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Mispah form (FAO: Lithosol) soil contaminated with >380 000 mg kg-1 creosote was co-composted with sewage sludge and poultry manure for a period of 19 months. The soil was mixed in a ratio of 1:1 (v/v) with wood chips to improve aeration. The soil-wood chips mixture was then mixed in a ratio of four soil-wood chips mixture: one compost material (sewage sludge or poultry manure). Compost heaps were set up in triplicate on wood palettes overlaid with double layers of nylon straw bags. Control experiments, which contained the contaminated soil-wood chips mixture without sewage sludge or poultry manure were set up in triplicate. Moisture, temperature, pH, ash content, C:N ratio of the compost and the creosote concentration of the soil was monitored monthly. The concentration of selected creosote components in the soil was also determined at the end of the incubation period. Temperature was observed to rise to about 60¦C in the poultry manure and about 58¦C in the sewage sludge within two months of incubation, while temperature in the control fluctuated between 15¦C and 35¦C throughout the incubation period. Creosote concentration in the control was reduced by 17 per cent at the end of the incubation period while concentration of creosote in the sewage sludge and poultry manure compost was reduced by about 99 per cent. The percentage rate of creosote reduction in the poultry manure compost was initially lower than observed in the sewage sludge compost. However, the reduction rate became similar towards the end of the treatment period. The concentrations of selected creosote components were reduced by about 100 per cent within the same period. Respiration of compost microorganisms and counts of microbial populations showed that microbial activity correlated with reduction in creosote concentration.

Jan 1, 2004

By R Herbert, M E. Malmstr÷, S U. Salmon

The study focuses on sulfide oxidation processes in unweathered pyrite-rich mine tailings from a soil-covered impoundment in northern Sweden. To simulate the oxygen-limited conditions in water saturated tailings two long-term column experiments were performed. Results are presented for the first 11 months of the experiments at room temperature. The tailings used in the experiments were of two different grain sizes (one coarse grained, in the range 0.02 - 0.6 mm with a grain surface area of 1.86 m2 g-1, and one fine-grained, in the range 0.0015 - 0.06 mm with a grain surface area of 10.00 m2 g-1) and with slightly different mineralogical composition. The S:Fe molar ratio in the leachates (1.0 - 1.5) indicates either that pyrrhotite is the main iron sulfide undergoing oxidation in our experiments, or, alternatively that the S:Fe molar ratio in the leachate is determined by pyrite weathering in conjunction with other processes releasing iron or immobilising sulfate. However, speciation modelling of the leachates indicates that ferrihydrite was close to saturation, suggesting that a ferric oxyhydroxide may have dissolved/ precipitated during the experimental period, thereby affecting the S:Fe molar ratio. Pyrite oxidation rates obtained from the two column experiments during æsteady-stateÆ were 2.25 + 10-13 and 8.45 + 10-14 mol m-2 s-1 in the coarse and fine tailings, respectively. Pyrrhotite oxidation rates, as an alternative, were 2.31 + 10-12 and 1.24 + 10-11 mol m-2 s-1 in the coarse and fine tailings, respectively. Natural microbial activity was confirmed in both the tailings and the leachates; therefore, the obtained oxidation rates are not strictly abiotic. It is thus concluded, based on the experimental results, that tailing heterogeneity will result in zones with different oxidation rates, related to their physical and chemical properties.

Jan 1, 2003

By W J. Bruckard

In Australia, ilmenite is produced from beach sand deposits. These deposits contain varying quantities of the valuable minerals zircon, rutile and ilmenite, along with a variety of other minerals. The valuable minerals are upgraded using combinations of gravity, magnetic and electrostatic processing techniques to yield high-grade products. One of the penalty elements present in ilmenite concentrates is chromium, present mostly as chrome spinels. This mineral is particularly difficult to separate from ilmenite, as the properties of the two minerals are very similar in relation to the traditional separation processes. While the deposits which have been treated in Australia in the past have had only low levels of chromium present, many of the newer deposits contain much higher levels of contamination and, thus, there is an increasing need to develop robust processes to separate chrome spinel from the ilmenite. Flotation, which is a unit process with the potential to solve this problem, has surprisingly received little attention for this application. The literature suggests the research effort to characterise the general flotation properties of chromite (chrome spinel with high chromium content) has been extensive and a substantial amount of data is available. However, the problem of separating a small amount of chrome spinel from a large amount of ilmenite requires a different approach to that from separating chromite from siliceous gangue, which has been the main commercial focus of chromite flotation in the past. Much of the work done has been fundamental in nature and the results obtained are fairly inconsistent, as is the testwork completed at batch and pilot plant scale. While these inconsistencies have been attributed to the variable composition of the mineral, there is little evidence supporting this premise. Just as chrome spinel and ilmenite have similar physical characteristics, the flotation response of both minerals is also similar. This paper reviews the literature relating to the flotation of chrome spinel and presents a considered approach to the development of a robust flotation method to separate chrome spinel from ilmenite.

