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By Petr Salak, Adrian St. John, Angelos Gakis

"INTRODUCTIONCrossrail will provide London and the South East of England with a world class, high-capacity railway to ease congestion on London’s public transport system and enable better access between business, residential and entertainment districts. At the heart of this £14.8bn scheme lies Farringdon Station, which will form a major interchange between Crossrail, Thameslink and London Underground. The £400m Farringdon Crossrail station was awarded to a joint venture of BAM Ferrovial Kier (BFK) in 2011, and comprises two ticket halls, two platform tunnels and a multitude of connecting tunnels all constructed using open face sprayed concrete linings (SCL). BFK appointed Dr. Sauer & Partners (DSP) as their specialist SCL designer. As the main contractor for both C300/C410, the Western Running Tunnels and Station Caverns, and also for C435 Farringdon Station, BFK were in a unique position to develop strategies to reduce cost, time and risk for the Client. One such innovation involved the continuation of the TBM drives for the running tunnels right through the length of the new Farringdon platform caverns. This created large diameter precast concrete pilot tunnels for the subsequent platform enlargement using SCL. The benefits of such an approach are described in detail by St.John et al 2015. BFK also introduced another alternative strategy to reduce cost, time and risk, by proposing to abandon the TBMs in the ground (“Turn & Bury”) permanently rather than recovering them through the ticket hall shafts. This de-risked the programme, enabling critical ticket hall construction to proceed unhindered. However, one consequence of this revised strategy was an area of increased underground congestion, with two SCL tunnels (RTE2 and ES2) and a bored tunnel (complete with abandoned TBM) in close proximity to a ticket hall excavation (see Figure 1) and a sensitive, listed railway structure. Faced with this complex geometry, and combined with challenging geotechnical conditions (Gakis et al 2015), the team chose to install enhanced monitoring to study the behavior of tunnel RTE2 and provide additional assurance that the tunnels were performing as predicted. Applying advanced 3D numerical modelling and taking into account the residual tensile strength of the SCL helped avoid unnecessary reinforcement installation in primary lining and/or early installation of a secondary lining in RTE2 prior to ES2 construction."

Jan 1, 2016

By Robert Henderson

SEVERAL very interesting articles have been written on the caving system of mining, but most of these papers have dealt separately with the slusher system or the chute and grizzly system. In this paper, a comparison will be made of the two methods, both of which are being used at the Climax mine. The mine is on the Continental Divide, in central Colorado, at an altitude of approximately 11,500 ft. above sea level. The long, cold winters, characteristic of the high Rockies, create a number of operating problems not usually found in other mining areas. THE ORE BODY The Climax ore body is just east of the Mosquito fault. This structural deformation has brought the granite on the east against the sedimentary rocks on the west. The ore occurs in an intensely altered granite and schist around a silicified granite core. The ore is in a highly fractured zone of moderate silicification. The principal mineral, molybdenite, is found in these fractures. Barren dikes of granite porphyry cut through the ore body and core with no apparent regularity. The grade of the ore is fairly uniform and gradually decreases toward the footwall core and also into the hanging-wall granite. The footwall and hanging wall are determined entirely by assays. The ore body is roughly dome-shaped, with the apex truncated at the surface. It is covered to varying depths with talus. A horizontal section of the ore body on the Phillipson level appears as an annular ellipse surrounding the footwall core. The ring of ore varies in width from 300 ft. to as much as 800 ft. The north-south axis of the core is about 1200 ft. and the east-west axis is about 1800 ft. In the ore zone, the rock is sufficiently strong to maintain reasonably large openings, and openings larger than haulage drifts usually have to be supported. The ore is cut by a great many fractures and shears and breaks into rather large blocks when caved, necessitating considerable secondary blasting. The waste that overlies the ore body breaks relatively fine in comparison with the size of the blocks of the caved ore and tends to run through the broken ore in the cave, causing dilution unless the draw is carefully controlled. MINING SYSTEMS For several years the ore was mined by the chute and grizzly system, a system that differs from that used in other caving mines only in the size of openings and equipment. In the chute and grizzly system, main haulage laterals were driven along the footwall core and also along the

