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In a gravity separation device, particle shape, size and density all play a role. The combination of these determines where each individual particle reports to in the viscous fluid, where particle crowding (solids to water ratio) also plays a role. To understand the performance of gravity separation devices in the heavy mineral industry, these particle characteristics need to be measured. There are various challenges in analysing particle density and particle size simultaneously for the purpose of quantifying gravity separator performance, not to mention particle shape. These analytical challenges include the high cost of high-density sink-float fractionation, toxicity of high-density sink-float media, inability of sink-float media to fractionate at densities greater than 4.0 g/cm3, and the time-intensive nature of these fractionations. The use of the detailed particle-by-particle output from Qemscan® particle mineral analyses (PMA) as a fast and cost-effective alternative is evaluated. The size and density outputs from the Qemscan® were employed to characterize the performance of a heavy mineral spiral concentrator as an example. Critical analytical requirements are to be addressed before the Qemscan® output data can be utilized. Keywords Gravity separator, performance, heavy minerals, Qemscan® particle mineral analysis

Jan 1, 2011

By J. Searra, M. Aboagye

"For new mining projects or for medium- to long-term areas of existing mines, drilling data is invariably on a relatively large grid. Direct estimates for selective mining units (SMUs), and also for much larger block units, will then be smoothed due to the information effects and the high error variance.The difficulty is that ultimately, during mining, selection will be done on the basis of SMUs on the final close-spaced data grid (grade control), which will then be available, i.e. the actual selection will be more efficient, with fewer misclassifications. However, this ultimate mining position is unknown at the project feasibility stage and therefore has to be estimated. This estimation is required because any cash flow calculations made on the basis of the smoothed estimates will obviously misrepresent the overall economic value of the project, i.e. the average grade of blocks above cut-off will be underestimated and the tonnage overestimated for cut-off grades below the average grade of the orebody. Similarly, unsmoothed estimates will be conditionally biased and will give even worse results, particularly in local areas of short- and medium-term planning or mining.This paper presents a case study of indirect post-processing and proportional modelling of recoverable resources designed for mediumand long-term mine planning at the Gold Fields’ Tarkwa Mine in Ghana. The case study compares long-term indirect recoverable mineral resource estimates based on typical widely spaced feasibility data to the corresponding production grade control model as well as the mine production. The paper also proposes certain critical regression slope and kriging efficiency control limits to avoid inefficient medium- to longterm recoverable estimates, and highlights the danger of accepting block estimates that have a slope of regression less than 0.5."

Jan 1, 2015

By A. A. Debrah

Artisanal and small-scale mining (ASM) plays a fundamental role in the national and local economies of resource-rich countries in Africa. As such, more effort must be geared towards fostering this type of economic activity to sustain livelihoods in rural communities. Across Africa, efforts to operationalize ASM based on a sustainable development framework have been quite tedious and often ineffective. This is because most of these frameworks lacked context and an understanding of the continuum from the artisanal through small-scale and up to the junior miner, grouped under the general term ?ASM?. Driven by the need to contextualize some of the definitional issues in this niche sector, and gain a clearer understanding of their peculiar challenges, this paper aims to contribute to the debate on ASM by considering two of Africa?s resource-rich countries ? Ghana and South Africa. Since South Africa has made greater inroads in protecting the interests of small-scale miners, we thus draw comparisons from both countries? approach to ASM, outline the lessons learnt, and provide recommendations regarding the ways forward.

Jan 1, 2014

By V. B. Teffo

The aim of the project was to experimentally determine the coefficient of restitution, and static and rolling friction of Eskom-grade coal for discrete element modelling (DEM). These parameters will be used in further DEM studies to simulate the flow and behaviour of coal during conveyance. For the simulation to yield realistic results, the parameters for the different coal shapes identified in the samples were determined. DEM uses spherical particles, and it was found that the difference between the DEM default parameters and those measured is significantly different, by as much as 97%, which illustrates the importance of experimentally determining these parameters. The spherical particles in DEM can be used to construct the various coal shapes identified, and the parameters measured can be used to simulate realistic results. The conveyance system can then be optimized accordingly.

