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The process of hydraulic transport through pipelines, worldwide well known for the transport of fine-grain solids is also being used to a greater extent in mining for the handling of primarily coarse-grain solid û water mixtures for the removal of, for instance û coal and waste mixtures after coal winning using hydraulic monitors, tailings for backfilling underground and washer tailings and flotation waste to the dump. A large-scale test unit was taken into service at the Bergbauforschung GmbH (Mining Research Company) in Essen-Kray, GFR, to further improve this technology. The unit was planned and built by Krupp Industrie- und Stahlbau of Duisburg in cooperation with Bergbauforschung GmbH. The paper deals in detail with the demands made on this test unit and the technical results obtained on it and with the planned experimental research work.

Jan 1, 1980

By Kanhaiya Mishra, Phanil. K. Mandal, S. K. Behera, P. S. Paul, Prashant Singh, C. N. Ghosh

The techno-economic and social benefits of cemented paste backfill (CPB) resulted in its wide acceptance by the mining industry. The Ordinary Portland Cement (OPC) remains the key binder but to diminish its economical constraints, suitability of alternate binders has been examined worldwide. The present study aimed to investigate the effect of partial replacement of OPC with fly ash on the CPB’s strength and to determine the most optimal mix to achieve the required strength (1 MPa at 28 days of curing) at the most cost-effective way using the Bayesian network (BN). The CPB mixes were prepared at 72 wt.% solid concentration with mill tailings (87–91%), OPC (6–13%), and fly ash (0–4%), and instantly after mixing, fresh (slump, bleeding, density) CPB properties were measured. The strength was tested at 7, 14, 28, and 56 days of curing and initially analysed through traditional model. The traditional models follow the aleatory principle and are considered not appropriate for geotechnical engineering. Hence, the BN model was developed and tested. The reliability of two classifiers in learning model structure was compared which gives Naïve Bayes as the highest reliable tool. The CPB’s strength is most sensitive to the OPC content. The most consistent mix(s) is mill tailings: 87–88%, OPC: 9–11%, fly ash: 1–4%. Adding fly ash at 89–91wt% mill tailings possesses high failure probability of the CPB. The collinearity test indicates that the fines percentage and chemical composition of CPB’s ingredients are highly correlated with its slump, bleeding, and strength development.

Jul 14, 2022

By Sarra Scherbakova, Alexander Kurkov, Andrey Egorov

"Deep concentration of iron ores involves extraction, in addition to magnetite concentrate, of hematite and martite concentrates and recovery of other valuable components. A flotation technology was designed specifically for finely disseminated hematite ore, allowing producing saleable hematite concentrate containing 64-66% Fe. The main process challenge is caused by the presence of other ferruginous minerals such as carbonates, silicates, and phosphates, and fine hematite and martite mostly concentrated in the -0.044 mm fraction. Main operations of the processing circuit are removal of slime, flotation to remove impurities (carbonates, phosphates etc.) with an anionic phosphorous collector in an alkaline environment, hematite and martite flotation using same collector in a slightly acidic environment to produce commercial iron concentrate. Tailings of wet magnetic separation of magnetite containing 24.5% Fe and finely disseminated hematite ore containing 51% Fe were processed to produce iron concentrates with a yield of 15% and 50%, respectively, containing approximately 65% Fe and 5% of silica. Hematite flotation tailings had the highest content of gold (up to 0.09 - 0.11 g/t), while in other products it never exceeded 0.02 g/t. Using a Knelson concentrator was able to recover 40 to 45% of the gold into 3.31 g/t Au gravity concentrate. The success of gravity concentration is due to high gravitational contrast of gold to mostly quartz and aluminosilicates in the hematite flotation tailings. In general, the proposed process can be used as a solution to the challenge of selective concentration of finely disseminated iron minerals."

