Search Documents

By T. Kagawa

Ever-increasing demands for non-ferrous metals owing to population growth and improvements in the standard of living in the world, under the constraint of limited availability of natural resources, would easily make us imagine that more efficient and sustainable developments in mining and smelting activities are required in the near future with due consideration for environmental concerns. In order to realize the efficient and sustainable development of mining and smelting activities and operations, the following issues are indispensable: (1) to investigate the up-to-date pollution control technologies of mine and smelter, (2) to promote those technologies, (3) to develop technologies which are environmentally friendly. This paper discusses the recent technologies for environmental conservation/prevention and their diffusion in the world, and also the key technologies required for the most desirable sustainable development.

Jan 1, 2007

By Mayumi Ito

Mineral processing of rare-earth ore from the Dong Pao deposit, Vietnam, was investigated. The ore contained bastnaesite, (REE (rare earth element)) CO3F, as the valuable mineral and barite, quartz, and fluorite as gangue. To establish a mineral processing flowchart, selective comminution, size classification, and magnetic separation were tested. Bastnaesite shows brittle properties and selective comminution was applied to improve the purity of products. The crushed products were classified by decantation, and the overflow product showed high purity of bastnaesite indicating that bastnaesite becomes fines by crushing and that selective crushing can be used to concentrate the bastnaesite. Magnetic separation for the sink product was carried out and the mag-product showed high recovery rates of bastnaesite, indicating that the combination of size classification and magnetic separation is effective.

Aug 1, 2013

By A. E. Reschke

Optimized mine planning of the Leeville Gold Mine, Nevada, necessitated the requirement for an average of 2540 tpd of high quality cemented rockfill (CRF) in primary stoping blocks and underhand cut and fill workings. To meet this requirement, a number of innovative design features were incorporated into the overall batching plant design. Foremost is the use of a dedicated skip to transfer crushed aggregate from surface down through the ventilation shaft to twin underground storage silos. Also, the use of a high shear colloidal mixer on surface to pre-slurry a cement/flyash binder with pipeline delivery to one of two separate underground CRF mixing stations was incorporated. Finally, the binder and aggregate are combined in a 9 m3 twin shaft compulsory type mixer before loading into teleram trucks for placement.

May 1, 2007

By A. Hosseini

Full-scale rupture tests were conducted on 508 mm (20 in) diameter API X-60 5.7 mm wall thickness pipes containing either axial semi-elliptical cracks or simulated corrosion defects with equivalent depths ranging from 20 to 60 percent thickness. In the case of the cracks, the failure pressures were predicted with reasonable accuracy when considering a level 3 failure assessment diagram (FAD) procedure based on API 579, as compared to level 2 FAD, since the mechanical properties were known. In both cases, the accuracy improved for shallower cracks. Considering simulated corrosion defects, the RSTRENG approach provided more reliable results on predicting the collapse pressure. For a similar depth and profile, the cracked pipes failed at a lower pressure than those containing simulated corrosion defects of equivalent depth.