Jan 1, 2001

By Y Peng, D J. Bradshaw, G Long

As copper reserves deepen, they are becoming increasingly less pure. Penalty elements such as arsenic, antimony and bismuth are occurring in increasing concentrations. Environmental issues with smelters are leading to increases in penalty rates for these elements such that some concentrates may become difficult to sell. Arsenic concentration in Rosebery copper concentrates during 2010 varied between 3000 and 8500 ppm. Arsenic levels above 5000 ppm can preclude import to some customer countries. Optical mineralogy has indicated that arsenic in Rosebery copper concentrates is most commonly associated with tennantite (Cu12As4S13) in solid solution with tetrahedrite (Cu12As4S13). In addition, silver reports to concentrate as freibergite (Cu12As4S13 ). A reduction in arsenic content will therefore lead to a concurrent reduction in the silver content.A literature review by Long, Peng and Bradshaw (2012) indicated that arsenic bearing minerals could be removed using potential controlled flotation. Some success in reducing the amount of arsenic had also been reported using either roasting or chemical depression using reagents such as lime and cyanide. Due to the location of the Rosebery concentrator adjacent to world heritage listed temperate rainforest, the use of a roaster or cyanide was immediately discounted. Lime is already used as a reagent in the Rosebery flow sheet, so could be considered as a potential arsenic depressant.In order to determine a method of separating out the high arsenic bearing minerals, two series of flotation tests were conducted. The first series investigated the effect of pulp potential on the recovery of copper, silver and arsenic minerals; a second series investigated the effect of pH, froth depth and regrind.CITATION:Long, G, Peng, Y and Bradshaw, D J, 2012. Production of a low arsenic copper concentrate at MMG Rosebery, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 347-354 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By H R. Watling, W van Bronswijk

Of the various copper sulfides in the EarthÆs crust, chalcopyrite is by far the most common. This mineral is also the most difficult to leach, with hydrometallurgical extraction processes characterised by prohibitively slow kinetics. While a number of processes, including leaching at atmospheric pressure and temperature, pressure leaching at higher temperature, and bioleaching of concentrates in stirred tanks and whole ores in heaps have been demonstrated to some extent, there is currently no full-scale unit process for the hydrometallurgical extraction of copper from chalcopyrite ores. Titan Resources NL operated a test bacterial heap leach at Radio Hill, Western Australia in 2001, to extract nickel and copper from a mixed sulfide ore that contained significant amounts of chalcopyrite, pentlandite and pyrrhotite. A proprietary mixed bacterial culture was used to inoculate the heap, and the nickel and copper were precipitated from the resultant solution as mixed hydroxides. Although 65 per cent of the copper from the Radio Hill heap was eventually recovered, the rate of extraction followed an unexpected trend, as the initial rates of copper recovery were much slower than column tests had suggested. It was postulated that this was due to dissolved copper precipitating as covellite or chalcocite in the most poorly aerated zones within the heap, ie furthest from the aeration pipes at the base. It was further hypothesised that the lack of oxidants (oxygen, ferric iron) and the likely inactivity of bacteria in these areas limited the re-dissolution of such secondary copper precipitates. In this study the above hypothesis was tested in column experiments using the same mixed sulfide ore and a synthetic leach solution. One set of columns was aerated and another set was operated without aeration. Results showed that precipitation of copper occurred regardless of whether the column was aerated or not. The results suggested that pH of the leach solution played a significant part in the precipitation of the copper. Further column tests revealed this to be the case. These results are described here in detail.