Jan 1, 2014

By K. Eckert, A. E. Sommer, S. Heitkam

One highly relevant challenge in flotation is the recovery of fine particles. Due to their low inertia, these particles are mostly pushed aside by rising bubbles. Consequently, they typically exhibit low probability for bubble-particle collision and thus, a poor recovery rate. In this work, the trajectories of fine particles in the vicinity of rising bubbles were investigated. The measurements considered a chain of millimetric bubbles in a rectangular container filled with deionized water. The flow field near the bubbles was measured by tomographic particle image velocimetry (TPIV), employing fluorescent tracer particles of 33μm diameter. Subsequently, trajectories of larger particles and bubble-particle collision events are recorded with 4D particle tracking velocimetry using a high temporal and spatial resolution. The results reproduce the well-known collision of particles on the leading edge of a rising bubble. Additionally, collisions on the tailing edge were observed in cases with a low Stokes number. The TPIV results demonstrate, that the high turbulent kinetic energy in the bubble wake allows particles to divert from the fluid streamlines and collide with the tailing edge of the bubble. The tailing edge collision probability increases with the bubble Reynolds number and decreased particle inertia. Overall, the investigation shows that the collision of fine particles in the bubble wake should be considered for the development of further collision probability models. Furthermore, the importance of turbulence on the fine particle flotation was demonstrated. Keywords: 4D Particle tracking velocimetry (PTV), tomographic particle image velocimetry (TPIV), fine particle flotation, collision probability, bubble-particle interaction, wake induced collision

Jan 1, 2020

By R Van Laeren, G Langerak, S Fityus, K Koosmen

Co-disposal of partly-dewatered mixed-plant rejects (MPR) from coal processing into spoil dumps is gaining favour amongst Australian coalmine operators in preference to conventional tailings disposal methods which require a tailings dam. Owing to their often low permeabilities and high degrees of saturation, MPR materials may be prone to the development of excess pore pressures when covered by spoil during routine dump advancement. The likelihood of dump failure will increase if these pore pressure regimes and the associated delay of shear strength development are not accounted for during waste dump design. To date however, little is known about these pore pressure regimes due to the complexities associated with their in situ measurement.This paper describes the in situ measurement of pore pressures in a cell of MPR placed within a coalmine spoil dump. Four vibrating wire piezometers (VWPs) were placed within a cell that was then filled with MPR and allowed to consolidate under self-weight before advancing an adjacent tiphead across the cell. The VWPs recorded the transient pore pressures during MPR placement and dump construction, and their interpretations are described within the context of laboratory measured parameters. Following analysis of the measured pore pressures, their implications on shear strength and dump design are also discussed.CITATION:Koosmen, K, Langerak, G, Simmons, J, Fityus, S and Van Laeren, R, 2015. Field measurement of transient pore pressures within co-disposed processing rejects placed in a spoil dump, in Proceedings Tailings and Mine Waste Management for the 21st Century, pp 63–72 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By D Osborne

Current mineral extraction and valorisation activities still have a strong focus on the ‘delivered product(s)’ which might be a bulk commodity such as metallurgical coal, hematite, bauxite, etc., and mineral concentrates such as metal sulphides, oxides, and carbonates; precious metals, such as gold, silver and PGMs, REEs or battery metals; halogen salt forming elements; or industrial minerals, which are non-metals including crushed rock, sand, and gravel etc. With the current focus on tailings disposal the pushback has been to seek alternatives to the currently ‘conventional’ disposal method via increasingly higher risk tailings settling facilities (TSFs) with a changing focus towards ‘dry disposal’. Whichever option is selected, the alarming trend is the increasing cost and ESG impacts. An emerging approach is to pose a focusing question ‘Is there anything else passing into the tailings stream and/or disposable waste streams which could be recovered and pass into the product revenue stream? All of the above ‘primary’ process streams offer opportunities for increasing recovery and offsetting some of the disposal costs. Some secondary opportunities are already being pursued, such as REEs and battery metals from base metals deposits, but other opportunities are still being overlooked, such as building materials and soil conditioning pellets. These potential ‘new’ product streams can also utilise other mining and domestic wastes such as smelting slags, glass waste, plastic waste etc. This paper will address these opportunities and provide examples including process approach, economics and some of the current barriers that exist, which are holding back progress on such initiatives.

