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By Ronald E. Hamilton

"An attempt to review the recent literature on comminution reveals at once that the subject has many ramifications and that the applications are so diverse that there are very few general statements which can be made.It. seems reasonable to start classification on the basis of size of material as Input to the comminution process. However, we find that this can only be done in a very imprecise manner.The initial step of breaking rock from its natural state of very large pieces in geological formations into sizes suitable for processing is accomplished by blasting. This step in the process has not been ex-amined in this project except to note that great changes have been made during the last decade or so in the explosives used, and in the method of handling them. We should determine what effect changes in mining methods can have on the need for crushing.The second step in the process is usually primary crushing, performed by gyratory or jaw crushers. The largest chunk which can be fed to a jaw crusher depends on the size of the machine, and should be larger than the largest fragment produced in the miring process. It is obvious that the larger the crusher, the less care required in mining to en¬sure the absence of rocks too large to be crushed. The optimum size of crusher to minimize overall cost of mining and reduction is not at all obvious, and no information on this point has come to light in this investigation.It is rather generally accepted that size reduction is most efficiently done in small stages, and that different machines have different opti¬mum size ranges of feed. For example, crushers are preferred for sizes above about 1"", rod mills for the range between 1"" and 0.1"", ball mills are preferred from 0.1""down to perhaps 10 microns, and vibratory mills, or jet mills, below 10 microns."

Jan 1, 1976

By J. E. Nesset, F. Rosenblum, J. A. Finch, R. Payant

ABSTRACT Under certain conditions of moisture and oxygen, sulphides can spontaneously self-heat. We tested a hypothesis that galvanic interaction between some sulphides can promote self-heating. Galvanic interaction is controlled by the rest potential difference between sulphides and particle size. A series of binary sulphide mixtures was tested. Individual sulphides and mixtures with low rest potential difference did not self-heat, but mixtures of high rest potential difference did. When testing pyrite-sphalerite mixtures, self-heating increased inversely with particle size, with the fineness of the pyrite governing self-heating, indicating the rate-limiting step is the reduction reaction. The results support the hypothesis that galvanic interaction contributes to sulphide self-heating. A possible mechanism based on hydrogen sulphide as an intermediate product is proposed. The understanding gained will be of interest to those involved in storage, shipping, and disposal of sulphides.

Jan 1, 2012

Project Located in northwestern British Columbia 70km West of Cassiar Highway Galore Creek Project and Road Access Corridors Designated For Mineral Development in Stikine-Cassiar Land Resource Management Plan (LRMP)Approved in 2000Strong BC Government, local community and Tahltan First Nation Support for Project

May 1, 2006

By C. Schuffeneger

"The design of SAG milling circuits for mining projects requires the incorporation of the energy consumption considerations and the evaluation between different flexibility-adding options to absorb the time-dependent variation in the ore lithology. In order to expand the focus of the engineering design, a higher level of deepness and quality in the baseline information must be achieved to elucidate all of the constraints on throughput, energy consumption and product quality generating a design optimization consistent with an integrated vision of the process. This paper presents a review of a collection of typical examples of the copper and gold mining project in the Chilean Andes, in which both process and energy efficiency criteria are considered when the focus is to deliver an integral solution that incorporate those areas in the design. The intensive use of energy in the SAG milling, the process continuity and the quality of the product are a function of the proper design and selection of grinding equipment and plant layout, aligned with the mining plan and associated with projected future expansions and production increases.INTRODUCTION Autogenous and Semiautogenous grinding technology, large diameter and short length mills, begins with the Handsel design. During the decade of 1930’s, Handsel-Hardinge studies led to cascade mill development. Nowadays, 80 years later, continuous improvements in design and technology have been conducted to a completely new generation of gearless SAG mills that have led to an increasing capacity at the concentrator plants. Considering the existing alternatives, SAG mills are the most implemented option for the grinding circuit today. This work presents a review of the SAG mill implementation parameters and circuit considerations and its correlations in the most recent projects in Chile with an ore treatment capacity above 50 ktpd since the decade of the 90’s up to now. From this period, the tendencies will be analyzed using project design data. Based on process design information, the recent SAG mill history in relevant projects related to the new generation of commercial SAG mills, technical associated considerations and performance will be discussed. Circuit selection based on energy utilization will be revised to bring up a more holistic perspective on the SAG mill circuit design concepts. As is now well known, the heart of every concentrator plant are the grinding mills. Higher demands on throughput, availability and power consumption optimizations require not only high performance electrical equipment but also a well selected grinding circuit. The energy efficiency issues associated to the circuits design will be reviewed."

