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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

Titanium dioxide (TiG ) is produced 2 commercially by two methods generally referred to as the "sulfate" route and the "chloride" route. The SCM Corporation, in supplying 15% of the world's demand for TiO 2, uses both of these processes. Of the two processes the sulfate route is generally regarded as the older and more established technology. Worldwide the sulfate route accounts for some 70% of total Ti02 production. However, in the U.S.A. the chloride process accounts for 80% of production. The processes vary in chemistry and raw material requirements. In the "sulfate" process finely ground ilmenite or titanium slag is digested with sulfuric acid to produce a porous "cake" vrnich is extracted with water and recovered liquor. The solution is treated with scrap iron to reduce any ferric ions to the ferrous state which is then crystallised and removed as ferrous sulfate heptahydrate. The liquor is concentrated by evaporation and hydrolysed to form the insoluble hydroxide which is filtered and washed free of impurities. This is converted to the dioxide by calcination at 800 - 1000¦C and milled to break up any aggregates. The "chloride" process uses a higher quality feedstock of natural or synthetic rutile and/or high grade titania slag. The feedstock is mixed with ground high-purity coke and chlorinated at 950¦C in a fluidised bed. The resulting crude titanium tetrachloride is purified by distillation and chemical treatment. The purified titanium tetrachloride is burned with oxygen or oxygen enriched air to give titanium dioxide and chlorine, which is recovered and recycled to chlorination. Titanium dioxide pigment from both processing routes undergoes similar finishing techniques. It is subjected to further stages of wet and dry milling, then coated with layers of inorganic oxides in order to enhance its properties for specialist applications.

Jan 1, 1986

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

THE slag from the blast furnaces having been granulated, a representative sample is taken to the sample mill. It is then dried, crushed through a 100-mesh screen, intimately mixed, and packeted for assay purposes.METHOD OF ANALYSISINSOLUBLEWeigh 1 grm. of the finely-pulverized slag into a 3t-in. casserole, moisten with a little water to prevent sticking, cover with a clock glass, and add 20 cc. of hot acid (1 HNOs, 3 HCI, 4 H2 0). Evaporate slowly to dryness on a hot plate and bake (without overheating) until all the gelatinous silica is dehydrated. Cool, add 20 cc. HCI (1-1), and boil gently for several minutes; filter into a 300-cc. beaker, cleaning out the casserole thoroughly with a rubber-tipped glass rod. Wash residue with hot distilled water and transfer it to a No. 1 porcelain crucible. Dry, ignite, cool in a desiccator and weigh as total insoluble matter. Dilute filtrate to 200 cc. and neutralize some of the excess HCI with NHa.OH. Warm, and pass H2S for 15 min. Filter and wash withhot water. Boil off H2 S, oxidize with 5 cc. HNOll , reduce bulk of the solution by evaporation. to about 100 cc., and transfer to a 250 cc. flask.SILICAFuse insoluble residue over a blast lamp in a platinum crucible with three or four times its weight of anhydrous carbonate of soda, for 10 to 15 minutes, or until decomposition is complete. Cool, and extract fusion in an 8-in. evaporating dish with HCI (1-1); rinse crucible and cover with distilled water, and evaporate to dryness over a water bath. Take up dry mass with 10 cc. strong HCI, and then.

Jan 1, 1922

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

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 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 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

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

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

The methods employed in dressing 'the tin ore up to the required percentage of purity depend mainly on the type of ore crushed.These may be briefly divided into two distinct classes-viz., q, free: milling ore and a complex milling ore. In the former the ore is practically free from the common metallic sulphides, such as pyrites, mispickel, blende, galena, etc., and 1!' clean concentrate of 70 per cent metallic tin can be directly obtained in one continuous act of concentration; in the latter, however, in order to get a good extraction, it is only advisable to primarily concentrate up to 30 to 40 per cent. tin oxide, the rest being the sulphides of the metals. This concentrate is then roasted, re-ground, and re-concentrated up to 70 per cent. metallic tin on a separate set of tables.In the complex milling ore the tin is intermixed very closely with the sulphides, so that it becomes necessary to collect all the mineral present in order to obtain clean tailings. Stamp mills are the most popular machines for the primary crushing in the district, but whether they are invariably the best for the purpose is rather an open question. No doubt for some of the very tough chloritic ores they are the most efficient, but for a brittle quartzose ore, such as the free milling proposition usually is, the writer thinks Oornish rolls would well hold their own, especially when the tin occurs coarse, as is so often the case with this class of ore. Of course, the great drawback to stamps is the excessive quantity of tin slimed and slime tin is one of the problems of tin-dressing-while rolls give...

