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By F. F. Roberto

Sulfuric acid has been demonstrated to be an efficient lixiviant for in situ extraction of copper in the past. It has also been proposed that microbial production (by iron-oxidizing acidophilic bacteria) of acidic ferric solution either in situ or ex situ could provide an economic strategy to generate an effective lixiviant for secondary copper sulfides like chalcocite and covellite. Copper heap bioleaching also takes advantage of this approach. One of the major environmental challenges for in situ copper recovery projects is the containment of acidic solution within the ore deposit, and ultimate restoration of the depleted ore body to its original state. A perhaps overlooked aspect of projected site operations is the possibility that transient acidic conditions and pumping of solutions from surface reservoirs into the subsurface could stimulate the growth of iron- or sulfur-oxidizing species of acidophilic microbes in the ore body, leading to additional generation of sulfate, acid, and ferric iron. This outcome would benefit extraction of copper in mixed oxide/sulfide deposits in the near term, but could complicate restoration activities. Possible impacts of pumping acidic solutions into the subsurface on microbes resident in, or introduced into the extraction zone will be discussed in the context of our understanding of copper heap bioleaching and acid rock drainage. INTRODUCTION As high-grade deposits of metals such as copper continue to decline, conventional mining methods have evolved to mine and process larger and larger quantities of ore. What is currently defined as ore today often contains less metal than the tailings of past operations. The throughput of current-day mines also leads to tremendous amounts of waste rock and tailings that may threaten the environment and communities around the mines, as has recently been experienced by 3 major tailings disposal facility (TSF) dam failures in the past 5 years. Community acceptance of large mine operations thus becomes more difficult to achieve—loss of the social license to operate. In response, the mining industry urgently pursues innovations that can fundamentally change the way in which metals are mined. In situ recovery (ISR) of metals represents one approach to a more sustainable mining industry, by virtually eliminating the need for a pit or underground excavation, eliminating the excavation and transportation of ore to a process plant and the ensuing noise and dust that is generated, and any need for waste rock dumps or TSFs. An ISR mine could have very little visible footprint beyond a network of wells, pumps, and piping to transport pregnant solution to a recovery facility, such as a solvent extraction- electrowinning (SX-EW) plant for copper. Some solid waste streams and/or storage ponds may still be necessary to treat and store bleed streams, PLS and raffinate, or as part of groundwater restoration activities. Thus capital and operating investments would be potentially much lower than that required for a conventional mine. In theory production could also be controlled to adjust to subsurface ore grade and match market demand. A not insignificant challenge of in situ recovery is the requirement for an efficient, low-toxicity, and ideally selective lixiviant for the metal that is targeted.

Jan 1, 2019

By M. Silva, G. Olivares

In a leach pile, the drainage layer or commonly known as the cover layer has been a matter of discussion due to the multiple phenomena or events that occur within it which impacts negatively the leaching process. All these events are directly related to the loss of draining capacity and irregular flow net due to an excessive compaction and entrapment of fines reducing the flow velocity which in turn cause more fines to get trapped. Over time the problem will continue to worsen rapidly and accelerating continuously like a geometric progression. When the cover layer gets to a certain level of deterioration, the only alternative to continue operating safely is to replace the entire cover material, a huge cost for any mining operation. TESRA has developed a solution to counteract the accelerated advance of the deterioration of the drainage layer. The solution consists of using large agricultural tractors coupled to a proprietary design of a heavy- duty subsoiler-ripper for deep tillage for loosening and breaking up the ground at various depth and with different legs spacing from 0.4 m or above. Also, if the ground still requires additional breaking, TESRA proposes the use of a rototiller that breaks any lumps to homogenize the ground to a certain particle size. With the subsoiler-ripper/rototiller combination equipment a layer of 0.5 m deep is properly treated to match closely the original conditions of the ground thus regaining the flow net as close as to the original seepage condition. Once the ripping is completed the subsoiler-ripper tool is replaced with another that is similar to a plow and thus continue with the installation of drainage pipes. The tool digs a Y-type profile trench allowing the pipe to sit precisely at the bottom of the trench, contributing additional lateral and lower support. This tool controls the installation depth of the drainage pipe at different inclination grades and also eliminating any irregularities that may block the flow of the leaching solution. Sometimes the obstructions block completely the flow of the solution. The main result of this methodology is the reduction of groundwater levels and at least 2% less humidity in the leaching gravel due to an increase in the draining capacity of cover. The methodology developed by TESRA applied during a couple of cycles of a dynamic leach pile has proven to be a real alternative to remedy and recuperate the cover close to its original conditions. The methodology requires only one large tractor and two accessories operated in different stages while replacing the more traditional and less efficient machinery such as the Bulldozer, Motor-grader, and backhoes.

