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By Bob A. Hamaguchi

The Trojan Tailings Pond at Highland Valley Copper developed in only 15 years from a biologically inactive water body into a productive lake with a well established aquatic ecosystem. This process was assisted by a few simple and inexpensive procedures including fertilization and introduction of essential organisms. The success at the Trojan Pond has encouraged the use of similar techniques in other water bodies on the mine site. This paper details: ??The history, water quality and thermal behavior of Trojan Tailings Pond ?Techniques used in fertilization; introduction of micro-flora (algae etc.), large aquatic plants, zooplankton and invertebrates ??Selection and sources of suitable transplant materials ??Rate and extent of vegetation establishment ??Biomass production ??Rates of organic sediment accumulation over tailings ??Bacterial activity in the sediments ??Improved water quality through the bioaccumulation of aqueous metals in plants and the subsequent bacterial binding or precipitation of that metal during decomposition in the tailings sediment. Robert A. Hamaguchi graduated with a Mining Diploma from the British Columbia Institute of Technology in 1971, followed with a Bachelor of Applied Science in Mineral Engineering from the University of British Columbia in 1975. Since graduation, he has worked for numerous open pit mining operations throughout British Columbia. In 1991, he joined Highland Valley Copper working as an Environmental Engineer responsible for government liaison, reclamation, decommissioning and other environmental duties. Heather M. Larratt graduated from the University of Calgary in 1978 with a B.Sc. in Environmental Biology (First Class Honors). She has provided drinking water management and reservoir design for numerous cities and irrigation districts in British Columbia for 20 years. Heather has also worked for 12 years in reclamation of mine water, including passive SRB water treatment systems. Mark R. Freberg joined Highland Valley Copper in 1991 after working at number of mining operations in British Columbia and the Northwest Territories. He graduated in 1979 with a Bachelor of Applied Science degree in Mineral Engineering from the University of British Columbia.

May 1, 2001

By P York

The study commenced in October 2012, to understand the basic ecology of the spotted-tailed quoll (Dasyurus maculatus) and to confirm whether the species utilise previously mined land on the Mt Owen mine site that was rehabilitated in 1996 to 2011, as part of their home range. The spotted-tailed quoll is a widely distributed marsupial carnivore that is listed as rare or threatened in every state in which it occurs. The species is a forest-dependent marsupial carnivore that occupies a wide range of habitat types (Long and Nelson, 2004).The first part of the study involved trapping a spotted-tailed quoll and fitting the individual with a combined global positioning system / ultra high frequency (GPS/UHF) radio-transmitter collar (Telemetry Solution – Quantum 4000E). The benefit of the GPS/UHF collar is that it records locations (or fixes) at intervals programmed by the user. The UHF component allows the researcher to locate the individual at a frequent interval, and the Quantum 4000E also has a remote download function, enabling the GPS data stored in the collar to be wirelessly transmitted between the collar and researcher. The benefit of this new technology is that the animal does not need to be repeatedly captured to retrieve the GPS data.Daytime searches are also undertaken using the UHF receiver to locate the individual and locations of daytime denning and latrine sites. The location of each den site is recorded using a hand-held GPS. Data downloaded from the GPS collar is then imported and mapped using computer software program ARCGIS (geographic information system), to overlay the data points onto aerial photography.Results demonstrate that threatened species will utilise rehabilitation areas post mining. The data presented will be used to develop a greater understanding of threatened species ecology in and around rehabilitated areas, which will assist in the planning and design of native forest rehabilitation areas.CITATION:York, P, 2014. Mt Owen Mine spotted-tailed quoll study, in Proceedings Life-of-Mine 2014 , pp 591–596 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By M Barkh, F X. Spachtholz