Jan 1, 2012

By D. J. Thorkelson

New data indicate Wernecke Breccia-associated iron oxide copper-gold (IOCG) deposits likely formed from moderate-temperature, high-salinity, non-magmatic brines. The breccias formed in an area underlain by a sedimentary sequence that locally contained evaporites (potential source of chloride and possibly sulfur) and was thick enough to produce elevated fluid temperatures. Metals (Fe, Cu, Co, U) were probably derived from host strata, transported as chloride complexes, and precipitated due to changes in fluid temperature and pressure during brecciation. These new data suggest that the spectrum of genetic models for IOCG deposits that typically invoke formation from magmatic or hybrid magmatic-non-magmatic fluids should be expanded to include those systems that formed in a non-magmatic environment. Modifications to the definition of IOCG systems are proposed that reflect the degree of involvement of magmatic and/or non-magmatic fluids and the nature of the mineralizing environment. A division into magmatic, non-magmatic, and hybrid magmatic-non-magmatic IOCG types is suggested. Typical magmatic end-member IOCG deposits include Lightning Creek and Eloise, Australia. Hybrid magmatic-nonmagmatic IOCG examples include Ernest Henry and Olympic Dam, Australia. The Wernecke Breccia and Redbank deposits are suggested as possible representatives of non-magmatic IOCG end members. Endmember magmatic IOCG deposits have similarities to some porphyry deposits, whereas non-magmatic IOCG end members share characteristic with some sediment-hosted Cu deposits, suggesting that the range of IOCG deposits may form a link between intrusive- and sediment-related deposits

Jan 1, 2007

By P. Wells

"In the past decade, there have been many developments in processing of nickel/copper sulfide ores. Improved pentlandite/pyrrhotite flotation selectivity with the use of new reagents, fine regrinding and Redox measurement (control) has led to increased pyrrhotite rejection and thus higher concentrate grades. Flash flotation has increased overall nickel and associated precious group metal recoveries. Recovery of both fine and coarse particles has been improved through an understanding of flotation cell hydrodynamics. The separation of copper from nickel in such ores has seen a similar level of development. From the late 80s to very recently, the use of flotation columns, the use of pyrrhotite depressing reagents and the improved understanding of the complex mineralogy of chalcopyrite, pentlandite and pyrrhotite have considerably increased copper / nickel separation efficiency. Copper concentrate with more than 30% copper and less than 0.5% nickel grade is now obtainable at 80-85% copper recovery from mill feed compared to 70-75% copper recovery with 30% copper and 1% nickel grade only a few years ago. Improved copper / nickel separation in milling operations has the potential to substantially improve the processing of the resulting concentrates and thus the overall return from the extraction of nickel and copper from sulfide ores.INTRODUCTIONNickel metal by itself or in alloys with other metals and materials makes a significant contribution to modern society. The relatively high price that nickel commands compared to other base metals is a reflection of the fact that, although abundant in the universe, nickel is sparingly distributed in the earth’s crust. There are two types of nickel deposits that are currently being exploited: complex sulfide ores and lateritic oxide ores. Lateritic contain no copper but in sulfide deposits, nickel– bearing minerals, principally pentlandite (Pn, [Ni,Fe]8S9), some millerite (NiS) and violarite ([Ni,Fe]S) are often found in association with copper-bearing minerals, mainly chalcopyrite (Cp, CuFeS2) and cubanite (CuFe2S3). The copper content of such ores can vary considerably from as high as 5-8% (in some Sudbury basin ores) to as low as 0.1% (in ores mined by Western Mining at Mount Keith, Western Australia and by Inco in Thompson, Manitoba). The valuable minerals Pn and Cp are almost invariably found associated with pyrrhotite (Po, Fe7S8) which along with nonsulphide rock has to be rejected to tails in the primary milling stage. The bulk concentrate from the primary stage is then treated to depress Pn and float Cp. A complete Cu/Ni separation at the mill would be highly desirable but this is not possible because of locking between a small fraction of the valuable minerals. As a result, Cu/Ni separation has to be carried out both during milling and during refining of the matte following smelting. At Inco, Jinchuan and in Norilsk, a chalcocite / heazlewoodite separation is made by flotation following slow cooling of the matte. In the present Sudbury Inco operations, Cu/Ni separation has not been conducted at the mill since 1991 and all Cu/Ni separation takes place at the matte processing plant. Because of the rising ratio of Cu to Ni in the ore, this strategy is under review and this paper describes some recent studies of circuits in preparation for possible installation of a copper / nickel separation circuit at Inco’s Clarabelle Mill in the Sudbury area."