Sep 1, 2024

By Ye Junjian

The feed ore is the concentrate of low intensity magnetic separation of pyrite tailings, containing TFe47.47%, S19.41% and SiOpyrrhotite. The exploratory tests showed that using a process of flotation followed by magnetic separation could respectively obtain a sulphur concentrate of S 31.08% with a recovery of 82.89%, and an iron concentrate of TFe 59.97% and S 9.52%with a recovery of 30.93%. The reason of iron concentrates with high sulphur grade was that the sulphur grade of flotation tailings was 7.03%. So enhanced flotation for pyrrhotite was investigated in this research. Results show that when the natural pH of ore pulp varied from 5.5 to 6.0, a reagent combination of FS, LH, and BJ displayed the optimum flotation performance for pyrrhotite. After comparisons made on different flotation processing, the previous flotation flow of ?one roughing, one cleaning, one scavenging system and middling reprocessing?, was found, among others, to have the best flotation effect, as the grade and recovery of sulphur concentrate were 31.08% and 82.89%, respectively, and sulphur content of the flotation tailings was 7.03%. The above results indicated it was difficult to raise further the effect of flotation desulphurization. The authors therefore suggested that oxidizing roast should be adopted following the magnetic process for the purpose of further desulphurization. Besides, flue gas containing SO2 8.25%. And the valuable minerals are magnetite and 2 generated from roasting could be reused for sulphuric acid production. Keywords: pyrrhotite, enhanced flotation, desulphurization

Sep 1, 2012

By Boden A

A procedure was developed for recovering metal values from copper-smelter fume. The fume is first leached to neutrality with dilute sulphuric acid to obtain a solution of zinc sulphate containing a low level of impurities. The tailings from this leach are treated with hydrochloric acid in two stages, in which tin and zinc are quantitatively leached and recovered as a tin concentrate and a zinc sulphatelzinc chloride solution.About 80 per cent of the lead sulphate is converted into lead chloride, which can be recovered as a pure product.Data on the effect of temperature, acid concentration, aeration, and solid/liquid ratios on the hydrochloric acid leach are presented, and the fate of the most important metals during the recovery process is described. Alternative treatments of the tailings from the neutral sulphuric acid leach were briefly investigated.

Jan 1, 1977

By Peter W. Willcocks

The Cyanide Regeneration Process (CRP) was developed by Golconda Engineering and Mining Services and installed at Golconda's Beaconsfield Gold Tailings Retreatment Project situated in Tasmania, Australia. The Tailings Project is a 1000 ton per day carbon-in-pulp treatment plant, feed for which consists of tailings material which is dredged from the Tamar River. The pulp is screened to remove debris, then cycloned to form two products. Underflow pulp is stored in a 350 ton bin and cyclone overflow passes to a thickener. The coarse pulp material is fed to the grinding circuit then rejoins the thickened slimes in the cyanide leach circuit. Gold is extracted by carbon-in-pulp techniques. The plant is sited in rolling, hilly country approximately 200 meters from the Tamar River. The residue disposal area is quite small consisting of four dams, totalling 42 acres. The regional climate is cool and wet with little evaporation; therefore, we anticipated a net increase in contained liquor, and subsequent storage problems. Our plan was to remove cyanide from the liquor prior to returning the detoxified solution to the river system. Such treatment was not required until midway through the second year of the project. Initial testwork on sample material indicated that destruction of cyanide by chlorine was the simple solution. Subsequent follow-up on actual plant liquor indicated a huge chlorine requirement due, we believe, to the substantial degradation of vegetable matter plus organic content from the marine environment. The cost of cyanide destruction could not be justified.