The mining industry faces significant environmental challenges in managing tailing ponds and wastewater treatment, including the risk of contaminated water, breaches of tailing dams, and energy-intensive treatment processes. Traditional methods involve pumping water to secondary treatment facilities for neutralisation, which is costly and poses operational and environmental risks. In response to these challenges, the Professional Mine Water Action Plan (PMAP) introduces a revolutionary approach that combines cutting-edge AI technology and environmentally friendly reagents for in situ neutralisation of acidic wastewater. PMAP leverages magnesium-based slowrelease reagents and AI-driven intelligent dispensing systems to optimise treatment processes within the tailing pond itself. The unmanned smart dispensing vessel (USDV) conducts bathymetry to gather data on water quality and hydrography, allowing the system to determine the appropriate reagent quantity and injection strategy. This approach transforms the tailing pond into virtual reaction columns, enabling individual and simultaneous water treatment while optimising efficiency. PMAP’s novel reagent demonstrates exceptional efficacy in removing critical metals like cobalt, nickel, and copper, as well as reducing harmful metals such as mercury and manganese to undetectable concentrations. Unlike traditional lime-based treatments, PMAP’s reagent prevents gypsum formation and reduces waste volume significantly, improving reagent utilisation and solid waste handling. Furthermore, PMAP technology offers benefits beyond environmental stewardship. By reducing water treatment costs and improving resource management, it enhances operational efficiency and aligns with broader corporate sustainability goals. Additionally, PMAP expedites permitting processes by eliminating the need for traditional water treatment infrastructure and mitigating risks associated with hazardous reagent transportation. Integrating PMAP as a pre-treatment for existing lime facilities provides further benefits, including optimisation of treatment operations, reduced reagent consumption, and the recovery of valuable metals from generated sludges. Overall, PMAP represents a groundbreaking advancement in mining wastewater management, showcasing the transformative potential of AI and environmentally friendly solutions to foster sustainable and responsible mining practices globally. Its inaugural project in the Yukon Territory aims to demonstrate its capability to remove contaminants from mine water, paving the way for widespread adoption in the industry.

Sep 1, 2024

By Yuexin Han

The wet preconcentration of haematite, comminuted by high-pressure grinding roller was studied, using the cylindrical ferromagnetic medium. The effects of various factors including rod diameter, rod gap and background magnetic field intensity (BMFI) on the performances of preconcentration for different size fractions and the force analysis of single mineral particles with different diameter in the separation zone were emphatically investigated. The results showed that as the rod diameter increased, the rod gap decreased, and the BMFI increased the tailings grade and the tailings yield reduced while the concentrates recovery raised that was regardless of variations in feed size distribution. For the -3.2+0.18mm size fraction preconcentration, the magnetic medium with a rod diameter of 4 mm and a rod gap of 7 mm at BMFI of 454kA/m is used to get a higher performance. For -0.18mm size fraction preconcentration, the magnetic medium with a rod diameter of 4 mm and a rod gap of 4 mm at BMFI of 263kA/m is used to get a higher performance. After the classifying preconcentration, at the optimum condition, the overall concentrate grade increases by 9.56 percentage points and the concentrate recovery is up to the 82.67% with a tailing grade of 10.68% and a tailing yield of 40.31%. The magnetic force, exerted on weekly magnetic minerals was proportional, to the specific susceptibility and square of BMFI. For the same BMFI, the magnetic capturing force on coarse particles increased, and that on fine particles decreased with the increasing of rod diameter. For the lean locked-particles, considering the gravity and viscous resistance of particles, rods with a big diameter were beneficial to the capture of coarse particles and had no influence on the capture of fine particles, while rods with a small diameter were good for the capture of fine particles and not suitable for the capture of coarse particles of lean locked-particles. Keywords: low grade haematite, high-pressure grinding roller, high intensity magnetic separation, pre-concentration