Jan 1, 1944

The compilation of the sampling nomogram is an essential tool in the optimization of a sampling protocol and allows the operator to track the change in relative variance of the sampling error (the Fundamental sampling error) during steps of comminution and mass reduction in the process of recovery of the aliquot for analysis. The data required for the compilation of the nomogram are centred on the sampling constant K and the exponent, alpha (a) of the nominal fragment size. Five methods for establishing K have been identified, but two of them, the duplicate sampling analysis method and the heterogeneity test method purport to provide such constant. The step by step procedure for each of these two methods was examined using broken ores from three mines, Mponeng mine, Kloof mine, and Lily mine. Two of these mines (Mponeng and Kloof) are typical Witwatersrand-type gold mines, and the third is an Archaean-type shear related gold deposit situated in the Barberton Greenstone belt. The results suggest that the factor K derived using the heterogeneity test is very accurate, but applies only to a single size fraction in the spectrum of comminuted materials. The calculation of gold grain sizes using this method appear to be too small and do not conform with the results from detailed mineralogical studies. The factors K and alpha derived using the duplicate series analysis are appropriate for use across a wide spectrum of comminution sizes and also provide realistic gold grain sizes, comparable with the equivalent circular diameter of gold grains identified in mineralogical studies of the ores.

Jan 1, 2010

The compilation of the sampling nomogram is an essential tool in the optimization of a sampling protocol and allows the operator to track the change in relative variance of the sampling error (the Fundamental Sampling Error) during steps of comminution and mass reduction in the process of recovery of the aliquot for analysis. The data required for the compilation of the nomogram are centered on the sampling constant K and the exponent, alpha (a), of the nominal fragment size. Five methods for establishing K have been identified, but two of them, the duplicate sampling analysis (DSA) method and the heterogeneity test (HT) method purport to provide such a constant. The step by step procedure for each of these two methods was examined using broken ores from three mines, Mponeng mine, Kloof mine and Lily mine. Two of these mines (Mponengand Kloof) are typical Witwatersrand-type gold mines, while the third is an Archaean-type, shear related gold deposit situated in the Barberton Greenstone belt. The results suggest that the factor K derived using the heterogeneity test is very accurate, but applies only to a single size fraction in the spectrum of comminuted materials. The calculation of gold grain sizes using this method appears to be too small and does not conform with the results from detailed mineralogical studies. The factors K and alpha derived using the duplicate series analysis, are appropriate for use across a wide spectrum of comminution sizes and also provide realistic gold grain sizes, comparable with the equivalent circular diameter of gold grains identified in mineralogical studies of the ores.

Jan 1, 2009

By Shadia J. Ikhmayies

"Nanocrystalline fluorine doped tin oxide (Sn02:F) thin films were produced by the spray pyrolysis technique (SP) on glass substrates at a substrate temperature T, = 450 ° C. The hydrated stannous chloride (SnClz.2H20) was used as a precursor and ammonium fluoride (NH4F) and hydrofluoric acid (HF) were used as the doping compounds. A comparison between the compositional, morphological and optical properties of the films obtained by using the two doping compounds was performed by using scanning electron microscope (SEM) observations, energy dispersive spectroscopy by X-rays (EDAX) and transmittance measurements. It is found that the films prepared by using HF have larger grain size, and higher fluorine content than the films prepared by using ~- In addition these films have higher transmittance and higher direct optical bandgap energy. These results are interesting for the use of Sn02:F as a forecontact in CdS/CdTe solar cells.IntroductionTin dioxide is an n-type, wide bandgap semiconductor (about 3.5 eV) and in the form of thin films, it is a transparent conducting material, characterized by high optical transmission (80- 90%) [l]. Sn02:F is the most widely used transparent-conducting oxide material. The transparent-conducting thin films of this oxide are used as electrodes in a variety of optoelectronic devices, as heat mirror coatings in 'smart windows' for buildings, in automobile and airplane windshields, in incandescent light bulbs, in flat-plate and concentrating solar collectors. Tin oxide and zinc oxide, in thin film configuration are technologically important materials because of their applications in photovoltaic cells [2].Among various dopants for Sn02, fluorine is often preferred, because fluorine doping results in films with high transparency and good conductivity [3]. By exploring the literatures, it was found that different sources of fluorine were used to produce Sn02:F thin fihns, among which are ammonium fluoride (NH4F) and hydrofluoric acid (HF) are the most used. Some researchers used~ [2, 4-7] and others used HF [3, 6-8] for doping Sn02 films. Gordillo et al.[7] used both of NH4F and HF and showed that the resistivity of thin films obtained by using HF was lower than that of similar films obtained by using NH4F by a factor of 2.2. They also found that the films obtained by using HF have more random orientations."