Jan 1, 1998

By E. B. Thornhill

Discussion of the paper of E. B. THORNHILL, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 104, August, 1915, pp. 1653 to 1670. D. B. HUNTLEY, Oakland, Cal.-About 15 years ago it chanced to be my lot to cyanide some mill tailings, assaying about $5 per toll in gold and, a few cents in silver. It was in southern Idaho, a. desert region where costs are nearly as large as in central Nevada. We knew that we had a lot of tailings that had been amalgamated and contained about 1 lb. of quicksilver per ton. We knew that we would dissolve some of the quicksilver when we cyanided them with say 0.25 per cent. cyanide solution. We figured (optimistically) that the cyanide would dissolve all of the quicksilver, and we thought to capture all of that, 1 lb. per ton; and we figured out profits as partly gold, partly silver, and partly quicksilver from the recovery in working those tailings. Actually, however, we did this: We had a battery of three large, old-fashioned silver-mill retorts. We retorted our zinc-dust precipitate, recovering about ¼ lb. of quicksilver per ton treated. That was only a loss of 75 per cent. of what we originally had, but still it was profitable by virtue of the conditions. I mention it partly to call attention to the crude old way of doing it, and partly to call attention to how chemical methods for recovering quicksilver have developed in the last 15 years.

Jan 12, 1915

By Robert G. Beverly

The presence of radioactivity in uranium mill wastes has resulted in somewhat unique waste disposal methods. In addition to the common problems of disposing of large quantities of solid wastes, neutralizing acids, minimizing dissolved heavy metals, and clarifying all liquid effluents, the uranium mill operator must sample and analyze liquid effluents for micro-micro quantities of radionuclides such as, radium-226, thorium-230 and lead-210. Special disposal methods or decontamination procedures are required to meet the stringent limitations established by the U. S. Atomic Energy Commission on the concentration of radionuclides in liquid effluents. Routine monitoring of the receiving stream is also normally required. In recent years there has been a growing concern about the radioactivity remaining in the uranium mill tailings, particularly at the sites of shut-down or dismantled mills. The State of Colorado, in cooperation with the industry, has adopted regulations for stabilization of inactive uranium mill tailings. The Mining and Metals Division of Union Carbide operates mills for the recovery of uranium and vanadium at Uravan and Rifle in Colorado and a mill in the Gas Hills of Wyoming for uranium production. The company also owns former mill sites on which mill tailings are located at Maybell, Rifle, and Slick Rock in Colorado and at Green River, Utah. The purpose of this paper is to describe the methods used by a few uranium mills, largely those operated by Union Carbide Corporation, to dispose of radioactive wastes, to discuss briefly the regulations under which the mills operate, and to present reference information on the subject.

Jan 1, 1968

By Abishev D. N, Balakirev V. F, Davydov G. I

The problem of recovery of cobalt products from poorly oxidised nickel ores (Kempirsaisk deposit) and pyrite concentrates with a low content of cobalt (Sokolovsk-Sarbaisky deposit) has not been solved to the present time in Kazakstan. The balance reserves of nickel ore of the Kempirsaisky deposit make up more than 35 million tons with a nickel content of 1.13 per cent and cobalt content of 0.05 per cent; besides there are 110 millions tons of out-balance ores (Anon, 1995). In a tailing dump of Sokolovsk-Sarbaisk Mining and Productional Association more than 320 million tons of tailings of wet magnetic separation have accumulated, which contain 7.2 per cent of sulphur, 0.022 per cent of cobalt, 0.012 per cent of nickel, 0.094 per cent of copper, 0.079 per cent of zinc, 16.45 per cent of iron, and from which a high quality pyrite concentrate car) he separated through flotation, containing 48 - 50 per cent of sulphur, 44.15 per cent of iron, 0.141 per cent of cobalt, 0.07 per cent of nickel, and 0.10 per cent of copper. The well-known technological flow charts for cobalt and nickel recovery from the tailings of sulphide-magnetite ores dressing (to give pyrite concentrate) and from oxidised nickel ores, are not profitable at the existing prices for metals, but are not worthless and require significant capital investments for the realisation of their value. The flow chart `sulphate roasting - leaching - solution processing' is used widely for combined processing of cobalt-containing pyrite ores and concentrates.

