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By Atul Bhatnagar, M. P. Sinha, Md Zahid Ahmad, Pallav Garg, Sanjay Kumar Singh, Shashi Ranjan, An Ecolab Company Nalco Water

"Froth Flotation is undoubtedly one of the most critical processes in coking coal washing plants. This work was carried out at Bhelatand Coal Preparation Plant of Tata Steel was with an objective of cost optimization in fine coal processing. The sources of raw coal for the plant include coal from underground mines of Jharia coalfield and procured open-cast raw coal. Coal from different sources vary in quality and possess different floatability characteristics.Synthetic collector is being used in plant as a flotation reagent. An alternate collector with new chemistry was developed for reducing total treatment cost without compromising product yield and quality. A series of laboratory tests were conducted to assess the performance of various sources of raw coal and plant scale trial was started. At first, the new generation coal collector was entirely substituted with the existing collector in the laboratory scale. However the plant scale trial data was not in accordance with laboratory results and indicated inferior flotation performanceThe issue was studied and data based conclusions were drawn that the new generation collector aids faster and uncontrollable kinetics. A decision was taken for co-dosing of the synthetic and new collector for the second phase of the trial. A detailed laboratory analysis was done again and few dosing ratios that predicted optimum results were selected for plant trial. The dosing ratio was optimised as per flotation yield and product ashes during the plant trial. The results showed significant improvement from the first phase and optimum dosing rates were decided based on plant observations and data collection. The influence of various process indicators including water quality, feed and product ashes and process yield was studied during all the trials and promising results were established. The co-dosing of both the collectors conferred cost advantage and enabled to achieve superior performance."

Jan 1, 2016

By G. Merle, David L. McKay, Elmar Randveer

"THE ENGINEERING AND TECHNICAL SERVICES department is responsible for supplying the following services:1. Administration and co-ordination of the work required for capital projects.2. Engineering, design and drafting services for inplant changes or improvements as requested by the operating• and maintenance departments.3. Rock mechanics.4. Preparation of plans and layouts required for efficient extraction of the ore as mutually agreed with the mine operating departments.5. Mine planning and long-term schedules in co-operation with the mine operating departments.6. Grade control (long term and short term) and the preparation of ore and waste removal schedules to maintain a uniform grade of mill feed. This also includes the calculation of metal reserves and metal production forecasts.7. Surveying, underground and surface.8. Operations research at the mine site.9. Industrial engineering.10. Chemical analyses and assays.11. Instrumentation.12. Environmental control.A staff of approximately 35 supervisors, engineers and scientists, and 80 technicians, supply the foregoing services. Consulting engineering and drafting firms are used to supply specialized services and technical personnel in-house at Timmins to handle overloads on capital projects under the direction of Ecstall engineers.The operating and maintenance departments depend on the engineering department to supply engineering, design and drafting services for any new facilities or for any planned alteration to existing facilities. To do this efficiently requires a high level of communication and co-operation with these departments. This has been a major factor in developing the present organization of the department shown in Figure 63.The organization is predicated on a strong central technical and engineering resource from which personnel can be assigned responsibilities for different projects, and in different areas, as the work load demands. Job satisfaction is very high because of the variety and challenge in the tasks. Additionally, participation in projects with production and maintenance personnel broadens each technical employee's knowledge of operating and maintenance problems."