Jan 1, 2013

By Rajendra Singh

Underground depillaring of contiguous coal seams/sections is a difficult problem, which becomes even more complex if the interburden/parting between them is thin. Conventionally, the depillaring with caving of contiguous seams/sections is done in descending order. However, if the parting is thin, the influence of the top sections/seams working over the standing pillars of bottom the sections/seams becomes extremely important for the safety and optimal exploitation of coal. Field investigations were carried out at Nowrozabad East mine where superimposed pillars of the Johilla top section were depillared above the developed pillars of the Johilla bottom section with a 3 m thick parting between the two sections. The competency of the 3 m thick parting (which consists of alternate layers of shale, sandstone and shaly sandstone) was improved by reinforcement (underpinning). The strata control investigations did not show any considerable instability problem of the 3 m thick reinforced parting and pillars of the bottom section due to top section depillaring. On the basis of this field investigation and a laboratory study on simulated models, simultaneous extraction of both the sections, along with reinforcement of the parting, was suggested for the geomining conditions of the Johilla top and bottom sections. Adopting the suggested simultaneous depillaring process, two sub panels of Johilla top and bottom sections have successfully been extracted without any strata control problem. This paper presents this case study along with the results of the strata monitoring conducted to evaluate performance of the underground structures during the field trial.

Jan 1, 2004

By Dziurzyn&apos, W. ski

This article outlines the results of a computer simulation reproducing the transport of a mixture of air and methane following the outburst on 25 November, 2005 in transport gallery D-6, bed 409/4, at the ?Zofiówka? hard coal mine. Based on a mathematical model, numerical method and boundary conditions from the disturbance caused by the outburst, the Ventgraph system of computer simulation programs was adopted for the study. We performed a computer simulation of the influence exerted by the outburst on the propagation of the air and methane mixture, and compared the results of this simulation with those of the actual event as registered by sensors of the mine monitoring system. A validation of the model allowed us to present temporal changes in the concentration of methane and air along flow routes. Keywords: mine ventilation, outburst of methane, numerical simulation, monitoring system.

Jan 1, 2008

An efficient and portable mathematical model has been developed for simulating heat transfer in six-in-line slag-resistance-heating electric furnaces for smelting sulphide ores to produce base metals and platinum group metals. This model is a steady-state one relating furnace conditions and performances to various control and input parameters. Some transient effects occurring in electric furnaces are neglected for computation efficiency. This article describes the model development and modelling results. The present model is capable of predicting: (i) temperatures at various locations in a six-in-line furnace, such as slag bath, matte bath, solid charge, freeboard space, freeze lining, cooling water, and air gaps between solid material components, etc. (ii) freeze lining thickness, (iii)smelting rate, and (iv) heat loss rate, etc. A typical feature of the model is that it is easily portable to different application platforms and sufficiently efficient with execution times less than a few seconds. Therefore, it is possible to apply the model for online prediction and control of heat transfer and freeze lining thickness in industrial electric furnaces.

Jan 1, 2011

By N M. Vegter, C B. Kohler, E H. O Meyer

Zincor operates four Lurgi fluidised-bed roasters at its roast-leach-electrowin refinery in Springs, South Africa. In 2006, Zincor?s four roasters had reached ages ranging between 26 and 38 years, while the expected life of this type of roaster is generally 20 to 25 years. Based on its condition, Roaster 3 was the first to be rebuilt in 2006. Roaster 4 is being rebuilt in 2007 and the last two roasters will be rebuilt over the next few years as necessary. The main damage to Roaster 3 was caused by buckling of the roaster steel shell. This caused the freeboard and dome areas to shift, resulting in cracks in the dome brick lining and crumbling of the bricks lining the freeboard. The steel shell buckled due to the combined effects of too-high operating temperatures and corrosion caused by the following two main factors: ? Thicker external insulation was installed during a previous maintenance shut which resulted in and increased steel shell temperature and exacerbated the normal metal fatigue; ? SO2/SO3 containing gas that had permeated through the brick lining, passed through its dew point during frequent planned and unplanned shuts over the life of the roaster. This caused acid formation between the refractory brick lining and steel shell, resulting in corrosion of the shell. The initial plan was to replace only the damaged areas of Roaster 3, but upon dismantling more damage was found and the roaster was rebuilt from the tuyeres upwards. During the rebuild, the opportunity was taken to modify the design of some of the brickwork to extend its life and make future repairs easier. This paper discusses the mode of failure of Roaster 3 and the lessons learnt during its dismantling and rebuilding.

