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Orebodies of the Renison Bell - Mt. Bischoff tinfield in Tasmania are stratabound, conformable lenses of massive cassiterite-bearing pyrrhotite, with underlying discordant, fault-fracture-controlled and more siliceous ore. The deposits occur in eugeosynclinal, trough-deposited sedimentary and mafic volcanic rocks which are stratigraphically equivalent to diff erentiated volcanic rocks that contain volcanogenic massive base metal sulphide deposits in a parallel arc to the east. Textures of the massive pyrrhotite ore suggest that it has undergone metamorphic recrystallization. Although the massive ore contains both monoclinic and hexagonal pyrrhotite, a study of a few samples from one orebody suggests that their distribution is stratigraphically controlled, with the monoclinic variety abundant only at the base. The distribution pattern of rare-earth elements in massive ore is similar to that of volcanogenic massive base metal sulphides in New Brunswick, and suggests that mixing of sea water with exhalative hydrothermal brines occurred during ore deposition. Tin is a common accessory element in many volcanogenic massive base metal sulphide deposits providing strong evidence that it, too, is transported and deposited by volcanic affiliated hydrothermal brines. Sedimentary rocks containing the massive orebodies are fine grained, thinly laminated, compositionally banded and of mixed chemical sedimentary-tuffaceous origin. They include finely recrystallized chert, carbonate-rich and pyritic varieties; and soft sediment deformation structures in these rocks suggest that dense sulphides deformed underlying, less dense chert prior to lithification. Close to ore the chert becomes massive and coarsely recrystallized, as does the carbonate, which is here strongly enriched in Fe and Mn. Apatite is present in pyritic, cherty and carbonate-rich beds and in the ore itself, in all of which it shows a similar pattern of rare-earth element distribution, suggesting that it is of marine sedimentary origin. These deposits have traditionally been considered as epigenetic replacements of favourable carbonate horizons by hydro thermal flu ids derived from later Devonian porphyritic intrusions. Some of these intrusions have associated tinbearing greisens. The above relationships suggest, alternatively, that, like massive base metal sulphides, the tin deposits were of original syn-sedimentary origin. Circulating hydrothermal brines generated during early Cambrian eugeosynclinal subsidence and tectonism rose through steep fault-fractures, there giving rise to the discordant orebodies, and were discharged on the sea floor, where the massive pyrrhotite lenses were precipitated with the associated cherts and carbonates. In some places, Devonian intrusions remobilized the earlier deposits, forming greisens and newepigenetic veins. Finally , the geologic selling of these deposits, in trough deposited sedimentary and mafic volcanic rocks, suggests another favourable environment for massive sulphide exploration which is quite different from the domal felsic volcanic environment so heavily favoured currently in North America

Jan 1, 1979

By P. Cheung, J. Gulyas, R. Clary

Over the past 50 years, the use of autoclaves in the metallurgical industry has evolved from final metal recovery (hydrogen reduction) and purification (sulphide precipitate oxidation) to treating whole ores and concentrates. Recent process developments, such as the use of autoclaves for laterite leaching, have necessitated higher temperature and pressure operations, which in turn have seen marked advancements in technology and manufacturing capabilities to keep in step with process demands. This paper will discuss the advances in technology in terms of design, materials and construction of autoclaves and liners, typical dip tube design and geometric layout. This paper will also address specific auxiliary equipment, such as acid and steam injection systems, and components, such as agitator seal systems.