Jan 1, 1996

By E. A. H. Tays, F. A. Schiertz

In 1893 the Anglo-Mexican Mining Company, Limited, purchased the Guadalupe and adjoining mines, and, after opening up the Guadalupe, constructed a 20-stamp mill, which was started in August, 1894. Although the results were extremely profitable (the ore being very rich and carrying much free gold) they were not satisfactory, because the tailings were found to average about $20 per ton in gold. Early in 1895 the late Mr. James E. Mills made for the company a very thorough study of the subject, and decided that the tailings could be treated most advantageously by the cyanide process. In view of this opinion, Mr. Henry R. Batcheller, of Boston, was sent down in the winter of 1895-96 to experiment on a working scale. Mr. Batcheller was successful; and, during 1896-97, built a cyanide-plant with a capacity for treating 1500 tons per month. This was run until May, 1899, when all the sands suitable for the process had been treated. The mill was stopped about the same time for lack of ore; but no tailings from the mill had gone to the dump since October, 1897, as the values they contained after that date were far too low to warrant further treatment. When the cyanide-plant was stopped there still remained on the dump about 16,000 tons of slimes, consisting of about 75 per cent. of clay-slimes and 25 of sand-slimes. The successful treatment of this very unfavorable material is the theme of this paper. Fig. 1 is a view of the hacienda, taken in 1897.

Jan 1, 1902

By J. Shawn Peterson, David S. Yantek, Adam K. Smith

Occupational noise-induced hearing loss continues to be one of the most pervasive health problems in the mining industry, despite over 25 years of regulation. One of the loudest pieces of equipment used in underground mining is the continuous mining machine. Noise sample data collected by the Mine Safety and Health Administration indicate that 42% of noise overexposures between 2000 and 2005 involved continuous mining machine operators. Previously conducted field and laboratory tests have determined that the continuous mining machine conveyor system is a dominant noise source. Loud impacts occur as the conveyor chain flight bars, used to move the mined aggregate, traverse their path from the top to the underside of the conveyor deck. Various noise control treatments have been applied to abate noise caused by the conveyor system. A durable polyurethane coating has been developed for the conveyor flight bars, resulting in a time-weighted average reduction of 3 dB(A) for an eight-hour work shift. In an attempt to further reduce continuous mining machine operator overexposures, a similar urethane coating has been applied to the tail roller component of the conveyor system. Laboratory results showed a 2 dB(A) reduction in sound power levels, but the component failed during underground durability testing. An outer steel sleeve has been added to the urethane coating of the tail roller to enhance wear resistance during mining. The urethane jacketed tail roller is the latest effort, combined with previous noise treatments, to bring the continuous mining machine into compliance with federal noise regulations.

By D. Denton, L. Buckingham

The dry grinding and leaching circuits of the Barrick Gold’s new Roasting Facility at Goldstrike Mine in Nevada, USA incorporate well proven technologies in a unique combination. During the design of the plant, every effort was made to utilize the best available technologies from the perspective of capital, environment and operating costs. Crushing is a two stage process using gyratory and cone crusher. Ore is stockpiled and then reclaimed to two parallel dry grinding mills. The double rotator dry grinding mills include a drying chamber, coarse and fine compartments and utilize a gearless motor drive (GMD) or “wrap-around”. In-circuit product transfer is accomplished with air slides and bucket elevators. Finished product separation is performed by both static and dynamic classifiers with the final product captured in baghouses. Recycle of unfinished product utilizes cascade chutes. The final ground product reports to two 2000 ton roaster feed silos. The two-stage oxygen flu-solids roasters with their associated gas train are the subject of a separate paper. The cyanide leach circuit is a standard carbon in leach system with circular swept screens for interstage carbon retention. The loaded carbon is sent to the existing Goldstrike autoclave plant for stripping and regeneration. Tails, after cyanide destruction, are deposited in the existing tailings dam with tails from the autoclave facility. Throughout the circuit, Clean Plant Design criteria were used to assure environmental compliance, ambiental acceptability and ensure worker health and safety. The commissioning and operation of the double rotator grinding technology has been very successful at Barrick Goldstrike.