Jan 1, 2010

By E. Wright

"INTRODUCTIONThe Ruttan Mine is located twenty-two km east of Leaf Rapids in northwestern Manitoba. The copper-zinc exhalative massive sulphide deposit is comprised of a series of individual and interconnected en echelon lenses which strike east 20 degrees north and dip south 70 degrees. The mine has been owned and operated by the Hudson Bay Mining & Smelting Company since July 1987. Sherritt Gordon Mines Limited began mining the orebody in 1973 from an open pit, but production has been entirely from underground since 1980. HistoryDuring the progression from open pit to underground mining, it was recognized that manual methods of ore reserve calculation were inadequate for responding to changing metal prices and evaluating alternative mine plans. The mine planning system that was developed is described in the paper, ""Computerization of Ore Reserves and Mine Planning Functions at Ruttan Mine"", presented by B.C.W. Haystead at the 1984CIM Annual General Meeting. At that time, programs for core logging, diamond drill hole plotting, grade interpolation, ore outline digitizing, and reporting had been developed by the Systems Department for use on Sherritt's UNIVAC 1100 mainframe computer. Telecommunications linked terminals at the Ruttan coreshack and office to the UNIVAC in Fort Saskatchewan, Alberta. The same mine planning system was also used at a second property.Operation of the mine planning system was expensive, and it was evident that increased use required a costly upgrading of the UNIVAC in order to maintain an acceptable service level to all users. The system was re-evaluated,' and it was concluded that: core logging was not easily adaptable to a new property, nor flexible enough for exploration; plotting, grade interpolation, and digitizing required considerable manual effort; and that the UNIVAC operating system was difficult to learn.It was recommended that a study should determine if: core logging could be redesigned and/or implemented on a microcomputer; interactive graphics be used to reduce the effort for functions requiring visual interpretation; and the creation of input processors would assist users in setting up parameters for guiding the programs."

Jan 1, 1989

By W. Dresler

"A major Canadian chromite deposit at Bird River, Manitoba is one of the largest known chromite reserves in North America but it exhibits low chromium contents and low chromium to iron ratios. Commercial beneficiation techniques permit these ores to be upgraded to concentrations which, however, remain inferior to most other ores and concentrates used worldwide f or production of high-carbon ferrochromium. Analyses for reduction of low-grade concentrates by submerged-arc electrofurnace techniques predict chromium contents between 45.9% and 51% for a high-carbon ferrochromium alloy. Relatively high over-all chromium recoveries between 79.5% and 91. 1% can be achieved. Alloy compositions for Cr, Fe, C and Si are determined for various burdens with low-grade concentrates. Consumption data per ton contained chromium are obtained for electric power, coke, concentrate, quartzite and limestone. The data are compared with those of high-carbon ferrochromium producers in South Africa. Economic aspects are discussed and compared with a South African reference study.IntroductionThe present ferrochromium demand in the western world is estimated at approximately 1.7 Mt with a consumption of approximately 0.36 Mt in the U.S.III and 0.03 Mt in Canada'-', A low estimate of the installed production capacity for ferrochromium in the non -communist world is 2.4 Mt. This production capacity of ferrochromium by far exceeds the demand, which is also evident from the fact that no shortage in supply of ferrochromium to the world markets was encountered in the past. Supply-demand comparisons therefore suggest that in the foreseeable future no new chromite or chromium producers should enter the world markets unless superior chromites or superior technologies are successful in competing with existing producers. Neither one of these conditions exist presently in Ca nada or the U.S., although essentially no ferrochromium is produced and no chromite is mined in these countries at the present time."

Jan 1, 1986

By Rory S. McLaren

"Cost per ton as it relates to mobile underground vehicle maintenance and repair is invariably astronomical. The reason: the mechanic replaced a ten micron filter with a 250 micron filter because it appeared to be physically identical; the purchasing agent bought ""Brand X's"" filter because it was 25% cheaper- the result, a filter life 30% less; the hydraulic pump was sent for repair at a cost of $2000.00 and failed shortly after installation; pumps are failing because there is no testing equipment available in the workshop.This paper will identify a number of the common pitfalls resulting in high costs, and offer solutions to bring these costs to within acceptable limits.The following key issues will be discussed:A. Genuine parts replacement versus ""Brand X""B. Evaluating outside repair shopsC. The influence of environmental conditions on optional equipmentD. Personnel trainingE. Servicing and testing equipmentIntroductionThere is an old English aphorism, ""Look after the pennies, the pounds will look after themselves"".This paper is aimed at the pennies.Genuine Parts Replacement versus ""Brand X""High interest rates and floundering mineral prices are forcing many equipment owners and operators to purchase ""Brand X"" or ""Pirate"" parts in order to survive. ""Pirate"" part manufacturers invariably offer lucrative discounts which are most appealing to the average purchasing agent whose responsibility it is to save money. Unfortunately, in most cases, there is a tradeoff involved; when the price drops so does the quality, which can adversely affect owning and operating costs. Unless totally obvious, ""original"" or ""pirate"" part metamorphosis requires scrupulous analysis by competent individuals.Hydrostatic transmissions have been used successfully for many year s in various industries. They found their way into mobile underground loaders in the early seventies and have since become very unpopular in mining circles due to their apparent ""inherent unreliability"" and near ""prohibitive"" repair costs. The big benefit of the hydrostatic transmission is compact component packaging which enables the machine to work in very small headings."

