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The purpose of this study was to investigate the effectiveness of froth fl otation for the recovery of gold from the gravity tailing stream of the former Lihir Gold Limited (LGL) operation in Ballarat. The P80 of the tailings was determined by screen analysis as 1 mm, with test work conducted on as received material along with samples ground to P80 sizes of 47, 78 and 134 microns. Laboratory scale batch flotation tests were conducted at both natural and elevated pH to investigate the impact of pyrite depression on gold recovery. Results indicated that although an increase in grade from approximately 0.7 g/t to 3 g/t was achievable at a natural pH and a P80 of 134 micron, gold recovery was low at just 35 per cent. Finer grind sizes were found to have an adverse effect on both grade and recovery. Future efforts will be focused on the sub 310 micron size fraction of the tailings, which was revealed to contain approximately 85 per cent of the available gold and may be amenable to cyanidation or ultra-fine gravity treatment.

Aug 1, 2010

By Kellett KR

Gold Mining at the Lancefield Mine was commenced initially in 1897 and continued intermittently through to 1956 when rising costs and mine cave-in forced closure. The mine was redeveloped in 1981 and a new-oxide treatment plant commissioned, followed by a new refractory treatment plant in late 1982. Both these treatment plants are located within the Windarra Nickel Project plant area.The refractory ore is processed by fluidised bed roasting of flotation concentrates. This is carried out in two series connected reactors to maximise arsenic and sulphur eliminations and to minimise reagent consumption. For gold recovery the zinc precipitation process was initially used, to be later replaced by carbon-in-pulp.

Jan 1, 1984

By Floyd J. M

The recovery of gold from refractory ores and concentrates can be achieved in a number of ways using roasting and leaching procedures. The Sirosmelt system introduces a new approach to the recovery of gold from these materials using smelting procedures. Laboratory studies were carried out to establish procedures for operating the Sirosmelt system for recovery of gold and associated valuable elements from an ore containing 17.5 g/t gold and a concentrate containing 22 g/t gold as well as substantial copper levels. Copper, iron, lead and matte were evaluated as collecting phases for the gold. The ore smelting procedure recovered 80-90% of the gold whilst smelting concentrate gave 98-99% gold recovery.

Jan 1, 1987

By M Gerards, M Steinberg, T Graham

The quality and quantity of Australia’s secondary deposits, which include tailings streams and tailings dams, are higher than most countries’ primary deposits. In addition, these secondary deposits have already been mined and processed, meaning that the recovery of high-grade product requires low capital and operating costs.The most widely used process for the recovery of these valuable fine and ultra-fine hematite iron particles is the wet high-intensity magnetic separation technique, originally developed by Dr Jones over 55 years ago. Many of the original machines are still working today.The aim of this paper is to demonstrate the capabilities in terms of grade and yield of using sophisticated magnetic separation modules. All results and figures are obtained from test work and design and engineering work with material from the Pilbara. In addition, consideration is given to dewatering of iron ore product and tailings.Two types of tailings have been identified for processing. Firstly, tailings that have already been deposited in a tailings storage facility, and secondly, live tailings streams from an existing wet iron ore beneficiation plant. In both cases, the technical and practical results are shown along with relevant flow sheets and plant layouts. Finally, this paper highlights the total predicted power and water consumption, capital requirements and projected operating expenses.The results show that the operating costs are significantly lower than the mining and processing operations that are required for an equivalent amount of iron ore. The recovery of iron depends on various factors but is within the 30–50 per cent range at an equivalent grade to the mainstream product. Another significant benefit is that the magnetic separation process is water positive, meaning that water is released rather than added in the process.The processing of the Pilbara’s secondary deposits will therefore result in an overall reduction in operating costs, an increase in product yield from a plant, extend the life-of-mine and offer environmental benefits by reducing the tailings footprint and water usage.The quality and quantity of Australia’s secondary iron ore deposits and tailings are such that many other countries would judge these as equivalent to primary deposits. Furthermore, these tailings are already mined, available and liberated to such an extent that the recovery of a valuable product is highly economical in terms of capital expenses and operating expenses.CITATION:Steinberg, M, Graham, T and Gerards, M, 2015. Recovery of iron ore fines and ultrafines from tailings by using wet high-intensity magnetic separation – JONES WHIMS, in Proceedings Iron Ore 2015, pp 191–196 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By P S. R Reddy, V N. Misra