Sep 1, 2012

By Alfred Annor

Deep mining requires innovative approaches in ground control technology and the search for new ground support materials including backfill. There is a need to look at mine waste materials that may be available in deep mines to assess their suitability for use as composite backfill. Composite materials are widely used in all areas of engineering because of their unique and often superior structural properties compared to those of the singular constituent materials. Over the past few years, there has been a growing interest towards the investigation and use of various types of composite backfill materials as special applications fill in mines. A research project is being conducted at CANMET- Mining and Mineral Sciences Laboratories, Natural Resources Canada, to investigate the properties of potential ground support materials for use in deep mines. As part of this investigation, composite backfill materials were prepared by combining waste rock aggregates and mill tailings in various ratios, to determine their optimal mix proportions for strength development. A few cemented rockfill and straight tailings paste backfill samples were also prepared and were used as controls for the study. Two sample sizes were used for the investigation in order to evaluate the scale effects on the material behaviour. A set of samples were cured at 44 EC to simulate the anticipated temperature conditions in deep mines. A second set of samples were cured at 23 EC to provide engineering reference data. The test results show that compared to either the cemented rockfill or the straight tailings pastefill material, the composite backfill samples were the strongest and stiffest of all the test materials investigated in this study. The composite backfill samples also produced the lowest void ratio and porosity values compared to the cemented rockfill and the straight tailings backfill material. This paper will present the results of the laboratory investigations to date. It will also outline the on-going work at CANMET - Mining and Mineral Sciences Laboratories to further assess the ground support potential and viability of using composite fills in deep mines.

May 1, 2003

By N. S. Voiloshnikova

Large-scale continuous ?Resin-In-Pulp? process tests were conducted at Irgiredmet. The object for testing was the sample of oxidized gold ore from one of Siberian deposit. Weak-base ion ex-change resin PuroGold was used as adsorbent. The duration of ?Resin-In-Pulp? tests was sufficient for a completing of three cycles ?Adsorption/Elution?. Gold in feed was 0.95 mg/L. Gold in solid phase of adsorption tails? 0.065 g/t. Gold in liquid phase of adsorption tails? 0.026 mg/L. Total gold in adsorption tails ?0.11 g/t. Gold recovery ?94.3 %. The use of regenerated resin had practically no effect on gold losses with adsorption tails. Alkaline cyanide solution (NaCN-20g/L, NaOH- 5g/L) was used as eluent for gold elution from loaded resin. Elution temperature ?60 °C. Number of eluent bed volumes passed ?10. Loaded resin analysis was following (g/ton): Au? 1700~1250; Ag? 400~460; Cu? 250~450; Zn? 5100~9600; Fe? 860~1200. Elution efficiency: Au? 95.3%~97.8%; Ag? 84.0%~88.0%; Cu? 84.0%~88.0%; Zn? 97.0%~99.0%. Although high zinc resin loading was shown zinc was eluted efficiently. The adsorption properties of fresh PuroGold resin, PuroGold resin after 3 cycle ?Adsorp-tion-Elution?cycles in comparison with AM-2B resin were tested. It was shown from adsorption isotherms that selectivity of PuroGold resin slightly less in comparison with AM-2B resin mainly due to its higher zinc adsorption properties. It was founded on base of test results that adsorption properties of PuroGold resin was practically constant during three ?Adsorption/Elution? cycles. Gold may be efficiently recovered by resin tested. There were no difficulties during gold elution from loaded resin and its regeneration. PuroGold resin was recommended for industrial use in RIP process. Process design criteria were developed and plant design is on progress at the moment.

Jan 1, 2014

By A. V. Myaskov, V. S. Valavin, K. D’Elia, I. O. Krylov, Vu Van Toan, T. I. Yushina

