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By J. T. Fyles, C. G. Hewlett

"THE TOWN OF Salmo, rail centre for the Salmo lead-zinc mining district, is in southeastern British Columbia 26 miles by road east of Trail an about the same distance south of Nelson. The belt in which the mines occur is a few miles east of Salmo and extends about 6 miles north of the town and 17 miles south and southwest to the International Boundary. The belt includes the H. B. mine of The Consolidated Mining and Smelting Company of Canada, Limited, the Jersey mine of Canadian Exploration Limited, the Reeves MacDonald mine, and the Jack Pot property of New Jersey Zinc Explorations Limited.Since 1949, the Jersey and Reeves MacDonald mines have produced most of the lead and zinc in the district. From 1949 through 1954 these two mines together produced a total of 233,134 ounces of silver, 87,659,399 pounds of lead, 215,708,856 pounds of zinc and 475,397 pounds of cadmium. This amounts to more than 80 per cent of the total production of lead and 95 per cent of the total production of zinc from the district since the first recorded production in 1906. Reserves at the largest properties amount to several million tons of ore containing 4 to 7 per cent combined lead and zinc. Several prospects are potential producers.Regional StructureThe main structural units in the Salmo area are outlined in Figure I, and the major rock units of the stratigraphic column are listed in the following table. The lead and zinc deposits are replacements in a highly deformed limestone formation near the base of the Laib group, and all mineralized zones are on the south and east side of structural unit 2 (see Figure I)."

Jan 1, 1949

By Jason Berenyi

Presentation Overview ? What is Potash? ? Where are the potash deposits? ? Current potash mines ? production, capacity and planned upgrades ? Saskatchewan Saskatchewan?s Potash Production Potential s ? Current Status of Potash Exploration ? Increasing Demand for Potash ? Saskatchewan global role

May 1, 2009

By K. Duckworth

The treatment of Turam data by conventional procedures does not permit the interpreter to make any estimate of induced current magnitudes. This problem may be overcome by using an alternative data treatment procedure. It is shown that the conventional and alternative procedures are equally satisfactory in terms of their ability to provide the location of line currents and estimates of conductor quality. A full description and evaluation of both procedures is presented in conjunction with field examples to show how the alternative procedure may assist interpretation when used to complement the conventional approach.

Jan 1, 1973

By E. Nolasco, J. Ouellet, T. Tikka

"Shaft bottom structural systems are used in underground mining to support transfer of personnel, equipment and ore to and from underground to the surface. This paper first summarizes the evolution of design philosophy from a concrete tower supported on a concrete plug to the current structural steel design arrangement with unlined shaft walls, clear shaft bottom area with only a concrete housekeeping pad. Then the paper reviews the design parameters, analysis and design, and construction methodology for this rather unique shaft bottom structural system consisting of three sets of hitch beams, with each having specific functional requirements. INTRODUCTION Shaft bottom structural systems are used throughout the world in underground mining operations to provide a means to support the transfer of personnel, equipment, materials, ore and waste entering and exiting a vertical shaft ready for transport to and from the surface via the hoisting system. The shaft bottom structural system, for the Service Shaft (SS) at the Jansen Mine, is designed to accommodate two skips, a personnel/equipment cage, counterweight, services, loading pocket, rope changing, arrestor system, crash (or catch) beams, access platforms, stairs and rope divider system. Stantec and BHP have engineered a unique structural steel shaft bottom system for hoisting system at a Potash Mine. The shaft bottom structural system connects to a fixed guide and cantilever bunton system, used in the vertical shaft (shaft barrel), above the shaft bottom. The geometry for the shaft bottom structural system, shown in Figure 1 a), features three sets of hitch beams, beams which span across the shaft excavation or barrel. At the top of the structural system are two upper hitch beams. These hitch beams support a 64 metre suspended rectangular braced tower (upper tower), complete with steel guides for the conveyances. The upper tower passes through the main station (potash or mining) level and then connects to the top of the arrestor (lower) tower structure via a sliding vertical connection. The potash level is supported by two hitch beams spanning the shaft and loading pocket excavations. The arrestor tower, a braced compression tower structure, is carried by a third set of hitch beams. Below the third set of hitch beams is a short suspended tower complete with access platforms, stairs and the rope divider system."

