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By O. S. Perry, D. F. Kidd

"BEAVERDELL IS IN the Interior Plateau of British Columbia, 25 miles east of Penticton, B.C. In 1897, small but rich silver-bearing veins were found on Wallace Mountain, 2 miles east of Beaverdell. Production from the camp, at first of a modest character, started in 1901 and has never ceased. Over 90 per cent of the values in the ores have been in silver. Remaining values present are in lead, gold, zinc, and cadmium.There have been four principal mines in the camp. A fifth is approaching production. The four mines lie within a northeasterly trending rectangle about 5,000 feet long and 2,000 feet wide, the fifth mine being at the northeast end. From southwest to northeast the mines, with their silver production to Sept. 30, l 955, are: -Wellington Mine - 1,463,948 oz.Sally Mine (includes Rob Roy claim) - 1,930,124 oz.Bell Mine - 3,460,815 oz.Lass Mine - 12,557,956 oz.Other production - 331,252 oz.Total, Beaverdell Camp - 19,744,095 oz.The camp is near the centre of an area geologically mapped in 19 l l by L. Reinecke for the Geological Survey of Canada (1). The rocks of the camp comprise:1. Beaverdell stock of porphyritic quartz monzonite;2. Westkettle batholith of quartz diorite;3. Wallace formation of tuffs and lavas.The Wallace formation borders the east edge of the camp, and has been intruded by the Westkettle batholith. The batholith is the host rock for the orebodies. The Wallace-batholith contact dips gently east. Some of the Beaverdell ores at depth, while in the batholith, are under Wallace formation mapped on the surface above.West of the camp, Reinecke mapped a stock one by one and one-half miles in size, of porphyritic quartz monzonite, which he considered much younger than the Westkettle quartz diorite (I). Recent petrographic work by White and Dolar-Mantuani (3) suggests that in the vicinity of the camp the Westkettle batholith is more quartz monzonite than quartz diorite, and little different in composition from the Beaverdell quartz monzonite. They suggest a consanguinity with the Westkettle batholith."

Jan 1, 1949

By A. Miura, R. Winarko, D. B. Dreisinger

Chalcopyrite leaching in acidic ferric sulfate media at ambient temperature exhibits slow kinetics and poor leaching efficiency, which is typically attributed to surface passivation. Recently, the presence of small amounts of iodide in the leaching media has been reported to significantly accelerate chalcopyrite dissolution. To further explore the behavior of chalcopyrite leaching in the presence of iodide, it is essential to determine the factors that control the leaching kinetics. In this study, we investigated the effect of initial solution potential, iodide concentration, particle size, and temperature on the kinetics of copper extraction from chalcopyrite. The key in the experimental setup was to minimize iodide loss by iodine evaporation and precipitation as cuprous iodide. The former was achieved by performing the leaching tests in fully sealed leaching systems. The latter was achieved by maintaining the solution potential higher than 420 mV (versus Ag/AgCl electrode), where formation of CuI precipitate is possible. The experimental results show that higher solution potential and higher iodide concentration had a pronounced positive effect on the rate of chalcopyrite dissolution. There was no effect associated with particle size, which might be due to the narrow size range tested. The positive effect of increasing temperature only manifested itself when the solution potential was sufficiently high. INTRODUCTION The main challenge of chalcopyrite leaching using ferric ions as the oxidant at ambient temperature is the slow kinetics and poor extraction. The reason for this leaching behavior has been attributed to passivation resulting from formation of insoluble passivating layers on chalcopyrite surfaces, which could be metal-deficient sulfides, elemental sulfur, polysulfides, and iron precipitates (Klauber, 2008). To overcome passivation, technologies using higher temperature and pressure have been developed to leach chalcopyrite concentrates, such as the CESL process (Mayhew, McCoy, Mean, & Woeller, 2013), FLSmidth Rapid Oxidative Leach (ROL) process (Mulligan et al., 2017), Galvanox (Dixon, Mayne, & Baxter, 2008), Albion process (Hourn, Turner, & Hourn, 2012), HydroCopper (Hyvärinen & Hämäläinen, 2005), TPOX (Dreisinger, 2006), and AAC/UBC process (Dreisinger, Dempsey, Steyl, Sole, & Gnoinski, 2003). Even though they are effective in accelerating copper leaching kinetics, those technologies typically have higher capital and operating costs associated with production of concentrates and application of autoclaves. This makes them unsuitable for treating low grade primary copper sulfide ores.