Jan 1, 2004

By H. Bouzahzah

Sulfidic tailings, subjected to atmospheric water and oxygen action, can generate acid mine drainage (AMD); which is the most important environmental problem facing the mining industry. AMD generating potential can be assessed by the static tests which consist of a screening method where the neutralization potential (NP) and the acid-generating potential (AP) of a given rock sample are determined. The AP and NP are determined separately in the different static test existing procedures which can be divided into two distinct classes: chemical and the mineralogical methods. For some chemical method, the AP is calculated on the basis of total sulfide sulfur content in the sample which may overestimate its AP if significant quantities of non-sulfide sulfur species are present (e.g. organic S in coals, some sulfate as gypsum, barite, etc.). However, the mineralogical method evaluates the AP by the sum of the individual sulfides participation on acid generation based on their concentration within the sample and can be more accurate than the chemical method. For AP calculation, the Paktunc (1999a) mineralogical static test takes into consideration all sulfide minerals occurring in the sample. However, it does not take into account the rates of acidity production of these sulfides submitted to weathering conditions. So, the present study focusses on the modification of Paktunc AP calculation by considering the acidity production rates of these minerals. To reach this objective, five base-metal sulfides (pyrite as reference, chalcopyrite, sphalerite, arsenopyrite, galena), diluted within quartz (25?75 wt.%) simulating mill tailings were subjected to a kinetic test to evaluate their AP. The results show that all the studied sulfides minerals generate acidity at different rates, so the Paktunc?s equation for AP calculation has been adapted accordingly.

Aug 1, 2013

By S. A. Scott

In 1913, potash mineralization was recognized at a depth of over 7,000 feet in an oil-well hole (Norcanol's Radville No. 1 well) drilled near the town of Raddlle, Saskatchewan. Although, at this depth, potash is not considered mineable by conventional underground mining methods, this discovery shed the first light on the potash deposits of Saskatchewan in much the same manner as oil-well drilling was responsible for the first discovery of potash in the Carlsbad Basin of New Mexico in 19:25. The discovery near Radville alerted the oil-well geologists and, as a result potash mineralization was recognized in the Verbata No. 2 well, near Unity, Saskatchewan, in 1946. The mineralization, which was at a depth ?of 3,466 feet, graded more than 20 per cent K20 over a thickness of eleven feet. This discovery, as its depth, thickness and grade were all amenable to mining,

Jan 1, 1963

By Y. Lizotte

"This paper examines the application of blasthole sloping to narrow vein mining in Canada, for ore deposits where widths are less than two metres. While the technique has now been utilized for more than ten years, it remains in the experimental stage and on-site testing is necessary for its adoption at new mining operations or in differ¬ent geological settings. It is intrinsically a safer method than other narrow vein methods in current use.The investigation describes the main operating parameters for mines recently surveyed. The criteria for method applicability are defined and the problems encountered are studied. It is concluded that blasthole sloping for narrow vein mining is a viable modern method which promises to gain popularity and that emerging fragmentation technology and blasting research can be used to further increase applicability of the stoping method."

Jan 1, 1991

By J. G. Whellock, J. W. Matousek

"In spite of their importance to the economics of many metallurgical projects, the values and technical aspects of precious metals losses in smelting operations are relatively undeveloped subjects. This paper attempts to provide a condensed overview of precious metals losses in pyrometallurgy, using some of the limited production and experimental data available. It is intended for three general groups of metallurgists - operators attempting to improve recoveries, cost-accounting engineers making estimates of metal losses in new or modified smelting plants, and research engineers and students.Distribution coefficients of precious metals, the ratios of metal or matte assays to slag assays, are convenient means of comparing smelting processes and the tendencies of the different metals to enter the slag. For silver in lead blast furnaces this ratio is approximately 250;for gold and the platinum group metals in submerged-arc, electric furnaces it is 900. The ratio for gold and PGMs in converters varies with the degree of oxidation of the charge. The over-all cycle-average for industrial installations is tentatively estimated to be 35.IntroductionThe magnitude of precious metal s losses in pyrometallurgical operation s is difficult to assess. The primary producers are generally reluctant to discuss the subject. The by-product producers (the large sulphide smelters o f copper , nickel, and lead) might be more willing to disclose information, but these have precious metals in their feeds in such dilute quantities that accurate accounting is extremely difficult. It is not until numerous processing steps have been taken t hat the precious metals become sufficiently concentrated to permit acceptably accurate sampling and analytical procedures to be followed. Thus, knowledge of the precious metals contents o f the intermediate mattes and slags may simply be unavailable."

Jan 1, 1990

By M. Palfreyman

"Most of the minerals employed in Canadian refractory production are imported; in fact, most of the refractories consumed in Canada are imported. Whereas this situation is partly due to a definite shortage of suitable raw materials, e.g., fire clays, it is possible that the potential of certain mineral deposits which could be usefully exploited has not been recognized. There are exceptions, notably magnesite and dolomite, from which refractories are produced in Canada.In many cases, known mineral deposits are located either in inaccessible areas or in areas where transportation would be a major economic factor. This paper lists some minerals and their respective known deposits which could be employed in the production of refractories, irrespective of economics."