Jan 1, 2019

By D. A. Barr

High-grade gold-silver mineralization associated with electrum and argentite was discovered in a quartz vein at the Chappelle property, 273 km north of Smithers, B.C., in 1969 following a regional geochemical reconnaissance program in the Cassiar-Omineca Mountains. Subsequent exploration resulted in the discovery of numerous quartz veins with associated precious-metal values in a belt 17 by 3 km in area centrally located within Toodoggone Group volcanic and sedimentary rocks of Lower to Middle Jurassic age and a window of Takla Group volcanic rocks of Upper Triassic age. More detailed investigations, consisting principally of 5677 m of surface drilling in 57 holes and crosscutting, drifting and raising on Chapelle Vein A, has outlined a high-grade gold-silver-bearing shoot with a strike length of 200 m, an average width of about 3 m and extending to an average depth of about 40 m below surface.

Jan 1, 1978

By Robert B. Parsons

"For the fourth time in five years, residents of northern Canada have seen a change in the tax treatment of housing and travel costs. The latest changes announced in December 1990 by the Minister of Finance are to become effective in 1991. When will the tinkering stop?IntroductionFor man y years prior to 1987, certain travel and housing allowances received by employees in remote areas were generally considered by Revenue Canada to be tax-free, even though some observers suspected that such generous treatment was contrary to the provisions of the tax legislation. In the mid-eighties, Revenue Canada decided to reconsider the tax treatment of the mining industry's long standing practice of providing housing and travel assistance to these employees. The tax- free treatment was continued by virtue of annual remission orders, but only pending the enactment of legislation which would tackle the issue head-on and layout the rules in black and white to be seen by all. The sequence of events since then may be chronicled as follows:February 26, 1986-Finance Minister Wilson's budget present s the framework for the future tax treatment of travel and housing benefits provided to residents in remote regions of Canada.December 22, 1986-The Minister of Finance announces detailed rules to fill in the earlier announced framework.January 1, 1987-The new tax rules come into effect.April 29, 1988-The Minister of Finance announces the establishment of a Task Force to examine the existing criteria for determining the eligibility of residents of northern and isolated areas for tax benefits.October 27, 1989-The Task Force presents its report.December 7, 1990-Finance Minister Wilson announces a new system of tax benefits for residents of Canada's northern and isolatedregions.January 1, 1991-The new rules become effective."

Jan 1, 1991

By F. Vasilikou

Commercial deployment of Carbon Capture Utilization and Storage (CCUS) requires field testing of a scale that can stress the geologic reservoirs. One such reservoir of interest consists of unminable coal seams that exhibit favorable characteristics and depositional environments and lower pressure and temperature than other, deeper, reservoirs. Such conditions can reduce compression costs while utilizing the action of adsorption that offers a more effective carbon dioxide (CO2) bonding than free storage or solution. To ensure the success of such tests, a number of parameters should be accurately determined, such as the seam geometry and stratigraphy, coal porosity and permeability parameters as well as optimum injection conditions. In essence the CO2 injection model should be calibrated for the reservoir characteristics using all available data. History matching calibrations are based on gas and water production from existing wells prior to injection. This paper will present reservoir models that were developed for a pilot--scale 907 tonnes (1000 tons) CO2 injection test that was performed in 2009 through one legacy coalbed methane production well in Russell County, Virginia, USA. The model incorporates a number of individual coal seams about 0.3 m (1ft) in thickness located at depths ranging from 300 to 700 m (1,000 to 2,200 feet). Model calibration was performed through history matching to prior production data, and subsequently the model was utilized to develop different injection scenarios. The developed model was used after injection to calculate CO2 plume distribution patterns that were monitored at the injection test. The paper will present model data and assumptions with a special emphasis on the effect of hydraulic fractures and the skin factor to the coalbed methane (CBM) production model.