Jan 1, 2019

By M. S. Ünal

Creation of 3D model of coal seams is a tedious and time consuming operation. Computer aided 3D solid modelling of coal seams can be performed in a fast and detailed manner. However, in case of having tectonically disturbed coal seams where faults may be observed, expert opinion is necessary to decide the geometry of the coal seam. The aim of this study is detection of fault zones with computers by using drill hole data. Firstly, a digital database is generated by using the drill hole data. Surface models of the layers are produced by applying triangulation method on coal seam roof elevations. The generated surface is divided into small squares and consequently dip and dip direction values are calculated for each square. Variograms for dip and dip direction are determined and maps showing the variation of dip and dip direction are generated by using geostatistical methods such as kriging and co-kriging. Examination of disconformities observed on estimation maps created by kriging and co-kriging clearly reveals the extension of faults. The method used in this study is applied to detect fault lines of a known mine field. 3D solid modelling of the coal seam has also been performed by using conventional cross-section method. The results of both methods were in good agreement. Identification of major faults by using the method introduced in this study could certainly simplify 3D solid modelling of coal seams.

Aug 1, 2013

By E. W. Johnson

"The current increase in gold price may presage a resurgence of gold milling plants in Canada similar to that of the thirties. Because the industry has been at a low ebb for almost two decades, future operating responsibilities will likely devolve upon personnel with limited cyanidation experience. My comments, which may seem elementary to experienced operators, are directed towards these industry newcomers to highlight operational features in cyanidation which forty years' experience has shown me to 0e of importance. A brief review of what actually takes place in a cyanide circuit may help give a better understanding of the topics to be discussed.These will include: (1) expose the gold (2) dissolve the gold(3) make solid/liquid separation(4) recover the gold from solution.1. Expose the GoldThe chemical key to gold recovery by cyanidation lies in the ready solubility of the metal in a relatively weak solution (0.04%-0.075% NaCN equivalent by weight) of calcium or sodium cyanide. Obviously, dissolution cannot start unless the cya1ide is in contact with the gold. As an analogy, the yolk of a hard-boiled egg can oe taken to represent the gold particle and the egg white and shell, the surrounding gangue or sulphide host. As the sl1ell and white must be fractured to expose the yolk, similarly the surrounding gorngue or sulphide host must be broken or fractured to enable the cyanide solution to contact the gold."

Jan 1, 1979

By A. Rahimi

The smelting plant at Sarcheshmeh Copper complex commenced working with two reverbatory furnaces, four pierce smith converters, two anode furnace and one holding furnace having a known capacity of 420 metric tons of anodic copper according to the initial design in 1982 (1361 Iranian calendar). During the period from the commencement date up until now, the smelting plant has undergone changes for improvement and enhanced production by substituting holding furnace with another anode furnace, installing the fifth converter and utilizing oxyfuel burners on the ceiling of the reverberatory furnaces. According to the different operational processes in reverberatory, converter and anode furnaces different types of corrosion mechanisms and degradation can be observed in the refractories of copper smelting furnaces. In the case of reverberatory furnaces, the corrosion and degradation of refractories may be factors such as hot corrosion by melt, corrosion attack by the constituents of burning fuel, degradation of sidewalk exposed to mechanical stresses of the feed charge and the hydrostatic pressure exerted by the melt inside the furnace, all of which are due to the unique operational condition governing the reverberatory furnaces environment, i.e. burning fuel atmosphere, charge concentration and combination, furnace temperature, calm melt condition, etc. In the case of converter furnaces, the most likely corrosion and degradation of refractories may be due to the chemical reaction between the melt and degradation of tuyere zones due to sever movement of the melt and mechanical strikes of the punch rod. Also the most essential corrosion reaction occurs around the mouth of the converter because of the mechanical strikes by the cleaning device and thermal heat shock. Other parts inside the furnace are corroded due to the reactions between the fayalite slag and the furnace refractory. In anode furnaces the degradation mechanism is mostly related to the gaseous corrosion (penetration of natural gas or liquid gas inside the refractory structure due to the long period of reduction processes and refining stages. Finally another important factor responsible for destruction of refractories in these furnaces is the copper oxide slag together with the fume constituents of burning fuels.