Jan 1, 2019

CAT offers three different methods to accomplish crowd & retract on an electric rope shovel Rack and pinion Rope crowd HydraCrowd The presentation will compare the three systems and introduce the benefits of the Hydracrowd

May 1, 2013

By Xian Jian Guo, Lei Li, Hua Wang

The distribution and behavior of As, Pb and Zn elements in the copper smelting process were investigated using thermodynamics. The differences between thermodynamics evaluation and practical smelting processes were analyzed. The main factors affecting the distribution of the elements are oxygen enrichment in blowing, copper matte grade and smelting temperature. In general, the distribution ratio of As, Pb and Zn to off gas increases with the smelting temperature and decreases with oxygen enrichment in blowing. Meanwhile, their distributions in matte decreases and that in slag increases with the increase of copper matte grade. There is some gap in distributions between thermodynamics evaluation and practical smelting processes. INTRODUCTION The copper concentrates usually contain a number of minor elements, the control of which in copper smelting processes is often a key to the quality of the anode copper produced and may also have a bearing on the overall economics of the process (Yang et al., 2017). Knowledge of the thermodynamic properties of the minor elements in slag is very important in the control of the slagging reactions (Takeda Y, Ishiwata & Yazawa, 2007). In addition, the environmental contamination caused by the minor elements associated with copper-bearing ores, such as As, Pb and Zn, has become more and more important (Wang, 2004). Arsenic, lead and antimony, which are all carcinogenic metalloid elements, are usually contained in iron-rich wastes, and theirs emissions are considered to be one of the foremost environmental concerns (Wang, 2004). It is important to elucidate the elements distribution of them in the reducing-matte smelting process. The distributions of As, Pb and Zn elements in the copper smelting process are dependent on many factors, including the temperature, raw material composition, reduction atmosphere, concentration level, and slag chemistry etc. (Chen L Z, Xu X, 2016). Generally, these minor elements mainly distribute in phases of gas, copper matte and slag in the smelting process (Shu, Wang, & Qu, 2016). The arsenic exists mainly in the form of sulfo-arsenides in the copper concentrate, and it will decompose and be oxidized in the smelting process (Chen & Lu, 2016). Part of the products, including AsS, AsO, As2 and As2O3, evaporates into the gaseous state, and some other part enters into slag phase in the form of As2O3 and As2O5 (Wang, Guo, & Tian, 2017). Meanwhile, the As2O3 in the slag phase can also be reduced to metallic As and combine with Cu, S, Fe forming Cu3As2, As2S3 and Fe3As2, and then enter into copper matte. During the copper smelting process, about 50% of zinc enters into slag phase in the form of ZnO and 45% exists in the copper matte in the form of ZnS. Little volatizes into the gas phase. Most lead exists in forms of PbS and PbO in the copper concentrate, and it is mainly distributed in copper matte in the smelting process (Zheng & Lin, 2015).