Jan 1, 2013

By P. R. Pereira

This paper presents a comparison between the simulations of a semi-empirical model and a theoretical model for the prediction of radial stress in rock. The results of the simulations were compared to the results of a field test. The comparison of the radial stress predictions shows that in the semi-empirical model use of the plane wave approximation as the basis for the prediction of the radial stress provides a much better result than the theoretical model. The theoretical model includes the correct elastic wave geometry and theory but fails twice, firstly by considering the charge as a point (unit) energy source therefore; and secondly by using a Delta function with a very sharp rise and decay time to simulate, with no distinction between non-linear and linear zones, the attenuation of the energy per unit volume at the shock wave front as the wave moves away from the borehole. The results of the semi-empirical model were within the same order of magnitude as the results from the field test whereas only one result from the theoretical model came close to the result of the field test. However, even this closer result for the theoretical should be taken with caution as there was no measurement in the field test made at this distance from the borehole. Finally, because of its potential portability, convenience, and cost-effectiveness the theoretical model can be a most valuable tool for mining and civil construction operators in the early design of blasting rounds and of the mining operation itself.

Jan 1, 1993

By H W. Wu

The objective of this paper is to examine some modern forms of auxiliary ventilation fabric ducting with streamline design fittings and the efficiency of construction practices. The most dynamic form of auxiliary ventilation is a moveable fan and ducting, which is easy to install and uses light weight fabric. The porous nature of some ducting construction results in leakage and loss of efficiency. Flexible fabric ducting is primarily associated with metalliferous mining and industrial material quarrying. Some applications are found in coal in stone/rock development and with use of dust scrubbers.Forcing auxiliary ventilation systems employing circular fabric ducting have been examined. Leakage of air from ducting is affected by ducting material and construction methodology (such as welded or sewn), quality of installation, number of joints, total length, pressure differences between the inside and outside and diameter. The best ducting system with welded construction can have up to four times less leakage as compared with some systems with sewn construction. It is also evident that the sewn type ducting will deliver less air to working faces and require more fan power, hence higher operating costs to run when connected to the same fan.A case study has also been undertaken to compare predicted operating costs of auxiliary ventilation systems connecting various types of ventilation duct fittings. The study also assesses the technical importance of streamline design with low aerodynamic resistance to reduce the shock losses of the duct fittings. It was found that the system connected with streamline designed fittings is able to provide some power cost savings and deliver more air compared with the system fitted with conventional fitting pieces.Using a ventilation ducting system with a lower leakage and frictional characteristic represents a sound solution to new challenges faced by the industry for increasing demand for better quantities and quality of air delivered to working faces; this is due to higher production schedules and stricter mine health and safety regulations within the Australian mining industry.CITATION:Wu, H W and Gillies, A D S, 2017. A comparison of auxiliary ventilation systems and their predicted operational performances, in Proceedings Australian Mine Vent Conference 2017, pp 207–214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

By T. A. Doyle

Column and batch test methods are used to simulate and model the leaching of various mining by-products. Batch tests are simpler, while column tests more intuitively represent the field conditions to be simulated. Projects that have included different types of leaching tests illustrate the strengths, limitations, and associated uncertainties of each. Column and batch tests have been used to evaluate arsenic solubility in leachate from treated copper smelter flue dust, to simulate the leaching of pit wall rock from a proposed gold mine, and to simulate the release of uranium from aquifer materials.

