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By S. Ranganathan

Production of ferro-chromium in the submerged-arc furnace is highly energy-intensive. Several process rates have been attempted to reduce the specific energy consumption in the production of this alloy. An exhaustive understanding of the thermo-chemical features of the production process can be used to reduce the specific power consumption significantly. The knowledge can be used to reduce the energy drain through undesired reactions in the process can be minimised. Plant trials have shown that specific energy consumption up to 100kWh/tonne of hot metal can be obtained by adopting this strategy. The process requires modification in operation practice and does not require any capital investment unlike other processes adopted for reducing power consumption. The principle of thermo-chemistry involved in this process development and results of plant trials are discussed in detail.

Jan 1, 2007

By Hagen G, Uhlig H

In the Federal Republic of Germany the thermal environment in underground coal mines has deteriorated progressively over the past years. The number and capacity of refrigeration plants have increased correspondingly. In recent years high capacity spray coolers have been used in addition to conventional cooling coils.

Jan 1, 1988

The nickel refinery at Yabulu consumes an average of 118,000 kg/h of low pressure steam in stripping operations to-recover ammonia and carbon dioxide from product liquor and ore tailings streams. The steam used in the stripping operations is produced from the 25 MW power generation plant at the refinery. The back pressure on the non condensing steam turbines is maintained at 60 kPa.Studies have been carried out to determine the optimum amount of steam required to be generated for stripping operations. When steam production is increased, the product liquor flow from thewashing thickeners can be increased. This results in improved washing efficiency of the thickeners and therefore improves the plant nickel recovery. A computer thickener wash model was used to predict nickel recoveries at varying steam consumption rates.The studies have shown that in most plant situations, it is not economical to-generate more steam than that already available from existing power production.The studies also highlighted the necessity to use all steam generated from power production at all times to maintain maximum nickel recovery.

Jan 1, 1985

By B P. Wrench, N R. Steward, C Wu

A gold mine in southern China has transitioned from an open pit operation to a fully underground mine. An overhand mining method is used which requires the mined stopes to be filled with backfill having an unconfined compressive strength of at least 0.35MPa. Tailings from the process plant is deposited into separate flotation and Carbon in Leach (CIL) tailings storage facilities. The tailings has a high fines content and the thickener on site could only produce a dilute flotation tailings at a solids content of about 30 wt%. This resulted in two undesirable outcomes: the density of flotation tailings in the Tailings Storage Facility (TSF) was significantly lower than planned so that the TSF could not store the life of mine tailings and that the backfill product had to include a high percentage of manufactured aggregate and increased cement content to achieve the required strength. This paper presents results of a tailings study which identified that the flotation TSF would be capable of storing the remaining life of mine tailings if an amended backfill product made up of 100% of flotation tailings could be introduced. A 100% tailings backfill with the addition of 8% cement was developed by classifying the tailings and using the coarser fraction in the backfill. This backfill allowed 60% of the flotation tailings to be placed underground and only the remaining 40% in the flotation TSF. A new thickener was designed and installed to thicken the tailings and rheological testing of the tailings, using a closed loop pilot plant, provided parameters to upgrade the underground reticulation system to deliver cemented backfill to stopes throughout the mine and to deliver thickened tailings to the TSF. The revised design was implemented and has resulted in considerable cost savings. CITATION:Steward, N R, Wrench, B P and Wu, C, 2018. Optimisation of surface tailings storage and tailings backfill – a case study, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 454–464 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

Optimisation of the Advanced Automation System at Inmet Mining's PyhSsalmi Mine, Finland

Sep 13, 2010

By Torben L. Edens

Norddeutsche Affinerie is the largest copper producer in Europe. Two sites are in operation for the production of refined copper: primary copper production is located in Hamburg and secondary copper production in Lünen. A total of just under 560,000 tonnes of copper cathodes is produced annually at the two sites. Primary copper production using the flash smelter furnace has been in operation in Hamburg since 1972. Over the years, the capacity of the flash smelter has been increased in three expansion projects from 120,000 t to almost 500,000 t of anode copper p.a. The anode sector consisting of two anode furnaces and a 24 mould casting wheel still features just about the same equipment as in 1972. The paper describes the most important steps taken to optimise operation which have put NA in a position to cast meanwhile more than 500,000 t of anode copper annually with the two anode furnaces and 24 mould anode casting wheel.