Jan 1, 1974

By J. G. A. Stevenson

GENERAL Location of Property THE Stirling zinc-lead-copper deposit is in Richmond county, Cape Breton, about a quarter of a mile southwest from Stirling P.O., and some 45 miles southwest from Sydney. The property is accessible by road for at least nine months of the year, and during the navigation period coastal boats maintain a service touching at the port of Fourchu, eleven miles from the mine. When flying conditions permit, the Cape Breton Flying Club, of Sydney, will carry passengers to or from the property, a service which has been particularly advantageous on two occasions when hospitalization was urgently required. The Company has mining rights on leases Nos. 73-74, 78-80, and 101-104, inclusive, and owns 166 acres at Stirling and one acre at Fourchu. The ore deposit is in lease No. 73. History The Stirling deposit was discovered in the 'nineties', probably about 1895. The original discovery is said to have been made in the bed of Copper creek, which traverses the ore-zone. For a number of years, indifferent development work was carried on with a view to copper production only, and it was not until1915-16 that attention was directed to the possible production of zinc. Previously, the ore had been considered too complex for metallurgical treatment. During the period 1916-24 several operators attempted to develop the property, among them being the Stirling Exploration Company, Stirling Mining & Smelting Company, New Jersey Zinc Company of New York, Hayden & Stone Company of Boston, and The Eastern Mining & Milling Company. Up to 1924, the work was limited mainly to trenching, stripping, ? diamond drilling, and shot drilling.

Jan 1, 1937

By E. Van Den Neste

A concise historical review is given of the Overpelt plant and the general principles of zinc hydrometallurgy are briefly outlined, together with a description of the new electrolytic zinc plant ore roasting process and, more particularly, the original Overpelt roasting process. A detailed description of the flowsheet of the zinc electrowinning plant, .with an annual capacity of 100,000 t, successfully put into operation in May 1974, is given. The treatment of the residues is performed following the goethite process ; the purification is continuous in two steps; the tank house is equipped with cells containing jumbo cathodes; the handling and stripping of the cathodes is fully mechanized. Concluding the description, some metallurgical and other features of the project are explained in detail.

Jan 1, 1977

By M. Dupuis

Recently, a new approach to controlling alumina feeding in a H.H. cell has been proposed [1, 2 and 3]. This approach, labeled In Situ aluminium cell control, is based on a correlation between the slope of the cell voltage during a no feed observation period and the dissolved alumina concentration in the bath. Once the dissolved alumina concentration has been estimated, a primary calibration surface uniquely relating the dissolved alumina concentration and the anode-cathode-distance (ACD) to the cell voltage is used to estimate the cell ACD. Once both the alumina concentration and the ACD have been estimated at the end of the observation period, a PID controller can be used to regulate the rate of alumina feeding using only the normalized cell voltage to calculate the error between the estimated dissolved alumina concentration and a target value. The In Situ aluminium cell control algorithm has now been tested with the Dyna/Marc [4, 5 and 6] cell simulator demonstrating that it should be possible to control a H.H. cell in this manner with less variation in the dissolved alumina concentration and other process variables producing a prediction of improved cell current efficiency and power efficiency.

Jan 1, 2011

By The Mill Staff

Introduction The Hollinger cyanidation mill commenced operating in June, 1912, and 49,779,978 tons of solids had been discarded by the end of 1950. The present tailing disposal plant has been in operation since 1924 and holds 39,762,250 tons. The tailing disposal plant is on the property of Kayorum Gold Mines, Limited, about a mile and a quarter south of the mill. The area available for tailing covers 640 acres and is roughly l ¼ miles long by ¾ mile wide. Tailing was first deposited on the McKay veteran claims, l mile long by 1/2 mile wide, which were later included in the Kayorum property. This site sloped from northeast to southwest, the slope amounting to 106 feet between the northeast corner and the low land to the southwest. The low land was wet and a considerable area was covered with muskeg up to 17 feet deep. It was drained by a brook running across the property in a southwesterly direction. Later, when more ground was needed, the low-lying mining claims adjacent to the south and west boundaries were acquired to allow for the extension of the plant. This entailed some extra difficulties and expense that would not have been encountered had the whole of the present site been acquired with the original purchase. At the present time, the top of the tailing deposit stands at two elevations, but eventually the tailing deposited on the extension to the original plant will be raised to the same height as the older deposit. Previous to 1924, all of the tailing was deposited on Low land and in lakes near the mil1. Between 1924 and 1926 an increasing tonnage was sent to the McKay claims until, in 1926, all of it was being sent there. In 1936, the area of the deposit was extended and shortly afterwards all of the tailing was deposited on the new site.