Jan 1, 1907

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

By Howard G. L

The gold treatment plant of Great Western Consolidated N.L. is at Bullfinch, Western Australia, 22 miles north of Southern Cross, which is 100 miles west of Kalgoorlie. A railway to Southern Cross connects the mine with the main Perth-Kalgoorlie line.The plant, which commenced operations on the 31st October, 1952, was, designed by Messrs. Alex: Stewart & Co. in consultation with the technical staff of Western Mining Corporation Ltd. to treat 1,000 long tons a day, made up of approximately equal parts of quarry oxidized ore and underground sulphide ore, by moderately coarse grinding (2-3 per cent plus 65 mesh Tyler) and straight cyanidation without straking. The design had to allow for future treatment of all sulphide ore at the rate of 40,000 tons per 28-day periodand to make provision for the possible later inclusion of preaeration treatment.The test work on .which design was based was done on representative development samples, the mine being divided into zones for this purpose, to check possible local variations from average treatment behaviour. Proposed mill feed mixtures were made up from time to time for complete ore dressing tests, which were conducted in the Western Mining Corporation laboratory at Kalgoorlie.The chief feature of the design was that the plant had to handle the extremes of very soft, poor settling and filtering oxidized ore and very hard, good settling and filtering sulphide ore. On one hand, crushers and mills had to be large enough to handle the sulphide ore and on the other, thickeners and filters had to be large enough to handle the oxidized ores.The design is marked by simplicity, the use of large units of equipment, and flexibility to cope with a variety of ores.

Jan 1, 1956

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

The Myally Beds, Surprise Creek Beds and Mount Isa Group form the upper part of the Proterozoic Western succession of rocks in northwest Queensland. The latter two units are partly equivalent, and are important for their base metal content at Mount Isa and elsewhere. The arenaceous Myally Beds lie unconform- ably or disconformably below the Mount Isa Group and Surprise Creek Beds; they contain four mappable units which were deposited on a broad, shallow shelf washed by east-trending currents. Emergence of the Myally shelf initiated a series of parallel, possibly intra- cratonic, troughs to the east and west. In the east, the Surprise Creek Beds were deposited between the emergent shelf and base- ment rocks, and consist of a shelf and trough facies divisible into 7 and 10 units, respectively. Trough sediments are continuous north-south, and show turbidite character- istics; algal dolomite characterizes the shelf facies. To the west in the Mount Isa trough, an orthoquartzite facies, the Warrina Park Quartzite, defines the base of the Mount Isa Group. Hinge zones between the shelf and trough environments are marked by east-west facies changes, and provide a focus for later major north-south faulting. However, some penecon- temporaneous faulting near Mount Isa is best explained as complex post-depositional faulting of a disconformable sequence. Subsidence and deposition in the Mount Isa trough continued for a longer period than in the Surprise Creek trough, allowing accumulation of lead and zinc deposits in the upper half of the sequence. The Surprise Creek Beds appear equivalent to the lower half of the Mount Isa Group, and are probably not prospective for lead and zinc. Turquoise and possible other secondary copper minerals are widespread in the Warrina Park Quartzite and equivalents in the Surprise Creek Beds, and further investigation of these formations seems warranted.

