Search Documents

By Heyes GW, Orrock GI

The fine grained complex orebody at Elura contains galena, silver and sphalerite in a matrix which is mostly pyrite; sulphides account for about 90% of the orebody. A differential flotation circuit is operated to con- centrate lead and silver in the first section and zinc in the second. The metallurgical performance to date indicates the need for improvement. The lead section has been sampled in detail to comprehensively assess its overall performance and to define problems associated with lead and silver flotation. The substantial amount of pyrite in the feed dominated the perfor- mance of each stage of the separation. It was responsible for extensive dilution of all concentrates and, as a consequence, for very large circulating loads. The rougher concentrate contained 68% pyrite some of which was rejected through the three stages of cleaning to 39% in the final concen- trate. However, this was still sufficient to hold the assay of this product below 45% Pb. Size by size behaviour shows that flotation rates of the valuable minerals, as characterised by their rate coefficients, were low through- out the circuit. There was a significant deterioration in these rates between rougher and cleaner I stages which resulted in large circulating loads between these stages. There was practically no recovery of coarse valu- able minerals to the final concentrate. The dilution of the final concen- trate by non-sulphide gangue was not a major problem. If the recovery of the galena and silver in the +38¦m size range could be raised to that of the - 381Lm size range, galena recovery would rise from 78% to 86% and silver recovery would improve from 46% to 59%. The Pb grade of the final concentrate would remain the same.

Jan 1, 1988

IN carrying out experiments to determine the amount of metal taken away by fume or fine dust the apparatus used requires to be strong and compact, as it is generaIIy necessary to work on a staging fixed at considerable distance above the ground and against a smoke stack or flue. It is essential that the test be carried out at a part of the stack welI away from any bends, to prevent eddies. The method described below has been used for estimating the dust losses from Dwight & Lloyd sintering machines, antimony drossing furnaces, pot sintering trials, etc., and has been founc! satisfaetory. It is welI to bear in mind that the longer the period over which dust coIIection is extended, the more valuable the result is likely to be. The rateof aspiration of the gases through the apparatus is kept as regular as possible by changing the filter thimbles frequently, and the rate of flow should correspond closely with the rate of flow of the gases up the stack.The apparatus required for carrying out a test is as folIows :-{a) Thermometers : two are required, one for taking the temperature of the gas in the stack and one for giving the temperature of the gas in the aspirating drum. If the temperature in the stack is over 400° C., it is advisable to use a pyrometer such as the "Bristol" the couple being standardized before use at about the temperature of the stack.(b) Pilot Tube: The foIIowing sketch (Fig. 1) iIIustrates the principle of the pitot tube. The readings are to be taken in a straight flue, or stack, preferably 20 diam. away from any bend. A, which is open, faces the air current and...

Jan 1, 1922

The object of this paper is not to describe any particular type of engine in close technical detail but rather to give a general picture of past and present practice in this important section of the industry. The subject involves the study of almost 600 years of coal mining history which has fluctuated from stagnation to extremely rapid progress.Up to the year 1300, coal was obtained from outcrops and exposed beds, but this easily worked source soon gave out, so by that time shafts had to be sunk to the coal seams, and coal mining proper commenced its career.These early shafts were very shallow, and winding equipment was not a matter for much concern. Raising coal was generally achieved by human labour, which was called "coal bearing." Coal was loaded into baskets. strapped to a man's back, and carried from the underground workings to the pit bottom. The unfortunate individual then climbed ladders to the top of the shaft, unloaded the basket, and descended the shaft for the next load. Men, women and even children, were employed as the mine expanded, and were generally bound to the colliery for life, performing daily work which would now be regarded as slavery. This primitive method of raising coal to the surface was slow, inefficient and highly dangerous, yet for want of a better method it continued until in 1602 it isrecorded as having been abolished in Wales, one of England's great coal districts. It was not until 100 years later, however, that it was abolished in the northern coalfields, and the colliers were freed forever from their dreadful life.WINDLASSThe first step towards winding coal was the introduction of the windlass. This was similar to the common windlass used over a well, and was operated by man power. The...