Jan 1, 1983

By Francis Breton, Lei Pan, Kun Liu

"Effects of Fe, Cu and Mg minor alloying elements on creep properties of 8xxx aluminum conductor alloys were investigated by conducting compressive creep tests at 100 ºC under a constant load of 69 MPa. The microstructural evolution was studied using an optical microscope and electron backscatter diffraction technique. The creep properties were characterized by the primary creep strain in the first stage and the minimum creep rate in the second stage of creep deformation. Results showed that additions of all three minor alloying elements (Fe, Cu and Mg) demonstrated a beneficial effect on the creep resistance but individual alloying elements exhibited different impacts on creep properties. The additions of Fe and Mg showed a strong impact on both the primary creep strain and minimum creep rate, while the Cu addition displayed a favorable effect on the primary creep strain but no impact on the minimum creep rate. Microstructure observation revealed that with increasing Fe contents, the volume fraction of fine intermetallic dispersoids gradually increased while the subgrain size and misorientation angle became smaller, providing strong barriers to the dislocation and grain boundary movements. The addition of Cu moderately reduced the subgrain size and misorientation angle and in combination of solid solution strengthening it promoted the retardation of the creep deformation during the primary stage.INTRODUCTIONThe 8xxx aluminum conductor alloys offer significant advantages such as low density, high conductivity to weight ratio and low price over copper conductors and become an attractive candidate in electrical engineering and building wiring (McQueen, Chia, & Starke, 1986; Birol, 2009; Engler, Laptyeva, & Wang, 2013). The mass resistivity of aluminum conductors is one-half of copper conductors. Thus, only one-half the weight of aluminum is required for equivalent capacities. Therefore, this unique property coupled with a relatively low price of aluminum offers significant economic advantage to develop aluminum conductor alloys for electrical applications in the electrical wire industry."

Jan 1, 2016

By J. R. Pronovost, P. M. Tiley, D. W. Watts

"In 1980, J.S. Redpath Limited were retained by Dickenson Mines Limited to increase hoisting capacity from 700 to 1000 tons per day. The hoisting capacity of the present skip-over-cage system could not be increased significantly through design modifications. The other constraint was a minimum of disruption to the current production. The Dickenson shaft was a three-compartment shaft from surface to the 8th level and four compartments before the 8th to the 23rd level. The solution to the problem was to remove the services from the shaft man-way compartment from the 8th level to surf ace by drilling a borehole close to the shaft and rerouting those services through the borehole to the 8th level. A service compartment would then be available from surf ace to the 23rd level.By providing a service hoist in this compartment, the skip hoist would then be free to produce the 1000-tons/day required production. This paper will describe the search for the required hoist and the friction hoist selected for the project, and trace the design modifications required to modify the hoist to suit the Dickenson application and satisfy the Department of Labour safety requirements.The paper has been co-authored by J.S. Redpath Limited and C.L. Tiley Associates, who participated in the design changes and eventual commissioning of the hoist.IntroductionIn the earl y part of 1980, J.S. Redpath Limited conducted an engineering study on the Dickenson Mines Limited No. 1 Shaft materials handling system. The object of the engineering study was to investigate the possibility of increasing the tonnage hoisted from the No. 1 Shaft complex by adding a service cage compartment and service hoist. This would permit utilization of the existing hoisting plant solely for production hoi sting. A further consideration of the engineering study was that any alterations had to be accomplished with little, or no, interruption to production."