Jan 1, 1973

By Adam J. Szladow

"INTRODUCTIONThe low calorific value and the high moisture and ash contents of run-of-mine (r.o.rn.) lignite limits the utilization of Saskatchewan lignites to thermal power plants in Saskatchewan, Manitoba and western Ontario. On the other hand, the low sulphur content and the low mining cost make lignites potentially attractive for new coal markets, if suitable processes can be found for upgrading. Economic upgrading to a higher quality fuel would make lignites more acceptable for industrial and utility boilers.This paper presents a summary of the process engineering evaluation of coal upgrading technologies conducted in support of a study of selected lignite upgrading processes. The summary presents the relative advantages and offers specific comments on the application of coal upgrading technologies for lignites. ResourcesLignite comprises the largest fossil fuel energy deposits in Saskatchewan. The total resources of immediate and future interest exceed 27 billion tonnes. The measured and indicated lignite resources of immediate interest are large enough to last for over 200 years at the current production rate of about 10 million tonnes per year (Table 1).Saskatchewan lignitedeposits are present in three different geologic formations: (I) the Ravenscrag Formation of Paleocene Age in southern Saskatchewan; (2) the Swan River Formation of Lower Cretaceous Age in central north Saskatchewan; and (3) the Judith River Formation of Upper Cretaceous Age in central west Saskatchewan. The deposits in the Swan River and Judith River Formation are not currently of economic interest.The Ravenscrag deposits are confmed to four lignite basins, namely: the Estevan, the Willow Bunch, the Wood Mountain and the Cypress basin (Fig. 1). The estimated resources of immediate interest in the four basins are over 7 billion tonnes (Table I). These resources include lignite seams at least 1.5 m thick, at depths of less than 45 m.The average quality of lignite varies relatively little between the four basins considering the distance between deposits (Table 2). There is, however, a progressive increase in moisture and ash content, and a decrease in calorific value and carbon content as one moves westward from the Estevan to Wood Mountain and Cypress basins. The Estevan area deposits are of higher rank despite the fact that they are stratigraphically younger than the Willow Bunch deposits (Cameron and Marconi, 1979).In upgrading lignite to a higher quality fuel, it is important to know not only the average proximate and ultimate analysis (Table 2) but also other lignite characteristics influencing the product quality . For example, Figure 2 shows the effect of lignite HIC ratio on the product equilibrium moisture for hydrothermal treatment of lignite. It is apparent that lignites with a higher HIC ratio, or high liptinite content yield much better quality product for the same hydrotreatment conditions. Other examples of lignite characteristics which are important in upgrading include distribution of sulphur forms and mineral matter, maceral composition and degree of surface oxidation."

Jan 1, 1989

By V. N. Mackiw, H. Veltman

"A process is described which permits the separate recovery of lead and zinc from complex low-grade sulphide concentrates such as are produced from the New Brunswick type of disseminated pyrite ores. This process calls for the preparation of a lead-zinc bulk concentrate from the original ore by flotation. The bulk concentrate is then oxidized under pressure in autoclaves in aqueous sulphuric acid solution, such as return electrolyte from zinc electrolysis. This yields a zinc sulphate solution and a residue containing lead sulphate, unaltered pyrite and elemental sulphur generated during the oxidation. After the oxidation, the leach slurry is briefly heated to above the melting point of elemental sulphur and then cooled. Thus, pellets containing pyrite and elemental sulphur are formed, and they are separated from the lead sulphate by screening and/or flotation. Laboratory tests have shown that about 95 per cent of the zinc and lead can be separately recovered from complex bulk concentrates. Process possibilities are discussed."