Jan 1, 2004

By J. Ling, L. Racine, P. Cousin, A. R. Laplante

"The problem of gold recovery by gravity from high density gangues is first discussed from a fundamental perspective. Three potential difficulties are presented: poor liberation and fine gold particle size, concentrate bed erosion by coarse dense particles, and slow percolation of gold particle through the high density gangue bed. Two case studies are presented to illustrate the impact of these three problems on circuit performance and the feasibility of gravity recovery. It is concluded that careful testing can determine the scope of the first problem and feed preparation can limit the extent of the second. The third problem is clearly the most serious and results in lower gravity recoverable gold recoveries than would be achieved with the same recovery effort on a low density gangue application.IntroductionSince the commercialisation of the Knelson Concentrator in 1980, gold gravity recovery has experienced a surge in interest, in Canada and the United States first, then Australia, Zimbabwe and Russia, and finally worldwide. Table I details number of units sold in various countries, which corresponds approximately to their importance as gold producers 1. Exceptions. such as the low rank of South Africa or the importance of Zimbabwe, despite the much higher gold throughput of the former, are generally the result of differences in the effectiveness of the local marketing strategies, as well as a reflection of the ""philosophy of recovery"" prevalent in the country.The impact of the Knelson has not been felt equally in all plants. Typically, a highly successful application falls into the following categories: (a) alluvial or hard rock recovery for which gravity is the sole recovery method, (b) hard-rock free gold milling with ores which have a very high gravity-recoverable gold (GRG) content, (c) low grade sulphide copper-gold ores which include flotation ahead of cyanidation and (d) free-milling gold ores ahead of Other applications can be successful, but are not as common. The ""explosion"" of centrifuge concentration, of which the Knelson is the most successful but by no means the sole model, has created a certain euphoria which can be summarized as ""gravity recovery will improve the profitability of gold milling, no matter what."" This presentation is about applications that are not so successful, or even downright unsuccessful. The rationale is not to down-play gold gravity recovery, but to set it in its rightful position: metallurgically, it is not appropriate all of the times; it can become inappropriate if poorly designed; and it can be metallurgically possible but economically unjustifiable. In this paper, this will be illustrated and explained rather than demonstrated, with the use of both industrial plant surveys and laboratory test work. Alternate gravity approaches will be suggested."

Jan 1, 1997

By Jean Rosaire

?GWIOP Location ?Climate and Physiography ?History ?Geology ?Deposit Types ?Mineralization ?Historical Mineral Resources ?Conclusions and Work Program

May 1, 2007

By Turgut Yalcin

"Recent years have seen a resurgence of interest in flotation cell design, and one of the areas where attention is being focused is flotation in centrifugal fields which aims at increasing the flotation rates by forced bubble-particle contact.A new cycloflotation device was developed at Laurentian University and tested on the nickel ore of Falconbridge Limited in Sudbury. In comparison with conventional flotation, cycloflotation yielded at least seven times faster flotation rates in terms of the mass of solids floated per unit time per unit flotation volume, while producing concentrate grades similar to those of conventional flotation. These results, which were obtained with a bench-scale ""cyclofloat cell,"" suggest that cycloflotation has the potential of setting a new direction in flotation technology.IntroductionFroth flotation is no doubt the dominant separation technique in the mineral processing industry. Since its inception early in this century, much research effort has been and is still being expended on this process. Until recent years, emphasis seems to have been placed on the chemical aspects of flotation, but there is now an increased awareness for the need to improve flotation machine performance and more and more attention is being directed towards this area. In a 1982 article on ""Flotation Machines"" (Young 1982) it is stated that "" ... by now the design of the machines used would have reached a high level of perfection, ... In fact, this is not the case ... "".The process of flotation in a flotation device may be considered to take place in several distinct stages as follows:"