Jan 1, 1982

By J. R. Pronovost, P. M. Tiley, D. W. Watts

"In 1980, J.S. Redpath Limited were retained by Dickenson Mines Limited to increase hoisting capacity from 700 to 1000 tons per day. The hoisting capacity of the present skip-over-cage system could not be increased significantly through design modifications. The other constraint was a minimum of disruption to the current production. The Dickenson shaft was a three-compartment shaft from surface to the 8th level and four compartments before the 8th to the 23rd level. The solution to the problem was to remove the services from the shaft man-way compartment from the 8th level to surf ace by drilling a borehole close to the shaft and rerouting those services through the borehole to the 8th level. A service compartment would then be available from surf ace to the 23rd level.By providing a service hoist in this compartment, the skip hoist would then be free to produce the 1000-tons/day required production. This paper will describe the search for the required hoist and the friction hoist selected for the project, and trace the design modifications required to modify the hoist to suit the Dickenson application and satisfy the Department of Labour safety requirements.The paper has been co-authored by J.S. Redpath Limited and C.L. Tiley Associates, who participated in the design changes and eventual commissioning of the hoist.IntroductionIn the earl y part of 1980, J.S. Redpath Limited conducted an engineering study on the Dickenson Mines Limited No. 1 Shaft materials handling system. The object of the engineering study was to investigate the possibility of increasing the tonnage hoisted from the No. 1 Shaft complex by adding a service cage compartment and service hoist. This would permit utilization of the existing hoisting plant solely for production hoi sting. A further consideration of the engineering study was that any alterations had to be accomplished with little, or no, interruption to production."

Jan 1, 1982

By Thomas J. Ferree

"There has been a major evolution of mineral processing technology in the past decade. The evolution has been driven by recognition of broad environmental changes which have impacted various areas of the world. Numerous government regulations have required many mining and mineral processing organizations to introduce changes to reduce the undesirable effects on the environment; many mining companies have voluntarily introduced new technologies and practices which surpass the limits imposed by the regulations. Surprisingly, some of the 'new' technologies are modifications of concepts developed decades earlier. Gravity separation technology is in this category.Mineral Deposits Limited, an Australian company, has been a world leader in the development of mining concepts, reclamation practices and equipment technology that is environmentally suited to today's requirements. This paper relates to equipment used for gravity concentration and electrostatic separation systems, including Reichert cones, spirals, and electrostatic separators. The basic devices and technologies were developed more than forty-five years ago, but use of the systems was quite limited until environmental and conservation practices were recognized and introduced. Spiral concentrators, for example, were patented in the mid-l940s by I.B. Humphreys in Denver , Colorado. The two types of spirals marketed (the Model 24A for minerals and the Model 24C for coal) remained essentially unchanged until expiration of the patents. During this early period most spirals were used in iron-ore wash plants on the Mesabi Range to recover fine iron which previously had been discarded. About the same time sand mining plants in Florida incorporated spirals to recover titanium minerals and zircon from lowgrade beach sands. The Labrador Trough opened new opportunities for spirals when hard-rock iron-ore deposits were developed; approximately 10 000 Model 24A Humphreys spirals were installed between 1958 and 1966. About 5000 spirals of this model are still operating in the same capacity, in spite of the fact that extensive improvements have since been made in gravity separation technology."