"The purpose of this research is the development of a technology for processing the concentration tailings of limonite ores at Kamysh-Burun Mining and Processing Complex into an iron- containing product for ROMELT process. In the concentration and metallurgical processing of iron ores, substantial amounts of waste are generated (tailings, slimes, slags, metallurgical dust, etc.) that can be converted into a commercial product if is properly processed. This article presents the results of experimental studies of the feasibility of processing iron- containing tailings of Kamysh-Burun deposit (slimes from Verkhne-Churbash storage facility (VCh sample) and dumped waste of Kamysh-Burun (KB sample)) to obtaining valuable components for further uses. Applying Mossbauer spectroscopy has been shown that the iron minerals in the studied samples are technogenic varieties of iron hydroxides. The most suitable technology is the magnetic roasting process, since iron hydroxides are easily transformed into a magnetic, easily recoverable form. Have been shown the advantages of a two-stage magnetic separation, in a strong field and with polygradient separators. Concentrating this material through magnetic methods makes it possible to increase the iron content to 20-21 %. Magnetic concentration of the deeper layers materials results in iron- containing concentrates with a total iron content of up to 58% whit a phosphorus content of about 2%. Reducing the phosphorus content is possible by blending upper and lower layers of the material at a ratio of 1:1. Thus, it is possible to process the material accumulated in the slime storage facility using the ROMELT process combined with preliminary three-stage magnetic concentration and disposal of tailings with the permissible iron content of less than 5%. Wastes from smelting will be slags suitable for use in construction industry. The Kamysh-Burun material preparation for ROMELT process should be performed applying gravity-washing with separation of the -0.4 mm fraction and additional blending for dephosphorizing the concentrate. The smelting will produce pig iron and slag."

Jan 1, 2018

Creating Value and Improving Environmental Performance by the Recovery of Valuable Products from Tailings from the Athabasca Oil Sands in Alberta, Canada

Sep 13, 2010

By Daniel W. Kappes

The 1.8 million tonne Itos tailings deposit is the result of sever­al years of processing by Comibol of the San Jose and ltos Mine ores at Oruro Bolivia. The historic flowsheet consisted of a flotation circuit in which silver-bearing galena and stibnite were concentrated. Approximately half the silver in the ore was not recovered, and the existing tailings contain slightly over 200 grams silver per tonne, in the form of complex sulfides of silver, antimony, lead, and other metals coating grains of pyrite. An extensive laboratory program was conducted to exam­ine conventional processing l)lethods such as cyanide leaching and regrinding followed by flotation. However, the silver-bear­ing sulfides are in intimate contact with the pyrite and the pyrite comprises 30% of the total tailings weight. Cyanide leaching recovers only 40% of the silver and the cyanide consumption is uneconomically high. Flotation produces an upgraded, but still uneconomic, concentrate. For this reason, chloride leaching had been evaluated by Comibol, and in the mid 1980's Comibol undertook an innova­tive test consisting of two 300 tonne field heap leaches using a cold chloride solution. The tests were partially successful, but could not be applied to the bulk of the tailings which are already ground to 50% minus 200 mesh. In 1992, Kappes, Cassiday & Associates undertook to develop a hot agitated chloride leach flowsheet, and is current­ly constructing a treatment plant under a turnkey contract from a Bolivian company, Baremsa, S.A. Using this process, 85% of the remaining silver can be recovered, with significant by-prod­uct credits from antimony and lead. The project features hot, agitated acidic chloride leaching, filtration, crystallization of lead chloride, and precipitation of metals using iron scrap. Arsenic is removed from the precipitate by caustic leaching, and minor metals by furnace refining. Antimony is recovered as the oxide by cupelation, and silver is poured as <lore from the cupel. Production rate will be 600 tonnes per day. Startup is scheduled for late 1995.

Jan 1, 1996

The Inglewood Gold Project consists of several titles in and around the Inglewood and Wedderburn Goldfields which were purchased from Strata Mining Corporation (now Excalibur Mining Corporation). The Inglewood goldfield has a historical production of approximately 160 000 tonnes at 23 grams per tonne. The company has reopened and developed the Maxwells reef, picking up a high-grade shoot (Max Shoot) identified and worked by Goldquest et al up until 1994. Production ceased on a low gold price, rapidly reducing vein size and a requirement for further development. Greater Bendigo Gold Mines Ltd (GBM) has undertaken the development and encountered the mineralised vein in an increasing size setting with high grades and the existing mine development has access to a further shoot to the north (Max North Shoot). The Maxwells reef also contains the Woolf shoot south of the Maxwells shaft and an inferred resource of ~20 000 ounces is contained from 60 m below surface. Best intersection 2.9 m at 50 g/t. The Maxwells reef area produced approximately 20 per cent of the Inglewood goldfields production and GBM is preparing to increase its exploration efforts across the rest of the field. A processing plant has been constructed at the Maxwells mine site, which is capable of treating tailings, hard rock and alluvial material utilising gravity recovery. The existing Maxwells tailings have been excavated and treatment commenced. It is intended to produce a sand product from the tailings for resale and to reduce the storage requirements per tonne. Several large water dams constructed by Goldquest in the 1990s are being re-engineered to enable storage of significant tonnages of tailings. A level of surface and underground production is planned for 2008 and this is to be expanded into 2009 and beyond. The long-term future at Inglewood is envisaged as structural repeats at depth below the current workings and a project plan to gain deep access will be prepared on successful drilling results.