Jan 1, 2004

By A. A. Nourbakhsha

During the last few years, low cement castable refractories found more application in industry, because of good low and high temperature mechanical properties.In the case of alumina based castables ,usually fume silica or reactive alumina are used as superfines in the matrix to form mullite at high temperature which depends to source of alumina . Mullite network formation improves the hot mechanical properties and thermal shock resistance of bodies. The mullitisation of andalusite (so called primary mullite) also produces a certain amount of rich silica glass, part of it being expelled out of the grain. This phenomenon restrict the application of andalusite low cement castable in pig iron ladle. In this research, firstly 8 pig iron ladles from ESC(esfahan Steel company) with different conditions selected and monitored during usage to determine of main factors affecting on corrosion of refractories. Results showed that tap to tap time and undischarged pig iron are the main ones.Therefore monolithic refractories such as castables which increase thermal shock resistance could be substituted. Regarding to ternary phase diagram of Al2O3-SiO2-CaO ,fumed silica was selected as fine particle.After optimization of particle size distribution by using of modified andreasen equation and rehological properties ,different mixture were composed with andalusite ,fume silica, reactive alumina,secar 71 ,SiC and kyanite.All aggregates system components were mixed dry for 1 min in hobart mixer .After water addition wet mixing was continued until getting flowing about 4 min.The ready mixes were poured in the moulds of of 25mm X 25mm X 140mm with vibrations .After demoulding the samples were cured for 24 hrs and dried at 110OC/18hrsand then fired at 1100-1500OC. Physical ,mechanical ,phase analysis(XRD) and microstructure studies (SEM)were carried out according to ASTM standards. Results showed that by addition of kyanite ,more mullite phase formed through the matrix but,because of expansion occured during this phase transformoration, some cracks formed and mechanical properties decreased.By SiC addition the silica film on the surface of SiC lead to improvement of mechanical properties at high temperature

Oct 1, 2011

By G. E. Wood

THE INSTABILITY of copper prices and their low levels since late 1974 are of great concern to both producers and, to a lesser degree, consumers of copper. The price levels portrayed by the "neutral long-term" price trend are assumed to have been enough to encourage sufficient investment to provide adequate world production capacity and yet not so high as to encourage development of excessive over capacity. The identification of the present level of this trend has taken on special importance for governments and for corporations active in the world copper industry. One approach to the possible determination of such a price is through the use of moving averages. In this way, the effects of relatively short-term price movements can be dampened, as can those related to business cycles. From such averaged price data, long-term trends can be sought over a selected time period. The choice of the past time period for which a price trend is sought may have a major effect on the result'. For example, in the period 1885-1973, the long-term price trend for copper in constant dollars showed a decline, whereas in the period 1919-1973 it was rising. For the purposes of discussion, an appropriate choice of period might be a relatively short and recent one so that the majority of the present cost factors would be present throughout. It should also be long enough to cover a number of full business cycles.