Jan 1, 1974

By B Menzel, M Wlasenko, J Phillips

Yandicoogina (Yandi) open pit mine is located in the eastern part of the Pilbara region of Western Australia. It hosts pisolitic iron-oxide mineralisation distributed within a meandering paleoriver channel. The deposit is characterised by clays occurring as lenses, accretions, desiccation joint infill and voids in the pisolite ore. The irregular shapes and highly erratic distribution of these clay pods represent a serious challenge for grade control at this mine.The current grade control approach used at the Yandi mine is based on sampling of production blastholes on a 6 m X 6 m spacing. Their good spatial coverage, tight spacing and low cost make blastholes a very convenient media for grade control and ore-waste definition. However the potential downside of this grade control approach is poor sample representivity. An alternative grade control approach is to drill and sample using reverse circulation (RC) holes distributed on a 25 m X 25 m grid. This approach has been tested in Yandi mine and compared with the blasthole sampling approach for sample representivity. The impact of increased drill spacing on mining block predictions has been studied using a conditional simulation technique. Blasthole and RC samples proved to be similarly biased, underestimating Al2O3 and SiO2 grades and overestimating Fe grades, when compared to full cone assays. These results show that using RC drilling for grade control at the Yandi open pit will not necessarily improve the representivity of the grade control samples. However, increasing the grade control spacing to 25 m X 25 m will lead to more than 12 per cent of the selective mining unit (SMU) sized blocks being misclassified. This is approximately a two-fold increase in the number of misclassified SMU blocks in comparison with 5 m X 5 m spacing.

Jan 1, 2007

By Moorrees C

Tests show how that temperature and rate of sparging to remove carbon dioxide determine the rate of precipitation and the chemical and physical properties of hydrated magnesium carbonates precipitated from clarified magnesium bicarbonate liquors derived from crude magnesia. At temperatures above 55°C the product is largely in the form ofhydromagnesite, Mg5(C03)4(OH)2.4H2O Between I5°C and 55°C the major product is nesquehonite, MgCO3·3H2O, although partial hydromagnesite precipitation can be induced by a high rate of aeration.As a result of the reduced time required to bring about greater than 95 per cent magnesium precipitation, the higher MgO content of the product, the ease of product recovery from the precipitation vessel, and the lower energy requirement for thermal decomposition of the precipitate to magnesia, it is preferable to precipitate hydromagnesite rather than nesquehonite. Moreover, the magnesia produced from hydromagnesite is somewhat more reactive than that formed from nesquehonite.Using precipitate calcination temperatures less than 1000°C, the chemical reactivities of the magnesias formed from both precipitates are greater than those of the crude magnesias derived from Savage River magnesite itself. The magnesias would be suitable for most uses developed for caustic-calcined magnesia. However, for refractory brick manufacture, it would be necessary to briquelle the caustic-calcined magnesia. adjust the CaO and SiO2 contents and sinter at temperatures in excess of 1600°C. This would bring about an increase in the bulk density and induce the appropriate chemical bonding reactions.Dead-burned magnesia produced from Savage River magnesite by the calcination/carbon dioxide route would have very low B2O3 contents, typically less than 0.0001 per cent.

Jan 1, 1985

By U Rehm

The biggest hard rock double-shield tunnel boring machine (TBM) in the world, with a diameter of 12.4 m, is currently excavating the North-South Bypass, which includes a crossing under the Brisbane River, Brisbane, Australia, to relieve the often congested road traffic in the city centre. Ten years ago a 10.7 m Mixshield giant helped to handle the Olympic Games in Sydney linking the airport with the city centre for an improved transportation of fascinated visitors. In Shanghai, currently the biggest TBM in the world with 15.43 m diameter presents excellent performance rates; and in Paris a convertible Mixshield could be quickly changed from earth pressure balanced (EPB) to slurry type in order to handle the very heterogeneous ground conditions. In Madrid a 15.2 m EPB giant finished its tunnel job four months ahead of schedule. Large-scale TBMs have proved their economic and innovative potential to create the current and future underground architecture of mega-cities. TBM technology has made further large innovative steps in the last decade. State-of-the-art is a TBM for very heterogeneous ground conditions comprising hard and soft rock with additional high groundwater pressure. The machine concept is adapted to the various ground conditions and allows for a quick and flexible application. Ground treatment by means of additives developed to a helpful tool in this connection. To meet the special challenges of urban tunnel projects of nowadays it is of very high importance, beside applying highest developed technology of TBM, to have a comprehensive technical support during the project phase and to develop positive relationships between all participants of the project which last longer than only for one project.