Jan 1, 2001

By Peter Dunn

London Mining Plc. is developing an iron ore project in Sierra Leone. The Marampa mine is a brownfield development effort, having been operated by others since the 1930?s until low iron ore prices forced closure. Peak production of the prior operations was approximately 2.5 Mt/a in the 1960?s. Following civil war In Sierra Leone, London Mining obtained the mining license in 2006 and commenced activity to redevelop the Marampa mining and processing operations. There are two ore deposits of interest for development as well as a resource comprised of tailings from previous operations. Phase 1 will be the mining of the tailings resource with a 1.8 Mt/a plant producing 66 per cent Fe sinter feed, expandable to 3.6 Mtpa. The first phase is currently under construction with start-up expected in the third quarter of 2011. Phase 2 will include development of the two ore deposits. These deposits are weathered secondary ore and the main primary ore resource. Phase 2 is currently under prefeasibility study and targets a final 16 Mt/a concentrate production in 2016. Outotec partnered with London Mining on testing efforts to develop the magnetic separation flowsheet using SLon VPHGMS machines to process the Marampa material. This paper explains the flowsheet development process for the weathered and primary ores commencing with sample characterization and detailed parametric investigations to evaluate SLon performance. The final flowsheet configuration is explained, including expected circuit recoveries and concentrate quality, clearly demonstrating the performance of the SLon VPHGMS circuit in iron ore tailings treatment applications.

May 1, 2011

By G. Naud, J. Robitaille, M. Buonvino, A. Veltmeyer, A. R. Laplante

"A B6 Falcon concentrator was tested on several systems: a synthetic magnetite-silica ore, a massive sulphide gold-copper ore, a fine gold-pyrite-silica flotation tail, and a high grade gold ore in a low sulphide gangue (cyclone overflow).The Falcon demonstrated its ability to achieve separation in all cases, but with varying degrees of success. For the magnetite-silica system, separation proceeds until the concentrate bed overloads, at a mass of 300 to 400 g of magnetite. Recovery increased with decreasing particle size, down to the 25-37 µm and in some cases even the -25 µm class. Bowl geometry (angle) and gangue size distribution were the most significant operating variables. Results with the massive sulphide ore were the poorest, with gold recoveries typically around 10 to 20% at very low upgrading ratios, 1: 1 to 5: I. Gold recovery from the fine flotation tail ranged from 30 to 70% at upgrading ratios of 5 to 15; actual plant application would require a continuous unit (Falcon C20). The high grade gold cyclone overflow yielded a wide range of gold recoveries, but a very consistent relationship between recovery and mass processed. A concentrate of 0.5% Au could be processed at a recovery of 50%.The paper also presents results of a C20 application to recover fine iron from a magnetic tail. The unit is capable of recovering approximately 70% of the fine liberated iron in a high density product, with modest upgrading below 105 µm, and virtually complete rejection of the + 300 µ m product."

Jan 1, 1993

By Dayle Denton, Lloyd Buckingham, Kenneth G. Thomas, Jacques McMullen

"The dry grinding and leaching circuits of Barrick Gold’s new Roasting Facility at Goldstrike Mine in Nevada, USA incorporate well proven technologies in a unique combination. During the design of the plant, every effort was made to utilize the best available technologies from the perspective of capital, environment and operating costs.Crushing is a two stage process using gyratory and cone crusher. Ore is stockpiled and then reclaimed to two parallel dry grinding mills. The double rotator dry grinding mills include a drying chamber, coarse and fine compartments and utilize a gearless motor drive (GMD) or “wrap-around”. In-circuit product transfer is accomplished with air slides and bucket elevators. Finished product separation is performed by both static and dynamic classifiers with the final product captured in baghouses. Recycle of unfinished product utilizes cascade chutes.The final ground product reports to two 2000 ton roaster feed silos. The two-stage oxygen flu-solids roasters with their associated gas train are the subject of a separate paper.The cyanide leach circuit is a standard carbon in leach system with circular swept screens for interstage carbon retention. The loaded carbon is sent to the existing Goldstrike autoclave plant for stripping and regeneration. Tails, after cyanide destruction, are deposited in the existing tailings dam with tails from the autoclave facility.Throughout the circuit, Clean Plant Design criteria were used to assure environmental compliance, ambiental acceptability and ensure worker health and safety. The commissioning and operation of the double rotator grinding technology has been very successful at Barrick Goldstrike."