Jan 1, 1951

By François Robert

"The Eastern Abitibi Sub province (Quebec segment), contains two main types of gold deposits: the common quartz-vein type and the rare pyritic type. Both occur in areas dominated by volcanic rocks and are spatially associated with major fault zones.Quartz-vein gold deposits consist of simple to complex networks of quartz-carbonate-pyrite veins in reverse shear zones and extensional fractures. They occur in areas dominated by mafic volcanic rocks and are commonly associated with small syn- to late-tectonic porphyritic felsic intrusions. Associated alteration, which postdates peak metamorphism, is dominated by carbonatization, sericitization and pyritization of the wallrocks. These deposits formed late in the geologicalevolution of the Abitibi.Pyritic gold deposits range from deformed massive sulphide lenses to zones of syntectonic pyrite-rich veinlets and associated pyrite disseminations. These deposits occur in environments dominated by felsic volcanic rocks that are typically converted to sericite schists near the deposits. The associated pre-metamorphic aluminous alteration is now characterized by metamorphic sericite, andalusite, kyanite and chloritoid, overprinted by a late retrogression of alumino- silicates. These deposits probably represent deformed volcanogenic massive sulphide deposits, but the timing of gold introduction in these is not fully understood.The differences in geological attributes reflect significant differences in the genesis these deposits: the pyritic deposits probably have a multistage genesis, including a pre-tectonic stage, in contrast to a single stage late-tectonic genesis for the quartz-vein deposits.RESUMELa partie Est de la Sous-Province de l’Abitibi( segment québécois) renferme deux types de gites d' or: ceux de type ""veine de quartz"" et ceux, plus rares, de type ""pyriteux"". Ces deux types se trouvent dans des secteurs particulièrement riches en roches volcaniques et mantrent une association spatiale avec des zones de failles majeures.Les gites de type ""veine de quartz"" correspondent à des réseaux de complexité variable de veines de quartz-carbonate-pyrite se développant dans des zones de cisaillement inverse et des fractures de tension. Ils se concentrent dans les régions dominées par des roches volcaniques mafiques et s’associent fréquemment à de petites intrusions porphyriques felsiques syn- a tardi-tectoniques. L’altération associée ace type de glte est postérieure à la phase majeure du métamorphisme et consiste en une carbonatisation, une sericitisation et une pyritisation des épontes. La mise en place de ces gites est tardive par rapport à l'évolution géologique de l'Abitibi.Les gites d' or de type pyriteux consistent en des lentilles de sulfures massifs déformes, des zones d' abondantes veinules pyrite uses syntectoniques, et des zones de pyrite disséminée. Ces gites se retrouvent dans des empilements contenant surtout des roches volcaniques felsiques qui sont transformées en schiste à séricine pres des gisements. L’altération a l’umineuse pre-metamorphique associée a ce type de glte a donné naissance, lors du metamorphisme regional, a la sericite, a 'andal ousite, a Ia lcyanite et au chloritoi""de; par la suite, les alumino-silicates ont été affectes par une alteration retrograde tardive. Ces gites representent probablement des depOts de sulfures massifs volcan gènes déformes, mais la chronologie de la mise en place de l'or dans ceux-ci est encore mal comprise.Les caractéristiques géologiques propres à chacun de ces types de gites reflètent des conditions de formation différentes: le type ""pyriteux"" montre une histoire génétique polyphasée, incluant une étape pre-tectonique alors que le type ""veine de quartz"" s’est développe au cours d’une seule phase tardi-tectonique."