Aug 1, 2013

By Lawrence D. Meinert

Abstract-Most large skarn deposits are zoned in both space and time relative to associated intrusions. Zonation occurs on scales from kilometers to micrometers, and reflects infiltrative fluid flow, wallrock reaction, temperature variations, and fluid mixing. The most spectacular examples of skarn zonation usually occur at the skarn-marble contact, where transitions between monomineralic bands can be knife sharp. Other small-scale examples occur in zoned veins and individual mineral crystals. Although, visually striking and scientifically interesting, in mineral exploration these small-scale variations are less useful than deposit- or district-scale zonation. In most skarn systems there is a general zonation pattern of proximal garnet, distal pyroxene, and vesuvianite (or a pyroxenoid such as wollastonite, bustamite, or rhodonite) at the marble front. As well, individual skarn minerals may display systematic color or compositional variations within the larger zonation pattern. Such patterns are reviewed for 14 well-studied examples of Cu, W, Sn, Au, and Zn-Pb skarns. In addition, many deposits have endoskarn or other alteration of the associated intrusion, and recrystallization or other subtle changes have occurred in the surrounding wallrocks. Copper skarns, such as Mines Gaspé in Quebec and Big Gossan in Irian Jaya, have high ratios of garnet:pyroxene and are zoned outward from the intrusion, to garnet, to pyroxene, to massive-sulfide replacement and vein deposits. Garnets in Cu skarn are Fe-rich and change from dark red-brown near the intrusive contact to paler brown, green, or yellow in distal locations. Pyroxenes in Cu skarns are pale and diopsidic near the intrusion, and become darker and more Fe- and Mn-rich away from the intrusion. Tungsten skarns, such as Salau and Costabonne in France and Pine Creek and Garnet Dike in California, have intermediate ratios of garnet:pyroxene, are more extensive vertically and along strike than perpendicular to the intrusive contact, and have zonation patterns commonly complicated by overprinting of metamorphic lithologies. In W skarns, garnet is commonly subcalcic and the pyroxene is Fe-rich, reflecting particularly reducing wallrocks or great depth of formation. Tin skarns, such as Dachang in China and Moina in Australia, also can have subcalcic garnet and Fe-rich pyroxene, but this reduced mineral assemblage typically is due to an association with reduced S-type granites. Tin skarns differ from most other skarn types in having a late greisen stage that may replace earlier Sn-bearing calc-silicate minerals, thus liberating Sn to form cassiterite. Many high-grade Au skarns, such as Hedley in British Columbia and Fortitude in Nevada, have low ratios of garnet:pyroxene and are associated both with reduced plutons and reduced wallrocks. Goldrich zones occur in Fe-rich, pyroxene-dominant, distal skarn. Zn-Pb skarns, such as the Yeonhwa- Ulchin district in Korea and Groundhog in New Mexico, have low ratios of garnet:pyroxene and generally form distal to associated intrusions. These skarns also are zoned from proximal garnet to distal pyroxene and pyroxenoid (bustamite-rhodonite), with significant zones of massive sulfides within and beyond skarn. Manganese enrichment of most mineral phases, particularly pyroxene, is characteristic of distal zones. Fundamental controls on skarn zonation include temperature, depth of formation, composition and oxidation state of associated plutons and wallrocks, and tectonic setting. Most W skarns form at relatively great depth, 5 km to 20 km, with extensive high-temperature metamorphic and metasomatic mineral assemblages. In contrast, most other skarn types are relatively shallow, <10 km and mostly <5 km, with limited, lower temperature metamorphic aureoles. Differences in oxidation state correlate well with different skarn zonation patterns, particularly garnet:pyroxene ratios and compositions, and can be used in both classification of and exploration for s

Jan 1, 1997

By Raymond S. Suglo

Increasing depths of mining coupled with the mechanization of most underground mining operations have led to growing complexities in the mine environments with the production of large amounts of gaseous and particulate contaminants. Over 100,000 lives have been lost due methane gas and dust explosions in coal mine workings in both the US and Canada since 1900 when the methane concentrations exceeded stipulated levels in some workings. There is therefore the need to constantly assess and evaluate the performance of the existing mine ventilation systems to maintain safe and acceptable mine environmental conditions in workings to meet the requirements of federal, state or provincial laws. This paper uses the results of continuous monitoring of methane concentrations conducted in selected development headings in coal mines in North America to assess the effectiveness of existing auxiliary ventilation systems in the control of methane gas concentrations in the headings. The results show that the average quantities of fresh air required to dilute, disperse and remove methane concentrations within set levels of one minute varied from 5.43 m3/s in Heading #11 (Condition 2) to 27.97 m3/s in Heading #10 (Condition 1) and this represented about 12 to 26% of the quantities required to dilute methane concentrations to safe levels within one minute under Condition 1 and 32 to 139% under Condition 2. The average dilution times in the headings studied were less than 8 minutes. The forcing efficiencies of the auxiliary systems in the headings studied ranged from poor to excellent implying that the auxiliary ventilation systems were capable of controlling the methane concentrations below statutory levels but may not be able to cope with large and unusual methane concentrations in the headings. This study is important to prevent methane gas and coal dust explosions in coal mines and thus avoid any loss of lives and property.