Jan 1, 2019

By P. Taskinen, A. Jokilaakso

The Flash Smelting Process for copper and nickel production has been one of the most successful new metallurgical process developments in the 20th century. The experimental data from laboratory scale investigations of studies has produced information and insight into the reaction mechanisms and kinetics of the flash reactions of copper concentrates. Advanced computer models coupling chemical reactions with a Computational Fluid Dynamics (CFD) software became possible with the development of commercial CFD software together with increased computing power. Currently, there are both experimental and CFD modelling, including also CFD–DEM coupling (Discrete Element Method), studies ongoing for investigating the kinetics of matte – slag reactions in the flash smelting furnace (FSF) settler. The ultimate goal of the modelling work is to develop a comprehensive FSF model for studying the behavior of changing raw materials, both primary and secondary, and distribution of their elements between the matte, slag, and gas phases. The use of User-Defined-Functions (UDF) is one way to digitalize metallurgical knowledge and experimental information to be used for supporting process experts and development of artificial intelligence applications in metallurgy. In this paper, a review of the development of the FSF reaction shaft modelling and experimental investigations, and the current work for the settler reactions’ experiments and modelling are presented. INTRODUCTION The Outotec Flash Smelting process (Bryk, Ryselin, Honkasalo, & Malmström 1958) has been used in the copper and nickel industry for 70 years with continuous support from experimental investigations by many research groups in academia, among others Jorgensen and Segnit (1977); Jokilaakso, Suominen, Taskinen and Lilius (1991), Chaubal and Sohn (1986), Kim and Themelis (1986), and in industry Mäkinen and Jåfs (1982); Kojo, Lahtinen and Miettinen (2009); Tuominen, Pienimäki and Fagerlund (2016). It has been developed also for pyrite and lead smelting (Asteljoki, Sulanto & Talonen, 1984). A schematic illustration of the FSF and other main parts of the process is depicted in Figure 1. Also, two and three dimensional physical water or gas models have extensively been used in product and process development (Lilja & Rajainmäki 1995). During the recent decades, computer based models have increasingly been adopted in the FSF research and development (Ahokainen & Jokilaakso, 1998); Ahokainen, Jokilaakso, Taskinen & Kytö, 2006; Hahn & Sohn, 1990; Jorgensen & Elliot, 1992; Zhou, Zhou, Chen & Mao, 2014; White, Haywood, Ranasinghe & Chen, 2015) to name a few, but also in supporting engineering and sales of the technology (Miettinen 2017). Even though the reactions of the sulfidic concentrate particles have been included in the reaction shaft modelling, the complete FSF model with both gas-solid/drop and molten slag/matte drop has so far been and still is too demanding for today’s computer hardware.

Jan 1, 2019

By Virginia Lawson, Adam Price

The rougher flotation stage is used to reject liberated gangue at a coarse size. Base and precious metals due to their high specific gravity tend to recycle in the grinding circuit and report to the flotation circuit at a finer size than the gangue from which it is being separated. Generally liberation levels of greater than 50% are achieved in roughing and thus a portion of the feed can be immediately upgraded to final concentrate. The proportion increasing as the amount of liberated mineral in the feed increases. Scalping out concentrate from the rougher can reduce the size of the subsequent regrind and cleaner circuits. Examples of alternative process design for circuits demonstrates that you don’t need large volumes to effectively recover all of the minerals – you need smart flotation. Several case studies will be presented where circuits can be significantly reduced in size by designing smarter – not larger. This paper describes case studies to test the amenability of the Jameson Cell to produce final concentrate from rougher feed. Metallurgical performance and plant design implications are discussed. Froth flotation has been used extensively for mineral concentration for over 100 years. Although the flotation cell design has changed, it has generally simply increased in capacity rather than increase in efficiency. Cell capacities of over 600 m3 are now being used in rougher flotation particularly for the lower grade ores where the amount of surface area for flotation is deemed less significant than the required volume. The plant design volume in rougher flotation comes from laboratory batch flotation tests and/or pilot plants. The volume required for the plant is calculated based on the laboratory residence time multiplied by a scale- up factor. Examples of laboratory to plant scale-up design are given in Dunne et al. (2010 part 2). The factors are usually between 2 and 3. A laboratory residence time of 10 minutes would equate to a plant design residence time of 25 minutes using a 2.5 factor. The volume of the required flotation cells can then be calculated using the design tonnage and volumetric throughput. This scale-up factor takes into account the differences in performance between batch flotation efficiency and full scale plant. The differences are likely related to inefficiencies of bubble particle contact in large cells as well as froth recovery, which is essentially complete in a batch cell. Flotation cells, such as the Jameson Cells, are smarter designs that improve bubble/particle collection, bubble/pulp separation efficiency and operate with high froth recovery rather than simply getting bigger. The Jameson Cell was designed on the principal of improving the collection efficiency of fine particles by separating the contact zone from the separation zone and having the bubbles generated by the plunging jet shearing and entraining air. This means the method of bubble generation is not impacted by the wear of mechanical equipment as it the case in other flotation devices. Jameson Cells make bubbles less than 600 micron in size under most operating conditions (Ding, 2016).