The iron ore deposits in India are rich in iron content but usually contains high alumina and silica. The current industrial practice results three products namely lump, fines and slimes. The lump and fines with or without beneficiation are directly fed to the blast furnace and sintering plant respectively while the super fines or slimes are presently discarded as tailings. Around eight million tonnes of iron ore slimes estimated as being 15 - 20 per cent by weight of run of mine ore containing ~48 - 62 per cent Fe, is discarded every year causing huge loss to the Indian iron ore industry. In addition, these slimes stored in massive tailing ponds pose enormous environmental and ecological problems. Extensive investigations have been carried out on three iron ore slimes generated at Barsua, Bolani and Meghahatuburu iron ore mines of India. The studies include physical, chemical, mineralogical characterisation and beneficiation studies adopting conventional hydroclyclone and wet high intensity magnetic separator. The characterisation studies have indicated that most of the gangue minerals are accumulated in finer fractions. The studies on hydrocyclone and magnetic separation had indicated that an iron concentrate of 61 - 65 per cent Fe could be obtained by recovering 59 - 78 per cent of the iron values present in the slimes. Depending on the nature of the slimes the alumina and silica levels have been brought down to 2.5 and 1.0 per cent respectively. The effects of different operating and design variables along with detailed characterisation studies have been highlighted.

Jan 1, 2002

Low-grade manganese ore from Sandur, India containing 28.2 per cent manganese and 20.9 per cent iron was roasted with sulphating agents, viz ammonium sulphate and sulphuric acid. The effects of temperature of roasting (300 - 800°C), time (15 - 120 minutes) and amount of sulphating agent (0.5 to 2.5 times stoichiometric amount) were studied. Under optimum conditions, it was possible to extract 92 per cent manganese with 0.5 per cent iron at a temperature of 700°C for 90 minutes duration with stoichiometric ammonium sulphate. Further increase in quantity of ammonium sulphate has marginal effect on manganese extraction.Similar studies were carried out with sulphuric acid, and an extraction of about 87 per cent manganese with similar iron contamination was achieved at a temperature of 700°C for 60 minutes duration with stoichiometric quantity of the reagent. Selective extraction of manganese with respect to iron is discussed.

Jan 1, 1996

By Paramguru R. K, Jena P. K

Increasing demand for copper, lead and zinc with continuous depletion of good grade ore reserves, warrants the need for utilization of lean and complex ores of these metals. There is worldwide effort to develop suitable processes for these purposes. The present paper deals with the chloride leaching of complex base metal sulphide concentrates of two Indian deposits. A two stage leaching process is suggested keeping selectivity of metal dissolution in view.In the first stage most of the lead and zinc are reacted with dilute FeCI3 (0.3M) solutions at a temperature ono°c where major portion of the lead remains in the residue and zinc reports in solution. After taking out lead chloride by brine leaching, the residue has been treated for copper dissolution. A concentrated FeCI3 leaching at boiling temperature as well as ambient aqueous chlorine leaching have been separately tried to dissolve out the copper from the residue. Based on the results obtained, a flowsheet for the proposed process is presented.

Jan 1, 1981

By Ranklin W. J, Swinbourne D. R

Lead blast furnace slags contain typically 10 - 20 mass per cent Zn. The zinc can be recovered by slag fuming, in which a powdered coal/air mixture is injected into a molten slag bath. Volatilised zinc is recovered as zinc oxide from the gas and processed hydrometallurgically to produce zinc metal. The ISASMELT continuous lead smelting process in operation at Mt Isa produces a slag containing about 12 per cent Zn which is currently granulated and stockpiled. This slag could flow to a further bath-smelting reactor for continuous zinc fuming. Rather than collecting the zinc as fume from such a process, one option would be to recover it using a lead splash condenser. The success of this operation would depend on the composition of the gas produced, in particular the concentrations of gases capable of reoxidising zinc vapour, viz, carbon dioxide and water vapour. The aim of this investigation was to determine the smelting conditions for a bath-smelting process such as ISASMELT or SIROSMELT which would produce a gas similar in composition to that from an Imperial Smelting Furnace by thermodynamic modelling of the process using the CSIRO Thermochemistry System. It was found that such a gas could be produced if coal injection and oxygen enrichment of the combustion air to at least 40 per cent 02 were used. However, the off-take gases must be conditioned by addition of methane to reduce the water vapour content to acceptable levels. An addition rate of ten per cent of the off-gas flowrate was found to be adequate. Addition of methane to the off-gas had the added advantage of cooling the gas from the bath temperature to a temperature suitable for normal lead splash condenser operation.

Jan 1, 1993

The extinction of zinc from steelmaking dusts has been studied with a laboratory scale reduction-vapour deposition (RVD) process in the absolute pressure and temperature ranges of 13 - 1300 Pa and 1333 - 1513 K respectively. This process consists of pre-reduction of iron oxide by carbonaceous agent, carbothermic reduction of zinc oxide under reduced pressure to produce zinc deposits and subsequent re-evaporation of the deposits to liquedify the zinc. In the RVD, the gas mixture of Zn and CO quenched from the hot zone is separated into Zn deposit and CO which is transferred to the pumping system. The Zn are a layer of fine metallic zinc powders accumulated on the cold wall and removed to the vacuum distillation chamber. The experimental works have been carried out according to the following items: carbothennic reduction of zinc oxide under vacuum, carbothermic reduction o zinc ferrite under vacuum, and 3. distillation of the Zn deposits. It has been concluded from these experimental results that the recovery of metallic zinc from the EAF dusts is successful by this process, and a new concept for a practical zinc winning process has been proposed.