Jan 1, 1990

By T. C. Rao, Deb Kanungo

"The results of tests carried out on a 3-in. hydrocyclone have shown the linear relationships that exist between: (1) the flow-rate of water in the feed to and the overflow product from the cyclone; (2) the flow-rate of solids in the feed to and the underflow product from the cyclone. The underflow size distribution of an operation has been found to be linear and its usefulness in evaluating particle distribution in the cyclone has been discussed. Effects of design and operating variables on the above relationships have been discussed. The performance of this 3-in. hydrocyclone has been compared with that of a 20-in. cyclone, the results for which are available in the literature."

Jan 1, 1973

"In many gold deposits, a reasonable amount of the gold that remains in the cyanide leach tailing is associated with pyrite. Some gold plants have found it economical to incorporate gravity circuits on the cyanide leach tailing to recover the gold bearing pyrite before subjecting it to fine grinding, then recycling the ground pyrite back to the cyanide leach circuit. For low grade gold deposits the loss of gold bearing pyrite is exacerbated as the broader process economics normally favour a coarse grind (> 150 ?m). An option that has received attention over the last decade is to recover the gold bearing sulphides in the milling circuit by gravity, similar to the widely-practiced gravity recovery of free liberated gold. The modelling and simulation of a gravity gold circuit in a milling flowsheet is relatively simple compared with a gold bearing pyrite circuit, as the recycle advantages and breakage functions for gold and pyrite are quite different. This current work represents an attempt to simulate the gravity recovery of gold bearing sulphides in a milling circuit that also incorporates fine grinding, using relatively simple simulation models. The simulation results are further explored with reference to the economics of this type of circuit.INTRODUCTION It is well known that gold is associated with pyrite in most sulphide gold deposits and that the amount of gold associated with pyrite will vary from deposit to deposit. Most of the gold associated with pyrite is liberated in the grinding circuit during the processing of a gold ore. However, some gold remains occluded in the pyrite and is lost in the final leach tailings. Little information is available in the literature on methods to enhance gold liberation from pyrite by specifically removing pyrite, either in the grinding circuit or from the final tailings stream, and grinding it finer. One of the earliest record of this approach is described by King (King, 1949) where pyrite present in Witwatersrand gold ores was recovered by circular jigs and Johnson barrels followed by fine grinding of the concentrate prior to cyanide leaching. Dominy et al., 1970, provides information on the use of Yuba jigs to recover pyrite from the discharge end of ball mills. Again, the jig concentrate was ground to a fine particle size before cyanide leaching to recover the liberated gold. Unfortunately, no information is provided in these articles on the additional gold recovery from implementing these circuits in a flowsheet. Martins et al., 1993, provides details of a gravity circuit using Reichert cones and spirals to recover gold bearing pyrite from a cyanide leach tailings followed by fine grinding of the pyrite concentrate in a Tower mill. The ground concentrate returned to the front end of the already installed leach circuit. Overall recoveries were increased by some 2% to 4% depending on the amount of gold and pyrite in the plant feed."