Jan 1, 1998

By Hugo E. Camargo

In an effort to reduce Noise Induced Hearing Loss (NIHL) in the mining industry, the National Institute for Occupational Safety and Health (NIOSH) is conducting research to develop noise controls for mining equipment whose operators exceed the Permissible Exposure Level (PEL). The process involves three steps: 1) Noise source identification (NSI), 2) development of noise controls, and 3) evaluation of the developed noise controls. For the first and third steps, microphone phased array measurements are typically conducted and data are processed using the conventional beamforming (CB) algorithm. However, due to the size and complexity of the machines, this task is not straight forward. Furthermore, because of the low frequency range of interest, i.e., 200 Hz to 1000 Hz, results obtained using CB may show poor resolution issues which result in inaccuracy in the noise source location. To overcome this resolution issue, two alternative approaches are explored in this paper, namely the CLEAN-SC algorithm and a variarion of an adaptive beamforming algorithm known as Robust Capon Beamformer (RCB). These algorithms were used along with the CB algorithm to process data collected from a horizontal Vibrating Screen (VS) machine used in coal preparation plants. Results with the array in the overhead position showed that despite the use of a large array, i.e., 3.5-meter diameter, the acoustic maps obtained using CB showed "hot spots" that covered various components, i.e., the screen deck, the side walls, the I-beam, the eccentric mechanisms, and the electric motor. Thus, it was not possible to identify which component was the dominant contributor to the sound radiated by the machine. The acoustic maps obtained using the RCB algorithm showed smaller "hot" spots that in general covered only one or two components. Nevertheless, the most dramatic reduction in "hot" spot size was obtained using the CLEAN-SC algorithm. This algorithm yielded acoustic maps with small and well localized "hot" spots that pinpointed dominant noise sources. However, because the CLEAN-SC algorithm yields small and localized "hot" spots, extra care needs to be used when aligning the acoustic maps with the actual pictures of the machine. In conclusion, use of the RCB and the CLEAN-SC algorithms in the low frequency range of interest helped pinpoint dominant noise sources which otherwise would be very hard to identify.

Jan 1, 2009

By S. D. Thompson

Several discrete-event simulation models have been developed for designing the conveying system capacities of underground coal mines. Most of these models require the use of a step-function approach to mimic the loading process at each conveyor loading point. Over the years, practitioners have generally employed one of three versions of the step-function approach. An important issue is how these three versions compare to each other in terms of simulated capacity requirements for belt conveyors. Another is-sue is how they compare to peak capacity design. This paper addresses these issues by statistically and quantitatively comparing their simulation results. The consequences and problems associated with using each version are also discussed and summarized.

Jan 1, 1997

By H. E. Urtubia

"BHP-Billiton’s Minera Escondida Limitada, in Chile, is one of the world’s largest copper producers. It treats a combined 10,500 tph of copper ore through its two concentrators: Los Colorados and Laguna Seca. The performance of the rougher circuit at each concentrator was measured and analysed using the JKTech flotation optimisation methodology. This involved developing a flotation model for each concentrator based on cell operating parameters (for example bubble surface area flux and froth recovery) and feed ore floatability characteristics. The rougher circuit performance was evaluated in terms of traditional measures, such as grade and recovery. The operation was also analysed using the measured cell operating parameters and ore floatability and the effect of these with regard to performance (grade / recovery) was assessed.This paper presents a comparison of the rougher circuit at each concentrator in terms of these parameters. The main differences in rougher circuit performance between the two plants were attributed to differences in feed ore floatability characteristics and froth recovery. The comparison highlighted opportunities where performance improvements could be made. The study also showed the usefulness of separating the flotation process into easily measurable subprocesses to further identify the different mechanisms occurring in each rougher circuit. Traditional measures like grade and recovery are useful for monitoring flotation operation but this data lacks the depth to determine the root cause behind operational issues. Determining the operating parameters and ore floatability provides the operating plant metallurgist better tools to assess and manage these issues."