Jan 1, 1967

By B. Simonian

"The Mount Milligan mine, located in central British Columbia, began operation in September 2013 with a design nameplate capacity of 62,500 dry metric tonnes per day, using a conventional SABC grinding circuit followed by bulk sulphide flotation. Over the last few years, many changes have been implemented in order to debottleneck and optimize the comminution circuit to meet and exceed the nameplate design capacity as well as improve the metal recovery of the flotation circuit. This paper discusses changes, grouped into mechanical modifications, projects, and operational changes. The implementation of these changes as well as ongoing challenges and potential solutions will also be discussed.INTRODUCTION The Mount Milligan mine began production in the fall of 2013. It is a conventional truck-shovel open pit gold and copper mine and concentrator with a 62,500 tpd design capacity copper-gold flotation processing plant. The average head grades for the life of mine are 0.18% Cu and 0.39 g/t Au. At full capacity, the mine produces an average 400 dry tonnes per day of copper-gold concentrate containing approximately 21.5% Cu and 50-60 g/t Au. Current mine life is 21 years. This paper will discuss a selection of the operational, engineering, and administrative changes that have been made at the site since the publication of An Update on Reaching Nameplate Throughput and Continuous Improvement at Mount Milligan (Yue, 2017). PLANT FLOWSHEET CONFIGURATION Run-of-mine ore is crushed through a 60-110 gyratory crusher to a top size of seven inches. Up to 100% of the gyratory crusher product is diverted to the secondary crushing circuit by a moveable gate. The material diverted to the secondary crushing circuit is fed to a vibrating double-deck screen. The screen oversize is placed on a surge pile and reclaimed by four vibrating feeders to feed two 1250 kW standard head cone crushers in open circuit. The screen undersize is combined with the crusher discharge and returns to the conveyor feeding the coarse ore stockpile along with primary crusher product that by-passes the diverter gate."

Jan 1, 2019

By Francis W. Gray

Introduction Extensive undersea coal-mining has developed on both the east and west coasts of Canada, and at this time some four million tons, or 25 per cent of the total output of Canadian coal, representing a pit-mouth value of $14,000,000 per annum, is coming from coal seams lying under the sea. Of the coal now being mined in Nova Scotia, 55 per cent comes from under the sea, and this proportion is likely to increase. The Province of Nova Scotia is at this date receiving in royalties paid upon coal ?removed from undersea areas a sum exceeding $400,000 per annum. The subject is, therefore, one of special interest to the Mining Society of Nova Scotia and the Canadian Institute of Mining & Metallurgy, and while this paper will deal chiefly with undersea coal-mining in Nova Scotia, an attempt has been made to summarize the development of mining practice in the winning of undersea coal in other parts of the world, with the hope that such a recital may assist in the assessment of local problems. The connection between the mining of coal under the sea and an insular or maritime geography is obvious in regard to Great Britain, Japan, Vancouver Island, Australia, and Cape Breton Island, and also in regard to the curious conformation of Chile; the tendency towards growth of a mercantile marine and naval eminence is equally obvious where coal is mined at tidewater. But, there is the further and less favourable significance that the mining of coal under the sea is associated with poverty rather than with abundance of coal reserves.

Jan 1, 1927

By C. H. Buck, A. J. Taylor

"In 2015 Vale’s Clarabelle Mill began the process change to enable a higher degree of Cu separation. The flowsheet modification was part of a complete divisional flowsheet change being implemented in Vale’s Sudbury Operations. This paper will centre on some of the Operational Readiness (OR) work, culminating in the commissioning of the circuit in March 2017. The emphasis will be on the implementation of an integrated test plan to prepare for the process flow change through the mill.BackgroundClarabelle Mill (CBM) began making a saleable Cu concentrate in 2006 with the installation of the Cu separation circuit. Typical separation rate was 35-40% of Cu in feed reporting to Cu concentrate based on a target Ni concentrate Cu:Ni target of 0.7-0.8. The remainder of the recovered Cu (55-60%) reported to Ni concentrate and was smelted through the Copper Cliff Smelter.In 2013 a capital project, Surface Facility Upgrade (SFU), was implemented to focus on matching the surface plants capacities with the future mine plan to maximize revenues. This was to be achieved by redistributing the metal flow, specifically Cu, through CBM to increase metal unit content through the Smelter. The amount of redistribution of Cu was determined to be based on maintaining a stable Cu to Ni ratio (Cu:Ni) in the Ni concentrate.The change in Cu:Ni ratio in the Ni concentrate would affect the Sudbury divisional flowsheet, and required process changes across the all of the surface plants. The detail that surrounds the flowsheet changes as a whole will not be discussed, but the outcome of the decisions plays a large role in the operational readiness interactions between the mill, its suppliers and its downstream customers."