Jan 1, 1977

By J. Drenda

The model for methane removal from the workings directly behind an advancing longwall as presented in this paper may be used for U-ventilated cross or length-wise systems of longwall advance, starting from the boundaries of the extraction field. In mines with high methane concentrations, Uventilated longwall extraction from the boundary of the extraction field produces a methane risk. The risk is manifested by a higher methane concentration at the end of the longwall, upper gate and incline. The methane in increasing concentrations, which is emitted mainly from the workings with caved areas, appears towards the end of the longwall and in the upper gate. The graphical model of the longwall provided includes both the workspace of the longwall and the caved area. Based upon a model of the rock distribution in the cave, a ventilation model comprising two standard air splits and large number of crosswise air splits has been constructed. Simulations indicate that if the shearer is not present in the longwall, the air distribution in the caved area is symmetrical. The presence of the shearer, both at the bottom and the top of the longwall, causes an asymmetrical air distribution in the cave, that is the asymmetrical flows in the goaf and longwall workspace and in the inflows and outflow between those regions. This phenomenon causes an increase in methane concentration at the end of the wall and at the upper gate. The methane removal method proposed for workings with a caved area directly behind the advancing U-ventilated longwall, starting from the boundary of the field, consists of the removal of methane-air mixtures from the caved area with the boom pipes. These booms may be of several meters long and may be extended if necessary. The insertion of the boom pipes into the cave towards the end of the longwall and connecting them with the general methane removal system creates negative pressure causing a change in the depression field, which is the aerodynamic potential field in the goaf area around the ends of the boom pipes. This will promote the extraction of the methane-air mixture from voids and the goaf cracks, and the area from which methane is being extracted may extend to the side of the cave. This is a positive phenomenon, which could result in the methane flow from the goaf into the airflow in the longwall workspace decreasing or even stopping. A reduction in the methane flow into the longwall workspace with extractions booms will be advantageous for the production continuity. Based on the floor load model for the rocks in the cave and the methane-removal boom load model, the value of the force required to move the boom behind the advancing wall has been calculated. Strength calculations for the pipe were performed using a finite elements method. The analysis results, as well as the comparison of the pipe material effort to the strength parameters of steel, which can serve as piping material (for the assumed load conditions), lead to a belief that the proposed boom construction satisfies the strength requirements.

Aug 1, 2013

By Bryan Tisch

Goal of GMGE ??To advance mine reclamation through the beneficial use of organic residuals for the sustainable establishment of bioenergy crops and other productive land uses Objectives ??To advance the current ?standard? of mine reclamation by demonstrating that mining lands can safely become agriculturally ?productive? through the use of organic residuals. ??To generate and compile sound scientific data that can be utilized by stakeholders to better assess the environmental impacts of utilizing these materials. ??To complete a manual of best practices regarding the utilization of organic residuals on mine lands. *Considerable effort to be placed on operations/implementation side of utilizing residuals ?rather than just pure research

May 1, 2011

"HPGRs operate in a wide array of applications in mining and minerals. These range from relatively coarse crushing of copper and gold ores at one end of the spectrum to the processing of fine iron ore pellet feed and fine industrial mineral products at the other. Traditionally these circuits involve the treatment of relatively moist materials in pre-grinding circuits; using single stage HPGRs in open circuit or as a closed circuit operation with screening or partial product recycle. More recently, HPGR is increasingly deployed in dry grinding in combination with air classification. The high throughput demonstrated by larger HPGRs has stretched the application towards coarser size reduction duties for a number of new projects. At the same time, the basic principle of inter-particle crushing for generating a product with a high fines proportion gives ground to investigate applying the technology further towards finished product duties. The advantages that HPGR brings are clear. Multi Stage grinding enables energy efficient size reduction across more of the comminution circuit. Primary HPGR serves for a coarse grinding duty, whereas a secondary HPGR is geared towards generating a fine product for downstream processing or even a finished product. This publication summarizes some of the features of staged HPGR circuits and discusses some examples of operating and projected applications.INTRODUCTION In most HPGR applications, the grinding units are used as a single process stage, either in open circuit, such as for pebble crushing [Rose 2002, Westermeyer 2000, Van der Meer 2015] or in closed circuit with classification [Van der Meer 2009, 2012] or product recycling [Van der Meer 2010]. Where a higher plant capacity is required, the HPGRs are commonly installed in multiple lines in a parallel arrangement [Parker 2005, Koski 201, Vanderbeek 2006]. HPGRs can take a wide range of feed size distributions, and grind this down to a product with a high fines content. In general view, there is a limit on the size reduction range that even an HPGR can provide. Grinding down from a very coarse ore (e.g. 70-100 mm) to a very fine product (e.g. 100 microns range) is stretching the ability of a single stage HPGR circuit. Fine grinding generally requires a classification stage that returns the classifier oversize to the head of the HPGR circuit. From one point of view, it can be demonstrated that when recycling a large proportion of fines and middlings back to the HPGR feed results in an effective feed with an overrepresentation of middlings and fines, which will eventually reduce the grinding efficiency and consequently result in an increased recirculation mass flow [Van der Meer 2014]. From another point of view the large recycling mass puts a strain or even a limit on the capacity of present-day conveyor capacity."