Jan 1, 2008

By M P. Schwarz

This paper describes modelling studies of the flotation cells used by BHP Billiton Nickel West Ltd in the slimes circuit at Mt Keith. The aim is to achieve a better understanding of the hydrodynamic limitations in the existing cells to improve nickel recovery. Computational fluid dynamics (CFD) simulations have been performed on the 100 m3 mechanically-stirred cells. The data produced by the simulations include velocity vectors, gas holdup distribution, solids distribution, bubble size, superficial gas velocity and turbulent parameters. These results have been checked and found to match hydrodynamic measurements at the plant. An analysis of the data found that one of the limitations in fine particle recovery was related to the air bubbles discharged with the tailing. The air in the discharge was caused by the high viscosity of the pulp. In the presence of fine particles, the pulp exhibits high viscosity with a yield stress. The increased drag due to high viscosity caused the bubbles to rise more slowly up the cell. This results in a significant hold-up of bubbles, some of which report to the tailing with associated loss of valuable minerals. A parametric CFD study of the operating variables shows that less air would be discharged in the tailing by decreasing the following: air flow rate,pulp density,stirring speed, andfrother addition. Simulation results indicate that these changes would cause a lowering of the air hold up in the cell especially near the discharge area. Predictions of the bubble-particle attachment rates under various operating conditions have also been calculated. Although it is possible to increase attachment rates by increasing stirring speed or by increasing pulp density, the amount of air discharged and the amount of nickel loss through the tailing will also increase. The project showed that CFD modelling is a powerful tool for studying hydrodynamic effects of bubble-particle interactions in flotation cells.

Jan 1, 2008

By Robert Ammon

THE milling practice at Mascot, at present, consists of dry crushing to 5/8 in., jigging, fine grinding, and flotation. The ore arrives at the mill from two mines, No. 1 mine shaft being located in the mill proper, and No. 2 mine shaft being situated about 2000 ft. east of the mill. No. 2 mine is served by a coarse-crushing plant and aerial tramway to the mill bins. In jigging the No. 2 mine ore, 50 per cent. of the mill feed is rejected by the jigs as a jig tailing ranging from 5/8 in. to 10 mesh; the balance of this ore, after removing a jig concentrate, is ground to flotation feed. In jigging No. 1 mine ore, approximately 80 per. cent. of the mill feed is rejected as a coarse jig tailing, and the balance is ground to flotation feed without the removal of a jig concentrate. The difference in the treatment is due to the fact that No. 2 mine ore carries a higher content of zinc than that from No. 1 mine. Of the total concentrates produced, 30 per cent. are made on jigs and 70 per cent. by flotation. At present, 1350 tons of No. 2 ore and 800 tons of No. 1 ore are treated each 24 hours. The recoveries for a typical week are as follows: Heads 3.55 Jig concentrates 61.30 Jig tails 0.80 Flotation concentrates 60.60 Flotation tails 0.09 Extraction 86.50 The associated minerals are a straw-colored sphalerite, dolomite, calcite, pyrite, quartz, and a small amount of greenockite, which occurs as a stain and sometimes small crystals along with sphalerite, but is comparatively rare. Pyrite is present in small quantities, usually less than 0.5 per cent.; quartz occurs in bands a few inches thick. A complete analysis of the ore is as follows:

Jan 9, 1924

By R. B. Stein

"Countercurrent decantation circuits(CCD) are often employed in hydrometallurgy to effect solids-liquid separation in combination with a barren solution wash to reduce losses of valuable solution component(s) from the tails. Bechtel Corporation uses its BEMSIM flowsheet-modelling PC computer program to simulate a CCD circuit where the ""dirty"" wash solution contains significant amounts of soluble components. An algorithm successfully models this condition and is simplified to a single expression when the wash solution is a ""clean"" solution containing no solubles. The CCD calculation module is compared to operating data from uranium and nickel-cobalt CCD circuits.IntroductionHydrometallurgical process schemes have often relied on countercurrent washing methods to reduce soluble losses of valuable product in the tails. Commonly used equipment to effect this operation is a CCD circuit of N thickeners in series, where a slurry is fed to thickener I and a wash solution to thickener N; clear pregnant solution overflows thickener I and washed tails exit as underflow from thickener N. CCD circuits of this type are typically employed when the solids tend to slime heavily and where a filter is not economic.A typical CCD circuit usually employs from two to five thickeners and can employ either a ""clean"" washing solution of, say, pure water or else the hydrometallurgical circuit may generate a ""dirty"" barren solution containing an appreciable amount of soluble component which may be advantageously returned to the process as CCD wash solution. Various authors have treated the case for clean washing solutions (Svarkovsky 1977, Merritt 1970). For the case of dirty wash solutions, an algorithm has been developed to enable efficient modelling by the BEMSIM computer program, because the need for an accurate method to calculate soluble losses is required when performing an economic trade-off analysis of the value of product lost in the tails as a function of the number of thickeners."

Jan 1, 1990

By James S. Browning

Recovery of fine mica from a sample of impounded tailing from the Dixie Mines, Inc., crushing and screening operations, Heflin, Ala., was investigated to determine if plant losses of mica could be reduced while econmically producing a commercially acceptable mica concentrate. Conventional wet-tabling and agglomerate-tabling methods yielded concentrates assaying about 95 percent mica from the tailing that contained 17 percent mica. Products containing about 98 percent mica and accounting for mica recoveries of 86 percent were obtained by flotation, using as reagents sulfuric acid, kerosene, and coco amine acetate (a mixture consisting chiefly of dodecyl and tetradecyl amine acetates). The results of the batch flotation tests were verified by tests in a small continuous flotation pilot plant.

Jan 1, 1963

By P M. Ashley

Antimony (Sb) is an element similar to arsenic that forms anions in surface waters, rather than cations like other metals. Little is known about natural mobility of antimony in the environment. However, recent studies have suggested that antimony has similar toxicity to arsenic. Current New Zealand drinking water guideline values are 3 parts per billion (0.003 mg/l) for antimony and 10 ppb (0.01 mg/l) for arsenic. Antimony occurs principally in stibnite (Sb2S3) in mesothermal vein deposits hosted in low grade metamorphic belts of eastern Australia and southern New Zealand. Many deposits have been mined historically, including one large deposit, at Hillgrove, NSW. Oxidation of stibnite and associated arsenopyrite and pyrite causes local acidification of natural waters but acid is readily neutralised by carbonates in mineralised zones and host rocks, and therefore waters are near-neutral (pH 6-8.5). Antimony dissolves readily in moderately oxidized, near-neutral waters. Experimental stibnite oxidation yielded up to 37 mg/kg (37 000 ppb) dissolved antimony, and mine and tailings dam waters at Hillgrove mine have up to 55 mg/kg (55 000 ppb) dissolved antimony. Natural stibnite can contain >5000 mg/kg arsenic in solid solution and therefore dissolution of stibnite, arsenopyrite and arsenian pyrite also releases arsenic into waters. Resultant dissolved arsenic concentrations are up to 3.6 mg/l (experimental) and up to 7.2 mg/l (mine and tailings dam seepages). Mine and tailings discharge waters have elevated antimony and arsenic where they emerge, but attenuation occurs by deposition of the metals on to amorphous iron oxyhydroxides which can contain >10% each of antimony and arsenic. Historic disposal of mineralised waste rock and tailings into the stream system at Hillgrove has caused contamination of sediments and waters with antimony and arsenic for at least 20 km downstream with effects traceable in sediments for 300 km. Other stibnite-bearing deposits are likely to have analogous environmental effects, particularly from historically-mined sites.