Jan 1, 2006

The ball level and pulp load play a key role in the operation of industrial scale ball mills. These two operating variables have a significant influence on the performance of the ball mill and any attempts to optimize these devices require manipulating either one or both of these variables. There are several models that can be employed to perform simulations to optimize the performance of industrial ball mills. However, most of these models require ball load and pulp load measurements as inputs to calibrate the model. The changes suggested by the simulator also provide operating regimes that involve moving either the ball level or pulp load and sometimes both of these variables. A ball mill model coded in JKSimMet was used in conjunction with the Sensomag to optimize the secondary ball mill at Waterval concentrator in South Africa. The results show that the perfect mixing ball mill model can be used in conjunction with the Sensomag to optimize the performance of the secondary ball mill.

Sep 1, 2012

By J Bower

The JORC Code does not specify a definition for ælevel of confidenceÆ. The common approach in the mining industry is to use a certain drill spacing or number of data samples to classify resources by the confidence categories. Another approach includes application of the quantitative geostatistical criteria for non-subjective and consistent definition of the resource categories. This approach, which has become increasingly popular in the mining industry, provides a quantitative measure of the resource estimation uncertainty. In this study, uncertainties associated with the different drilling grids at the Weipa bauxite deposit (Queensland, Australia) have been modelled using conditional simulation methodology. Criteria used at Weipa for defining resource categories are as follows: mineralisation can be classified as a Measured Resource if tonnage and grade of its volume representing three months of production can be estimated with a precision error not exceeding ¦10 per cent (at 95 per cent probability, ie two standard deviations); and mineralisation can be classified as an Indicated Resource if grade and tonnage of its volume representing annual production can be estimated with a precision error not exceeding ¦10 per cent (at 95 per cent probability). Several different grids have been created by subsampling the drill hole database. The studied variables include Al2O3, SiO2, thickness of the bauxite seam and location of the footwall contact. The variables have been modelled by sequential Gaussian simulation. Based on this study the following drill spacings have been recommended: Measured Resources û 150 m by 150 m, Indicated Resources û 300 m by 300 m.

Jan 1, 2009

By S. J. Macnaughton

The Olympic Dam deposit in South Australia contains 2% copper and 0.06% uranium, which are present as bomite, chalcocite, chalcopyrite, uraninite, coffinite and brannerite. The copper is principally recovered using flotation. The flotation tailings, which contain the bulk of the uranium, and typically 0.3% copper, are leached using sulfuric acid and an oxidant at atmospheric pressure. This paper describes the results from a detailed investigation of the leaching behaviour of this ore. The influence of acid strength, Fe3+ concentration, redox potential and temperature on the dissolution of uranium and copper minerals was studied. Experimental data were obtained from extensive batch leach tests. Mineralogical studies, including SEM analysis of individual mineral grains, were also carried out. The maximum extraction of uranium determined experimentally was in good agreement with that predicted from SEM determination of the proportion of refractory brannerite in the ore. The leaching rate of copper was found to be independent of acidity over the range 20 g L-' to pH 2.0, but was strongly dependent on temperature and Fe3+ concentration. Uranium extraction showed a weak de endence on acidity, Arrhenius temperature dependence and a complex co-dependence on F>+ concentration and redox potential. A kinetic model has been developed, which predicts copper and uranium dissolution. This model will be incorporated into a more comprehensive model that can be used to predict leach performance and reagent consumption for specific ore samples.

Jan 1, 1999

By Gallet M, Watsford R. M S

The 1989 feasibility study of the chessy, france, copper, zinc, pyrite and barite orebody was based upon a classical flowsheet. The flowsheet included, three stages of crushing, conventional rod and ball mill grinding in series, and sequential flotation of four products utilising conventional mechanical flotation cells. To improve the profitability of the project the partners commenced a search for ways to reduce the projected capital and operating costs of the proposed development. This paper presents the improvements in the projected capital and operating costs as a result of design and concept changes. The major changes included: ò the simplification of the crushing section with the adoption of semi autogenous grinding (SAG) milling of primary crushed ore. ò the deferral of pyrite and barite flotation circuits until market conditions were positive. ò the simplification of the flotation circuit by the use of column flotation cells for combined rougher/cleaner duty. ò the purchase of second hand equipment. The result of these changes was to reduce the capital cost estimate of the plant from US$ 21 million to US$ 11 million and operating cost estimates from US$ 18.90 to US$ 13.60 per tonne of ore for a decrease of approximately ten per cent in potential revenues.