Jan 1, 1994

By R. J. C. MacDonald

"The first recorded discovery of placer gold in Canada was on the Chaudiere River in Quebec's eastern townships in 1832 and the first major production came from placer deposits in British Columbia in 1858 (Verity 1974). Placer deposits were the major source of gold until the late 1890's, although lode deposits were found in Nova Scotia in 1862. In 1896 the Klondike Gold Rush commenced with the discovery of gold by Henderson and Carmack in the Yukon. The peak was reached in 1900 when over one million ounces of gold valued at approximately $22.3 million were recovered. With the exhaustion of the easily accessible deposits, more capital intensive operations were necessary such as those initiated in 1906 by the Yukon Gold Corporation. This giant dredging operation continued to produce until 1966. Today only a few small operations remain.The development of Canadian lode gold mining began in earnest around 1900 when the railroads were opening up the interior and western part of the country. At this time many discoveries occurred, particularly where mineral outcroppings were near surface. Tremendous impetus for exploration and the reopening of former producers followed the elimination of the gold standard by Great Britain around 1931, and more importantly the United States in 1933, with subsequent stabilization of the price of gold at $35 per ounce in 1934. Prior to this the price of gold was basically unchanged since 1717, approximately, $20 per ounce (Verity, 1974; Green, 1973).The depression of the 1930s adversely affected most industries, however for the gold producers it was a very profitable time. Prices of materials and equipment declined and labour was cheap and plentiful. Interest in gold mining was further stimulated in 1936 when a three year tax exemption was declared for all new metal mines. In 1938 a decline in the value of the Canadian dollar relative to the U.S. dollar benefitted the industry. The second world war provided yet again further incentive for gold production and in 1941 Canadian gold production reached a high of approximately 5.4 million ounces. This figure was not surpassed until 1991 when approximately 5.6 million ounces were produced (Natural Resources Canada, 1996 )."

Jan 1, 1998

By D. Lindeman, R. M. Nendick

"From a technologist's point of view, there have been few fundamental 'breakthroughs' in the process of gold extraction since the turn of the century. However, the industry has made a great deal of progress in its response to the problems that have arisen: a changing demand for products, lower grade and more refractory ores, inflated labour and materials cost and the increasing control of governmental legislation. The trend in North America and South Africa to alleviate these problems has been to adopt a common policy to process larger tonnages of lowgrade ore through more efficient milling circuits. Here, however, is where the similarity ends.The purpose of this paper is to emphasise the different principles of practice employed by extraction plants in North America and South Africa to achieve the same common goal, namely gold bullion.2.0 INTRODUCTIONTwo very important aspects of extraction metallurgy for which the metallurgical engineer is responsible are plant design and plant commissioning.In North America and South Africa the importance of these aspects is due to the fact that both are related to the financial success of the project of which they form part. Good design is imperative if the capital invested in both the plant and the whole mining undertaking is to reap maximum profit. Furthermore, good design leads to a carefully planned and well executed start-up procedure resulting in bringing the plant into production on time, at designed efficiency and within budget.Due to the competent nature of the quartz pebble conglomerate and quartzite ore on the South African Witwatersrand, the characteristics of the ore type are not much different from one mine to the other. Thus a significant operating knowledge and experience has been built up over the years and there is considerably less risk in selecting a gold extraction process route."