Jan 1, 1993

By Todd S. Harris

"IntroductionThe western region of North America has long been considered a dynamic market supporting such diversified industries as lumber, agriculture, cement, and electronics. In recent years however, the region has also been the focus of a great deal of activity in the extender and filler mineral s business from companies seeking to position themselves to take advantage of the large demand for these materials both here and abroad.Extender and Filler MineralsExtender and filler mineral s include 15different products used in numerous end uses. The nine most important extenders and fillers are listed in Table I. These products range in price from under $50 a ton to over $4,000 a ton and basically represent the following three different types of minerals:Particulate products-Calcium carbonate; Kaolin; Silica.Platy minerals-Talc (Montana, Canada); Mica.Fibrous minerals-Talc (California, Eastern U.S .A.);Wollastonite.What distinguishes these products from other industrial minerals? For one thing, an important feature of extender and filler minerals is that they are used primarily in such consumer non-durable good s as paper, paint, plastics, sealants, and rubber which traditionally do not react as violently to swings in the economic cycle. In addition, these minerals are sold principally on key physical properties, usually brightness, particle size and shape, oil absorption, and hardness. As part of the physical minerals segment of industrial minerals, extenders and fillers are traditionally the most stable group and typically the most profitable. Studies by C.H. Kline over the past decade on extender and filler minerals has revealed historic average profitability (expressed as a percentage on sales) nearly double that of most other industrial mineral s.The demand for these minerals in the western region (Fig. 1), closely follow s the areas of greatest population concentration. These markets include pulp and paper, mainly in the Pacific Northwest, and paint, around Los Angeles and San Fran cisco. Other significant markets for extender and filler minerals in western North America include joint cements, carpet backing, and plastics, mainly PVC pipe."

Jan 1, 1986

By R. R. Potter

"IntroductionAlthough iron formations were recognized in western New Brunswick about 1820, commercial production was not achieved until 1848. The Jacksonville deposits near Wood stock were developed first, and ore was treated at a blast furnace constructed 5 km to the southeast on the Saint John River. Several charcoal burners were also constructed. Pig iron was produced sporadically until 1884, when operations were finally suspended. There was little further interest in the area until 1953, when Stratmat Ltd. initiated an ambitious exploration program. At the same time, the company carried out research on an electric-reduction process to produce metallic iron and ferromanganese. The technical feasibility of this process (Udy process) was confirmed, however the high energy costs were prohibitive and work ceased in 1957. Some metallurgical research was carried out by Mandate Refining Ltd. about 1970 and more recently by the New Brunswick Research and Productivity Council.The Years Prior to 1848New Brunswick became a separate province in 1784, shortly after the influx of a large number of Loyalists from the United States. Man y of these people sett led along the Saint John River, including the area of excellent farmland near the present town of Woodstock.There were indication s that some of these earl y settlers recognized iron formations in the vicinity of Woodstock about 1820, however the nature of the deposits was not known until 836 when Dr. C.T. Jackson, a geologist commissioned by the Stat e of Maine, investigated the mineral potential of the area. Jackson (1836) described the ""ore"" as compact red hematite yielding 44 per cent pure metallic iron, and suggested that sizable tonnages may be present. It was about this time that an extensive limonite deposit was found in north-central Maine, near Mt. Katahdin. Other iron deposits were found in the Nictaux area of Nova Scotia, and a few years later at Londonderry."

Jan 1, 1983

By A. C. Bonnier

"INTRODUCTIONI have been asked to make available our method of determining the ""Live Inventory"" within a gravity type thickener. The ""Live Inventory"" being that tonnage compatible with the optimum underflow density that satisfies the process requirements.One prerequisite of the calculations is the proper determination of the final pumpable underflow density. A further requirement is that, prior to shutdown, the thickener has been operated properly, particularly that it has not been operated for storage capacity.Until recently I had few chances to check the method. However, I was most fortunate to have had a recent opportunity covering several months on a 300 ft diameter thickener at a plant in the U.S.A., wherein the method was tested and verified many times. The method proved accurate in that our calculated tonnage checked with the actual dried physical inventory of the drying facility.At the precise moment that the feed to the thickener was shut down, a per cent solid determination was made at the critical place and the tonnage in ""Live Inventory"" determined. Let us say that at the moment the thickener U' Flow solids was 20%. The thickener then was allowed to cperate until the U' Flow solids dropped below 20%, at which time the thickener was deemed to be relieved. Anything much less than 20% was not suitable to the subsequent centrifuge operation.Most large metallurgical thickeners have a configuration of three slopes at the thickener ""bed"". These slopes are dictated by the raking blades and are shown on the manufacturers' erection drawings. The first slope is from the outer periphery and generally takes up a distance of about 60% of the radius, the second slope generally taking up the remaining 40%. Within this 40% area, there is usually another slope taken up by the cone on ""B"" type thickeners or an actual steeper inner slope on ""C"" type thickeners."