Jan 1, 2018

By F. Fournier Dit Chabert, F. Tancret, J. François Castagné, R. Le Gall

During the repair of turbine blades by a laser cladding process, it is observed that interdendritic cracking occurs in the heat affected zone (HAZ). Auger electron spectroscopy, performed on sample broken at high temperature, reveals interfacial sulfur segregation, which is known to be one of the main causes of microcraking in the HAZ during welding processes. Moreover, thermal gradients generated by the process are responsible for important thermomechanical stresses. In the present work, the finite element code Comsol Multiphysics is used to simulate the laser cladding process in order to determine both the thermal field and, using an elastoplastic law, the stress field inside the component. The originality is that the temperature history calculated in the thermomechanical model is extracted then injected in a segregation model in order to calculate the local segregation behavior. This model can be eventually used to predict the risk of cracking as a function of process parameters and alloy chemistry.

Jan 1, 2006

By R. S. James, D. C. Peck, P. T. Chubb

"The East Bull Lake gabbro-anorthosite intrusion is one of several Early Proterozoic layered mafic intrusions occurring along the Southern Province - Superior Province boundary in the Elliot Lake - Sudbury region, Ontario. The base of the East Bull Lake intrusion comprises abundant plagioclase cumulates (Leucogabbronorite Zone) hosting two distinct styles of sulfide mineralization: (1) disseminated magmatic sulfides, and (2) disseminated to massive, structurally controlled sulfides. Both types of mineralization are enriched in Pd, Pt, Au, Cu and Ni. The host rocks to the sulfide mineralization display a range of interesting features including nodular textures, dendriticpegmatitic mafic pods, inclusion-rich zones (pyroxenite, anorthosite and granitic inclusions) and magmatic breccias. Disseminated chalcopyrite, pyrrhotite and pentlandite occur in concentrations of up to 100/0 within erratically distributed gossan pods near the base of the intrusion. The sulfides were originally precipitated from intercumulus liquids, but were subsequently recrystallized and locally remobilized along fractures during the major period of metamorphism and deformation that affected the East Bull Lake intrusion. The base- and precious-metal tenor of the structurally controlled sulfides can be explained by extensive remobilization and reconcentration of the pre-existing magmatic sulfides in the vicinity of a major shear zone. Both styles of sulfide mineralization locally contain ppm abundances of POE + Au and are generally Pd- and Cu-rich, although both display large variations in Pd:Pt and Cu:Ni ratios. The magmatic sulfides are more strongly enriched in POE and contain up to 1470 ppm Pd in 100% sulfide. Several factors, including source characteristics, magmatic processes (crustal assimilation, fractional crystallization) and post-magmatic interaction with hydrous fluid, have influenced the present distribution of POE, Cu and Ni in the East Bull Lake intrusion.The magmatic sulfide occurrences from the East Bull Lake intrusion compare favorably with marginal-type deposits from rift-related intrusions (Noril'sk-Talnakh region, Siberia; Duluth Complex, U.S.A.) in terms of their distribution, but are most similar to reef-type POE deposits with respect to their base- and precious-metal geochemistry. The combination of high POE tenors and large tonnages that is observed for the POE-Cu-Ni mineralization from the East Bull Lake intrusion suggests that a good potential exists for the presence of economic concentrations of Pd, Pt, Au, Cu and Ni in the intrusion. On a global scale, the East Bull Lake intrusion is part of a suite of Early Proterozoic plutons, many of which contain significant volumes of stratiform and/or marginaltype POE-Cu-Ni mineralization. The East Bull Lake intrusion is also one of the oldest mineralized intrusions derived from a sub-Superior Province mantle source region that has given rise to several significant magmatic POE-Cu-Ni deposits during the period 2.5 Ga to 1.1 Ga."