Jan 1, 1992

By R. Biss

"Niobium or columbium, as usually called in North America is becoming a more and more important metal in the industrial world because of its attractive properties.Eventhough discovered in the early nineteenth century, it was not until the 1930's that its use became more widespread. The discovery of large commercial deposits in the late 1950's and the demand of industry for steel with high strength and toughness further promoted its use. To assure its adequate supply new or-e processinr technologies had to be developed. So, beside the traditional gravity concentration, the flotation process was employed for the partial upgrading of pyrochlore concentrate at S8ve (Norway) plant in 1953. The flotation process was equally used at St-Lawrence Columbium plant in the province of Quebec, Canada. The new plants at Araxa and Catalao in Brazil and at St-Honore, Canada, which supply almost of all niobium used today in the non-Communist world, also adopted the flotation process. The processes using long chain aliphatic rnonoarnines, diamines or sulfonates as collectors proved to be difficult especially on weathered types of ores or on those having a substantial content of silicate gangue minerals. They almost completely failed in case of alkaline intrusive rocks as syenites, nepheline syenite, ijolites and pyroxenites, composed almost entirely of silicate minerals. A lot of effort was devoted all around the world on research of new, more efficient flotation processes. So in 1982, Les Services T.M.G. Inc. (Niobec), sole producer of pyrochlore concentrate in North America, developed the direct pyrochlore flotation process from carbonatites. In 1983, at about the same time, the Quebec Department of Natural Resources Laboratories and Les Services T.M.G. Inc. announced a successful recovery of niobium Minera1s from a low grade nenheline syenite ore."

Jan 1, 1984

A series of fault slip seismic events have occurred at Garson mine. The events have occurred during development advancement exposing faults, after the extraction of single stopes due to mining induced stress change, and due to time effects after delays of many days during which no mining had been conducted. This paper discusses the study completed consisting of a global or holistic approach aimed at mitigating some of the risks associated with the structurally controlled rockbursts at Garson. Background Garson mine is located in the south east area of the Sudbury Basin (Ontario, Canada). Mining occurs in two main ore bodies (the #1 Shear and the #4 Shear) which are copper-nickel-sulphide and occur within shear zones that (as well as the ore bodies) strike approximately east-west and dip steeply south. From the north (footwall), lithologies progress (Figure 1a) from norite (NR), to greenstone (GS), and finally to meta-sediment (MTSD). A generally WNW-ESE striking, bifurcated sub-vertical dyke of olivine diabase (OLDI) cuts the series. The current active mining block at Garson is between 4700L and 5100L (levels denote depth in feet below surface).

May 1, 2009

"Air Products approached LAC Minerals in 1987 to initiate trials with pure oxygen in the leach circuit at several LAC facilities in Canada. This paper gives an overview of oxygen use and covers specific operating data from the Doyon and Terrains Aurifères mills near Val D'Or in northwestern Quebec. LAC also uses oxygen at the Macassa mill in Kirkland Lake, Ontario.In addition to the LAC facilities, oxygen is currently in use for test purposes at the Holt McDermott mine in Kirkland Lake, Ontario and the Camflo mine near Val D'Or.There are two requirements for oxygen in the cyanide leach mill. Firstly, oxygen is required in the familiar gold dissolution reaction as follows:(1) 4 AU (s) + 8 NaCN (aq) + 02 (aq)+ 2H 20(aq) - 4NaAu(CN) 2(aq)+ 4Na OH(aq)Of the reagents on the left side of the equation, the availability of gold is fixed by the ore in question. Obviously, water is available in excess, 1eaving cyanide and oxygen as the limiting constituents in the process. Most mills operate with an excess of cyanide, which is fairly closely monitored. Dissolved oxygen levels, however, are not normally watched very closely and find their own level based on the temperature, the availability of compressed air and the ability of the ore to consume oxygen.Typical D.O. levels are in the 3-8 ppm range, however levels less than 1 ppm are not uncommon for refractory ores.With pure oxygen, dissolved oxygen levels as high as 30 ppm can be achieved. (Figure #1). This greatly accelerates the kinetics of the leach reaction and may also shift the reaction further towards completion thus increasing gold dissolution slightly."