Jan 1, 2002

By L. Flett, S. Woods, C. Deredin

"Following a significant reduction in world Imperial Smelting Furnace (ISF) capacity from 2003 to 2008, a major optimization initiative was successfully carried out on the lead upgrading and bulk (zinclead) flotation circuits in the Brunswick Concentrator.This paper describes the process improvements that have taken place in order to reduce the quantity of bulk concentrate:•,Increase in lead upgrading rougher flotation circuit capacity•,Installation of a closed-circuit bulk regrind utilizing MillPebs and 150 mm Cavex hydrocyclones•,Installation of two additional cleaner flotation stages on the bulk circuit•,Optimization of the manipulated variables in the lead upgrading and bulk circuitsINTRODUCTIONThe Brunswick Mine (Xstrata Zinc Canada) is located near Bathurst, New Brunswick and is one of the world’s largest lead/zinc mines. Milling of ore commenced in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore from the No. 6 open pit mine. The planned closure of the mine will take place at the end of the first quarter in 2013. The ore is a complex massive sulphide, consisting of approximately 20% valuable minerals, 60% pyrite and 20% silicates.A simplified flowsheet for the concentrator is given in Figure 1. The overall flowsheet involves SAG milling, secondary ball milling, selective flotation, and dewatering by thickening, filtering and drying. The complexity of the flowsheet evolves from the requirements for fine grinding and the production of four concentrates: zinc (sphalerite), lead (galena), copper (chalcopyrite) and bulk (sphalerite plus galena).Primary hydrocyclone overflow feeds the copper-lead aerator-roughers for flotation at pH 9.0 with soda ash. Plant xanthate (80-20 sodium isopropyl and potassium amyl) and AP407 are used as collectors in the copper-lead circuit. The rougher concentrate is directed to the copper-lead regrind circuit and the tailings are fed to the secondary grinding circuit. The secondary hydrocyclone overflow from the secondary grinding circuit feeds the copper-lead rougher-scavenger banks. Rougher concentrates are reground to a P80 of 28 µm prior to three stages of copper-lead cleaning. The first stage of cleaning is done in an open circuit with the tail reporting to the cleaner scavenger. The cleaner scavenger concentrate reports to the regrind circuit and the cleaner scavenger tail combines with the rougher/scavenger tail to feed the zinc circuit. The second and third stages of cleaning are performed in closed circuit with the cleaner tail being recycled to the copper-lead regrind circuit. The average final concentrate grades 23% Pb, 3% Cu and 704 g/t Ag, and is fed to the copper separation circuit."

Jan 1, 2013

By K. C. Dean

The Bureau of Mines, following passage of the Solid Waste Disposal Act of 1965, was authorized to conduct research on methods for utilizing, stabilizing, or reclaiming mineral wastes. This report summarizes laboratory research conducted by the Bureau and cooperative field studies made with the mineral industry to reduce the environmental impacts of air and water pollution deriving from mineral wastes. Utilization research showed that both dry-press and steam-cured building bricks and ceramic tile could be made from copper mill tailings and other discarded mineral wastes. Sponge iron was produced from steel mill and amination industry wastes, and valuable mineral products were extracted from such diverse sources as flue dusts and uranium tailings. Mineral wastes were combined with agricultural and/or sewage wastes to make artificial soils that have the potential of resolving several disposal problems by production of a useful commodity. Physical, chemical, and vegetative stabilization methods were evaluated, singly and in combination, for reclamation potential on copper, uranium, lead-zinc, and asbestos tailings. Suitable methods were developed in the laboratory for overcoming major adverse conditions of these wastes such as salinity, heavy metal toxicity, acidity, basicity, lack of nutrients and organics, and unstable surfaces. Field testing of these methods was performed on a variety of mineral wastes. Costs for the various stabilization and utilization-procedures developed were also determined.