Jan 1, 2019

By Aleksandar N. Nikoloski, Tendekayi Tapera, Glen P. O’Malley

Chalcopyrite (CuFeS2) and enargite (Cu3AsS4) are among the most important copper sulfides. At present, many known deposits of these minerals are uneconomic to process because they are of too low grade, insufficient size or too remote to be cost effectively treated with conventional technologies that can deal with their refractory nature. Therefore, developing an effective, low cost extraction process for recovery of copper from them is important. The research presented in this paper will deal with potentially cost effective strategies involving catalysed processing, studied under conditions relevant to recycle heap leach conditions. The present account summarises some of the main findings from the research program carried out at Murdoch University between 2014 and 2018 that has investigated the effects of the common leaching parameters, major and minor species present in the leach system and additives on the rate and extent of copper extraction from low-grade ores, in which primary copper sulfide minerals such as chalcopyrite and enargite were the dominant copper minerals. The study involved batch leaching tests, detailed structural and mineralogical characterisation and thermochemical modelling under conditions relevant to heap leaching. The results have demonstrated that silver in particular has a strong catalysing effect on the extraction of copper from ore under the leach recycle conditions and, at an adequate dosage, can effectively overcome surface passivation, enabling almost complete extraction of copper from chalcopyrite as well as enargite. The aim of this paper is to present the effect of the studied additives and to briefly elaborate on the mechanism of the silver catalysed leaching process. In addition, the presence of certain secondary agents was shown to lower the critical dosage of silver and promote extraction. The data from the study allowed the development of a theory for the silver catalysed reaction mechanism, which enables further optimisation and cost-effective application of this strategy in practice. INTRODUCTION Researching the extraction of copper from chalcopyrite is important as it is the most abundant copper mineral in the earth’s crust. The majority of ores containing this mineral are processed by flotation to produce a concentrate, followed by smelting and electrorefining. However, most reserves available nowadays are of low grade or limited in size, and therefore not suitable for this conventional approach. Hydrometallurgical processing via heap leaching with or without the aid of bacteria has been shown to be potentially economical. Chalcopyrite is refractory though and suffers from slow leaching kinetics and poor overall extraction, so understanding the mechanism of how it leaches has received significant attention by researchers working in this field over the past 50 years. The refractoriness, which is the main issue with the hydrometallurgical approach, is generally described as passivation of the chalcopyrite surface whereby a surface coating is formed that blocks the reaction interface or wetting of the chalcopyrite, limiting the copper extraction. Exactly what this layer is made of or how the issue could be overcome has been the subject of intense study and debate for many years. One of the limitations of the work to date is that much of the research focus has been on typical initial leaching conditions, while there has been relatively little research into how the leaching is affected under recycle solution conditions where certain metal ions (Fe, Mg, Al) have often accumulated in the liquor. This occurs because the leach solution exiting a heap contains many species from the dissolution of gangue minerals and typically this solution is returned to irrigate the heap, once the copper has been removed, leading to an accumulation of these species.