Jan 1, 1989

By Charle Lawrence, Murray E. Blanchard

"A significant addition to Canada's commodity export transportation capabilities is being built by Ridley Terminals Inc. on Ridley Island in the Port of Prince Rupert on British Columbia's north coast.Site preparation began in early 1982. By the end of 1983 the terminal will be receiving coal from two major new mines being developed in B. C.'s massive $2.5 billion northeast coal development.The RTI terminal will have initial capacity to transfer 12 million tonnes of coal a year from railcars to storage and to ships. Capacity will be doubled to 24 million tonnes as market opportunities are developed.The challenge of building this $220-million facility in less than 24 months, the benefits Canada has received from the work the project is generating, and the long-term faith in Canada that this project represents, are all part of the excitement which surrounds the Ridley Terminal and northeast coal developments.RIDLEY TERMINALS INC. (RTI) is building a very significant addition to Canada's commodity export transportation capabilities. The facility, on Ridley Island at Prince Rupert, on the north coast of British Columbia, is a key link in British Columbia's northeast coal development which in itself is a massive undertaking involving investment of some $2.5 billion. The entire project is making a substantial contribution to Canada's economy at a time when such contributions are sorely needed.The development is creating employment, not only in British Columbia where its various parts are being created by mining companies, railways and others, but also across the country where suppliers are manufacturing a wide variety of good s and equipment for the development's participants.The RTI terminal is no exception. It is a critical cog in this mega project's large transportation system and its construction is contributing substantially to job creation and to other economic activity, in the Prince Rupert area and in many other parts of British Columbia and the rest of Canada."

Jan 1, 1983

By Sigmund J. Andersen, Eva A. Mørtsell, Jostein Røyset, Randi Holmestad, Oddvin Reiso, Yanjun Li, Øyvind Paulsen, Calin D. Marioara

"With the purpose of increasing the extrudability of lean alloys, we reduce levels of the alloying elements Mg and Si in a 6xxx extrusion alloy. To avoid sacrificing strength, we investigate the compensating effect of smaller amounts of lithium and germanium, alone and in combination. Recent studies by atomically resolved high angle annular dark-field scanning transmission electron microscopy (HAADF-STEM) together with density functional theory (DFT) calculations have demonstrated that while Li replaces Mg, Ge substitutes Si in the columns in the hardening ß’’-phase. Several studies have shown that Ge replaces Si in other precipitates as well, and that both Li and Ge promote precipitation and can improve macroscopic properties, like strength and thermal stability. The scope of the current work has been to investigate and document the combined effect of the two alloying elements. Of the investigated alloys, the one containing both elements displayed highest hardness, corresponding with a significant increase in precipitate number density. We used HAADF-STEM to determine the extent of Li and Ge in the precipitate structures.INTRODUCTION The two most important alloying agents, with respect to age hardening of the 6xxx alloy series, are Mg and Si. In general, 6xxx alloys are formed into their final shapes by rolling or extrusion, before being age-hardened at elevated temperatures around 180°C. The hardening effect during this stage is due to the formation of metastable, semi-coherent, precipitate needles along the three < 001>Al directions generally following the precipitation sequence (Marioara, Nordmark, Andersen, & Holmestad, 2006): Supersaturated solid solution ? atomic clusters ? Guinier-Preston-zones (pre-ß”) (Edwards, Stiller, Dunlop, & Couper, 1998) (Marioara, Andersen, Jansen, & Zandbergen, 2001) ? ß” (Andersen, et al., 1998) (Hasting, et al., 2009) ? ß’ (Vissers, Huis, Jansen, & H. W. Zandbergen, 2007), U1 (Andersen, Marioara, Vissers, Frøseth, & Zandbergen, 2007), U2 (Andersen, Marioara, Frøseth, Vissers, & Zandbergen, 2005), B’ (Vissers, Marioara, Andersen, & Holmestad, 2008) ? ß, Si (stable)"