Jan 1, 1945

To provide suitable backfill material for the cut and fill mining method being used to extract ore from the C.S.A. mine, flotation tailings are deslimed in single stage cyclones. The sand fraction is stored on the surface in unagitated storage bins before being pumped underground, while the slime portion is thickened and pumped to a residue dam. Design and operating details of the sand preparation and storage sections are given.The initial preparation of a stope for filling, consists of laying percolation and decantation drainage pipes on the sill floor. Preparation for fllling of subsequent lifts consists of rearing up the ventilation pipes at the ends of the stope, preparing the decantation system, and building a barricade or wall. The fill is carried to, and discharged into, the stope through 4 in. diameter pipes and filling of any section is carried out 24 hr/day normally without supervision. The fill is retieulated underground to the 1200 ft level through rubber lined pipe installed in the upcast shaft. Boreholes of 3 in. diameter drilled from the1200 ft level convey fill to the 1800 ft level. The mine drainage system is basically designed to handle the water associated with hydraulic filling. A system of primary and secondary sumps has been installed to handle the percolation and decantation drainage from stapes and to allow settling of fines before the water is fed to the main pump station.A programme of investigation into the properties of the emplaced fill is in progress. Results to date indicate that the fill as presently placed and drained, will not cause any problems in the mining of vertical orebodies by cut and fill methods for lifts in excess of 300 ft.

Jan 1, 1969

By J J. Cilliers, K Hadler

The potential of air rate profiling to significantly improve the flotation performance of a bank of cells has been shown by numerous studies. To date many profiles have been evaluated; balanced, increasing, decreasing and combinations of these. An increasing air rate profile, where the air rates are increased proportionally to each cell down the bank, has been typically found to yield improved flotation performance compared to other profiles. Air rate profiling has two distinct components; determining the correct total air rate for the bank, and determining the correct distribution of that air. This can be done simultaneously by finding the optimal air rate to each cell in the bank in turn through air recovery optimisation. Air recovery is the fraction of the inlet air which overflows the weir of a cell as unburst bubbles. The peak in air recovery (PAR) is a robust measurement of optimal performance for the operating conditions in the cell, and sets the PAR air rate. In this study, plant trials of different air rate profiles and profiling based on PAR, were evaluated. The total air rates to the bank for all profiles tested were approximately equal, allowing direct comparison. The highest cumulative mineral recovery was achieved by operating at the PAR air rates for each cell down the bank. Different distributions of the same total air rate to the bank gave the same cumulative upgrade ratio, but lower cumulative recoveries. In this work the upgrade ratio is defined as the ratio of the concentrate grade to the feed grade. The increase in cumulative mineral recovery, at the same upgrade ratio, in this study was greater than 33 per cent. Air recovery based optimisation of air rates to banks of cells significantly improves flotation performance as it simultaneously determines the optimal air rate to the bank, and the optimal distribution of that air.

Jan 1, 2009

By E Rubio, N Morales

Currently, mine plans are optimised by using several criteria as objective functions, like profit, life-of-mine, concentration of some pollutants, mining costs, confidence level or ore resources, with consideration of constraints related to production rates, plant capacities and grades. Whilst this approach is successful in terms of producing high value production schedules, it uses a static opening sequence of the drawpoints and therefore the optimisation is made within the level of freedom left by the original opening schedule, and as a result it is far from the true optimal value of the project. This paper presents a model to optimise the sequence of drawpoint opening over a given time horizon. The emphasis is in the precedence constraints that are required to produce meaningful operational sequences considering the exploitation method (panel or block caving), physics considerations and logical rules. Further on, while it applies the standard approach of maximising net present value (NPV), it considers other targets for optimisation, like the robustness and constructability of the plans. Finally, it applies the model to real data to obtain feasible plans by means of this model. The first tests were done varying the production capacity of the mine.

Sep 26, 2011

By S Hapugoda, A Edenton, M Adam, R J. Holmes, J R. Manuel, R Smyth

Lump ore is one of the key ferrous feeds used in the blast furnace ironmaking process, normally accounting for ten to 20 per cent of the blast furnace burden. Like other feeds, the characteristics of lump ore have an important impact on the stable operation and productivity of the blast furnace. Lump ore is usually assessed chemically, physically and mineralogically to establish its likely blast furnace performance. The physical characteristics of lump ore are strongly related to the chemistry, mineralogy and texture of the ore involved. In this paper, the blast furnace ironmaking process is reviewed, with particular emphasis on the key quality requirements for the ferrous burden material. A variety of key physical characteristics of lump ore measured using standard and non-standard methods are then described and subsequently correlated with the chemical and mineralogical characteristics of the ore wherever possible. Finally, the impact of these physical characteristics on blast furnace operation is assessed.