Jan 1, 1986

With the burgeoning opportunities for new gold mines to be developed the opportunity now exists for a reassessment of the optimal processing path for these gold ores. Projects have traditionally followed the path of whole ore cyanidation with either Carbon In Pulp (CIP), Carbon In Leach (CIL) or Merrill Crowe. The challenge ahead is whether these traditional processing methods are the best option for stakeholders, or to go beyond CIP and use a combination of currently available technologies that can be used to provide gold recovery at a significantly reduced capital cost and minimal deleterious impact on the environment. Either stand alone or combinations of gravity concentration and flotation, can produce a gold bearing concentrate that can then be fed directly to a batch or continuous intensive cyanidation unit. Treatment of a gravity/flotation concentrate by intensive cyanidation may be processed by the InLine Leach Reactor (ILR) developed by Gekko Systems of Australia. The advantages of the technology combining gravity/flotation followed by intensive cyanidation are significantly reduced capital costs; improved safety due to reduced cyanide inventories; the majority of tailings un-contaminated by cyanide; less energy required due to an opportunity for an increase in grind size û reducing green house gas emissions; detoxification of concentrate tailings can minimise tailings disposal and containment requirements.

Jan 1, 2004

By John D. Bardill

A cooperative project between the Bureau of Mines, Department of the Interior, and the Montana School of Mines, investigating results of hydraulic systems for transporting sand slurries for emplacement, was begun in August 1955. In such circumstances, the Bureau provides funds for a fellowship leading to a Master of Science degree as well as available instruments, tools, supplies and equipment. The School provides the ore-beneficiating laboratory with its equipment for use as a pilot plant. The success of a hydraulic backfilling method employing mill tailings depends largely upon the physical characteristics of the tailings. One of these characteristics is size distribution, especially in the finer size ranges. When being emplaced underground, almost impervious layers of slimes separate from some types of mill tailings containing a high percentage of slimes and settle on the top of the sand fill. These coatings may be so dense that the water in the slurry cannot percolate through them for proper drainage. Successive layers of fill in a stope, each with an impervious coating on top, sometimes build up a hydrostatic head, creating pressure on the bulkheads. This may cause breaks in the lining material, inducing sand runs and flooding the level and sumps below. Deleterious slimes flocculated and disseminated throughout the sands should help prevent the formation of impervious slime layers, increasing percolation and settling rates.

Jan 1, 1960

By W. A. Stickney

The Bureau of Mines conducted beneficiation and smelting studies on siliceous offgrade manganese samples to indicate effective methods of manufacturing silicomanganese from materials that cannot be economically beneficiated to meet General Services Administration (GSA) specifications for stockpiling. Laboratory and pilot-plant flotation tests were conducted on 2 samples, 1 from the waste-tailings dump of the Domestic Manganese & Development Co , Butte, Mont., and the other from the Nettie mine in the Butte-Philipsburg area. Smelting tests were made on the combined concentrates from the pilot-plant flotation tests and on part of the tailings sample The flotation scheme found to be adequate in laboratory tests and later in the pilot-plant tests consisted of flotation by a fuel oil-tall oil emulsion in an acid circuit (pH )-1-.8 to 5.0). The oil-emulsion reagent flocculated the manganese minerals and selectively floated them from the siliceous gangue, In the pilot-plant flotation test on Butte-mill tailings 90.1 percent of the manganese was recovered in a concentrate that analyzed 32.6 percent Mn, 4.8 percent Fe and 28.1 percent Si02. In the test on Nettie mine ore 90.8 percent of the manganese was recovered in a concentrate that analyzed 36 6 percent Mn, 5.5 percent Fe, and 26-4 percent Si02.

Jan 1, 1956