Jan 1, 2017

By F. D. Ball, D. E. Gemmell

"Coal mining in New Brunswick dates from the year 1639, when, as today, coal was produced from the area surrounding Grand Lake near the geographic centre of the province. Current production by N.B. Coal Ltd. is approximately 453,000 tonnes per year and comes from a single, thin Pennsylvanian age seam. The history of development and understanding of the geological setting of New Brunswick coals is discussed, beginning with the work of A. Gesner in 1839. The lithostratigraphic sub-divisions of the Pennsylvanian strata proposed by various workers are summarized and the present hypothesis that only biostratigraphic subdivision is possible is detailed. The Pictou Group, which includes the New Brunswick coals, consists of repetitive sequences of sediments consistent with flu vial and upper delta plain sedimentation. The development of coal mining in the province has progressed from horse-powered stream-bank stripping operations prior to the late 1800s through to the present use of open stripmining techniques utilizing eight walking draglines of various sizes. During the intervening years the introduction and eventual phasing out of underground mining occurred with the near termination of the coal mining industry in the province prior to 1973. The coal seam worked by N.B. Coal Ltd. averages 46 cm in thickness, has never been found to be more than 76 cm thick and has both high ash (12.4 -31.1%) and sulphur (4.9-7.8%) content. The wide variety of coal beneficiation studies and tests completed on New Brunswick coals have lead to the conclusion that a washed New Brunswick coal product could be expected with 10 to 12% ash, 5 to 6% sulphur, and 30,225 to 31,388 1/gm. Recoverable coal reserves, including those in the area licenced to N.B. Coal Ltd. and those licenced, although not yet in production, to both the St. Lawrence Cement Co. Ltd. and Provincial Holdings Ltd. at Lake Stream and Beersvil/e, total 60 million tonnes. At present, coal in New Brunswick is removed from prospecting and staking and all existing and future coal exploration and development will be administered by agreement under the Mining Act. Based on the results of a recently completed drilling project undertaken by the Department of Natural Resources, seventeen coal exploration blocks which carry minimum work commitments will be offered to the New Brunswick Electric Power Commission on a sixty-day first right ofrefusal basis. Those blocks not taken up by the Power Commission will be offered under the same commitments to the private sector."

Jan 1, 1985

By G Anders

The mucking out of sub-level headings by slusher and by Autoloader is described, and the performance and costs of both methods are compared. The introduction of the Autoloader is taken as an example to illustrate the meth-od used at the mine for evaluating the suitability of new equipment. The consecutive steps in the process are covered in detail -from the preliminary cost analysis and the preliminary estimate of savings and advantages, and the review of these data after a field investigation, through the elimination of dubious factors, and detailed time and performance studies at the property, to the final assessment of the new equipment and an evaluation of the initial forecast. The Autoloader in Operation. 1198 Introduction AT Mattagami Lake Mines Ltd., the mine workings are laid out basically on a pattern of 60-ft.-wide stapes alternating with 60-ft.-wide pillars, with stope lengths varying from 80 to 200 ft. and stope tonnages generally between 80,000 and 200,000 tons. Stapes are drilled off from sub-levels at 60-ft. intervals, with the

Jan 1, 1965

By K. Yamaguchi

"This Precious metals content copper scrap is treated with a copper smelting process, and copper and precious metals are recovered. The precious metal lost in the slag is increasing with increasing of treated scrap amount. To determine the portion of the precious metals chemically dissolved and that associated with the mechanically trapped matte in the slag will be a key factor to improve the recovery of those metals. The distribution of precious metals between copper matte and FeOx-SiO2 slag phases is determined at 1300 oC and a fixed partial pressure of SO2 of 0.1 atm for the matte grades from 40 to 70 mass% of copper. The distribution ratios show a tendency to increase when the grade of matte is increased. In addition, the solubility of platinum and rhodium in the FeOx-SiO2-CaO slag under the air and the solubility of platinum in the FeOx-SiO2 slag equilibrated with a pure platinum or the Pt-Cu alloy were determined at 1300oC in the range of oxygen partial pressure from 10-9 to 10-7. The solubility of platinum and rhodium in the slag tends to increase with increasing content of copper in the slag and oxygen partial pressure. INTRODUCTIONThe stringency of the supply and demand of platinum group metal (PGM) intensifies by localization for resources and increase of the demand for PGM, furthermore, smelting and the recycling of PGM in the nonferrous smelting process take an important role. Scrap containing a precious metal for the purpose of collection of PGM is treated by copper smelting recently. With increase of the treatment amounts, loss of PGM in slags is the serious problem.In considering of PGM loss in slags, knowledge of the distribution behavior for PGM between slag and matte phases or slag and copper phases and the PGM solubility in slags is important. The distribution ratios of silver and gold between the FeOx-SiO2 slag and the Fe-Cu matte were reported by Roghani (Roghani et al., 2000), Shinha (Shinha et al., 1985) and Takeda (Takeda et al., 1993) However, there is no data for the PGM distribution ratio. On the other hand, the PGM solubility in slags was reported by Sano (Sano et al., 1997) However, there is very few report for the solubility of PGM in the nonferrous smelting slags."