Jan 1, 2007

By C. S. Chenga

Sustainable development is currently an extremely relevant concept worldwide, especially in developing countries like South Africa. Mining and metals companies are also directly affected by the sustainable development agenda. The Mining Charter?s requirements as well as new legislation (Mineral and Petroleum Resources Development Act) expect, among others, for mines to support infrastructure and poverty eradication projects in the communities surrounding the mines. The aim of this paper is to, against the background of a qualitative study of mining communities in the North-West Province of South Africa, put forward some critical factors for the sustainability of social development projects in the mining sector.

Jan 1, 2006

By J. Thompson

Paper written on project work carried out in partial fulfilment of B. Eng (Mining Engineering) degree A study into the stability of strike gullies at TauTona mine was to be carried out. In reaching the conclusions to follow, a three-stage process was followed. ?A literature study as well as discussions with management and workers alike to gain their inputs and perceptions. ?An ESH Department survey into the history of transgression from proper mine standard in TauTona strike gullies. Unfortunately records were found to be scarce and obscure and the importance of better inter-departmental recordings cannot be stressed enough. ?An underground survey of 11 pre-selected gullies distributed in both the UCL and LCL reef horizons. This stage is to carry the greatest weight in the points to be brought forward in this report. The process highlighted and ranked the following areas as the most critical role players in the deterioration of strike gullies: ?Poor blasting practice ?Sub-standard siding depth and lead-lags ?Split sets ?Off line gully development ?Adverse geological features. The following has been suggested to the mine to help reduce strike gully instability. ?Where possible, maintain lead-lags below 7 m?else increase the siding depth to suit. ?Decrease the gully width standard to 1.2 m (especially in ?high risk areas?). ?Discourage drillers from drilling side holes in gully face blast pattern. This will allow for time in which tendon support can be kept up to date. ?Ensure split sets are installed 1 m behind the gully face so as to prevent opening up of fractures. ?Utilize the centre split set in the 1:2:1 pattern to maintain the proper gully line. ?When adverse conditions prevail, it is necessary to apply a special support plan not only to the panel face, but also (just as importantly) to the gully as well. Shotcreting of the gully hanging is a suitable long-term consideration for providing the additional a real coverage needed. List of Abbreviations [CL Carbon Leader M/O Mine Overseer UCL Upper Carbon Leader FOG Fall of Ground LCL Lower Carbon Leader H/W Hanging Wall ERR Energy Release Rate F/W Foot Wall WDL Western Deep Levels ESH Environmental Safety and Health ]