Jan 1, 2019

By Jocelyn Koshinsky

The uranium milling process at the Key Lake Operation of Cameco in Northern Saskatchewan, Canada results in the leaching of untargeted metal(loid)s such as arsenic, iron and nickel that are found at high concentrations in the ores. To mitigate the environmental impact of the leached As, Fe and Ni in the mill waste solution (raffinate), the pre-discharge mine tailings (rafffinate and residue) are neutralized with slaked lime, resulting in the formation of secondary As-, Fe- and Ni-bearing mineral phases which serve to control their concentrations to very low levels in the tailings porewater. As part of an ongoing investigation into the long-term evolution (10,000 year) of arsenic mineralogy in the Deilmann Tailings Management Facility (DTMF) at Cameco?s Key Lake Operation, it has become necessary to characterize secondary As-bearing mineral phases during the neutralization process prior to the discharge of the tailings and compare these to in situ As-bearing mineral phases in the tailings facility. This is primarily to determine if changes have occurred over time to As-bearing mineral phases in the tailings facility as well as to investigate whether alternate neutralization techniques may have a beneficial effect on the stability of mineral phases formed. Thus a bench-scale model was developed and preliminary testing conducted to simulate the mill neutralization process (i.e. neutralization of synthetic raffinate solutions with incremental amount of slaked lime to terminal pH values). The resulting ?tailings? (precipitates and supernatant) were then separated and analyzed. XRD analysis of the precipitates shows that the Fe occurred mostly as ferrihydrite. This was complimented with results from Raman spectroscopy which further indicated that the As appeared adsorbed onto the ferrihydrite. Results of geochemical modeling of the bench-scale test compared well with supernatant profiles from ICP-MS analysis.

Aug 1, 2010

1.Overview of the Ministry of Energy, Mines and Petroleum Resources (MEMPR) 2.First Nations Perspective 3.Revenue Sharing Framework 4.Economic and Community Development Agreements (ECDA?s ) 5.Questions? The Ministry of Energy, Mines and Petroleum Resources (MEMPR) MEMPR manages the responsible development of British Columbia?s energy, mining and petroleum resource sectors. Through the promotion of teamwork and positive working relationships with our clients, MEMPR facilitates a climate for thriving, safe, environmentally responsible and competitive energy, mining and petroleum resource sectors. Works closely with others in Government (e.g.. Integrated Land Management Bureau) to engage First Nations in consultation processes as well as revenue sharing negotiations.

May 1, 2010

THIS article describes the origin and details of a hydraulically operated switch to control the operation of an automatic underground pumping system. This system was installed in December, 1939, on the 2,350-foot level at No. 5 shaft of the Buffalo Ankerite mine. The pumping unit consists of an 8-stage centrifugal pump direct-connected to a 150 h .p., 2,200 volt motor. Pumping capacity is 200 gallons per minute under a 1,315-foot head and with a flooded suction. The control consists of a 200 amp., 2,200 volt oil-immersed starter with thermal overload protection, thermal protection on the pump, and a flow switch that is bypassed for 30 seconds after the pump starts. The 30-second time lag is operated by a relay mounted on the panel. There is a swing check valve in the discharge line of the pump which originally operated the flow switch by means of a connected outside lever which closed the switch as the check valve opened.

