really achieved until they were widely implemented in Moly circuits in early 1983. The use of columns in Moly plants has since left a legacy of information, both published and unpublished. The data collected since that time has been used primarily used to establish column operating parameters and to determine upgrading potential within the circuit.Byproduct Moly generally occurs with porphyritic copper ores in concentrations less than 0.1% MoS2. As such, piloting a new moly ore to produce, say 5 Kg/Hr MoS2, in a slurry approximately 1 Lt/min (at 10% solids), requires a moly plant feed of about 570 Kg/Hr, which translates to a mill feed rate of 200 to 800 Tonnes per day (depending whether you deal with Chilean or Canadian ores). Piloting Moly for any lower tonnage would reduce the accuracy of the testwork.Thus a need exists to collect operating data in order to establish the existence of process relationships that could be used for feasibility and design purposes.The object of this paper is to expose a preliminary relationship that could be used for the purposes described above and to apply the findings to a preliminary design of a moly cleaning circuit flowsheet."

Jan 1, 1990

By Jan E. Nesset, James A. Finch, Frank Rosenblum, Rebecca Payant

"Under certain conditions of moisture and oxygen, sulphides can spontaneously heat, a phenomenon known as self-heating. In this paper, we test a hypothesis that galvanic interaction between some sulphides can promote self-heating. Galvanic interaction is controlled by restpotential difference between minerals and their surface area of contact (particle size). In order to decrease rest-potential, four sulphides were tested: pyrite, chalcopyrite, sphalerite and galena. Two series of samples consisting of mixtures of two sulphides, each representing a mass fraction of 50 %, were run in standard self-heating tests. The first series comprised five mixtures, two with low rest-potential difference (pyrite-chalcopyrite and chalcopyrite-sphalerite) and three with high rest-potential difference (pyrite-galena, chalcopyrite-galena and pyrite-sphalerite). The second series was performed on a pyrite-sphalerite mixture at four particle sizes (80 % passing 850 µm, 300 µm, 75 µm and 38 µm). The first series showed that the individual sulphides and the mixtures of low rest-potential difference did not self-heat but the mixtures of high rest-potential difference did self-heat. The second series showed that self-heating increased inversely with particle size (increased specific surface area) and revealed that it was the fineness of the pyrite (high rest-potential sulphide) that governed the self-heating effect, indicating the rate-limiting step is the reduction reaction on pyrite (cathodic mineral). The increase in self-heating with high rest-potential difference and increasing particle fineness supports the hypothesis that galvanic interaction contributes to sulphide self-heating. A possible mechanism based on the hydrogen sulphide hypothesis is proposed where the reduction reaction at the cathodic mineral surface is ferric to ferrous. The understanding gained will be of interest to those involved in storage, shipping and disposal of sulphide minerals.INTRODUCTIONSulphide mineral ores are the major source of most base metals such as copper, zinc, nickel and lead. Once the ore is extracted, the sulphide minerals are liberated by size reduction (crushing and grinding), concentrated by flotation, and the concentrates sent for further processing to extract the metal (smelting and refining). At the various stages from mining to concentrate shipment and tailings storage, the sulphides are exposed to a variety of environments, which under certain conditions of air (oxygen), humidity and temperature can cause them to heat. Since no external heat source is involved, this pyrophoric process is known as spontaneous heating or, more commonly, self-heating. The self-heating of sulphides is associated with oxidation reactions under moist conditions. If the heat generated is greater than the heat dissipated, the sulphides will self-heat and the material temperature will rise. This can make working conditions potentially dangerous as oxygen is consumed from the surrounding air and significant quantities of sulphur dioxide (SO2) can be released (Ninteman, 1978)."

Jan 1, 2011

By Seppo Haapamaki

"Since it started production in 1970, Anderson Lake Mine has experienced several radical changes in mining methods as a result of the physical characteristics of the ore zone along with the need for improved productivity. This paper examines the transition from longhole open stoping to mechanized cut-and fill then to vertical crater retreat at Anderson Lake Mine. The reasons for the need to change mining methods is analyzed taking into consideration the impact of such criteria as improved productivity and ground conditions and how these criteria affected the over-all economic viability of the mine. IntroductionAnderson Lake Mine is one of five producing copper mines in the Snow Lake area and is situated approximately 3 km southeast of the townsite of Snow Lake (Fig. I).The mine production rate at Anderson Lake is 700 tpd and all ore is trucked from the mine to the mill located approximately 2.5 km from the mine site . The mine is serviced by a three-compartment shaft sunk to a depth of 3300 ft with level spacings at 175 ft and 150 ft intervals from the 600 ft level to the 3000 ft level. .The shaft was collared in 1965 and, a year later, the hoist and compressors were installed. By 1969, a crusher system had been built and actual production began in 1970. During 1978, production was suspended for the year to allow for 430 ft of shaft deepening."