May 1, 2001

By G. A. Van Wielingen

"The conditioning of natural gas to meet contract pipeline specifications involves in the majority of cases in Western Canada, the appropriation of substantial capital expenditures for treating and processing installations. It is therefore of vital importance to the owners and producers of natural gas to arrive at an appreciation of the economics pertaining to the production and surface handling of their gas reserves at the earliest possible stage of field development. The treating and processing of natural gas often results in considerable changes in composition and related physical characteristics between the wellhead and the point of delivery of the saleable gas. This paper presents a general outline of engineering calculations which can be applied in the pre-determination of the effects that surface handling of gas will have on the composition and physical characteristics of sale-able gas. Emphasis has been placed on the calculation of hydrocarbon dew-points, heating values and shrinkages prevailing at the various stages in the surface handling of natural gas. It is believed that advance information indicating the approximate ranges of these physical properties will facilitate process selection and will provide a basis for the estimation of capital and operating costs of process installations."

Jan 1, 1960

By W. G. Miller

An overall study was conducted to evaluate the properties and processes influencing the rate and magnitude of consolidation for oil sands fine tailings produced using different extraction processes. As part of this overall study, consolidation tests using slurry consolidometers were carried out for caustic and non-caustic fine tailings. The influence of a change in bitumen extraction process (caustic versus non-caustic) on consolidation properties, namely compressibility and hydraulic conductivity, was determined, as was the effect of adding a coagulant (calcium sulphate [CaSO4]) to caustic fine tailings. For fine tailings originating from two different oil sands ores (Ore A and Ore B), results were presented in terms of variation in average void ratio with time (settlement), void ratio with effective stress (compressibility), hydraulic conductivity with void ratio and coefficient of compressibility and hydraulic conductivity with sodium adsorption ratio. Findings indicated that chemical characteristics of fine tailings (water chemistry, degree of dispersion) do not have a significant impact on their compressibility behaviour and have only a small influence at high void ratio (low effective stress). The biggest advantage of non-caustic fine tailings and treating caustic fine tailings with coagulant is an increased initial settlement rate and slightly increased hydraulic conductivity at higher void ratios. Thereafter, compressibility and hydraulic conductivity is governed by effective stress.

Aug 1, 2010

By W. G. Miller, J. D. Scott, D. C. Sego

An overall study was conducted to evaluate the properties and processes influencing the rate and magnitude of consolidation for oil sands fine tailings produced using different extraction processes. As part of this overall study, consolidation tests using slurry consolidometers were carried out for caustic and non-caustic fine tailings. The influence of a change in bitumen extraction process (caustic versus non-caustic) on consolidation properties, namely compressibility and hydraulic conductivity, was determined, as was the effect of adding a coagulant (calcium sulphate [CaSO4]) to caustic fine tailings. For fine tailings originating from two different oil sands ores (Ore A and Ore B), results were presented in terms of variation in average void ratio with time (settlement), void ratio with effective stress (compressibility), hydraulic conductivity with void ratio and coefficient of compressibility and hydraulic conductivity with sodium adsorption ratio. Findings indicated that chemical characteristics of fine tailings (water chemistry, degree of dispersion) do not have a significant impact on their compressibility behaviour and have only a small influence at high void ratio (low effective stress). The biggest advantage of non-caustic fine tailings and treating caustic fine tailings with coagulant is an increased initial settlement rate and slightly increased hydraulic conductivity at higher void ratios. Thereafter, compressibility and hydraulic conductivity is governed by effective stress.