Jan 1, 2018

By Anthony A. Zoobkoff

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

May 1, 2013

By V. C. Pandolfelli, L. Paulino, C. E. Zangiacomi

The continuous improvement in aluminum industry over the last years has increased metal output at lower cost. The potlining plays an important role in this regard due to high cost involved and more recently in its environmental aspects. In order to increase potlife and thus minimize the aluminum cash cost, a systemic approach is needed, through investigating the main cathode failure mechanisms by using a pot autopsy as a powerful tool, and then making the necessary chances to achieve extended potlife. In addition, preheating and pot start-up may also damage the cathode if not performed with an appropriate technique or even if the lining materials are not specified correctly, among other factors. All these aspects will be presented and discussed in this paper. Increased potlife, especially now that the current is reaching values up to 300kA, will allow aluminum plants to match their goals more efficiently at lower costs.

Jan 1, 2006

By R. G. Helberg, S. M. Rebelo

"Copper and cadmium are removed from the zinc sulphate solution in primary purification by cementation reaction followed by filtration. The solids are sent to the cake leach section where zinc and cadmium are recovered from the copper rich slurry. The copper solids from cake leach are filtered, bagged and sold. In order to achieve a lower cadmium copper cake, an altered process flow is suggested. It is proposed the primary purification be split into two steps for individual removal of copper and cadmium. INTRODUCTIONCopper cake is a product of primary purification, also known as copper and cadmium removal. Copper and cadmium are removed by cementation reaction in four stirred reaction tanks. The slurry is then filtered through Shriver manual plate and frame presses with the solution advancing to the next stage of purification, hot zinc purification (HZP). The solids from primary purification are further processed at the cake leach section where zinc and cadmium are separated from the copper rich slurry. Cadmium and zinc are dissolved in the cold cake leach tanks using spent electrolyte. Cadmium is recovered from this cadmium rich solution by cadmium sponging, leaching and electro-winning circuits. The residue from this cold leach step is further leached using spent electrolyte at a temperature of 85°C. The copper rich solids from hot cake leach are filtered on a drum filter and bagged in 1 ton bulk bags for shipping to buyers.Current assays of copper cake are shown in Table 1. The goal is to produce a copper cake product that is below 0.5% cadmium."

Jan 1, 2012

By A. G. B. Hayes, C. L. Renzoni

"AbstractThis paper provides a rough outline of economic expectations for the near future and then goes on to discuss their effect on the expected demand for metals, using zinc, copper, nickel and aluminum as detailed examples.IntroductionMINING, as you are all painfully aware, is a cyclical business. We believe that it will remain so as long as there are economic cycles.The fortunes of the industry are tied to those sectors of the economy where decisions can be postponed until more favourable times: automobiles, housing, appliances, inventory building and capital spending. Over the past two years, while U.S. GNP fell by 3.8%, automobile production declined by 31 %, from 9.7 million units in 1973 to 6.7 million in 1975. Housing starts plummeted some 43% from 2.1 million units in 1973 to 1.2 million in 1975. Little wonder that U.S. copper consumption fell by 37% and zinc consumption by 38% over the same period.With the rest of the world's economies moving in unison with that of the United States, it is also not surprising that times have been tough for mining.Arthur Burns, chairman of the U.S. Federal Reserve, contends that each recession sows the seeds of its own recovery. The worse the recession, the better the subsequent recovery. If this is true, then the lot of the mining industry should improve significantly over the next two to three years."

Jan 1, 1977

By Güney

"An investigation carried out to determine the tendency of coals to heat spontaneously in terms of time-temperature curves coupled with the resulting oxidation products in an adiabatically controlled laboratory setup is described. Tests were made at differing moisture conditions, including oxidation and also oxidation with either adsorption or desorption of water vapour on coal.From the experimental evidence, there is reason to believe that the oxidation of coal, although an important factor, is not the predominant one in the development of spontaneous heating. In addition to the temperature increase in coal and the gaseous products evolved due to auto-oxidation, the process is accelerated by adsorption and retarded by desorption of water vapour on coal."