Jan 1, 2012

By R. K. Collings

The annual production of mineral and mineral-based wastes in Canada is conservatively estimated to be 400 million tons. Over 90 per cent of this amount is waste rock or mill tailings from the mining and mineral processing industry. The remainder are slags and ashes from the metallurgical industry, and sludges from the chemical industry. Because of impurity content, remote location, or both, most wastes have limited usefulness and, indeed, are a liability in that they must be transported to waste disposal sites and contained therein. Certain mineral wastes do, however, have potential value. Some contain small percentages of metals that may be economically recoverable with price increases; others consist of mineral material that may be used for such purposes as road base, construction aggregate, railroad ballast, and in the manufacture of bricks or blocks. Increased interest has been shown in mineral wastes in recent years. This paper reviews the types of mineral wastes available in Canada, briefly describes a number that have potential value, and notes possible uses and related research by CANMET and other organizations.

Jan 1, 1978

By Y. Wang

In this study, homogeneous (unseeded) precipitation of nesquehonite (MgCO3.3H20) by the reaction of MgC12 with (NH4)2CO3 in supersaturated solutions at 30 °C was investigated. Factors which influence the precipitation of MgCO3.3H20, such as initial concentration, temperature, stirring speed, titration speed, equilibration time, have been studied. SEM images show that the initial concentration and titration speed have significant effect on nesquehonite's crystal morphology and particle size. In addition, stirring speed and equilibration time also have some influence on its properties. With the optimization of operating conditions, the crystals can grow up to a length of about 40µm and a width of 5µm, indicating good filtration properties. High purity nesquehonite obtained in this study was calcined to produce highly pure MgO at 800 °C as shown by XRD results. The novel process may be used to separate Mg and Li from high Mg/Li ratio brines in Qinghai, China.

Jan 1, 2007

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

ABSTRACT: As part of an overall study of properties and processes influencing consolidation of oil sands fine tailings resulting from different extraction processes, self-weight consolidation tests were per-formed. Fourteen 2 m and 1 m high standpipe tests were instrumented to monitor the rate and magnitude of consolidation of seven fine tailings materials. Tests provided valuable information about the variation of consolidation related to differences in fine tailings material properties that result from a change in bitu-men extraction process (caustic versus non-caustic). In addition, the effect of adding a coagulant to caustic fine tailings was measured. Test results are presented and discussed in terms of consolidation rate long-term volume decrease, compressibility, and downward drainage. Non-caustic fine tailings had fasterinitial consolidation, which is important in tailings ponds for reducing storage capacity and returning decant water at a faster rate to the extraction plant. Similar results were obtained for caustic fine tailings with a coagulant addition. However, long-term settlement of the different fine tailings appears to con-verge to a similar value over time because it is then governed by effective stress.

Jan 1, 2011

By Xing-Liang Sun, Xian-Min Han, Zheng-Guo Zhu

The study on the material properties of frozen soils in China was being developed quickly in the past decades of years, because of the increase of the engineering constructed in frozen soil regions on the Qinghai–Tibet Plateau, and numerous research results were achieved mainly about strength and deformation behaviour and constitutive models of frozen soils. Studies have shown that the deformation procedure of frozen soils is irreversible; therefore, it is more appropriate to research mechanical properties of frozen soils by means of damage mechanics theory. In this paper, an elasto-viscoplastic anisotropic damage model is presented by introduction of Perzyna viscoplastic equation into analyzing the deformation of frozen soil. In addition, an anisotropic damage evolution equation is obtained by means of extending non-viscous damage to viscous damage. Triaxial tests at different strain rates are chosen to illustrate the application of this model, and good agreement between experimental data and numerically simulated results is obtained. The calculated results indicate that, under the same conditions, the damage at higher strain rate is less than that at low one. So the phenomena are explained from damage mechanics that the frozen soil has larger elastic modulus and shear strength at higher strain rates, and explained from mechanism that the emergence and development of micro-cracks in the body of frozen soil can be to some extent retarded at higher strain rates.