Jan 1, 2010

By Niels Verbaan, Ernesto Bourricaudy, Ronald Molnar, James Brown, Sunil Jayasekera

"SGS Minerals commissioned an internal R&D program to develop in-house expertise to design, set up and operate REE separation pilot campaigns at its Lakefield, Canada site. A team approach was used that included expert consultants in the field of REE separation as well as in-house expertise supported by the on-site Analytical Laboratories.A mini pilot scale SX circuit was designed and constructed consisting of several modular banks of 30 stages each. About 80 stages of mixer-settlers were used to separate Pr and Nd from La and Ce contained in a synthetic REE chloride stream simulating the raffinate from an initial La-Nd|Sm-Lu+Y separation.Two pilot campaigns were operated totalling 25 days of 24/7 operation with 99.4% plant availability in the final 20 day campaign. Purities of > 99.9% in the Pr-Nd strip liquor and > 99% in the La- Ce raffinate were achieved and simultaneously sustained in the second campaign for prolonged operating periods. Key to controlling and adjusting the circuit was the rapid assay turnaround (as low as 2 hours) provided by the on-site SGS analytical lab. This permitted the operating team to modify conditions and observe the effect of changes relatively quickly considering the large number of stages. A portable XRF analyser was also used to track evolving assay trends in the circuit. The small physical size of the circuit enabled longer operation with a limited amount of sample.This paper discusses the design of the circuit, operating conditions, some of the key operating factors and controls and results obtained.INTRODUCTIONSGS Minerals commissioned the design of and built a solvent extraction (SX) pilot plant for the separation of rare earth elements (REE). The plant consists of a battery of 120 mixer-settlers set up in a dedicated piloting facility.The purpose of the programme was to develop capabilities and in-house expertise in REE SX separation technology. The program aimed to establish operating procedures, sampling regimes, and analytical methodologies and protocols for this type of process/plant. A team comprised of SGS personnel and several external consultants with prior industrial and REE separation piloting experience was assembled to achieve these objectives."

Jan 1, 2016

By Thom Sedun

No sector of the mining industry is more closely tied to local communities than aggregates. Noise, dust, traffic, visuals aesthetics and environmental impacts all fall under the purview of neighbours, local governments and nongovernment organizations. It is no surprise that aggregate operations may not be welcome neighbors in many communities. In British Columbia (BC), the regulatory environment for aggregate operations includes a local component of municipal bylaws and land use-planning in addition to the usual suite of federal and provincial statutes that apply to its sister mining sectors. Additionally, under common law aggregate extraction is not considered to be a land use subject to land use zoning, and some aggregate processing activities can be regulated by zoning. Often traversing multiple local government jurisdictions, aggregate producers find themselves negotiating not only local traffic, but also a mosaic of regulatory environments. There is an opportunity in this present situation to facilitate the application of current technology and practices in the mitigation of many community and environmental concerns associated with aggregate operations. Moving a bulk commodity is a volume-based business, and operators often don?t have time for matters that don?t directly or obviously affect the bottom line. Regulatory agencies often play multiple and complex roles with respect to their industry, not the least of which is the fiduciary role. Balancing how and when to assist an industry is often a difficult decision. For the aggregate industry in BC, the Ministry of Energy and Mines saw an opportunity to assist not only aggregate operators, but also regulatory agencies, local governments, and the general public. Through investing in the Aggregate Operators Best Management Practices Handbook (handbook) project, the Ministry of Energy and Mines (MEM) believes that it will realize a return of upheld community values, enhanced environmental protection, improved regulatory compliance, reduce disputes and a stronger provincial aggregate industry. It is the intent to publish the handbook electronically through the MEM website later this year.

May 1, 2002

By Engineering Staff

"This paper briefly describes the surface mining operations at Smoky River Coal Limited (formerly McIntyre Mines Limited, Coal Division) near Grande Cache, Alberta. Geology, mining methods, equipment, manpower and reclamation and environment topics are discussed. Introduction Smoky River Coal Limited, a wholly-owned subsidiary of McIntyre Mines Limited of Calgary, Alberta, produces a high quality, low sulphur, low volatile bituminous coking coal from its minesite located at west central Alberta, approximately 20 km from the town of Grande Cache and 360 km west of Edmonton (Fig. I). Most of the coal property is contained within a 30 km long by 20 km wide block lying in a northwest-southeast direction in mountainous terrain. Elevations vary from 920 m above sea level ·at the plant·site on the Smoky River to 1,600 mat the upper levels of current and proposed mining operations. A total of 11 coal seams have been identified. Only two seams, No. 4 Seam and No. 10 Seam are currently considered of major commercial value. These seams are recovered by both surface and underground mining. Approximately 650Jo of the plant feed coal is obtained from a conventional shovel/truck surface mine; the balance is mined underground by room-andpillar methods. Background In 1921, McIntyre Mines Limited acquired some 26,300 hectares of coal leases at Smoky River. Since then the leases were explored intermittently by different companies including McIntyre. In the mid-l 960s exploration activities increased and several feasibility studies were completed. Development of the area was undertaken by McIntyre in 1968 and, in August of the same year, a long-term contract was signed with a consortium of Japanese steel mills for delivery of 2 x 106 t of clean coking coal for 15 years commencing in 1970. Production was focussed on underground mining, using longwall methods."