Jan 1, 1971

By John H. Nicholls

In this paper, Imperial Oil Limited reports briefly on the drilling by contract rigs of twenty-nine slim holes in the Virden, Smiley, and Souris Valley areas. The investigations showed that slim-hole penetration rates, as a result of careful control of drilling variables, can approach those achieved with 7 7/8-inch bits. It is concluded that slim-hole drilling is acceptable for development or exploratory work. Imperial also reports on the completion techniques and equipment that have been developed on twenty-seven tubingless 2 1/2-inch casing completions. Indications are that the method is economically desirable where excessive gas or water production is not anticipated.

Jan 1, 1958

By Harold S. King

TWO summers ago I had the privilege of visiting No. 18 colliery at Glace Bay. Underground, we travelled along the haulage-way behind a powerful electric locomotive. The motor in this locomotive began to smoke badly and a carbon tetrachloride type of extinguisher was used several times during the trip. At the time, I noted a strong odour of phosgene. On my return to Sydney, I spoke to Mr. Harold Gordon, Assistant to the General Manager of the Dominion Steel and Coal Corporation, and warned him of the danger of using this type of fire extinguisher in confined spaces, since it is a well known chemical fact that phosgene, a dangerously toxic gas, is formed by the decomposition of carbon tetrachloride. This danger does not seem to be widely recognized, so it seems advisable that a more general warning be published. Attention was first drawn to the danger by the death of two men in the United States Navy Department subsequent to breathing fumes from carbon tetrachloride used in a fire, and when an employee of the Bureau of Mines also was overcome while fighting a fire with carbon tetrachloride liquid. These accidents led to an investigation of the decomposition products of carbon tetrachloride by a group of workers in the Research Laboratory of the Pittsburgh Experimental Station, U.S. Bureau of Mines, the results of which were published in the October, 1920, number of the Journal of the .Franklin Institute (1). Their experiments were conducted by two methods. The first consisted in applying the liquids to actual fires and -to hot metal in a closed room of 1,000 feet capacity, where the gaseous products, mixed with air, could be retained and analyzed. The second method was to pass the vapour:s in air through heated tubes of iron or quartz, where the humidity of the air and the temperature could be controlled.

Jan 1, 1939

By J. -F. Lepage, M. A. Tomlinson, K. K. Murray

"The Éléonore mine, located in the James Bay region of Northern Quebec, generates a gold bearing flotation concentrate rich in pyrrhotite. Direct leaching of the finely ground concentrate results in high cyanide consumptions and poor recoveries, so effective pre-treatment to passivate the pyrrhotite is required before leaching. Following a number of process upsets at the mine, test work was performed to better understand the requirements for passivation so a more effective operational strategy could be implemented. This paper outlines the testing and implementation of the new strategy at Éléonore.ÉLÉONORE CONCENTRATOR The Éléonore mine is located 800 km north of Montreal, in the James Bay Region of Quebec, Canada. The ore, containing gold disseminated in sulfide minerals and free fine gold in silicate gangue, is crushed in a two-stage crusher plant before being ground in a single 9,000 kW ball mill. The ball mill is in closed circuit, with about 25% of the cyclone underflow passing through a gravity plant with two, 48” Knelson concentrators. Gold recovery in the gravity circuit varies between 15 and 20%, with gravity concentrate leached daily in an Acacia intensive leach reactor. Cyclone overflow, at a target P80 of 65 microns, flows to the flotation plant where the gold bearing sulfide minerals, mainly pyrrhotite and arsenopyrite, are recovered to concentrate with a mass pull of about 5%. The flotation concentrate, typically containing between 80 and 120 g/t of gold, is thickened and finely ground using a 1500 kW Isa Mill. The finely ground concentrate is diluted, to generate a slurry with about 32% solids by weight, before feeding the leach plant. Initially the concentrate slurry is pre-treated with oxygen and lead nitrate, to passivate the sulphide minerals before cyanide is added for gold dissolution. Gold is recovered from the leached slurry using carbon-in-pulp (CIP). The flotation tailings are also thickened and gold is recovered in a dedicated leach-CIP plant."

Jan 1, 2017