Jan 1, 1993

By James DG, Turner JH

This paper describes laboratory testwork for the beneficiation of a medium grade iron ore, using Wet High Intensity Magnetic Separation (WHIMS). Factors leading to the selection of machine type for this method of separation are discussed and the design of the WHIMS section flowsheet of an iron ore beneficiation production plant are described including methods to achieve operational flexibility and adequate protection of the WHIMS separating chamber. The results of the commissioning and operation of the production plant over several years are described, together with experience developed in plant operation. Data on recovery of hematite over the full range of size fractions treated is presented and compared with data in the literature. Potential methods of improving plant performance are discussed. KEYWORDS: Iron ore beneficiation, plant design, wet high intensity magnetic separation.

Jan 1, 1993

By Blaskett K. S

The physical properties of monazite and the minerals with which it is associated in east coast beach sands are studied and discussed. Possible means for improving recovery are examined. In particular, it is suggested that the practice of splitting the heavy mineral concentrate into two parts as a preliminary to treatment in separate "rutile" and "zircon" mills can be modified in a simple fashion by the addition of an extra magnetic separation stage so as to aid monazite recovery.INTRODUCTIONBeach sand deposits on the east coast of Australia are a major source of rutile and zircon, and supply 80 per cent or more of world requirements. The deposits also contain a small amount of monazite, averaging about 0.5 per cent of the total heavy minerals. Yet despite a steady price in recent years of £150-160 per ton, monazite production has remained relatively small.On the basis of rutiIe production, potential monazite production on the east coast was about 1,500 tons in 1956, but only 102 tons of high-grade concentrate was produced. The main deterrent to increased production has probably been the capital cost of the necessary equipment. Separate electrostatic and magnetic equipment and separate table concentration and drying constitute a considerable outlay for a by-product contributing 4 to 5 per cent of the revenue from rutiIe.It is the purpose of this paper to discuss monazite concentration generally and to suggest modifications to flowsheets, which may facilitate monazite production while maintaining the main objective, simple and cheap means for maximum rutile production.

Jan 1, 1958

By McCarthy P. L

With production at Olympic Dam scheduled to commence in June 1988, it has been necessary to install a complete underground ore handling system capable of removing two million tonnes of material annually from the mine. The system comprises mucking from stope drawpoints, truck haulage to underground storage, primary crushing, conveying and hoisting. On the surface, material is conveyed to stockpiles at the Metallurgical Plant. Mining of the numerous special excavations within the system was incorporated into the mine's ongoing development program prior to the underground construction phase. Utilisation of a decline from the surface enabled versatile construction equipment to operate at the underground sites as well as simplifying the transportation and delivery of construction Materials.

Jan 1, 1988

Recovery of Pentlandite and Chalcopyrite from Tailings by Bi-Cyclone Powerful Flotation Column

Sep 13, 2010

By Armoo K, Tran T

Cementation has been widely used in several hydrometallurgical processes for recovering gold and silver from process liquors. The process is electrochemical in nature, in which gold and silver are precipitated from the solution using less noble metals such as zinc, iron and copper. This paper summarises the current practice of gold cementation from cyanide liquors using zinc powder. Various developmental studies on the cementation of gold and silver from various systems (chloride and bromide) are also reported.

Jan 1, 1991

Recovery of Precious Metals from Chloride Solution by Magnetite The development of environmental-friendly lixiviants as alternatives to cyanide and the increasing recycling activities to recover precious metals from waste materials have prompted researchers to find efficient methods and alternative sorbents to recover precious metals, such as Au and Pt, from aqueous solutions. This study explored the use of magnetite (Fe3O4) as a sorbent to recover Au and Pt from chloride solutions. Magnetite is a semiconductor and has the capability of transferring electrons both within the solid state and across the solid-liquid interface and has the ability to reduce metal species on its surface. The sorption of AuCl4- and PtCl62- from NaCl solution on commercial grade magnetite powder was investigated by batch-sorption experiments. The effects of pH, contact time, NaCl concentration and precious metal concentration on the recovery were studied. Sorption experiments revealed that magnetite exhibited selectivity towards other metals with the strongest affinity for Au. Both Au and Pt recoveries showed similar pH dependence curves, with peaks at pH 6 - 7. At this pH range, a maximum of 4.4 ¦mol Au/g Fe3O4 and 3.0 ¦mol Pt/g Fe3O4 were recovered after 24 h at an initial metal concentration of 0.05 mol/m3. Increasing the initial precious metal concentration increased the Au and Pt uptake amount by magnetite. It was also observed that the recovery decreased with high NaCl concentrations. The SEM images of the magnetite particles after the treatment showed that gold, which was confirmed by the EDX analysis, agglomerated and were deposited on Fe3O4 surface.