Jan 1, 1985

By H. G. Stephenson, C. W. Gregory

"This article describes the methods used for pillar extraction in the Wilson Seam at Canmore Mines, Alberta. The Wilson Seam is extremely gassy and varies in thickness between 7 and 30 ft, with an average of 12 ft. The gradient varies from 5 to 30 degrees, with an average of 15 degrees. The seam was developed using Lee Norse miners to form pillars which the company intended to extract using the same machines. Before pillar extraction commenced, an abandoned and inaccessible mine 160 ft above the Wilson had to be drained of 360 million gallons of water. When this was done pillar extraction commenced. Conditions in the mine slowly deteriorated and after 6 months the decision to withdraw the continuous miners was taken. Slusher hoists were installed to replace the continuous miners, with the result that roof conditions and safety steadily improved."

Jan 1, 1972

By I. M. Gordon, J. Blackshaw

"Run-of-mine ore is reduced to 4 per cent on 4 mesh (3/ 16 inch), using four stages of dry crushing. The mill feed is then reduced to 2 per cent on 48 mesh and 60 per cent through 200 mesh, using one stage of ball milling in cyanide solution. During the subsequent treatment about one-tenth of the tonnage is removed as a table concentrate and ground to 77 per cent through 325 mesh. The size of the mill tailing is 66 per cent through 200 mesh.Four stages of ore reduction began when the dry crushing plant at No. 26 Shaft came into operation at a maximum rate of 440 tons an hour on July 31, 1941. A detailed description of the crushing plant was published in C .I .M. Trans., Vol. LVI, 1953.One-stage ball milling in cyanide solution came into full operation on January 11, 1938. The ball milling plant had a capacity of 4800 tons daily when reducing mill feed at 2. 7 per cent on 1/ 4 inch to 54. 3 per cent through 200 mesh. Primary grinding to 60 per cent through 200 mesh commenced May 22, 1945, when the feed was 3 per cent on 4 mesh. Under these conditions the capacity of the ball milling plant is estimated at 4300 tons daily. A description of the 11 Hollinger Crushing and Grinding Plant"" by D. C. McLaren was published in the Canadian Mining Journal, June 1944.The concentrate grinding circuit using a 7 x 15 low discharge ball mill and a 19-foot diameter bowl classifier with a 6-inch centrifugal pump to close the circuit began operating in the ball milling plant on October 24, 1939."

Jan 1, 1965

By Zhao-shu Sun, Mao-lin Li, Rui Cui

"Fankou Mine produced pyrite concentrate (PC) from the tailings by rougher and scavenger flotation before 2010, with the averaged S grade of 40.54% at recovery of 79.34%. The economic value of PC increases with its purity. When the S grade of PC feed to sulfuric acid plant is over 47%, the pyrite cinder can reach a high Fe grade of 62% which can be directly sold as iron concentrate. On this background, engineers in Fankou Mine set the PC with S grade of 47% as target.The direct way to reach that goal was to clean the rough PC Lots of cleaner flotation tests were conducted and the balanced strategy was to produce two ranks PC with the S grade of 47% and 38% at the recovery of 44.20% and 36.17%, respectively. The dual-product flowsheet was successfully industrialized in 2010. In order to further improve the quality of PC, gravity separation was introduced into the flowsheet. The numerical simulation and tests were performed to evaluate the beneficiation and classification performance of hydrocyclone as it treated the rough PC. It was shown that the underflow of hydrocyclone could addressed the S grade of 47%, and the finer size overflow could also be enriched to PC with S grade of 47% by flotation, due to the contained pyrite with well liberation. That is to say, with the ingenious addition of hydrocyclone to treat the rough PC, single PC with S grade of 47% could be achieved. The new flotation-gravity separation flowsheet was commissioned to produce the high quality PC at the end of 2013. The averaged S grade is 47.20% at recovery of 82.26% in 2014. The two steps flowsheet improvement maximizes the enterprise benefits and leads to simple utilization of Fe in pyrite in downstream industry."