Jan 1, 1945

By J. L. Bowlby

OWING to the pressure of routine analyses at the laboratory during the past year, work of an investigational nature on water problems has been greatly curtailed. Favourable reports by the boiler inspectors indicate that the operators are giving satisfactory attention to feed-water treatment at the several . power plants. Some of the operators prefer lime and soda, others the phosphate and soda treatment. Springhill is still on the same commercial treatment as they were using last year, and satisfactory results are being obtained. This individual preference for a particular treatment, although it may be different for feed waters of similar nature, indicates that so long as proper conditions of alkalinity and concentration are maintained within the boiler, it is immaterial what chemicals are used in the treatment. This statement refers, of course, to the low-pressure boilers. In high-pressure boilers, a treatment such as lime and soda cannot be satisfactorily employed, on account of the precipitated calcium carbonate unduly increasing the amount of suspended matter. Only one case of impending trouble was reported during the year. That was at Seaboard, shortly after the new unit was placed in operation. Dissolved oxygen, considerably in excess of the permissible 0.030 cubic centimeters per liter, was detected at various points on the feed line. It was suspected that a de-aerating heater was being operated at too low a temperature, and chemical tests, made before and after the necessary adjustments, proved that this was the source of the trouble. While it has been impossible to give mine waters the attention they require, progress can be reported in the work carried out by the Mechanical Department. Past investigation of the action of mine water on various acid-resisting alloys is now bearing fruit, and it can safely be said that repair work on mine pumps has been reduced by nearly one-half.

Jan 1, 1938

By J. S. Scott

"The Quemont property comprises 640 acres lying directly north of and adjoining the Noranda (Horne) mine. During 1926 to 1928, United Verde Extension Mining Corporation sank a shaft 235 feet and did 3,000 feet of development work on the 215-foot level. Between 1929 and 1930, Quemonr Mining Corporation deepened the shaft to 926 feet, and did further lateral work and drilling on the 500- and 900-foot levels. This work was confined to an area close to the south boundary and met with little encouragement.Following a period of idleness a magnetic survey was made in the winter of 1944, followed in the summer of the same year by some 6,000 feet of drilling. The first successful hole was drilled in February, 1945, from the ice of Osisko lake.To the time of writing, 131,000 feet of drilling, surface and underground,have indicated a large tonnage of gold-silver-copper-zinc ore, and although some knowledge of the geology and structure is emerging, it is too early to draw definite conclusions. The following brief description deals only with those structural features that are reasonably well indicated."

Jan 1, 1954

By D. MacGREGOR, VERA TUREK

The St. Canul and Ste. Scholastique orthoquartzite sanrjstone deposits are located within the Potsdam Formation near Montreal. The sandstone lies on Precambrian basement and is Cam/3rian to Ordovician in age. The Potsdam Formation shows two distinct stages of development. One created a deepwater facies represented by impure, limey or shaley sandstone; the other produced a shallow facies composed of grey, highpurity sandstone. The shallow-water orthoquartzitic sandstone at St. Canul is being mined and beneficiated by lndusmin Limited. The operation manufactures a wide range of silica products and supplies industrial areas of Quebec and Ontario with material for the glass, foundry, silicon carbide and ceramic industries.

Jan 1, 1984

By Unkown

The Province-wide road ban on movement of heavy loads has had the effect of curtailing or stopping development in several areas. In the Leduc area, the discovery well has been restricted in flow to only a fraction of normal due to limited storage space, but drilling at Imperial Leduc Nos. 2, 3, and 4 has been continued for the past few weeks. The No. 2 well is within 100 feet of the producing horizon and should be completed within the next few days. According to reports, it has been logging slightly lower, structurally, than the discovery well, but, regardless of the outcome, it has been an important test to determine westerly cross-section of the Leduc structure. No official reports have been made available of any tests made during the drilling of the well, but it is understood that promising oil and gas shows have been indicated in higher formations. Actual drilling operations have not been commenced by any of the other companies holding leases in the immediate area but it is expected that, when road bans are removed and equipment is assured, several wells will be commenced. Announcement has been made that Imperial Oil, Limited, will drill a well near Battle lake, about thirty-five miles southwest of the Leduc field. Location of the well has been made following an extensive seismograph survey of the area, with total depth of the test well expected to be in the 8,000- to 8,500-foot range. Drilling will be done with the heavy-duty rig last used at the De Winton test. Imperial Oil have abandoned hope of oil production in the Provost area after drilling nine wells, but will retain their interest in the gas field that has been developed. Of the wells drilled, four indicated gas flows ranging from 1,060 to 13,200 M.c.f. per day.