Jan 1, 1990

By Joe Holmes, Peter Langlois

"Falconbridge’s Raglan Concentrator, located at the northern limit of Quebec’s Nunavik Region, has been in operation since December 1997. Originally designed to process 800,000 tonnes per year of high grade nickel ore, plant capacity is currently approaching 1,000,000 tonnes per year. Optimization of the ABC grinding circuit has proven to be the most challenging component of this capacity increase. Concurrent flotation circuit modifications have allowed concentrate grades to increase from 16 to 18 percent nickel, while maintaining recoveries. Concentrate drying problems encountered during the first 15 months of operation have been addressed through equipment upgrades. The challenges associated with surface deposition of filtered tailings in an arctic environment are now better understood, and detailed design of the progressive reclamation cover is underway.GRINDINGCircuit OverviewOre from open pit and underground mining operations is introduced to the underground ore handling system via surface and underground truck dumps. Two underground bins with a combined capacity of 4000 tonnes provide roughly a day of live capacity for ore storage. Ore from the two bins is conveyed to a 150 tonne surge bin immediately ahead of the grinding circuit in the concentrator.The Raglan grinding circuit has a conventional ABC configuration (Figure 1) with a design capacity of 100 mtph (800,000 tonnes per year). The 24’ AG mill is powered by a 3000 hp motor with variable speed drive. AG discharge is sized on a 8’x16’ single deck vibrating screen, with screen oversize recirculating to a 5 %’ short head cone crusher (400 hp) before returning to the AG mill. Screen undersize is combined with ball mill discharge to feed up to 5 15” Krebs hydrocyclones. The ball mill is a 14’x 21’ overflow mill driven by a 3000 hp motor at 73% critical speed. Target Pgo for the grinding circuit product is 65 microns"

Jan 1, 2001

By P. J. Mehrfert

"Recovery of mineralized coarse particles is notoriously low in conventional concentrators that treat copper porphyry ores. The application of coarse particle flotation techniques has the potential to achieve comparable metallurgical performance at significantly coarser primary grind sizings. ALS Metallurgy has been conducting both internal research and commercial test programs to investigate the potential of Eriez HydroFloatTM technology on copper sulphide ores. Test data indicates that significant improvements in the recovery of copper sulphide bearing particles between 200µm to 850µm can be achieved using Hydrofloat technology compared to conventional flotation techniques. This paper summarizes the results from a selection of test programs that includes recovery by size data from both conventional flotation and HydroFloatTM tests. An explanation of the apparatus and test procedure will be given. Mineralogical analyses of the test products and characteristics that limit metallurgical performance will be discussed.Description of Apparatus and Test ProtocolThe HydroFloatTM laboratory cell is a clear plastic column with an internal diameter of 14 cm. Water and air are introduced to the cell through a circular manifold located in the lower portion of the cell. The cell has a conical bottom which directs slurry to a pinch valve that controls the discharge of tailings. Feed solids are introduced to the upper portion of the cell in slurry form. The cell is operated in a manner such that the manifold water creates a teeter bed zone of suspended solids in the middle portion of the cell. Tailings exit the bottom of the cell at a high pulp density, and most of the introduced water overflows the top of the column and is collected in a launder. Particles with a hydrophobic surface may attach to air bubbles that percolate through the teeter bed, rendering them sufficiently buoyant to be transported to the overflow launder.Three significant features of this arrangement enhance the potential to recover coarse, low grade particles by flotation techniques. Firstly, the hindered settling conditions created by the rising column of water allow coarse particles to become buoyant once attached to a modest sized air bubble. Secondly, the flow conditions within the cell are relatively quiescent compared to a conventional flotation cell that utilizes an agitator. This allows particles that have limited hydrophobic portions of their surface, due to a low content of exposed sulphide mineral, to maintain a somewhat fragile attachment to air bubbles long enough for successful recovery to the overflow. Finally, the absence of a froth phase at the top of the cell allows particles attached to air bubbles to immediately exit the vessel, reducing the potential to drop back into the pulp phase as bubbles burst or coalesce."