Jan 1, 2016

By M. Melashvili

"The SART Process was introduced to the gold mining industry in the late 1990s and several commercial plants were built over the following 10 to 15 years. Most of these plants have operated well and enjoyed plant performance in line with expectations based on the underlying chemistry of the process. However, two of the plants have consistently suffered from high reagent consumption and poor product quality, which has motivated a re-examination of the process fundamentals.The primary product in the SART Process is synthetic chalcocite, Cu2S, and the stoichiometric requirement for sulphide addition is therefore 0.5 moles per mole of copper in solution. It is shown here that when sulphide is added in excess of this amount (generally with the express purpose of lowering residual copper in solution) it not only fails in its intended purpose, but can have a negative impact on the settling and filtration characteristics of the copper sulphide product, as well as the final product quality. Moreover, these problems are exacerbated the higher the selected operating pH in the pH 3 to 5 range, and the longer the residence time of slurry in this pH range.Freshly precipitated copper sulphide compounds are known to be very fine and slimy, with poor settling and filtration characteristics. It is therefore important to engineer the SART Process in such a way as to encourage growth in the average particle size of the precipitate crystals. This was the rationale for recycling a portion of the thickener underflow to the primary reactor when the SART Process was developed. However, the theory of crystallization predicts that compounds of extremely low solubility, such as chalcocite (Cu2S: KSP = 10-48), are unlikely to experience growth in crystal particle size through heterogeneous nucleation and are more likely to grow through aggregation. On this basis, the SART practice of recycling a portion of thickener underflow to the primary reactor in the commercial plants is unlikely to lead to crystal growth via nucleation, and could in fact promote breakage of the aggregates that formed in the thickener. Recycling thickener underflow could therefore have the opposite of the desired outcome. The result of testwork that is reported here lends credence to this theory and questions the wisdom of recycling in the SART process."

Jan 1, 2016

By A. Romero, M. Carvalho, D. L. Millar

"The number of underground mines replacing diesel by battery powered mobile equipment is growing worldwide. The latter produces locally zero emissions and less heat, improving air quality in the working areas, and can potentially reduce energy costs due to lower ventilation requirements and in the face of higher fuel prices. However, there is a lack of systematic approaches to evaluate the potential economic benefit of switching from diesel mobile technologies to electric ones. This paper presents a method to answer the question of whether a mine with known operating conditions should invest in battery-powered electric equipment to reduce its total costs. Furthermore, this method provides the optimal number of electric and diesel equipment in relationship with their impact on the ventilation system and the total costs of running the mine from the energy perspective. A case study of an underground mine in Northern Ontario (Canada) is presented to validate the method proposed. The scenarios studied include the specific climatic conditions of the region, and the possibility to use heat recovery for the ventilation system. Overall, results show that the use of electric mobile plants is cost competitive, especially so when electricity cost can be lowered through load-shedding strategies.INTRODUCTIONOf the multiple advantages of using electrical vehicles (EV) underground in mining, the potential increase in air quality and comfort is the most significant (Paraszczak, Svedlund, Fytas, and Laflamme, 2014). Until recently, fully battery-powered electrical underground heavy duty machinery (i.e., loaders or trucks) were not available, so that tethered and trolley assisted vehicles were the only electric alternatives to diesel powered vehicles (Bétournay et al., 2011). Today, several commercial battery-based alternatives are available (General Electric Fairchild, 2015; RDH Mining Equipment, 2015; Paraszczak et al., 2014). Operational details can be found in Ewing (2015). The main battery technology used (Lithium Iron Phosphate, LiFePO4, or LFP) was already anticipated in Bétournay et al. (2011) as a good candidate for these types of vehicles. More details on these batteries, along with state of the art and future perspectives can be found in Zubi, Dufo-López, Carvalho, and Pasaoglu, (2018)."

Jan 1, 2019