Jan 1, 1947

By H. C. Gunning

The Privateer is the major mine of the Zeballos mining camp; with a total production from May 1937 through the first half of 1946 of $5; 610; 423 from 139; 254 tons of ore. This includes a small production from the connected Prident mine. The entire operation was closed from October 1943 to November 1945.

Jan 1, 1954

By Z. D. Hora

Commercial production of barite commenced in British Columbia in 1940; shipments have been recorded from ten deposits, of which four are particularly significant: The Brisco, Silver Giant, Parson, and Mineral King mines. Parr of the barite production is a by-product of base metal deposits in which barite is a major gangue mineral. About one hundred occurrences are known. COMMERCIAL production of barite in British Columbia dates back to 1940. Production volumes rose and fell and reached 39,900 tonnes in 1972. The main producer is Mountain Minerals Limited . There are two principal areas of barite deposits in British Columbia: ( I) The southeast part of the province along the Rocky Mountain Trench , which is the source of all production to date. (2) Stratabound occurrences in northeast British Columbia in the Gataga-Ketchika rivers and Muncho Lake areas which are not yet developed. Of the approximately one hundred reported barite occurrences, only ten have recorded commercial shipments, some of insignificant amount. So far, four deposits have proved to be important producers: Brisco, Silver Giant, Parson and Mineral King. The Mineral King property is situated near the confluence of Toby and Jumbo Creeks, 35 km southwest of lnvermere. The lead-zinc-copper-iron-barite deposits, as described by Fyles (1959, 1962), replace Proterozoic Mount Nelson dolomites of the Purcell system on the limb in the trough of a complex, tight syncline. The mine operated from 1954 to 1967 and 2,097,900 tonnes of baritic ore that was processed for base metal concentrates and also produced 22,680 tonnes of by-product barite. Since 1970, some 136,000 tonnes of barite have been recovered from the old tailings pond. The deposit is now considered depleted. The Silver Giant mine is located 10 km northwest of Spillimacheen. Barite and lead-zinc sulphides replace Cambrian Jubilee limestone near their contact with the overlying Ordovician McKay slate. Mineralization is located in the nose of an overturned plunging anticline. Between 1951 and 1957 800,000 tonnes of baritic ore were processed for lead and zinc concentrate. After 1959 barite was also produced , first from open pit, and later from reworked tailings. Up to 1979 production of barite totalled approximately 180,000 tonnes. Two other producing mines in the area are descri bed by G.R. Manson (this volume). Large, bedded stratabound barite deposits are reported from northeastern British Columbia. Six major shale-hosted Industrial Minerals in Canada lead-zinc-barite occurrences are known in the Devonian Gunsteel Formation. These are: Driftpi le Creek, Mount Alcock, Cirque, Pie, Fluke and Elf. In addition, numerous occurrences of strati form blebby or nodular barite and locally massive white crystalline barite are known from the Driftpi le Creek and Ak ie River districts in Gunsteel shale and from the Kechika, Road River and Ware map areas from Silurian siltstone . The size of these deposits is impressive; the baritic horizon at Mount Alcock is reported to be 25 to 30 m thick , and dri ll ing results from the Cirque deposit indicate reserves of 30 mi llion tonnes of baritic lead-zi nc-si lver ore. Similar bedded barite deposits are also reported from Midway area, near the Yukon border.