Jan 1, 2004

Zircon is extensively used as moulding sand in the production of steel and high-duty iron castings. To achieve a satisfactory grade of zircon concentrate, the zircon particles may need ,to have surface coatings removed, as such coatings can hinder electrostatic and magnetic separation of zircon from associated minerals. Xn addition the presence of surface coatings on the zircon. particles can cause sinterins and fusion of the zircon to the east metal resulting in "burnt-on" sand or fritting of the casting. "Burnon" necessitates cleaning of the casting by grinding or chipping, which is expensive and results in loss of definition on the casting face.Hudson (1966) has pointed out ,that some coatings on mineral sands are lightly-adhering clays which can be removed quite readily by agitating the sands in water for a few minutes-thus their removal constitutesno problem. However, the heavy gingery coloured coating which adheres strongly to the surface of the particles in some deposits, and which has generally been considered to be a form of iron oxide (Baker, 1962; Hudson, 1966), cannot be removed effectively by attrition cleaning in water (Hudson, 1966). Such coatings limit the usefulness of these concentrates in the foundry industry.An investigation of strongly-adhering coatings on "ginger zircon" concentrates from the Yoganup deposit of Westralian Sands Ltd has shown that the stain consists of a cemented mixtvre of hydrated iron and aluminium oxides and kaolin minerals, and that it can be removed by a relatively low cost treatment.IDENTIFICATION OF THE STAIN BY XRDX-ray powder diffraction (XRD) patterns were obtained from small samples of stain scraped, individually, from the surface of hand-picked zircon and ilmenite crystals. These scrapings were mounted on quartz fibres and exposed to CoKa radiation in a 57 mm Debye camera. To avoid heavy fogging of the x-ray films, a substantial amount of the iron oxides was preferentially leached from the scrapings (before mounting) with an ammonium oxalate-oxalic acid solution (Rooksby, 1961). The residual material gave a fairly well resolved powder pattern (Table 1). A...

Jan 1, 1975

By D Burgess

Since the days of mandatory high cost pilot plant testing in 6 ft + 2 ft SAG mills autogenous griding/semi-autogenous grinding (AG/SAG), mill specific energy is now calculated by various empirical formulae and pilot testing carried out in small laboratory mills. Barratt (1989) described a method using Bond Indices than adding a single factor of 1.25 to arrive at the AG/SAG specific energy. In 1991 the Canadians developed the SAGPower Index (SPI), test using 2 kg of material in a 300 mm diameter mill. This test has been further developed by Starkey, Hindstrom and Orser (2006). This test now includes a Bond Ball mill work index test.The Macpherson test is still being offered by some testing facilities. This dry test requires at least 180 kg of material.Morrell (2004) formulated an AG/SAG specific energy generator based on a drop weight index Value (DWi). This DWi value is related to the Ab parameter and the material specific gravity. Morrell reconfigured the BondÆs Third Theory formula to cause extra power requirements to be inputed. This technique incorporates the material specific gravity (SG). The AG/SAG specific energy values produced using this method are susceptible to higher values of crushing and rod mill work indices.This paper offered describes a more simple method to generate AG/SAG specific energies. This programme incorporates Bond work indices, the JKMRC parameter Ab measuring impact breakage and the parameter ta measuring the abrasion rate, then employs a series of efficiency values nominated for each transition size (AG/SAG T80). These efficiency values have been compiled over the years from actual operating plants and documented reports.CITATION:Burgess, D, 2012. A method of calculating autogenous/semi-autogenous grinding mill specific energies using a combination of bond work indices and Julius Kruttschnitt parameters, then applying efficiency factors, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 37-44 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