Jan 1, 2017

By Jörg Fischer-Bühner

Conventional 22ct Au jewellery alloys are comparably soft in the as-cast and soft-annealed state and hardening by cold-working is the only way of giving strength to the jewellery. The paper first focuses on hardening of 22ct jewellery alloys by conventional means, i.e. solid solution hardening with base metals (aiming particularly at an increased as-cast hardness) as well as precipitation hardening by alloying with base metals beyond the solubility limits and subsequent age-hardening heat treatment. Co-additions of ~2% turned out to result in most pronounced solid solution hardening up to ~100 HV1 in an investment cast & quenched state. Subsequent heat treatments, consisting of solution annealing at 850°C for 1h followed by water quenching and ageing at 300°C ? 400°C for ~ 1h lead to pronounced age-hardening with a peak hardness of ~ 260 HV1. In contrast, 22ct Au alloyed with Ni, Fe or Sb display no age-hardening effect within the same range of heat treatment parameters. The 2nd part of the paper reports on some preliminary trials to enhance the hardening effect of the base metals by additions of suitable elements, which possess a strong tendency to phase formation with the added base metal during ageing treatments. It is shown, that the age-hardening behavior of 22ct and 990/995 Au alloyed with Co or Ni can be remarkably influenced by additions of Sb. A total of only 10wt%0 of CoSb or NiSb-additions lead to a peak hardness of ~ 150 HV1 for a 22ct Au and ~ 100-120 HV1 for 990 Au , which is obtained already after very short ageing times (20-40min) at 300-400°C after solution annealing at 700°C / 1h / water quenching. It is suggested that the age-hardening mechanism in these alloys consists in precipitation of intermetallic CoSb- and NiSb-phases.

Oct 1, 2003

By E. V. Manlapig, K. R. Rahal, J-P Franzidis

"The Floatability Characterisation Test Rig or FCTR is a highly instrumented, mobile flotation pilot plant built by Baker Process (now EIMCO) for the AMIRA P9 Project for use in flotation plant modelling and simulation studies. With the Mintek Floatstar system controlling the level in every cell, and the facility for full stream sampling on all streams, the FCTR is a unique tool for producing data of sufficient quality for flotation plant modelling. The plant was commissioned at Lonmin's Karee and Eastern Platinum concentrators in 1999/2000, and completed a number of test work programs to investigate the effect of power input and regrinding on metallurgical performance. Lonmin's experience with the FCTR has caused it to build two of its own to undertake further projects on its various mines. This paper describes the design of the FCTR and the features that make it a unique research tool, and the benefits derived by Lonmin through their experience with the unit.INTRODUCTIONAs part of the Australian Mineral Industries Research Association (AMIRA) P9 project, a new model for flotation has been developed, in which the overall recovery, R, of the jth floatability class of the ith size class of a particular mineral, is given as (Savassi, 1998):"

Jan 1, 2002

By Euler De Souza

This paper presents a study conducted to assess sillmat behaviour during undercut mining and to provide a comprehensive understanding of the effects of arching and sillmat thickness on stability. The design study investigated the effects of such parameters as stope width and height, closure strains, orebody geometry and inclination, wall roughness, fill properties and cure time on the stability of cemented sillmat during undercut mining. Arching effects have been quantified based on centrifuge modeling studies for the different mining conditions. Observations have indicated that the degree of arching is a function of stope geometry, wall roughness, fill cohesion, and wall closure. Arching effects increased with decreasing stope width, increasing backfill cohesion and frictional resistance, increasing wall roughness and with wall closure.

May 1, 2002

By P. W. Godbehere

"INTRODUCTIONMagnetic mill linings have been developed from an invention by Mr. Bertil Brandt, the originator of the Skega rubber lining. The OREBED magnetic mill lining is a concept in which rubber covered strontium ferrite permanent magnets are arranged on the inside wall of a mill shell so that they attract magnetically susceptible material be it ore, media or tramp, to form a protective layer over the magnets and to serve as an operating surface with indefinite life.Noranda Mines Ltd. through its associate Norcast Division of Noranda Metal Industries formed a joint venture with the inventors to market magnetic mill linings in North America. For the past twelve months trials have been undertaken in conjunction with Orebed Inc. to test and develop these linings for grinding mill applications at the Horne Concentrator. Two independent milling circuits are currently being operated, one treating a siliceous gold ore and the other a copper slag.This paper presents the results of the magnetic lining test programme to date."

Jan 1, 1983