Jan 1, 2018

By Ian Sherrell, Michael Pirie, Riddhika Jain

"Conventionally, direct reduced iron (DRI) mixed with char from rotary furnace discharge is cooled down before separating the char via magnetic separator. The DRI must then be reheated for further downstream processing (i.e. smelting). Outotec has developed the PyroMagTM specifically to conserve thermal energy during this separation process. DRI is cleaned by feeding hot particles directly after reduction onto a rotating surface just above a stationary magnet assembly. This allows hot feeding of the clean DRI to the downstream equipment which renders the entire process more energy efficient.Continuous testing of the PyroMagTM was performed to verify design and separation. Because the use of DRI/char was not practical for test use, the pilot plant testing was conducted using a magnetite/sand mixture heated up to and beyond the Curie temperature of magnetite. The separation was efficient for both cold and hot processing with recoveries varying from 98.6% to 99.74%. As expected, the recoveries decreased to as low as 83.99% as the temperature was increased beyond the curie temperature of magnetite (585°C). Long term separation was thermally efficient and sustainable.INTRODUCTIONMagnetic drum separators are widely used in the mineral processing industry. These low to high intensity magnetic separators are used to concentrate various magnetic material, such as iron ore, and less magnetic materials, such as ilmenite and garnet from silica and other contaminantsThere are multiple classes of magnetism based on the effect that each material displays while in and out of a magnetic field. In a simplified fashion, which is more appropriate for mineral processing, the magnetic susceptibility of a mineral is used to define it as being ferromagnetic, paramagnetic or nonmagnetic. In this respect, minerals with varying degree of magnetic susceptibility can be selectively separated with different types of magnetic separation equipment in different stages of a process flow sheet. The lower strength magnets are employed early in the process to remove highly magnetic fractions. Later stages utilize greater magnetic fields for capture of less magnetically susceptible material."

Jan 1, 2016

By Wayne D. Goodfellow, William M. Luff, Steven J. Juras

Brunswick No. 12 is a vent-proximal massive sulfide deposit that formed above a feeder pipe with associated hydrothermal alteration. Because of the replacement of bedded sulfides deposited above the hydrothermal vent, individual sulfide lenses show the following vertical and lateral zoning outward from the vent complex: massive pyrite-pyrrhotite-ehalcopyrite core interlayered pyrite-sphalerite-galena (pb-Zn ore zone) - bedded pyrite. The feeder zone which underlies the Cu-rich core consists of an anastomosing network of pyrite-pyrrhotite-chalcopyritequartz veins with variable and generally minor contents of ferroan carbonate, Fe-rich chlorite, sphalerite, galena and arsenopyrite. Hydrothermal alteration of felsic igneous rocks hosting and adjacent to the feeder pipe is represented by an inner zone (zone I) characterized by sulfide stringers, Fe-rich chlorite, muscovite, quartz, and ferroan carbonate that formed at the expense of feldspars and what was probably glass; outer zones (II and III) are represented by Mg-rich chlorite and muscovite formation, and the albitization and silicification of K-feldspar. Accompanying these alteration trends is a marked decrease of Na + Ca, and an increase in Fe + Mg + C03 in zone I, and increases in Mg + Na in the outer zones. The classic zoning of metals within the sulfide lenses, the formation of a feeder pipe below the vent-proximal sulfide core, and the hydrothermal alteration of felsic volcanic rocks cut by the feeder pipe all support a vent-proximal origin for the Brunswick No. 12 deposit. Furthermore, restriction of sulfide stringer mineralization to the footwall sequence, the folding of the stringer zone by an F, fold, and the difficulty of long distance mass transfer of metals from the sulfides into the host rocks all argue against a tectonic origin for the stringer zone. All of the evidence indicates that the feeder pipe formed during the ore forming hydrothermal event, but has been distorted and in some areas detached from the overlying massive sulfides during subsequent deformation.