Jan 1, 2002

By R. G. B. Pickering

The purpose of this report is to chronicle the interaction between Anglo Platinum and Sandvik Tamrock during the development and application of Long Hole drilling as a viable, practical method of narrow reef mining. Appendix I identifies the people from Anglo Platinum and Sandvik Tamrock who participated in the project. Experiments in long hole drilling and blasting were conducted in a number of mines during the late 1950s and early 1960s. Some of the advantages envisaged were: ? The face would require no daily examination ? No barring down or immediate face support required ? No sockets to clean ? Only a few well drilled holes rather than large numbers of short blast holes ? Avoidance of poorly-marked holes, inaccurate drilling and improper charging ? Better supervision and better face advance. Holes 50 mm in diameter were drilled parallel to the stope face by pneumatic rock drills from advance strike gullies. Holes were charged with amon gelignite cartridges. All holes were traced with detonating cord to ensure propagation of the explosive along the entire hole length. Rock was blasted onto a rolling scatter pile. A number of reasons have been cited for the failure of this mining method. The two most important were: ? problems in controlling hole deviation which was exacerbated by the fractured ground ? difficulty of charging up holes in fractured ground, due to relative movement of rocks along fracture surfaces over the hole length. The idea of drilling a few long holes parallel to the face rather than a large number of short blast holes perpendicular to the face was an attractive option to mechanize narrow reef mining. This particular project started in 1998. Initial trial The Union Section of Anglo Platinum purchased a Solo H 606 RA long hole drilling rig from Sandvik Tamrock in 1997. In order to evaluate long hole drilling as a basis for face mechanization. The Solo series of drill rig was selected because it combines high drilling performance and good drilling accuracy with high versatility with a compact size. The mining method proposed by a consultant to Anglo Platinum was based on the current trackless mining method practised by Union Section. The trackless mining method is better described as a trackless development and transport method. Reef strike drives were installed approximately 40 metres apart on dip using trackless equipment. Reef was mined using hand-held pneumatic rock drills and scraper winches and the excavation was supported by sticks and crush pillars at the top of the panel. Rock scraped from the panel was loaded by Lad Haul Dumps (LHDs), trammed to an inclined haulage, and loaded into trucks for transport to the surface. It was intended that the Solo would drill from the bottom advance strike drive to the top advance strike drive. Blasted rock would still have to be scraped from the panel. It is also difficult to understand how well the support would have survived. In the initial motivation it was stated that drilling accuracy was of the utmost importance. Previous work by Sandvik Tamrock had shown that, in other parts of the world, 20 metre long holes could be drilled with a deviation not exceeding 100 mm provided that: ? holes were collared with low thrust pressure ? operating percussion pressure was kept low -often not ? exceeding 10 MPa ? drilling was to be with tubes, not rods. In early 1998 initial trials with the Solo H 606 RA were inconclusive; attempts were made to drill 40 metre long holes from one advance strike drive to another. Difficulty was experienced in measuring hole straightness and accuracy, there was no survey of the holes, and no blasting was done to expose hole barrels. A long hole mining system was then defined, with a corresponding face mining system, which led to the design of a test programme for the Solo drilling rig. Earlier experience at Telfer The proposed face mining system was based on a mining method that is practised at Telfer Gold Mine in Western Australia. Telfer has a series of domed reefs that dip at angles of less than 20° to more than 40°. The ore bodies are all thin, with widths from 0.6 metres to 0.8 metres. Figure 1 shows the face layout. Experience at Telfer indicated the following: ? The spacing and size of the dip pillars should be based on the rock mechanics of Union Section ? The length of the face is determined by the ability to correctly position the blast holes. Positioning of blast holes is a function of hole straightness and hole direction. Blast holes in Telfer were drilled 64 mm diameter with retrac bits from Sandvik. Extension steel was T38 by 1.8 metre lengths. A guide rod was fitted immediately behind the bit. To achieve greater accuracy' operators drill at relatively low percussion and thrust pressures. Results show that very good accuracy is achieved in 20 metre stopes and only fair accuracy in longer stopes. In 34 metre long stopes 10-15% of the holes are not sufficiently accurately positioned and have to be re-drilled. All blast holes are

Jan 1, 2006

By D. Schreuder

Conditions in the South African market relating to atmospheric emission control has changed dramatically over the last 10 years. Companies have become more responsible, environmentally aware, and have to comply to more stringent environmental limits set by legislation. This paper describes briefly the current best available atmospheric pollution abatement technologies relating to particulates, acid gases, NOx, acid mists, aerosols, mercury, dioxins and furans. Mention is also made as to the best practices for product work-up and waste product disposal.

Jan 1, 2006

By A. Bothma

The Impala No. 1A ventilation shaft is located approximately 15 kms north of Rustenburg and is accessed via the R510road from Rustenburg to Northam/Thabazimbi. This situation is shown on the location plan in Figure 1. The geology at the site consists of an upper layer of black turf of about 1 metre in thickness, grading into weathered norite. The depth of weathering is generally less than 5 metres. Zones of deeper weathering may occur along fracture zones associated with faults and dykes. Groundwater occurs mainly within the weathered/fractured aquifer system. The ventilation shaft is a 5.1 m diameter upcast shaft ,and forms part of a system of ventilation shafts to complement the existing infrastructure as mining progresses down the new twin decline extension of Impala 1 shaft. The principal reason for these shafts is to lessen the risk of an extended RAW and also to counter the effects of the heat build-up as the mining goes deeper, with the resultant effect that the existing cooling centre (fridge plant) becomes more remote from activities. The No. 1A shaft (upcast) is positioned approximately 1,2 km on down dip from the main shaft and the No. 1Bshaft (downcast) approximately 1,2 km further. Geological information was used from a borehole 1,2 km away on strike from the upcast shaft. The first shaft (1A) is a 5,1 m diameter hole to be fitted with two 2,2 MW fans, whilst the second shaft (1B) is 4,1 m diameter to accommodate a 7MW fridge plant.