Jan 1, 1991

By B Cook, M Karakus, J Light, E K. Chanda

The aim of this research is to develop a model through which the appropriate paste recipe parameters, for example solid content and cement content, can be derived given the stope specific requirements such as stope dimensions, required strength and curing time. A large number of samples and uniaxial compressive strength (UCS) tests were conducted at the Angas Zinc Mine (AZM) paste plant and testing facility. Samples were varied in tailings solids content, cement content and sand content. The results of this testing were modelled to determine the UCS relationship to each of the parameters. The output of this modelling was an easy-to-use program that outputs the design mixture given the stope dimensions and required stope backfill-curing time. This allowed flexibility within the mining schedule for curing times of cemented paste backfill stopes. The optimised schedule using the stope specific mix design, when compared to the standard 14-day curing time used previously, resulted in a smoother weekly mine production schedule. For the modelled 12-week period (one quarter) the 14-day curing time would result in an eight-day shutdown of the mill. The cost for each shutdown is in the order of $125 000/d. Due to the unique nature of each mine’s tailings, this exact outcome would not be directly applicable to another paste fill site. However, the process in this report could be followed for another site to achieve a system similar to the output of this project. This would ultimately lead to a more optimised backfill and stoping cycle and a more optimised mine.CITATION:Chanda, E K, Karakus, M, Light, J, Cook, B and Harris, A, 2014. Optimisation of the paste fill cycle at Angas Zinc Mine, Strathalbyn, South Australia, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 259–264 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By J Hartog, V Beehan, D Beeson

Treating an ore blend from five different deposits at Newgold’s Peak Gold Mine (PGM) has always been a challenge due to the inherent variability between each of the orebodies.Operating the processing plant at PGM with a ‘normal’ level of process control and a good operations and metallurgical team has allowed the processing plant to achieve above the name plate design at 102 t/h and recoveries in the order of 92.8 per cent for gold and 88.4 per cent for copper.With a company value and culture of striving for continuous improvement, PGM implemented the Manta Cube® system, an advanced process control system, for the semi-autogenous grinding (SAG) mill, gold leach circuit and tailings thickener over a 12-month period throughout 2012 and 2013.The overall benefits of implementing the Manta Cube® system has led to an increase in mill throughput of at least two per cent, improved SAG mill liner life, a ten per cent decrease in cyanide consumption and a reduction in flocculant of at least 30 per cent with no detriment to the gold and copper recoveries.The entire project had a payback period of less than two months, and the combined benefits resulted in an overall increase in revenue of at least $5 million per annum. This is quite significant for a processing plant the size of PGM.This paper describes the methodology, overall results and economic benefit of applying the Manta Cube® system on-site at Newgold’s PGM.CITATION:Hartog, J, Beehan, V, Karageorgos, J and Beeson, D, 2014. Optimisation of the Peak Gold Mine’s processing plant through advanced process control, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 335–344 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

Optimisation of the Performance of Flotation Circuits Using a Genetic Algorithm Oriented by Process-Based Rules

Sep 13, 2010

By S. H. Castro

The El Teniente molybdenite plant (Chile), with a nominal capacity of 3,200 tpd of bulk Cu-Mo concentrate, produced approximately 3,385 metric tonnes of Mo during 1998 by using the phosphate Nokes process (LR-744 reagent). This is the only Chilean moly plant which traditionally has depressed copper and iron sulfides with LR 744 reagent, which is a hydrolysis product of P2S5 and NaOH. El Teniente's ore Mo grade is 0.02%Mo and is concentrated to 0.45%-0.50% Mo in bulk Cu-Mo flotation. Subsequently, molybdenite is floated by selective flotation to depress Cu and Fe sulfides, such as chalcopyrite, bornite, digenite and pyrite, to produce typically a 49%-50%Mo molybdenite concentrate. An optimisation program during the last two years was undertaken to improve molybdenite recoveries. Current molybdenite plant technology at El Teniente and use of phosphate Nokes in the process process is reported. Use of phosphate Nokes reagent has increased average Mo recovery from 70.5% in 1977 to a mean 74.7% for the first 4 months of 1999. The global Mo recovery, including bulk flotation, incresed from 42.3% in 1997 to 46.8% in 1998, with a projection of 55.5% for 1999. Mo production per year has increased from 3,181 tons in 1997 to 3,385 tons in 1998 and a projected 4,000 tons for 1999.