Jan 1, 1992

By C. M. Acuna, M. Sherrington

In high temperature processes thermodynamic equilibrium is assumed and therefore chemical potential of elements, among the condensed phases, have to be the same. In such situation impurities distribute accordingly and their fractional distributions are affected by the masses of the coexisting phases. It has been claimed Teniente Converters perform quite well for smelting high impurity concentrates, especially in respect to volatilization and/or slagging. Production of high matte grade near white metal composition, requiring moderate oxygen enrichment in the process air, results in a huge mass of gas and therefore a high fraction of impurities report to it. Similarly, by recycling different sources of slag, flue dust or other materials, the mass of the final slag is increased and so do the masses of impurities reporting to the slag. Nevertheless, above a certain level the recycling of smelter byproducts is no longer feasible and the cycle has to be halted, unless higher impurity mattes are accepted. Although the process might be continued during converting and refining, lengthy and expensive measures have to be taken if suitable anode copper for electrolytic refining is to be achieved. In the present study samples of high arsenic matte grade and slag were obtained on a pilot scale. The matte samples, on the order 75%Cu, were equilibrated with slag at varying ratios in magnesia crucibles under CO/C02 atmospheres at 1523 K. After 24 hours equilibrium matte and slag were rapidly solidified and analyzed by copper, sulfur, iron and arsenic. Furthermore, industrial samples of matte and slag, in the order 75%Cu at 1523 K, were obtained via a special quenching sampling device to determine the association of copper and suspension of matte and copper in slag. Based on the results, the distribution coefficient of arsenic between slag and matte was evaluated along with copper species in slag and the fractional distribution of arsenic, in the coexisting phases, was determined as a function of the ratio mass of slag to mass of matte.

Jan 1, 2007

By W. G. Hatch

Introduction Classified equipment originally installed in 1923 in the Britannia mill consisted of eighteen Model C Dorr Simplex classifiers, each in closed circuit with a pebble-mill. These machines were only three feet wide and of light construction, so that their raking capacities were necessarily limited, the maximum being about 20 tons per hour for each classifier. On conversion from pebble-mills to ball-mills, in 1924, attempts to increase circulating loads with higher rake speeds resulted only in overflowing a coarser product to the flotation cells. The elevator type of classifier was developed locally in 1935 and during that year two units, each having a sand carrying capacity of from 30 to 40 tons per hour, were installed. With such loads, a material increase in the grinding capacities of the ball-mills was noticed, and, after further development work on the elevator classifier, fifteen of the small Dorr classifiers were replaced by units of this type, capable of handling approximately 75 tons of sand per hour, with later alterations increasing this capacity to well over 100 tons per hour. In the meantime, the three remaining ball-mills had been placed in circuit with Dorr FX classifiers, one of which was six feet wide and the others five feet wide. Impartial and exhaustive tests conducted on the Dorr FX versus the elevator classifier showed the average results on these machines at Britannia to be substantially the same. A large amount of testing was carried on to determine the effect of circulating load on the output of the grinding units. This work indicated a rapid increase in the grinding rate as the ball-mill feed increased up to about 60 tons per hour, after which a relatively slow increase took place up to 100 tons per hour. No benefits were derived from higher circulating loads. A 300 to 500 per cent circulating load, depending on the nature of the ore treated, is considered to be within the economical range for the grinding of Britannia ores.

Jan 1, 1946