Jan 1, 2001

By M. Tangstad, B. Monsen, H. Midtgaard, R. Ishak, M. Syvertsen, I. Solheim

Charcoal is an interesting reducing agent because it is produced from growing wood which is renewable and does not contribute to global warming, provided that there is a balance between the felling of timber and growth of trees. In the biocarbon project the possibilities for using charcoal is investigated as well as other biological based and renewable reductants for ferroalloy production. The background for the project is that the Norwegian ferroalloy producers wish to contribute to reduction of CO2- emissions. This paper focuses on the use of charcoal instead of coke for the production of manganese alloys, and the consequences of doing so. Results from practical SiMn smelting experiments on a pilot scale have previously been reported, where only 12-14 % Si in the metal was obtained when charcoal was used. Recently we have attained satisfactory 18 %Si also for charcoal, but good results are indeed dependant on good furnace performance. There is no tradition for using charcoal for manganese production in Norway. A change from using mainly coke to the use of some charcoal will demand new routines for furnace operation, raw material and off-gas handling. Hence, characteristic properties of charcoal, which may cause change of furnace operation, have been measured and are compared to coke. The following results are presented: ?Thermal abrasion strength together with CO2- reactivity; important parameters for the performance of the upper part of the furnace. ?Coke-bed resistance; it will increase with the use of charcoal (a review of new results). ?Slag reactivity; a relevant method for testing slag reactivity of industrial reductants is of interest. The suggested method can serve for comparing slag reactivities towards graphite.

Jan 1, 2007

By M. M. D. S. Vemba

Paper written on project work carried out in partial fulfilment of B.Sc. Eng (Mining Engineering) degree Engineering refers to the application of mathematics and other sciences to logically develop solutions to problems. However, in opencast mining, due to the randomness of operations, the use of mathematics is often neglected. Instead, practices that have proven successful elsewhere tend to apply. At the other end of the scale, mathematical models have been developed to identify and quantify all the operational parameters involved in the loading and haulage of ore. These are then used to optimize the operation. Optimization in this context, refers to the reduction of cost through effective utilization of the available time and resources. Considering these mathematical models, the loading and transport system at SMC (Catoca Mining Society) will be assessed with a view to its optimization. The discussion will be introduced by considering the mine background and other general information, which will highlight the main operational variables.

Jan 1, 2004

The effect of composition on electrical conductivity (resistivity), viscosity and liquidus temperature of the slags encountered in electric smelting of PGM containing suphide concentrates is reviewed, along with the relationsip between slag composition, slag-matte separation and metal losses to slag. Optimization of slag composition at furnace temperature is attempted on the basis of opposite and parabolic trends of the viscosity-conductivity relationships of the slags. The importance of surface tension of slags is emphasized using available and estimated information which are grouped into a useful correlation. These values are coupled with or incorporated into slag viscosity plots to illustrate the correlation between the two factors, which influence slag-matte separation as well as refractory attack. The electrode in the slag-resistance furnace is the central factor in the design and operation of the furnace by its role of conveying the smelting power. The obvious parameters are the size and positioning in the slag bath, which are defined by slag resistively, furnace resistance, cell constant and power input. The flow of heat generated by the electrode current determines the energy distribution and furnace dimensions creating viscosity profiles, which in turn affect flow conditions. Thus the flow of slag and matte are discussed in conjunction with the furnace geometry, stirring action of the electrodes, buoyancy and electromagnetic effects. The action of the electric furnace as a settler is discussed. The rectangular furnace is considered as a channel reactor in this analysis, and the action of the electrodes is described to indicate a balanced settling rate in the area of electrode action. Discussion of the design and production capacity of the electric furnace, with reference too perating current, power input, furnace dimensions and electrode size, are based on considerations derived mainly from slag properties such as its resistively, viscosity and liquidus temperature.

Jan 1, 2004

By D. H. Diering

The paper deals with the technologies required for opening up anew idealized ultra-deep level mine. First the overall philosophy of accessing ore at ultra depths is discussed. Then the individual technologies which will be required are explored. Finally a detailed overview on one of the key technologies, namely tunnel support, is given. Because this project is still in its infancy, few definitive answers are provided; instead each of the envisaged technologies are briefly discussed, with emphasis on those which require the element of speed for rapidly accessing the ore body.