Jan 1, 1984

By B. Karnath, K. Guha, S. Tandon

At Jhagadia Copper Ltd (India), Copper is produced through secondary route using Top Blown Rotary Converter (Kaldo Furnace). The basic raw material being treated includes Matte-slag Reverts, Copper Cakes, dross & Cement, Oxide Ore, Low S bearing Concentrates & various metallics along with high grade copper scraps. The plant has capability to refine & produce Blister grade copper from Kaldo furnace as well as PS Converter . Blister is fire refined in Anode furnace & electro refined in Refinery to produce cathode copper. Our product , Copper Cathode , has already been established in the market as its quality is better than 'LME Gr'. The paper deals with the experiences encountered during commissioning & stabilization of the plant.

Jan 1, 2007

By J. E. Armstrong

"Pinchi Lake and Bralorne Takla mercury mines; Canada's only producing mercury mines during the past war (1939-1945); occur within the Pinchi Lake mercury belt. This belt is in central British Columbia; and is approximately 150 miles long extending from southeast of Fort St. James northwest to the headwaters of Omineca river.During the field season of 1937 cinnabar was discovered by J. G. Gray of the Geological Survey in Permian limestone on the north shore of Pinchi lake, where the main showings of the Pinchi Lake mercury mine were later developed. Following the publication of Gray's report claims were staked in the spring of 1938, and later that year were optioned to the Consolidated Mining and Smelting Company, and in June 1940 a reduction plant with a rated capacity of SO tons a day was brought into operation. In 1943 the plant had a capacity of more than 1,000 tons a day. The grade of ore treated was reported to average S to 10 pounds of mercury a ton, and production during the years 1940 to 1944 exceeded 4,000,000 pounds. The Bralorne Takla mercury mine, which is 90 miles northwest of the Pinchi Lake mercury mine, was brought into production in November 1943 with completion of a 50-ton mill. During the next 10 months more than 100,000 pounds of mercury were recovered, and the grade of the ore was reported to average 15 to 20 pounds of mercury a ton. During the war, production from these two properties was far in excess of Canadian requirements, and Canada was able to supply the United Kingdom and United States with part of their need.s. Both properties ceased operations during 1944, and, due to an excessive supply of mercury, have not yet reopened. The Pinchi Lake mercury belt lies along a northwest-trending major fault zone called the Pinchi fault zone (Fig. 1). This zone is approximately 150 miles long and varies in width from 200 to 1,000 feet . It forms the northeastern edge of a northwest trending fault belt up to 10 miles or more wide in places. In general the northeastern margin of the Pinchi fault zone represents the contact between closely folded, stratified Permian rock:s on the southwest and Mesozoic formations and Omineca intrusions on the northeast. It seems probable that the fault zone marks the site of major thrust faulting from the southwest and that the Permian rocks moved up with respect to the Mesozoic formations. Intense faulting occurs in the Permian rocks within the fault zone. There the more important faults trend north and northwest, dip steeply west and southwest, and may join a major low-angle thrust fault at depth."

Jan 1, 1954

By P. Henry

Hydrometal provides an adequate technology to reintroduce lost metal units into the industrial cycle, thereby saving resources and energy, while keeping the environment cleaner. Two case studies of our integrated hydrometallurgical recycling facility (Hydrométal) in Belgium are discussed.(I) Recycling of Cu-Co cement generated at the second purification step of zinc sulfate solution. (II) Recycling of lead complex materials from the lead industry.

Jan 1, 2008

By A. J. Rigby

Anode vessels are often relined with refractory every one to two years and a typical reline time can take between 16-20 days. Modem refractory designs using dense, corrosion resistant brick applied directly to the vessel shell in the high wear areas around and above the tuyeres and at the slagline can yield a lining life of up to 4-5 years before a complete reline is necessary. Installation methods using brick delivery on roller conveyors through an endwall access port and complete anchoring of the brick to allow for lining self-support can take days off the installation time. It is possible, therefore, to reduce the refractory relining period by 50% and increase lining life by 100% compared to typical anode repair schemes. Cu2007

Jan 1, 2007