Jan 1, 1993

By Rodrigo Madariaga, Tracy T. Morris

The practice of stripping copper starter sheets from titanium blanks to be used in a conventional electrolytic tankhouse has been around since the advent of Electrometallurgy. Historically it has been done manually, utilizing many people to produce in some cases thousands of starter sheets per day. Some other conventional tankhouses have successfully converted to using mechanical means to strip starter sheets or have made the switch to permanent cathodes and although robotics have been around for quite some time they have not been used to strip starter sheets until recently. The Amarillo Copper Refinery began using robotics to strip several years ago and the details to how this came about and how the robots work will be described. INTRODUCTION ASARCO LLC (Asarco) was organized in 1899 as the American Smelting and Refining Company. Originally a consolidation of several lead-silver smelting companies, it evolved over the years into an integrated producer of primary non-ferrous metals, including copper, lead and zinc and associated co- products (Marcosson, 1949). Grupo México, S.A.B. de C.V acquired Asarco in 1999. Today, Asarco is a vertically integrated producer of primary refined copper and associated co-products such as gold, silver, molybdenum, selenium and tellurium. Asarco operates three open-pit mines and mills and two solvent extraction-electrowinning (SX-EW) plants all in Arizona plus a copper smelter in Arizona and a copper refinery in Texas. Figure 1 illustrates how the copper metal comes out of the ground at the mine and ends up as a refined product at the refinery The copper refinery in Texas is known as the Amarillo Copper Refinery (ACR) and is located in Amarillo, Texas. It was built in 1975 to replace 3 other aging refineries and was designed to produce 550,000 tons (500,000 MT) of refined copper per year using conventional starter sheet technology. At the time it was commissioned it was the largest tankhouse in the world at just under 1/2 miles long under one roof.

Jan 1, 2019

By Adrian White

Agenda Overview of pastefill at Kidd Mine Pipeline wear and the need to monitor Steel Loop monitoring Steel Borehole monitoring Monitoring pipe linings

May 1, 2013

By Steve R. McCutcheon, Cees R. Van Staal, John P. Langton, Leslie R. Fyffe

Abstract -The volcanic-dominated Tetagouche Group in northern New Brunswick contains numerous massive-sulfide deposits and intimately associated metalliferous sediments. The massive sulfides are generally hosted by Llanvirnian-Llandeilian volcaniclastic sedimentary rocks of the Nepisiguit Falls, Patrick Brook, and Boucher Brook formations. These formations are intercalated or closely associated with rhyolitic-dacitic flows and/or domes, and minor continental tholeiites of the Flat Landing Brook Formation. Pyroclastic rocks are relatively scarce in the silicic volcanic complex. Tuffites and epiclastic sandstones, of the Nepisiguit Falls and Patrick Brook formations, respective­ ly, generally occur along the fringes of the silicic volcanic complex. The flows and tuffites are intruded by quartz- and feldspar-phyric rhyolite (porphyry) dykes, sills, and domes. Although the volume of the silicic volcanic complex is large, evidence for a previously proposed collapsed-caldera model is lacking. On the basis of host-rock type and stratigraphic position, the sediment-hosted massive sulfides have been divided into three groups: the Brunswick-, Caribou- and Halfmile Lake-type deposits. Syngenetic feeder pipes and asociated alteration zones are generally not present beneath the Caribou- and the Halfmile Lake-type deposits and, if present at all, are poorly developed underneath the Brunswick-type.