Jan 1, 2008

By M Thomas, M Caccetta, C Laukamp, A Hacket, M Jones, A Rodger, M Verrall, R Hewson

A mineral systems approach was used to interpret precompetitive airborne mineral mapping geoscience data for iron oxide copper gold (IOCG) exploration in the Starra area, Eastern Succession, Mount Isa Inlier. Selected minerals were chosen to map possible source rocks/fluids as well as intersecting structural/fluid pathways that could have transported mineralising fluids into domains of contrasting rocks/fluids where there was potential for physicochemical gradients to precipitate metals. Ferric and ferrous iron contents and white mica abundance were useful mineral criteria to identify theMount Dore Granite of theWilliams-Naraku Batholith as a potential source of æoxidisedÆ fluids rich in Fe, K and potentially Cu-Au. The Mount Dore Shear Zone (MDSZ) intersects the western margin of this granite with the highest contents and changes in composition of the white mica are peripheral to this contact within the granite. This together, with the fact that the entire N-S length of the MDSZ in the study area (~16 km) is abundant in white mica are evidence of K-rich fluids had preferentially moved along this pathway. To the south, the MDSZ intersects a domain comprising massive ironstone bodies as well as numerous graphitic horizons. This interpreted æreducedÆ domain is well mapped using the æopaqueÆ mineral products, as well as by white mica, which is pervasively developed and extends 2 km either side of the MDSZ fault into the meta-sedimentary sequences. The chemistry of this white mica also changes progressively away from the MDSZ, becoming more Si-rich (longer wavelength) towards the contact with massive ironstone bodies in the west and the graphitic units in the east, both of which represent sites of known Cu-Au mineralisation. Beyond these contact zones of interpreted æoxidised fluidsÆ with reduced rocks/fluids, there is either a rapid change to Si-poor white mica and/or no white mica developed, ie major æchemicalÆ gradient change. These mineral mapping results provide evidence for a direct link between possible source rocks and sites of IOCG metal deposition as well as extending the associated alteration zonation which had previously been mapped/published at the metre-scale to the kilometre-scale. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Derham DW, Newcombe LA

The operations of the Mount Morgan gold tailings retreatment plant were significantly improved in 1984 by the desliming of the CIP plant feed, which reduced cyanide consumption and increased percent gold dissolution. Subsequently, the development of slimes treatment, incorporating gold leaching from slimes solids by copper cyanide complexes in the process water, and gold adsorption on to carbon in an environment lacking free cyanide was successfully implemented. Mount Morgan was initially mined for gold commencing in 1882 on the auriferous ironstone cap to the sulphide ore body. During the history of mining and processing at Mount Morgan the operations have developed through numerous phases of gold and copper production, operating almost continuously through until the final closure in late 1990. The final phase of processing was the retreatment of tailings, predominantly flotation tailings accumulated since the 1930s, for the recovery of gold and minor silver. Tailings retreatment commenced in 1982 with the commissioning of a 10 000 tonnes per day CIP plant to treat some 40 million tonnes of tailings at an average grade of 1.08 g/t Au. The initial operations were hampered by high cyanide consumption and poor gold dissolution on the mixed siliceous/pyritic tailings. The successful development and implementation of the desliming of the CIP plant feed in 1984 (as described by Parsons and Hampshire, 1986) reduced cyanide consumption by 40 per cent, and raised gold extraction by 13 per cent points by virtue of significantly higher dissolved oxygen levels. The finer portion of the reclaimed tailings were deslimed by eighty 100 mm diameter AKW cyclones, with the desliming cyclone overflow thickened in a Warman 30 in diameter thickener for water recovery and slimes disposal. Operation of the desliming plant involved the discarding of a slime fraction representing about 15 per cent of the reclaimed feed. Although gold-bearing, the high content of cyanicides and oxygen consumers within the slime negated any conventional treatment of this stream with free cyanide for gold recovery. The natural lixiviant behaviour of the soluble copper cyanide complexes for gold, which was extremely evident at Mount Morgan, directed efforts in this area. Enhancement of gold leaching with a deficiency or near absence of detectable free cyanide, but with significant levels of copper cyanide complex from recycled process water were examined. Adsorption of this solubilised gold onto carbon in a similarly difficult environment required close examination.

Jan 1, 1993