Jan 1, 2006

By C. L. Pillar

INTRODUCTION The location, size, character of the ore de- posit and its adjacent formations will deter- mine the mining system best used for its ex- traction. The mining system chosen will have to be engineered to suit the ore deposit and the people that will extract it. What can be done in the Philippines cannot always be accomplished in the upper Northwest Territory of Canada. Mining operations developed under a free enterprise system are planned to achieve a satisfactory return on the capital required to place them into production. Hence, the choices made in planning and developing the mining system must be dictated by sound economics. The mining system planned must favor: 1. Maximum ore extraction of the deposit with minimal grade dilution. 2. The amount of capital expenditure used to develop and equip the mine must be in harmony with the financial policy of the group yet be sufficient to provide a reliable mining system. Excessive frills and extras that are not required for efficient, safe production should not be provided. 3. Mine production costs are determined by the amount of men, material, equipment and energy used in the production process. These must be kept to minimal limits in order to insure dependable, low cost production. Extras that later are needed in the production operation can be added as required by the operator. Basically block caving is a mining system that is normally used to extract deep seated, massive, low grade deposits of copper, molybdenum and iron where stripping ratios for open cast mining are economically prohibited. The caving system is limited in its selectivity. It is a mass production type of underground ore extraction that under favorable conditions is one of the most efficient, low cost mining systems in the world today. Usually block cave mining systems are large. They are capable of producing ore at a rate of 10,000 to 100,000 tons/day from large deposits at costs that are comparable with many open cast mining systems burdened by high waste to ore ratios. Block cave mining systems differ basically in the plan of the area layout, the method of undercutting and the method of draw. The balance of the mining system is the mechanical transportation of the drawn ore to the surface, the mine supply, drainage and ventilation systems. This paper deals mainly with the methods of draw. The method of draw to be used in a block caving system dictates the development plan of the mining area, is highly influential in the method of undercutting to be employed and the supply and ventilation system to be used. There are three basic methods of draw commonly used in today's operations. They are: Gravity draw Slusher draw LHD draw The type of draw planned will have a major affect on: Capital expenditure and time required to bring a deposit into production. Productivity, efficiency and cost of the production operation. Percent recovery of the ore deposit in both tonnage and grade. Degree of production reliability Fundamentally the type-of draw used must be compatible with the physical characteristics of the deposit to be extracted and the financial position of the mining operation. The type of draw chosen must work to advantage with the size, configuration, depth, strength and fracture pattern of the deposit and its adjacent formations. It must favor the financial protection of the planned mining operation. A Gravity Method of Draw requires a finely fragmented, easy flowing product from the caving action of the deposit. To obtain the desired economic advantage the ore deposit must also have sufficient area and height of column to permit minimal development cost on a per ton basis. The gravity system re- quires the maximum amount of development openings and uses the minimum amount of mechanized equipment of all the caving systems. It is unskilled, labor-intensive and is highly sensitive to size of ore fragmen-

Jan 1, 1981

By A. A. Rafiei, J. A. Finch

A previous study comparing n-pentanol and the polyglycol Fl50 showed that the two reagents could give the same gas holdup but with different size bubbles. The finding implies that bubble rise velocity depends on the surfactant (frother) type. A study using bubble swarms supported the frother type effect but bubble interactions were a possible confounding factor. This study resolves the question by measuring the rise velocity of single bubbles of 0.8-2.5mm diameter over a distance :::: 300cm in n-pentanol and F 150. It is shown that at typical frother concentrations used in flotation that the rise velocity in n-pentanol remains comparable to that in water alone while in Fl50 the velocity is significantly reduced. The results confirm the suspected impact of frother type. It is illustrated that in n-pentanol the rise velocity does not reached terminal; while in Fl50 it does and this is the source of the velocity difference. Advances

Jan 1, 2009

By L. A. Roe

Calcium carbonate and talc fillers both find use in a multitude of Industries. Each filler has many commendable properties, and often compete for the same market. Quality controls for various grades are exacting and complicated. A detailed comparison of the physical and chemical properties of these two fillers is tabulated and the mining and processing of the ores is described.

Jan 1, 1980

By M. Potesser

In the primary and secondary copper industry the anode furnace is the most important process step in the refining procedure to reach an enhancement in value and quality of the copper by a selective oxidation and reduction. By combining the advantages of the conventional tilting hearth anode furnace and the benefits of the drum type anode furnace a new aggregate was developed and patented. The cross-section of the furnace is elliptical and the whole unit is tiltable. During the melting period, the contact surface between the burners and the bath or the scarp is enlarged - an increased energy input is the consequence. For the refining period the furnace is rotated, whereby a high bath level is guaranteed. Due to the longer residence time of the gas in the copper melt, not only the gas consumption is reduced but also the efficiency/quality of the refining process is improved. This investigation shows a metallurgical comparison between the new anode furnace concept with the elliptical profile/cross section and the conventional, tiltable hearth anode furnace as well as the drum type anode furnace. The CFD modeling of the currents in these different furnaces delivered helpful data to compare the geometry of the different furnaces. The investigation covered the refining of blister copper from primary metallurgy, from secondary metallurgy as well as from a mixture of both copper melts. The paper underlines the possibilities for a reduction of the tap-to-tap time and for an increase in production capacity, which can be achieved by applying the new furnace concept in comparison to the conventional furnaces.