Jan 1, 2009

By Andrews B. S

The sinter plant at the Sulphide Corporation's Cockle Creek works produces a feed for the company's Imperial Smelting Furnace. The feed contains lead and zinc oxides and all the fluxes necessary for furnace operation. The quality of this feed is a major factor in furnace performance and the considerable flexibility of operation possible at the sinter plant has been used in efforts to achieve optimum operating conditions.Operating data is processed daily and information gathered from this source has led to three of the four improvements described.These improvements have resulted from elimination of gas recirculation, reduction of sinter returns sizing and alteration of ignition conditions and main layer fuel loading.

Jan 1, 1964

By J Hadjigeorgiou

The stability of open stopes in underground hard rock mines is influenced by several factors including the shape and size of the excavation, surrounding stresses as well as the degree of jointing. This paper presents a design methodology for stope analysis in jointed rock masses. The proposed approach relies on geological structural data to construct realistic three-dimensional joint network models. The stability of an open stope is determined based on a limit equilibrium analysis.

Jan 1, 2000

By J J. McNee and, J Crusius, T F. Pedersen

The use of water covers is a globally-applied method for minimising acid rock drainage (ARD) and metal leaching. Although the subaqueous disposal of potentially acid-generating materials has been shown to effectively limit ARD, under certain scenarios, metal remobilisation can occur at circum-neutral pH. The following paper presents data from four case studies in which metal mobility from submerged mine waste resulted in degradation to water quality in the overlying water cover. Subcentimetre resolution profiling of dissolved constituents across the tailings-water interface (dialysis arrays), in conjunction with mineralogical information, were used to elucidate the controls governing metal mobility. The data illustrate four pathways of metal remobilisation, including: the release of zinc via the oxidative remineralisation of metal-rich organics at the tailings-water interface; enhanced remobilisation of arsenic as a result of anthropogenic eutrophication; the release of copper resulting from the pH-driven dissolution of treatment sludges; and the release of radium from uranium mill tailings due to the dissolution of radiobarite [(Ra,Ba)SO4] by sulfate-reducing bacteria. The results highlight the conclusion that the chemical stability of submerged tailings reflects both: the solid-phase characteristics of the tailings material; and the physical, chemical and biological characteristics of the depositional environment. In particular, the data illustrate that rigorous characterisation of both the waste solids and receiving environment must be performed in order to assess their compatibility. The results have relevance to the management of tailings repositories including constructed impoundments, pit lakes and natural lakes.

Jan 1, 2003

By J Krnmmelbein, T Raab

In Lusatia, lignite mining causes major disturbances on the landscape scale. Reclamation procedures are applied to recover the reconstructed landscape for different land use options. In this study we investigate soil physical properties of an agricultural reclamation/recultivation site directly after site restoration and the development of selected parameters within the first years after site restoration depending on varying recultivation strategies.The sandy substrates used for recultivation stem from depths of several meters. They are unstructured, free of recent organic matter and have high contents of calcium carbonate. During reconstruction and most of all during surface levelling with heavy crawlers, the substrate?undergoes high mechanical stresses. This leads to partly severe soil or substrate compaction, depending amongst others on substrate water content and mechanical stresses during site construction. We investigate the effect of different organic soil amendments, different crop rotations and partly subsoiling on the development of soil structure. We compare results of laboratory measurements directly after site construction before any reclamation measures had been applied to results we gained after three and four years of recultivation. The results show that after the earthworks are completed, the site is strongly heterogeneous in relation to mechanical stability, permeability for air and water, pore volume and bulk density. The results after three and four years show that the heterogeneity of soil physical parameters decreases. The selection of suitable crop rotations superimposes the affects of other reclamation measures, eg varying organic fertilisers or deep loosening.CITATION:Krnmmelbein, J and Raab, T, 2012. Reclaimed lignite mining areas A73 in Lusatia - Development of soil physical parameters of an agricultural site within the first years after site reconstruction, in Proceedings Life-of-Mine 2012 , pp 117-122 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