Jan 1, 1999

By P Woodward, T Ly, N Muhamad

OZ Minerals' Prominent Hill copper-gold concentrator is located 650 km north-west of Adelaide in the Gawler Craton of South Australia. The concentrator was built in 2008 and commenced commercial production in early 2009. The Prominent Hill concentrator is comprised of a conventional grinding and flotation processing plant with a design capacity of 8 Mtpa. The flotation circuit includes six rougher cells, an IsaMill for regrinding the rougher concentrate and a Jameson cell preceding the three stage conventional cell cleaner circuit.Since its commissioning, the Prominent Hill flotation circuit has been the focus of a series of projects aimed at continual performance improvement. This has been achieved by a systematic approach of project identification, evaluation and implementation which has led to a progressive improvement in flotation performance. Innovations that have been successfully implemented include aspects of cell design and operation, and reagent suite optimisation.Improvements to the mechanical cells initially focused on the installation of a modified rotor-stator design in the rougher circuit, as recommended by Outotec. Isolated cells were fitted with this technology and proven to provide a benefit for copper recovery before the design was trialled in subsequent cells along the bank. Based on this success, further modifications were made to other aspects of the rougher and cleaner cells.Laboratory trials conducted at Prominent Hill involve a screening process identifying potential reagents from a list of products, which are then rigorously tested in the laboratory before one is selected for trial in the concentrator. Where these reagents have demonstrated a statistical improvement in performance they have typically been incorporated permanently into the flotation circuit. This paper discusses the process of continual improvement at the Prominent Hill flotation circuit and the optimisation achieved by the implementation of key initiatives.CITATION:Woodward, P, Muhamad, N and Ly, T, 2013. Optimisation of the Prominent Hill flotation circuit, in Proceedings MetPlant 2013 , pp 431-444 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

By Mark Weidenbach

OZ Minerals owns and operates the Prominent Hill copper-gold concentrator in South Australia, built in 2008 and commissioned in early 2009. The concentrator operates via semi autogenous grinding in a thirty four foot ThyssenKrupp Polysius mill followed by a twenty four foot Polysius ball mill in closed circuit with hydro cyclones, designed for an ore throughput of eight million tonnes per annum. The main aspects of designing the circuit are discussed along with the issues faced in the early stages of operation. A review of the optimization strategy applied to the milling circuit such as cyclone optimization, advanced process control strategy implementation and reline management to improve availability is then discussed highlighting areas of significant process improvement that have seen the plant exceed nameplate throughputs above nine million tonnes per annum.

Jan 1, 2012

By Brett Triffett, Mark Weidenbach, Christopher Treloar

OZ Minerals owns and operates the Prominent Hill copper-gold concentrator in South Australia, built in 2008 and commissioned in early 2009. The concentrator operates via semi autogenous grinding in a thirty four foot ThyssenKrupp Polysius mill followed by a twenty four foot Polysius ball mill in closed circuit with hydro cyclones, designed for an ore throughput of eight million tonnes per annum. The main aspects of designing the circuit are discussed along with the issues faced in the early stages of operation. A review of the optimization strategy applied to the milling circuit such as cyclone optimization, advanced process control strategy implementation and reline management to improve availability is then discussed highlighting areas of significant process improvement that have seen the plant exceed nameplate throughputs above nine million tonnes per annum.

Jan 1, 2012

By A Jakowetz

OxianaÆs Sepon Copper Operation (part of the combined Sepon Gold and Copper Operation) is located in Savannaket Province, in southern central Lao PDR. The site is relatively remote, being some five hours inland from the Laos/Thailand border crossing on the Mekong River. The copper processing plant was designed to treat 1.3 Mtpa ROM ore to produce 60 000 tpa of copper cathode. Commissioning of the processing plant was completed in mid March 2005 with ramp up completed in January 2006 with 5260 tonnes of cathode harvested in this month. The focus of this paper is on SeponÆs unique grinding circuit. The entire grinding circuit is operated at highly acidic conditions, adding a number of different operational issues over and above those normally associated with grinding circuits. It became evident early in the ramp up, that reaching design mill throughput of 170 dry tonnes per hour with the original design would be challenging, coupled with the realisation that ceramic grinding media consumption had been underestimated. In order to lift the overall throughput of the processing plant, a significant continuous improvement process has been implemented within the grinding circuit to lift this key unitÆs production capacity. This presentation will focus on the key areas that have been addressed in the 12 month ramp up period and highlight areas for future work. Results to date have shown this improvement process to be very successful, with throughput rates in excess of 200 tph regularly achieved.