When, in 1985, Marty McFly is accidentally sent back to 1955 in a DeLorean ‘time machine’ in Back to the Future, he encounters young versions of his parents, and must make sure they fall in love or he will cease to exist. This Conference celebrates 40 years since the inaugural Mill Operators’ Conference in Mt Isa in 1978; over this period, great technological advances have been made in almost all aspects of mill operation, resulting in considerable improvements in grades and recoveries from ever-more complex ores. However, in 2018, the mining industry is facing challenges that might have been difficult to predict in 1978 (or were they?). The computer revolution, the information revolution, and the bio-technology revolution have all contributed to the increase in operating efficiencies, but, combined with the social revolution, they also call into question the existence of the mining industry as we know it. This presentation looks back at the ‘way things were’ in the 1970s, and at what was then considered to be the ‘basics,’ and attempts to sketch what the mining industry might be like in 2058. As Back to the Future Part II demonstrates so vividly, it is much easier to review the past than predict the future. It is difficult to know what form the mining industry will have in 40 years’ time. Nevertheless, as this presentation will argue, it is important to take action now, to ensure that it (like Marty McFly) does not cease to exist. Some courses of action are proposed, including going back to the basics of a very different past.*This is an abstract only. No full paper was prepared for this abstract.*CITATION:Franzidis, J P, 2018. Back to basics or back to the future?, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 3–4 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By M J. Ireland, D P. Horne, S J. Simpson

The Sinclair Nickel Project, wholly owned by Xstrata Nickel Australasia (XNA), includes an underground mine, a concentrator, a village, an all-weather airstrip and associated mine infrastructure. The operation is located 70 km north of Leonora in the prospective and historically productive eastern Goldfields region of Western Australia. Xstrata Nickel has successfully developed the Sinclair deposit into a 350 kt/a operation since its discovery in November 2005, with contained nickel metal production to date in excess of 10 600 t, of which 5500 t is from the underground operation.The narrow, continuous, medium-grade Sinclair nickel sulfide deposit accessed by the Phoenix Decline has a shallow 15¦ - 20¦ plunge that is approximately 1000 m long down dip and open at depth to the north. The flat plunge and low-grade nature of the orebody presents challenges regarding the amount of horizontal waste development that can be economically excavated, which in turn affects ventilation, egress, mine services and diamond drilling.The mineralisation is hosted between an ultra mafic hanging wall and a basalt footwall and is being mined using a bottom up, sublevel open stoping method. The stopes are retreated to a central access, filling voids in the process with waste mullock where possible and using cemented mullock fill where end exposure is required. Tight filling the flat dipping hanging wall has proved to be challenging given the angle of repose for the waste mullock fill is steeper than the angle of the hanging wall.Reconciliations of mine performance against the original feasibility study show positive changes and operational optimisations enabling Sinclair to achieve its mining targets in a safe, sustainable and economic manner. This paper discusses the mining operations associated with the first two years of the Sinclair Underground Nickel Project, providing an overview of the various issues encountered and the lessons learnt in the process.CITATION:Ireland, M J, Horne, D P and Simpson, S J, 2012. Overcoming mine design and production challenges at the Sinclair underground nickel project, in Proceedings Narrow Vein Mining 2012 , pp 37-48 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By C J. Berry, P H. Briggs

Assessment of metal mobility, acid rock drainage production, and toxic effects from weathering of mineralised materials (ie mine-waste piles and outcropping alteration/ore zones) is key in evaluating and mitigating metal mining sites. Many metal- and acid-generating materials at these sites are sulfur minerals (sulfides to sulfates) with differing reactivities, weathering characteristics and metal contents. Our working hypothesis predicts that data on the types and quantities of sulfur minerals, coupled with comparison of their sulfur isotopic compositions to that of aqueous sulfate in the surrounding environment, can be used to evaluate primary pathways of weathering reactions at mining sites. We developed a sequential extraction scheme to separate the phases of sulfur in mineralised materials. Operationally defined phases, including monosulfides, disulfides, elemental sulfur, water-soluble sulfo-salts, acid-soluble sulfates, and resistant sulfate phases such as barite, are quantified gravimetrically and/or instrumentally. Several sets of experiments were designed to test the effectiveness of this extraction scheme. First, a variety of ore sulfides diluted in an inert matrix was analysed with the scheme. Results showed that molybdenite and arsenic-bearing sulfides do not behave as predicted from just their sulfur oxidation states, indicating that samples containing these minerals need special treatment for complete phase recovery. Second, a natural mine-waste composite standard was analysed by the scheme. The sum of the phases was equal to the total sulfur concentration analysed independently. Each phase had a similar, but distinct sulfur isotopic composition. The results were verified with a laboratory prepared æproxyÆ having the same mineralogical composition as the standard, but with each phase having a very different sulfur isotopic composition. The sulfur isotopic compositions of the recovered phases from the æproxyÆ showed that the scheme effectively separated the sulfur with only minor cross contamination of phases. The speciation and isotopic analyses were applied to both mine-waste piles and naturally occurring alteration zones at nine mine sites. A case study from one site comparing the sulfur isotopic composition of minerals in an alteration zone and dissolved sulfate in proximal streams is presented. This case study shows the application of our data to differentiating sources of sulfate in the surrounding environment.