Sep 13, 2010

Recovery of Rhenium from Mineral Raw Materials of Kazakhstan

Sep 13, 2010

By Parsons AE, Lewis IE

Selenium is a minor constituent of the raw materials processed by the Cockle Creek smelter of Sulphide Corporation Pty. Limited. Certain proposed raw materials contain higher levels of selenium, and these will create a problem with regard to the composition of liquid effluent discharged from the smelter. A method has been developed for removing selenium from the smelter circuit by the sulphation of the fume collected in the Acid Plant electrostatic-precipitators. It is planned that by treating 10-20% of the total dust arisings there will be a substantial reduction in the levels of selenium reporting in effluents when treating selenium-bearing raw materials. The selenium is recovered in elemental form.

Jan 1, 1978

Cyanidation tests by bottle agitation were conducted on an oxidized lead-silver ore containing approximately 12% Pb and 500 g/t Ag. Over 60% of the silver was leached when the ore was ground to 100% passing 0.15mm, whereas 50-60% of the silver was leached when the ore was only ground to 100% passing 0.5mm. Slurry handling and dewatering at the fine grind size were very difficult due to the high clay content of the ore, however this problem was greatly reduced at the coar- ser grind size. The silver recovery at the fine grind size was very dependent on cyanide solution strength with stronger cyanide sol- utions (containing more than 0.1% NaCN) re- sulting in lower solution silver contents. This was thought to be due to the precipi- tation of leached silver by zinc, which was also leached from the ore by the higher cyanide levels. There was a critical zinc level in solution of approximately 1000ppm, above which the silver recovery fell markedly. Cyanide and lime consumptions for this ore were high (typically 5 kg/t and 40 kg/t respectively), hence the processing of this particular ore by cyanide leaching is un- economic at current silver prices.

Jan 1, 1987

Some design and feasibility studies have been made into possible methods of economic recovery of sulphur from smelter gases emitted at Western Mining Corporation's Nickel smelter at KalgoorIie in Western Australia. The smelter discharges some 330 000 tonne/yr of sulphur dioxide in a stream of approximately 6 per cent by volume of S02.Three possible recovery methods have been considered:1. The manufacture of sulphuric acid using the contact process.2. The concentration of sulphur dioxide to a pure form using dimethyl aniline as an absorbing medium.3. The reduction to elemental sulphur using natural gas as the reducing agent.Design studies were carried out by final year chemical engineering undergraduate students at Royal Melbourne Institute of Technology (RMIT) in 1981 under the supervision of the author and other teaching staff of the chemical engineering department. These studies indicate that sulphuric acid manufacture is the economically preferred recovery route.Markets for sulphuric acid include several potential mineral processing applications and the established fertiliser industry.Sulphuric acid is currently manufactured in a number of plants located at coastal regions of Western Australia using imported sulphur. The sulphuric acid is then used for superphosphate or titanium dioxide pigment manufacture at adjacent plants.The possibility is considered of manufacturing superphosphate at Kalgoorlie from locally produced acid obtained from the smelter gases and from phosphate rock railed to the site from the coast. The advantages of such a scheme include improved scale and technology of operation compared with some existing coastal operations, use of a waste material to replace imported sulphur, and avoidance of costs and hazards associated with railing large quantities of sulphuric acid.

Jan 1, 1983

By Grinbaum B, Stevens G. W, Dalton R. F

Many hydrometallurgical processes produce large amounts of acid waste. This is illustrated by the copper S/X-E/W process which produces a bleed from the electrolysis tank house in the form of a concentrated sulphuric acid stream. Currently, the bleed stream is necessary to remove impurities such as iron that decreases the efficiency of the electrowinning process. However, this comes at a high financial cost owing to large quantities of make-up sulphuric acid required. A process has thus been developed with Zeneca Specialties (UK), ICI Australia and Gunpowder Copper Co at the University of Melbourne for the recovery of sulphuric acid from copper tank house bleed liquor. The recovery process uses the method of solvent extraction. Solvent extraction has the advantage of being a well-established technology in the mining industry and will reduce the environmental impact of the S/X-E/W process, as all additional chemicals can be completely recycled. A simplified diagram of the process is shown in Figure 1. Tank House

Jan 1, 1997