Jan 1, 1988

By A. N. Veltmeyer

"INTRODUCTIONTeck-Corona Operating Corporation's David Bell Mine is located 35 Km east of Marathon, Ontario. In conjunction with its 1000 tonne per day gold concentrator, the company operates an effluent treatment plant for the treatment of mill, mine and site effluents prior to environmental discharge. This paper provides a brief overview of treatment strategies and methods, employed to detoxify final effluent.WATER MANAGEMENTThe tailings impoundment basin, located 3 Km south of the plant site, provides storage capacity for all process effluents and site run-off. Basin supernatant (reclaim water) is recirculated to the plant for reuse as process water, in order to minimize fresh water consumption. As site run-off and limited fresh water intake exceed natural losses due to evaporation, reclaim water in excess of process requirements is passed through the effluent treatment (ET) facility to control tailings basin level by providing regulated discharge of environmentally acceptable effluent. (Figure 1)PASSIVE TREATMENTTo reduce treatment costs and burden on the ET facility, reclaim water is fed to the plant to take maximum advantage of natural destruction processes occurring in the tailings basin.These include: volatilizationphotodecompositionbiological oxidation and precipitationstorm water dilutionConcentrator recovery techniques (Figure 2) are directed toward maximizing gold recovery while limiting the addition and extraction of contaminants requiring disposal and eventual treatment.Currently ore ground to 90% -200 mesh is pre-aerated at pH 9.0 with lime for 18 hours prior to 30 hours cyanidation at pH 9.5 with low cyanide ion excess (70 ppm).Pre-aeration passivates cyanicides prior to cyanide contact. This reduces cyanide consumption and the formation of cyanide complexes. In conjunction with recovery requirements low pH is maintained to retard the dissolution of stibnite into its antimony and sulfide components."

Jan 1, 1989

By J. S. Roberts

"Campbell mines has been in operation since 1948. Part of the process consists of roasting of a flotation concentrate which contains arsenopyrite.At present operation level the milling rate averages about 850 t.p.d. and about 48 t.p.d. of concentrate is roasted. The gas emissions at present contain between 3 and 3.5 t.p.d. of arsenic trioxide, about2 - 2.5 t.p.d. of particulate matter and between 19 - 25 t.p.d. of sulphur dioxide. The particulate is extremely fine calcine dust and it assays about 4 oz. Of gold per ton.Until 1974 the gas was discharged untreated through a 125' stack.UntreatedThe decision was made in 1972 to erect a gas treatment plant, construction ran through 1973 and operation began in 1974. The cost of the installation was about 1.3 million dollars.On investigation the process technology, it was apparent that the plant operated by giant Yellowknife for the same purpose was the model we should study. I would like to acknowledge the fine cooperation we had from giant, the plant that emerged is very similar to theirs and incorporates some improvements suggested by them and based on their operating experience.The plant is designed to remove arsenic and solid dust and it has no provisions to remove sulphur dioxide.We depend upon the 125' stack to assure adequate dilution. Because of the nearby airport the stack cannot exceed 125'. A stack burner is provided that can ""y used to heat the stack gas and carry it higher into the atmosphere when there is an inversion condition.The removal of arsenic from hot gas is fairly simple. The sublimation point is about 425°f, a few degrees above this point the arsenic is a gas and a few degrees below it condenses into a solid form. The arsenic could have been removed by simply dropping the temperature below the condensation point and collecting all the solids in a baghouse. This would have removed the possibility of recovering the gold in the calcine dust and would have produced an impure and more bulky arsenic material to be placed in safe storage. The decision was made to collect the solid dust in a hot electrostatic precipitator operating above the sublimation point and subsequently lower the temperature below the condensation point and collect the arsenic in a baghouse."

Jan 1, 1976