Jan 1, 1997

By T. Tamura

It is known that the CaO-FetO-P2O5slag system has a wide liquid miscibility gap at steelmaking temperatures and this two-liquid slag phase has many advantages such as high activities of both FetO and CaO and high amount of P2O5. The phosphorus distribution ratio between the two-liquid slag and molten steel was confirmed to be higher than that between CaOsat.-FetO-P2O5slag and molten steel. The two-liquid slag may be useful for dephosphorization of molten steel and reduction of slag volume. Furthermore, the phase equilibrium of two-liquid CaO-FetO-P2O5slagsaturated with MgO was measured to determine the applicability of the two-liquid slag to dephosphorization in steelmaking process. Keywords: slag, refractories, solid solution, two-liquid phase, dephosphorization, steelmaking, phase equilibrium.

Jan 1, 2004

By R. T. Jones

Slags have been produced as by-products from pyrometallurgical processes for thousands of years, and have generally been regarded as waste. Many of these slags are no longer seen as residues merely to be discarded, but as potential economic sources of metals, or as useful by-products in their own right. Slag cleaning can enhance the recovery of valuable metals in existing pyrometallurgical processes, or can produce economic value from recycling old slag dumps. For example, a process has been developed and implemented industrially for recovering cobalt from copper smelting slag in a DC arc furnace. The relatively non-leachable behaviour of glassy slags allows them to be used as a very safe means of disposal of metallurgical waste streams, including many that are classified as hazardous. The high temperatures, reducing conditions, and good mixing that can be achieved in a DC arc furnace enable the effective functioning of this environmentally beneficial means of disposal. Lead blast furnace slag can be rendered safe for disposal by treating it using the Enviroplas process, which strips most of the hazardous lead from the slag (and traps the rest in an insoluble glassy slag) and recovers economic quantities of zinc by absorbing (in a lead splash condenser)the zinc vapour that is fumed off in the furnace. Another example can be seen in the recovery of a chromium- and nickel-containing alloy from stainless steel plant dust, while producing a disposable residual slag. The examples presented here illustrate that it is possible to develop processes that are both economic and environmentally beneficial. Keywords: DC arc furnace, plasma, slag cleaning, cobalt recovery, zinc fuming, Enviroplas, jigging, re-melting fines, environmental, economic

Jan 1, 2004

By R. Minango

Solvent extraction of Cu from acidic leach, using an oxime extractant, e.g. LIX 984 or Acorga is a well-known and establishedprocess. As the process advances, Cu is extracted to the solvent and is replaced by H+in the aqueous phase. At high acidity the distribution coefficient of Cu between the solvent and the aqueous phase decreases, and eventually a relatively high concentration of Cu remains in the raffinate, regardless of the number of equilibrium stages. In many processes this phenomenon is immaterial, as the raffinate returns to the leaching and no copper is lost. But when the copper raffinate is used for processing of additional metals, minimal concentration of Cu in the raffinate is essential for an effective process. The recommended process includes three steps: SX-1?standard extraction of Cu with oxime. The resulting raffinate contains about 6?8 g/lCu and 50 g/lH2SO4. SX-N?Extraction of most of the acid into an amine-based solvent, and stripping it with a base, e.g. aqueous ammonia, caustic, Na2CO3etc. The extraction removes no metals, but decreases the concentration of acid from 50 to10 g/l. SX-2?Extraction of the neutralized solution once more with oxime, to remove the residual copper. Such a process enables one to get down to 100?150 ppm Cu in the final raffinate. This process was developed and tested in the R&D facility at Chambishi Metals plc, by a joint team of Chambishi & Bateman, using Cognis LIX 984N and real solutions from the plant. In the laboratory the equilibrium isotherms for SX-1, SX-2 and the neutralization were developed. Computer simulation of the results showed that the optimal configuration is two stages for both extraction and stripping for SX-1. For SX2, 2?3 stages of extraction and two stages of stripping are needed. The stripping stages may be common to both SX-1 and SX-2. The neutralization of SX1 raffinate was tested both by conventional precipitation with lime and by extraction with alamine-based solvent extraction and stripping with a solution of 10% NaOH. For the SX-N only one stage was needed for both the extraction and stripping. The bench scale pilot plant was run for both SX-1 and SX-2. The results were very similar to the simulated ones, and the configuration of 2E2S for SX-1 and 3E2S for SX 2 delivered the required raffinate with less than 150 ppm Cu.

Jan 1, 2005