Jan 1, 1992

By P. J. Urso

"This paper discusses the mining problems encountered at the Fording River coal mine to date, and how they have been solved. All of the production now comes from surface mining, but underground hydraulic mining is planned and briefly described. Other facets of the operation, including markets, equipment and personnel, are covered. IntroductionFording Coal Limited was formed in 1968 to mine and market metallurgical coal resources held by the company in the East Kootenay region of southeastern British Columbia. The company, at its Fording River operations, holds approximately 48,000 acres of coal lands, having an estimated 2.0 to 3.0 billion long tons of metallurgical coal in place. (Although not a point for further discuss ion in this paper, the company also holds approximately 1.0 million acres of mineral rights. These consist primarily of coal in Alberta, potash and coal in Saskatchewan, and minor industrial mineral holdings of limestone, salt and bentonite in the three western provinces.)Fording Coal is owned 60% by Canadian Pacific Investments Ltd. and 40% by Cominco Ltd. Inasmuch as Cominco itself is owned approximately 54% by CPI, it is seen that the effective CPI ownership of Fording is about 82%. The company, originally staffed and managed by Cominco, has been independent since January, 1978, with its own board of directors, management, staff and employees.The company's head office is in Calgary, Alberta. The Fording River operations are located approximately 45 miles north of the Crowsnest Pass, near the Alberta/British Columbia boundary, only 95 air miles and 200 road miles from Calgary.A contract was signed in 1969 with the Japanese steel mills to provide 45 million long tons of metallurgical coal over IS years, commencing in 1972 at an annual rate of 3 million long tons per year. Despite some start- up difficulties, the mine and plant facilities achieved their rated capacities in 1975, producing and shipping at the 3-million-Iong-ton level. The coal specifications are shown in Table I."

Jan 1, 1980

By W. C. Ferguson

"This paper could be more appropriately titled ""The Application of Vacuum Filters to Underground Minewater Treatment. in a previous paper co-authored by the writer (1) the history and development of minewater treatment at the Sudbury District Operations of the International Nickel Company was outlined. It is the intention of this paper to dwell more specifically WIth the the application of the vacuum filter to a problem found in most, if not all, underground mining operations. In the preparation of this paper the writer has made liberal use of the illustrations and text of the paper previously noted.Last year we heard Mr. Nemeth discuss the basic mechanisms involved in vacuum filtration (2), and today Yr. Cunliff has exposed us to the selection of the filter media (3). I will now briefly review the various vacuum filter designs that generally lend themselves to metallurgical dewatering applications.DRUM FILTERSThe revolving drum-type which filters on its outside periphery is illustrated in Fig. 1. This is the Oliver or Drum filter and represents the first commercially available continuous vacuum filter. The original design dates back more than 50 years, but with the exception of the method of discharge, is identical to the units being marketed today. This filter is generally applied to slow filtering slurries. The cake may be discharged in several ways; strings, scraper blades, rolls or by passing the filter cloth over a small diameter discharge roller (belt discharge).THE DORRCO FILTERThe revolving drum type which filters on its inside periphery is illustrated in Fig. 2. Here the slurry lies in the drum and the cake discharges through a chute or conveyor. This type of filter is most often applied to fast-filtering slurries.THE DISC FILTERThe revolving filter which filters on the vertical plane of the disc is referred to as the Disc or American-type filter. This represents the most inexpensive design in terms of cost per square foot of filter area. This filter is widely applied, but is probably more suitable for slower draining pulps."