By C M. Rebagliati, J G. Payne, K Roberts

The Pebble deposit, located in south-west Alaska, is one of the largest Copper- Gold-Molybdenum (Cu-Au-Mo) porphyry systems in the world, with a current resource of >8 billion tonnes which contain 73.7 billion pounds of copper, 87.1 million ounces of gold and 4.16 billion pounds of molybdenum. The Pebble deposit formed at ~90 Ma, during emplacement of subduction-related hornblende granodiorite porphyry intrusions emplaced into Jurassic-Cretaceous sedimentary, volcanic and subvolcanic intrusive rocks of the Kahiltna Terrane. The Pebble West and Pebble East Zones represent two thermal and fluid centres in a single large magmatic-hydrothermal system. The West Zone extends from surface to about 500 m depth, has thin zones of oxide and supergene mineralisation overlying hypogene chalcopyrite mineralisation, and is amenable to open pit extraction. The East Zone extends to at least 1700 m depth, is higher grade, occurs beneath an eastward-thickening cover of younger sedimentary and volcanic rocks, lacks oxide and supergene mineralisation, and is being evaluated for underground block-cave mining. Most Cu-Au-Mo was introduced with early K-silicate alteration and related quartz-sulfide veins. In the East Zone, a structurally-controlled overprint by advanced argillic alteration significantly increased Au and, to a lesser extent, Cu grades. Exploration potential at Pebble and its surrounding district remains high. Several other porphyry, skarn and epithermal deposits have been discovered in the district, but only limited exploration has been conducted. The eastern, higher grade part of the Pebble deposit remains open in several directions, with significant potential for resource expansion by ongoing, aggressive drill exploration. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Hinkley JT, Smitham JB, Nicol SK

Historical investigations of mineral flotation have highlighted the many factors which affect flotation. Some of these are related to the mineral, its degree of liberation (however defined), solution chemistry and flotation reagents. The complexity of the interactions between these factors often leads to uncertainty over whether the final grade/recovery results obtained in either laboratory or pilot scale tests are in fact the maximum recovery at a specific grade. The reason for this is that there has not previously been a general technique for establishing the floatability of an ore. This study describes the further development of the 'tree' flotation technique for minerals other than coal, and in particular its application to iron ores. Oxides have been chosen as a system of intermediate complexity between the simple coal system and sulphide minerals to illustrate the principles and develop the concepts. The 'tree' flotation procedure consists of the progressive flotation to completion of successive froth products, followed by reflotation to remove entrained material. It therefore represents an 'equilibrium' approach to floatability rather than a kinetic approach. By ordering the various products and tails in order of increasing grade, a cumulative curve of grade and recovery can be constructed against which an isolated batch or plant result on the same ore can be compared. In this investigation, a Jaspilitic iron ore was tested at various grind sizes to determine the 'best' flotation grade-recovery that could be obtained with a given set of reagent conditions. The separation characteristics of a process are best measured on a system which contains particles which behave as 'middlings' with respect to the process in question. For this reason a natural ore was chosen in preference to model systems in order to relate more readily to plant situations.

Jan 1, 1993

IT is generally accepted that for transport of coal underground from different collecting stations, the system of endless rope haulage with remote control is not only one of the safest and most economIcal, but is also one of the most'efficlent in dealing with output.When the coal measures lie in regular planes, the work of laying down such a haulage presents no difficulties. When, however, the dip of the measures is Irregular, or dislocated by faulting, the management is faced with the problem of constructing vertical curves of such a radius that the sag of the rope will correspond with the road bed, and thus obviate the danger of the rope rising and throwing the skips off the track.The direction of the main roads for haulage purposes being ascertained from the plan of the royalty to be worked, the faults or dislocations in the seam are determined for these roads with the greatest possible accuracy by means of bore holes from the surface.From the data thus obtained the haulage road can generally be so graded that the rope will keep on the floor during haulage operations, although the grades may in places be synclinal.The writer confesses that he knows of no accurate formula for calculating the sag of the rope under varying tensions and so determining the safe radius of the curve, and he had to depend upon past experiences under somewhat sImilar conditions.The accompanying sections illustrate a case where the writer has recently installed an endless rope underground approximately 5000 ft. in length, from the foot of a tunnel in which an endless rope delivers the output to the screening plant, and having to negotiate, first a 400-ft. downthrow fault, and later a 32 ft. upthrow.