Jan 1, 2006

By C M. Ramotsabi, F Bornman, W Erasmus

Spiral concentrators are gravity based separation devices used for the preparation of coal, iron and heavy mineral ores (Wills, 1992). Gravity concentration remains the main separation method for fine iron ore and is used extensively for treating many minerals. Gravity concentration methods separate minerals of different specific gravity by their relative movement in response to gravity and one or more other forces, the latter often being the resistance to motion offered by a viscous fluid such as water or air (Wills, 2006).The main operating variables for a spiral concentrator are the feed rate, the slurry concentration, the wash water addition and the position of the splitters used to separate the tailings, middlings and concentrate streams (Sadeghi, Bazin and Renaud, 2014). Wash water is an important control variable used to wash away entrapped light minerals from the concentrate stream (Burt, 1984; Bouchard, 2001). Sadeghi et al concluded the spiral operation to be sensitive to the wash water flow addition and resultant economic trade-off. Increasing the wash water flow rate increases the concentrate grade at the expense of the iron recovery.The paper addresses the effects of dripping wash water onto the concentrate stream of a spiral processing fine iron ore. The dripping wash water system offers a low flow alternative and could possibly provide a good balance between grade and recovery of valuable minerals without using excessive water. The Multotec SC20HC/7 WW spiral was used for the experimental test work. The spiral was tested at varying solids concentrations; the wash water rate and the splitter positions were kept constant.The use of wash water at 30 per cent, 35 per cent and 40 per cent solid concentration improved the Fe grade in the concentrate by about 1.5 per cent, six per cent and eight per cent respectively. The additional SiO2 rejected by the wash water is approximately three per cent, nine per cent, and 11 per cent at 30 per cent, 35 per cent and 40 per cent solid concentration respectively. Addition of wash water is shown to have minimum effect on grade at 30 per cent solid concentration, and Fe upgrades obtained with and without wash water were found to be similar. Notable upgrades were observed at 35 and 40 per cent solid concentration with wash water. The results obtained suggest that the additional SiO2 rejection by wash water increases with an increase in solid concentration. This suggests that the benefit of using the drip system is notable at higher solid concentration. It is believed that at lower solid concentration SiO2 particles are reasonably free to move around and are possibly less entrapped than at high solid concentration.The Fe grade/recovery and solid concentration relationship with wash water showed a non-linear correlation. The Fe grade increases with an increase in solid concentration up to a somewhat optimum point and then drops as solid concentration increases further. The recovery of Fe on the other hand follows a converse relationship with solid concentration, the recovery decreases from 30 per cent to 35 per cent solid concentration and then increases again at 40 per cent solid concentration. The increase in recovery at higher solid concentration could possibly be explained by the fact that the mass yield to concentrate also increases with solid concentration. Since recovery is a function of both grade and mass yield, the higher recovery at 40 per cent solid concentration is likely due to the high mass yield obtainable at 40 per cent solid concentration.The drip system could therefore possibly offer a safe guarding against grade and solid concentration variations without using excessive water, and possibly ensure that a product within specification is maintained.CITATION:Ramotsabi, C M, Erasmus, W and Bornman, F, 2015. Optimisation of wash water on an iron ore spiral to improve grade, in Proceedings Iron Ore 2015, pp 175–184 (The Australasian Institute of Mining and Metallurgy: Me

Jul 13, 2015

By B Davis, A H. Benzer, A N. Mainza, C M. Evertsson, L M. Tavares, M M. Hilden, M S. Powell, C Rule, G Asbj÷rnsson

The Mogalakwena North Concentrator of Anglo Platinum Operations was built with the capability to allow for future projected expansion. Several research groups, currently working with Anglo American on collaborative research over a range of projects, were brought in to review the existing operation, with a view to releasing some specific bottlenecks and providing specialist design input to the potential plant expansion.Based on a thorough review of the high pressure grinding rolls (HPGR) and crushing circuit, a wide range of de-bottlenecking opportunities were identified. This was followed by a series of surveys to provide performance measurements and to model the circuit in detail. From this effort a full simulation of the plant has been developed, including updated models of the HPGR, crusher, bins and screens. These models have been implemented in the researchersÆ Model Development Kit (MDK) version of JKSimMet, and in a prototype dynamic simulator based on the Matlab Simulink software.The observations, measurements and simulations have provided guidance and quantified the benefits of some proposed circuit modifications. Based on these findings a number of upgrades are being planned, with the aim of ensuring full utilisation of the installed equipment. The circuit demonstrates a potential expansion of over 30 per cent through minor improvements, and a 70 per cent increase in capacity with the addition of some strategic equipment.The HPGR-crushing upgrade project demonstrates the potential benefits of nurturing disparate Research groups into a strong inter-disciplinary technical team capable of supporting operational needs while developing future technology.CITATION:Powell, M S, Hilden, M M, Evertsson, C M, Asbj÷rnsson, G, Benzer, A H, Mainza, A N, Tavares, L M Davis, B, Plint, N and Rule, C, 2012. Optimisation opportunities for high pressure grinding rolls circuits, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 81-94 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012