Jan 1, 1971

By L. Klatzel-Mudry, D. J. Nikols, J. D. Hughes

"The Geological Survey of Canada Paper 88-21, A Standardized Coal Resource/Reserve Reporting System for Canada, is referenced by National Instrument 43-101 for the preparation of resource and reserve estimates on coal deposits. However, for all disclosure of mineral resources or mineral reserves for coal, issuers are required by NI 43-101 to use the equivalent mineral resource or mineral reserve categories set out in the CIM Definition Standards and not the categories set out in Paper 88-21. “Because the framework is based on Canadian deposits and their tectonic regimes, the Canadian Securities Administrators believe it is not reasonable to apply Paper 88-21 to foreign coal properties. The “Resource Links” at the end of this volume provide website links to guidance documents for preparing coal resource and reserve estimates in Australia, South Africa, and the USA.”This report discusses definitions, concepts and parameters used to determine coal resource and reserve quantities, and provides a framework to facilitate consistent categorization of coal quantities found within various depositional and tectonic regimes in Canada.Coal deposits, or portions thereof, are first categorized in terms of their geological complexity <Geology Type), and probable recovery method (Deposit Type). Methods for determining quantification parameters including seam thickness, areal extent and bulk density are outlined for each geology/deposit category. Within each geology/deposit category, coal resources are further divided, based on their assurance of existence, into measured, indicated, inferred and speculative subdivisions and, according to their feasibility of exploitation, into immediate and future interest subdivisions. Coal reserves within each geology/deposit category are subdivided according to whether they are within or outside active mining areas, and may be reported on in place, recoverable and saleable bases.A wide cross-section of the coal industry in Canada contributed to the development of this reporting system. To ensure that the parameters and concepts expressed in the system continue to reflect technological and economic realities in the future, the Coal Association of Canada has established a committee to facilitate periodic reviews."

Jan 1, 2003

By L. Beiza, L. Velásquez-Yévenes

Chalcopyrite is by far the most abundant copper sulphide on Earth, but it has proved to be problematic in its processing by leaching, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface during leaching. This study is focussed on the electrochemistry of chalcopyrite dissolution in acidic chloride system, which has recently gained attention as a feasible alternative to conventional sulphate media for chalcopyrite leaching. It is well understood that chalcopyrite dissolution from mineral grains depends strongly on the solution redox potential at the mineral surface as well as the composition of the product layer that may form on its surface. The chloride ions in solution support the Cu(II)/Cu(I) couple as the primary oxidant, as they can form complexes with the cuprous ions formed, increasing the dissolution rate. Additionally, these ions increase the potential range where chalcopyrite can be oxidized. The chemical reaction that occurs on the surface of chalcopyrite in chloride solutions in the presence of cupric ions will be characterized and the coulometry of the reaction will be reported to understand what is happening on the surface of chalcopyrite. Electrochemical experiments have been carried out to investigate the behaviour of mixed potential by open circuit potential (OCP) measurement for 30 minutes using sulphuric acid at pH 1 under nitrogen at 25⁰ C. The effect of chloride was analysed at 0, 20, 50 and 100 g/L and cupric ions from 0 to 50 g/L. The potentials obtained were then used to carry out chronoamperometry tests for a period of 30 minutes in absence of cupric ions. The mixed potential did not show a significant difference with an increase in cupric concentrations in the absence of chloride. Increased mixed potentials and current densities were observed at higher chloride and cupric concentrations. Chronoamperometry tests at those potentials were used to construct a voltammogram trace which has the same trend as a normal cyclic voltammogram but with much lower current densities. INTRODUCTION As a result of the depletion of copper oxides, the leaching of copper sulphide minerals has become the focus of many studies. Chalcopyrite, (CuFeS2) which is the most abundant of copper minerals, has proved to be challenging to process via hydrometallurgy, mainly related to its refractory nature, slow kinetic rates of dissolution and the formation of an inhibiting product layer on its surface, during leaching.

Jan 1, 2019

By R. D. Brown

"When elephants fight, it&apos;s the grass that gets trampled". It is becoming increasingly evident that the federal and provincial governments are awakening to the fact that their vigorous scramble for a "fair share" of mineral revenues which kicked off in 1974 has done considerable harm to the Canadian mining industry. Recent public statements made by government officials, particularly at the recently held annual convention of the Prospectors and Developers Association, would indicate that there is a general consensus in the government sector that changes must be made to the mining tax system. However, there is no evidence as yet that there is any agreement among the various governments as to what form these changes should take , who should make them , or when. It would seem to be worthwhile at this time to reflect on where the Canadian mining industry has been vis-a-vis taxes and "where it is at ". This article is based on a recent address given by R. D. Brown, of Price Waterhouse & Co., Toronto, at the annual convention of the American Institute of Mining, Metallurgical and Petroleum Engineers in Denver , Colorado.

Jan 1, 1978