Jan 1, 1927

The face of project evaluation is changing continuously, making it ever more critical to get a clear understanding upfront of the objectives, decision-making processes and criteria of the project proponents. Requirements are usually significantly different for emerging æjuniorsÆ compared with long established major mining houses and companies in developing countries compared with western companies. Requirements are also different for different types of evaluations, such as due diligence studies compared with feasibility studies of æownedÆ projects and there are often key misunderstandings around the use of the term æbankableÆ. Further, it has become increasingly important with increasingly remote and lower quality resources to understand risk relativities such as infrastructure compared with the core processing facilities. This has also driven the need for more trade-off studies to determine the best balance between critical issues, such as optimisation of the interface between mine and processing plant, capital versus operating cost and automation versus labour. In every project evaluation, it is also critical to consider emerging trends and changing requirements and standards such as sustainability and community interface. Finally, the inescapable drive to reduce time to market has driven changes such as increased use of more standardised designs and innovative approaches to the transition from project development to execution. This paper begins with a brief review of the different objectives and risks of each project development stage to establish a framework for subsequent discussion of the different requirements of different types of companies in different locations, different types of evaluations and different types of projects and project components. This is followed by a review of changing requirements and standards and some innovative developments in sustainability and community interface are presented. Finally, the paper outlines a number of developments in the area of operational readiness planning that focus on minimising the time to achieve maximum return on investment.

Jan 1, 2009

The perception of the mining industry in the Australian community has changed considerably over the last four decades. This has been partly brought about by changing values and expectations of the Australian community as well as by global factors. The legal environment for the Australian mining industry has undergone significant changes. Among them, the High Court's Mabo decision which confirmed native title rights as part of the common law and environmental legal developments are dealt with particularly. It is suggested that the Australian legal environment is not necessarily the prime cause for Australian mining companies taking their investment off shore. This is a summary on how native title impacts on land tenements and some of the major issues that emerged since the `Mabo decision' and the Native Title Act 1993 are discussed. Environmental legal developments have resulted in liability facing directors and senior management in compliance matters .and in operating mining and other businesses because of changes in community expectations. In the next decade it is suggested there will be many drivers for more changes in the legal environment. Some of these are touched on including the changing roles of the Commonwealth and State Governments, the High Court as -law-makers and the changing relationships with indigenous people in Australia and in overseas countries. The changing legal environment for mining demands a national approach in many areas on the proposition that resource based industries are part of a national industry. The uniform law approach among the States has been taking place but this needs to further. An appropriate task force should be forward to review the entire legal structure and regulations affecting mining on a national rather than a State basis . The mining industry will need to work with expectations of a changing community which have powerful, but sometimes contradictory, interest groups.

Jan 1, 1997

With the advent of the resources boom in the late-1960s most of the larger scale resources projects in Australasia were commissioned on an æowner miningÆ basis, where the mine owners owned and operated the primary mining equipment. Then, in the 1980s with the arrival of the gold mining boom in Western Australia and elsewhere, a change occurred û most of the mining activities were carried out on a contract mining basis. Over recent years, however, there has been a marked trend amongst larger Australasian open pit mining operations back towards owner-controlled and operated mining operations. The reasons for this are many and are addressed in the paper, but centre on the concept of the owner being prepared to assume more of the mining risk in exchange for the advantage of not having to pay a contractorÆs profit margin. Where mining contracts are entered into, the older æschedule of ratesÆ format is being replaced with less adversarial and more entrepreneurial contract alternatives involving operating cost and risk sharing in a spirit of æfair dealingÆ. Leasing and maintenance issues have also been influenced in recent years by the advent of more flexible contract arrangements with equipment suppliers. Group purchasing has also had some influence for project operators. For many mine owners, the optimal risk management profile involves a practical offsetting of the assumption of mining risk on the one hand, against the removal of litigation risk and potentially reduced operating costs on the other. Potential pitfalls may arise where equipment size and type requirements vary with time or where manpower management, mining flexibility, job skills or safety exposure factors are relevant. These issues are commented on in the context of some well known examples, and an update is provided on contract mining versus owner operation in large-scale Australasian open pit mining generally.