Jan 1, 2003

By Davey KJ

Complex sulphide ores are characterised by complex texture of the valuable minerals. Any assessment of the performance of differential flotation circuits treating such ore should include analysis of the effects of poor liberation of the valuable minerals. Selected samples were taken from balanced size by size surveys of the flotation section of a concentrator treating an ore body containing fine grained galena and sphalerite in a matrix of iron sulphides which is mostly pyrite. liberation measurements were made using the QEM*SEM system on four or seven size fractions. The measurements indicated that less than 40 per cent of the galena and less than 60 per cent of the sphalerite in the flotation feed was liberated. The final concentrates were largely composed of composites of valuable minerals and iron sulphide however there was a small but significant fraction of liberated iron sulphide. Using the sized mineral flows obtained from the original balances and the liberation measurements, a satisfactory liberation balance was obtained. This enabled the effect of liberation on the flotation performance to be examined; not only for the entire circuit but also in the rougher, scavenger and cleaning sections. In the zinc circuit the effect of a regrind mill could be assessed. Recovery of galena in the lead circuit was limited by the recovery of composite lead particles. Recovery of liberated galena exceeded 95 per cent in all size fractions while the recovery of composite galena exceeded 85 per cent below 38 ¦m but fell sharply above 38 ¦m. Examination of the behaviour of the liberated and composite galena in the various sections of the lead flotation circuit revealed high recoveries of both liberated and composite galena particles in the roughing and scavenging sections but very low recoveries in the final two cleaning stages. Similar trends to those observed in the lead circuit applied in the zinc circuit although the performance of the liberated sphalerite also fell off at sizes greater than 38 pm in this case. Thus the zinc recovery was limited by the flotation of both the liberated and the composite sphalerite. The effect of regrinding was to increase the recovery of sphalerite mainly by reducing the amounts of sphalerite in less liberated sizes rather than by increasing the liberation of the sphalerite within a given size Fraction.

Jan 1, 1993

One common public perception of mining is that financially it is an extraordinarily risky business. Another and contrary view, in certain circles, is that price risk and other financial risks can be relatively easily removed or shifted by a number of schemes or devices. In reality, beyond the orebody discovery stage, mining need not be any more risky than most heavy industrial businesses. Nevertheless, as stated by Rappaport, (1986): Risk is a parameter of central importance in establishing the economic value of any asset. The level of risk is determined both by the nature of the firm's operations and by the relative proportions of debt and equity used to finance its investments. These two types of risk are commonly referred to as 'business risk' and 'financial risk' respectively. The term financial risk is used somewhat more loosely herein to recognise the reality that any risk inadequately addressed in the design, construction and operation of a mining project can have significant financial consequences; usually negative. For instance, a specific area of business risk in mining is often called technical risk which definitely has severe financial consequences if not managed correctly. The various risk shifting, hedging and risk management tools available all have their place and range from merely useful to almost indispensable. However, in the long-term, financial risks are maintained at acceptable levels by the application of sound, more fundamental business management principles. There are many with a vested interest in perpetuating a perception that the conceptual foundations of sound management practice are extremely complex. In reality, most management concepts are prosaic, but their application requires incredible discipline and hard work. There is no magic formula or 'silver bullet'. It appears that a company whose activities are restricted to mining can analyse the range of risks specific to mine development and operations more clearly than companies in which mining is just one of many businesses. The almost universal failure of large oil companies in the mining business is probably the best example of lack of any clear comprehension of fundamental business characteristics and risks. Conversely, in regard to risk management, choosing to be in other business as well as mining does nothing to reduce the risks of the mining business. The individual shareholder can probably make such 'diversification of portfolio' decisions better than most management teams. It could be argued that diversification across a number of disparate businesses may actually increase overall corporate risk due to lack of management concentration and focus. These differences between the perceptions of financial risks and the realities in mining need to be further elaborated in relation to some uncompromising business concepts.

Jan 1, 1994