Jan 1, 2006

By R. G. Lundquist

The idea that fragmentation for mineral processing results from some tradeoff between increased drilling and blasting and increased load on the primary crusher is well known. In a previous paper (Lundquist and Konya, 1963) the authors reported on the results of a laboratory scale study which attempted to quantify this tradeoff in energy terms. If we look only at the production of new surface area, blasting is 1700 times as energy efficient as crushing. On an energy cost basis, crushing is 57 times as expensive per unit surface produced as blasting with dynamite and more than 700 times as expensive as blasting with ANFO. The comminution process at an open pit operation starts with the drilling of blast holes and ends after the material has gone through the fine grind circuit. Operators usually break the process into several separate operations and do their cost analysis based on this approach. Often drilling and blasting is considered as one system and crushing and grinding are considered as one or two separate and independent systems. Operators will independently optimize each system. The cost for the drilling and blasting system is obvious in operations which use contract drillers to do the drilling and either contract the blasting to a distributor or do the blasting themselves. The cost of rock breaking in this manner is readily available since the contract driller will supply a monthly drilling cost to the operation and either the distributor who is contracting the blasting or the distributor who is supplying the powder to the blasting operation will also supply a monthly billing. In either case, the costs can easily be seen on a monthly or per ton basis. They are not difficult to compute even when drilling and blasting are done entirely in-house.

Jan 1, 1985

By B Wentzinger

This paper presents the results of state parameter estimates in an active coal boiler ash storage facility using the most recent methods currently in use for cone penetrometer interpretation. State parameters estimates are also compared with the critical yield stress ratio approach, which is proposed as an additional liquefaction potential screening method. Recommendations are provided regarding additional in situ tests and laboratory tests that may be undertaken to improve the state parameter assessment and data produced for use in subsequent stability or liquefaction studies.

Jul 1, 2021

Diamonds have been known in Australia for many years, especially in association with alluvial gold and tin occurrences. Exploration for primary sources in the Kimberely area of Western Australia has been carried out since 1968, culminating in the discovery, in 1979, of the Argyle Kimberlite No 1 (AK1) kimberlite pipe. Various techniques have been used in the search for diamonds such as geologic interpretation, aerial photography and Landsat, soil geochenical techniques, geophysics, and stream sediment sampling. Stream sediment sampling, a method developed in South Africa, using the detection of indicator minerals derived from kimberlites in gravel samples, has had wide application and success in Australia. Nevertheless, although Australian diamond bearing kimberlite and lamproite rocks contain some of the classic indicator minerals, their concentrations are generally much lower than in South Africa and Russia. Despite this paucity of indicators, stream sediment sampling is still the best known and most commonly used method of diamond prospecting in Australia. Small samples are treated by hand but larger samples require the use of diamond pans, jigs or heavy media separation plants. Each of these methods has advantages and disadvantages. The final concentrate is passed over a grease table or subjected to x-ray sorting.

Jan 1, 1984

By Michel Grenier, David A. Young, Brent Rubeli, Mahe Gangal

This work forms part of a study being conducted into the feasibility of constructing a prototype Load Haul Dump (LHD) vehicle for operation in underground Canadian mines. A series of in-laboratory assessments were performed to determine the relative contributions of a Deutz F4L912W (54 HP- conventional vehicle engine) and a Deutz F2L 2011 (31 HP hybrid vehicle engine) to diesel emissions concentration during mine transient duty cycles and steady-state operating modes respectively. Results indicate significant reductions in the quantity of gaseous and particulate emissions with the application of the hybrid system and a high-efficiency ceramic element filter. This paper describes the engine test protocols and the results of this laboratory study.

Jan 1, 2006

By A. Mainza

It is common practice in the mineral processing industry to use hydrocyclones for particle classification. However, classification in the UG2 platinum circuits using the hydrocyclone poses major challenges due to the differences in density between silica and chromite, the two major components of this ore. Silica, the PGM carrying component, has an average density of 2.7 and the barren chromite component has an average density of 4.5. When conventional and flat bottom cyclones are used there is a misplacement of particles due to differences in density leading to inefficiencies in the classification by size. As a result of the density effect, coarse silica reports to the overflow, resulting in loss of recovery, and fine chromite reports to the underflow, resulting in loss of milling capacity and unnecessary production of fine chromite. A classifier based on a different concept was required in the classification of UG2 platinum ore. The three-product cyclone, which is a hydrocyclone with two concentric vortex finders to produce three distinct products, has been tested in the UG2 ore application and indications are that misplacement of particles due to differences in the component densities can be minimized through the use of this unit. Industrial and pilot plant trials were conducted and indications are that the three-product cyclone can be installed to selectively produce a middlings stream that can be screened using Pansep screens to provide a screen oversize, which is predominantly coarse silica that can be preferentially reground to recover the PGM values contained in the coarse silica.

Jan 1, 2005