Jan 1, 2002

The decision to contract or owner mine should be independent of, and should be made before, definitive estimated cost comparisons could be available. There is substantial potential risk of estimated cost comparisons being flawed due to difficulties in compiling comparable estimates for differing business objectives, errors and omissions or because of significant bias. In the current economic environment, escalating cost pressures potentially lead to increased incidence of disputes in mining contracts. The apparent cost of contractorÆs profit and overheads has tended to become more significant to owners. Economy-driven trends towards higher levels of performance and commercial risk for contractors, and the potential for more adversarial, difficult management; with attendant disputation between contractors and owners, have encouraged development of so-called ærelationship contractsÆ û partnering or project alliance contracts. The paper discusses the issues, circumstances when owner mining or contract mining is favoured and perceived advantages and disadvantages of contract mining. The implications of owner-perceived core business and core skills and outsourcing are also discussed. Techniques of introducing objectivity into the decision to owner or contract mine using matrix analysis or benefit comparisons are described. A disciplined procurement process should deliver a decision to owner or contract mine independent of, and prior to, any definitive comparison of estimated costs. A generic procurement process model is provided as a basis for discussion. Risk is discussed generally and as a significant factor in comparison of estimated operating costs. Consideration of risk and any prudent cost treatment are promoted as necessary pre-requisites to any comparison of estimated costs. The paper develops the concept that the decision to owner or contract mine should be based on real benefits that become obvious in the early stages of the development, particularly the procurement process; and before definitive estimated cost comparison can be available. In the context of this concept estimated cost comparisons tend to be confirmatory. Problems with comparing estimated ownerÆs operating costs with tendered contract prices are also discussed, particularly recovery of capital cost of equipment; and the owner minerÆs trade-off of operating risks for the residual contractorÆs margin attributed to overheads and profit. The paper concludes that sole reliance on a comparison of estimated ownerÆs operating costs with tendered contract prices in deciding to owner or contract mine is questionable and likely driven by a need to satisfy specific in-house customers.

Jan 1, 2003

By P Kumar, J Safarian

Understanding how ore interacts with flux particles at elevated temperatures to create molten slag is crucial since it governs the dynamics of a chemical reaction. This study explores the smelting behaviour of pre-reduced Nchwaning manganese ore when combined with lime, with the objective of examining the evolving interaction between pre-reduced ore particles and lime over time. The research sheds light on the interaction between solid and liquid and the phases that emerge during this process. To achieve this, a sessile drop furnace was employed to rapidly heat the materials positioned adjacent to each other on an alumina substrate and to observe the smelting process as it unfolded over time. This method allowed for the direct observation of the melting temperatures and the flux-ore reaction progression rate, and the potential disruptive events that might occur. By comparing the molten interfaces of the fluxed materials at various time intervals, this study provides insights into the relative rate of slag formation from the two materials. The results indicate that the main slag formation initiated at approximately 1400°C and continued to advance with time, with complete mixing occurring around 1500°C. The possible phases formed were identified using Scanning Electron Microscopy and modelled using Fact Sage thermodynamic software. In addition, the iron particles in the pre-reduced Mn ore were separated and settled from a rich MnO-containing slag. It was found that the separation of molten iron droplets from the slag depends on the rate of solid MnO particles dissolution into the adjacent slag phase.

Jun 19, 2024

By A Murphy, D Lauder, J Miranda

Minera Alumbrera Limited operates the Alumbrera concentrator, treating a nominal 80 000 tpd of porphyry gold/copper ore. It is located in Argentina, 1200 km north-west of Buenos Aires in Catamarca Province, at an elevation of 2550 m above sea level. Treatment facilities consist of crushing, SAG and primary ball mill grinding circuits, rougher flotation, ball mill regrinding, gravity recovery and cleaner flotation. It is one of only a small number of concentrators in the world, which uses Jameson Cells alone for cleaner circuit flotation. Plant design was started in June 1995, with commissioning commencing in July 1997. Following a successful ramping up to design tonnages and recoveries, a number of options were reviewed to improve gold and copper recoveries. This paper concentrates on a detailed size by size flotation recovery project, commenced in May 1999. Utilising optical microscopy, scanning electron microscopy and analysis of surface chemistry it has allowed losses by size fraction to be quantified and options for enhanced recoveries to be developed. Coupled with extensive circuit surveys it has also provided an insight into previously unknown properties of Jameson Cell operation. Among these are unique residence time and bubble formation reactions to reagent and pulp flow variations, which has permitted a number of novel changes to the operation to maximise capacity and recovery.

Jan 1, 2000