Jan 1, 2001

By T. Guan, G. Fu, X. Zhou, L. Deng

"One kind of tailings from gravity separation of polymetallic mine contains 0.5–0.8% SnO2. The proportion of –0.037 mm size fraction of the tailings is 37.12%, and the distribution rate of SnO2 is 60.32% in it. For this tailings, two processing technology was carried out. One was felt-shaking table gravity separation in which the concentrate with a grade of 30% SnO2 at a recovery of 10% was obtained. Another was flotation-gravity separation in which the concentrate with a grade of 51.20% SnO2 at a recovery of 51.69% was obtained, but the dosage of collectors was as much as 990g/t. Therefore, the experimental research on the influence of grain size composition on the cassiterite recovery were conducted. The results show that: (1) the flotation recovery of the cassiterite distributed in +0.043 mm size fraction is only 10%, flotation recovery is not suitable for it; (2) a portion of cassiterite distributed in –0.01+0.005 mm size fraction could be recovered by flotation, but the tin flotation performance will be worsened rapidly and the dosage of collectors will be increased due to the increase of the content of –0.010 mm size fraction; (3) flotation processing is not suitable for –0.005 mm size fraction of the tailings. According to the results, a new technology has been carried out, that is discarding tailings by gravity separation-hydrocyclone desliming-flotation-gravity separation processing flow. The processing index is close to that of the flotation-gravity separation processing flow, and the collector dosage is decreased to 360 g/t by this processing technology. And, it has been achieved that the cassiterite slime is recovered economically and reasonably.INTRODUCTIONAs the density of cassiterite is higher than that of gangue, gravity separation is considered to be a low cost, pollution free, and effective process for cassiterite recovery (Wills, 2009). However, cassiterite is a kind of fragile mineral, over crushing is easy to take place and a large amount of slime is inevitably produced in the crushing and grinding of cassiterite. For the cassiterite slime exists in –0.043 mm size fraction, the efficiency of gravity separation decreases sharply. The production practice shows that, for the fine cassiterite, the recovery of tin exists in –0.043 mm +0.010 mm size fraction is about 13% through shaking table gravity separation, and the tin exists in –0.010 mm size fraction is almost impossible to be recovered by gravity separation. If flotation is adopted to recover fine cassiterite, the problems of high reagent dosage, high cost, environmental pollution, and poor economic benefit would take place. Therefore, the recovery of cassiterite slime is a worldwide problem. A key way to increase the utilization of tin resource and increase the economic benefit of tin mines is developing suitable beneficiation process for fine cassiterite."

Jan 1, 2016

By Paolo Toscano, Dave Hall

"This paper discusses the project history, process design and layout of the Rainy River Gold Project mill. The circuit consists of a 55 x 83 primary crusher, 15 MW SAG mill and 15 MW ball mill. Both mills are twin pinion with LSS and VFD drives. A pebble crusher is located in the process plant with bypass and stockpile options. Gravity concentration followed by 30 hours of leach residence, carousel CIP, carbon recovery, cyanide destruction and tails disposal in a 1,428 hectare tailings management area (TMA).The Rainy River Gold Project is located approximately 65 km northwest of Fort Frances in north western Ontario. The project is scheduled to be commissioned and reach commercial production in mid-2017. Expected cost is $US877M and is comprised of an open pit mine, process plant, tailings management area and supporting infrastructure. The process plant is designed to produce gold/silver Dore’ bars.The project has been developed on the following timeline:• 1967 – First exploration• 1994 – 17 zone discovered• 2004 – 433 Zone discovered• 2005 – Rainy River Resources Ltd. purchases property • 2006 - ODM zone discovered• 2007 - Cap Zone discovered• 2008 - First NI-43-101 Resource completed• 2012 - PEA completed & Intrepid Zone discovered• 2013 - New Gold acquires Rainy River Resources• 2014 - Updated feasibility study completed & Environmental Assessment report submitted• 2015 - Detailed Engineering >95% complete, EA approved and construction commencesThe property is accessed by provincial highway 600 and is within 17 km of 230 kV power line and railway sidings. Key project criteria are summarized as follows:"

Jan 1, 2016