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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

By J. A. Organiscak

The Bureau of Mines recently investigated dust generation and methane liberation during half-and full-web cutting at a double-drum shearer longwall face. This study was conducted to ascertain the health and safety aspects of wider web shearing at longwalls. Results from these underground tests showed that deeper web cutting reduced dust generation and methane liberation per ton of coal mined, while significantly improving shearer production rates. When cutting with the airflow (head-to-tail cut pass) full-web cutting decreased dust generation per ton by 16 pct at the shearer and 6 pct at the tailgate, with an increase in shearer production rate of 72 pct. Methane liberation per short ton also deere~sed by 36 pet at the shearer and 22 pct at the tailgate. Full-web cutting against the airflow (tail-to-head cut pass) decreased dust generation per ton by 41 pct at the shearer and 37 pct at the tail-gate over half-web cutting, with an increase in shearer production rate of 61 pct. Methane liberation per ton also decreased by 35 pct at the shearer and 41 pct at the tailgate. These improvements for full-web cutting are a reflection of improved cutting efficiency (less dust and methane produced from cutting) and less new face area exposed per ton of coal mined (less methane liberated from face per short ton of coal mined).

Jan 1, 1988

By B. E. Halbert, K. T. Himbeault, S. L. Fernandes, R. L. J. Phillips, B. J. Topp

The Beaverlodge uranium mining and milling facilities operated from 1952 to 1981 with about 94% of the ore extracted from the main underground mine and 6% from smaller satellite deposits. Decommissioning work occurred from 1982 to 1985 involving periods of shutdown, salvage and reclamation. Transition phase monitoring, leading to eventual delicencing commenced in July 1985. Over the last 15 years, discharge from the tailings management facility (TMF) and a fresh water stream, impacted during the operational phase by tailings spills, has improved in water quality for most parameters of concern. Loadings to the environment of three key contaminants (radium-226, total dissolved solids, and uranium) have consistently been less than during the operational phase with radium-226 having the greatest variability. Outstanding environmental issues associated with the recovering drainage system formerly used for tailings disposal, are being addressed in an enhanced environmental monitoring program to commence in 2000. Changes in water chemistry and the natural re-introduction of aquatic organisms are issues of concern. In the meantime, application for release from selected satellite areas commenced in 1999. This paper reviews the issues which have arisen during the transition phase, outlines how they have and are being addressed, and provides a comparison of original and current predictions of the recovery process under way at the former Beaverlodge mine site.

Jan 1, 2000

THE following paper is a description of the plant and operations of the above company in their recovery and treatment of the quartz tailings deposited as final residues by the various mining companies. at Waihi, Karangahake, and Waitekauri into theOhinemuri River. During the past few years the silting of the Ohinemuri River has increased to such an extent as to cause considerable interest and discussion, and has lately been the subject of a most extensive inquiry by a Royal Commission, to consider the best means of dealing with what is generally spoken of as a public nuisance.Ever since the commencement of mining in this district, which was opened as a goldfield in 1875, the various mines operating have been discharging their tailings into this river, it having been proclaimed a sludge channel in 1895. Up to the year 1900 the quantity of quartz crushed was not large, probably not exceeding 500,000 tons; but since that date the annual tonnages discharged have increased enormously, and at the present time would total from 550,000 to 600,000 tons per year. Records show that up to the end of 1910 at least 5,000,000 tons have been crushed, for an actual yield of bullion considerably exceeding &#163;10,000,000 in value, and all the valuable tailings from...

Jan 1, 1911

By R. Sprycha, J. Szczypa, W. Janusz, J. Neczaj-Hruzewicz

INTRODUCTION Because large aqueous suspensions of various fine solids originate during the beneficiation of ores, the Mining Industry constantly requires efficient and feasible dewatering methods. Among the techniques currently in use (1,2), filtration (vacuum, press) and centrifugation are fairly expensive and thus are limited to use with more valuable materials. Furthermore, installations of reasonable size produce only moderate outputs. Therefore, tailings from ore dressing plants are usually treated by sedimentation or settling techniques with a tailings dam as the ultimate sedimentation basin (3). Apart from growing environmental and economic questions, the solid product obtained from the tailings dam basin has in most cases, a rather high moisture content and a persistent semi-fluid consistency. Thus, it is difficult to transport and use further. On the other hand, the solids may be utilized as a filling material, fertilizer, etc. providing they are dry enough and properly granulated. Demands for an improved technique are met to a great extent, by the technique presented in this paper. AGGLOMERATION OF FLOCCIJLATED SLIMES Origin and Qualitative Description of the Process Phenomena of consolidation and pellet formation in flocculated fines were first observed by Yusa and Gaudin (4) while studying rheological properties of flocculated kaolinite. On this basis, methods of separating liquid from solids have been worked out (5-7) and many possibilities for application have been suggested (5,6). Yusa termed the technique "wet pelletization&apos;&apos;.

Jan 1, 1980