Jan 1, 2015

By John C. Wilmot, Robert E. Brewer, Royce J. Smith

In March 2003, Phelps Dodge started up an innovative pressure leach facility for processing copper concentrates at the Bagdad mine in Arizona. This paper describes the development of the concentrate leaching plant at Bagdad and the operational results from start up to present. Optimization efforts and an approach to mineralogical modeling of the concentrate to calculate oxygen utilization are also presented.

Jan 1, 2004

By R. Bashir, J. F. Hatley

ABSTRACT Northern Saskatchewan boasts some of the world’s largest known high-grade uranium deposits. The successful mining of these deposits cannot be accomplished without overcoming technical challenges, however, which include high grade of the uranium ore, specialized mining methods to deal with groundwater at high pressures, and poor ground conditions. This paper discusses the various mechanisms of inflows at three of Cameco Corporation’s sites in northern Saskatchewan. The risk of inflows is quantified in terms of unique challenges due to hydrogeological conditions, rock mass integrity, and uncertainty in geological conditions. Strategies in the case of an inflow are also briefly described.

Jan 1, 2013

By A. Alfantazi, E. Asselin

"The corrosion of commercially pure titanium (CP-Ti) was studied from room temperature up to medium temperature (~160°C) under conditions relevant to medium temperature chalcopyrite leaching processes. The corrosion rates were measured by immersion and electrochemical tests in sulfuric acid. The influences of chloride, cupric and ferric ions on the corrosion rate and behavior were also investigated. Corrosion of the CP-Ti did not account for the total current density measured by linear polarization tests. The corrosion rates of CP-Ti under medium temperature leaching conditions were surprisingly high. The corrosion rate measured in ferric containing solution at 160°C was approximately 20 mm/yr.INTRODUCTIONTitanium exhibits excellent corrosion resistance in many aggressive environments because of the formation of a very chemically stable, highly adherent, and protective oxide film on its surface (Mitsuo et al., 2002). Because titanium is the only metal available at a reasonable cost that resists the high temperature corrosive slurries in many leaching processes, it is often chosen as autoclave liner material where it is subjected to severe conditions of acidity, temperature and pressure (Grauman and Say, 2000).A number of distinctly different hydrometallurgical processes for the pressure leaching of chalcopyrite have been developed and patented by various companies and they can be classified into low, medium and high temperature regimes (McDonald and Muir, 2007). For example, chalcopyrite concentrate may be processed at medium temperatures (115-175°C) and pressures of 800-2200 kPa (Dreisinger, D., 2006; Sahu and Asselin, 2011). The lixiviant may contain 20 to 55 g/L sulfuric acid, 8 to 12 g/L chloride ions, 10-25 g/L Cu2+ (Jones, 1995, 1997, 1999), and less than 1 g/L Fe3+ (Defreyne et al., 2005). The amount of these additions is determined by the grade of the ores and details are shown in Table 1."

Jan 1, 2012

By F. Siciliano, D. Bai, S. Yue, A. M. Elwazri

The influence of niobium addition and very low temperature (`cool') deformation on the microstructure and properties of low carbon microalloyed steels was investigated. In this work, the specimens were heated at 1200°C and held at this temperature for 20 minutes. Some specimens were aged at 400°C for times ranging from 10 minutes to 10 hours followed by air-cooling. Others were subjected to deformation at 400°C (`cool deformation') prior to aging. It was observed that the proportion of acicular ferrite was increased as the amount of Nb was increased. It was found that cool deformation increased both yield and tensile strengths significantly compared to that of the aged only specimens.

Jan 1, 2005