Search Documents

By T. T. Chen, W. Petruk

The mineralogy of the feed, concentrates and tailings from the Heath Steele mill in New Brunswick was studied in detail to determine the liberation and recoveries of the major ore minerals, and to characterize the occurrence and distribution of the trace elements. The ore milled during August 1977 contained an average of 4.34 wt % Zn, 1.25 wt % Cu, 1.64 wt % Pb and 90 gr/ton Ag. The main trace elements in the ore are silver, cadmium, indium, cobalt, antimony, bismuth, arsenic, mercury, tin and gold. Silver is carried in solid solution in tetrahedrite-tennantlte, galena, Ag-Sb sulfosalts and Pb-Bi-Sb sulfosalts. Approximately half of the silver is recovered in the lead and copper concentrates and the remainder is lost in the tailings and Zn concentrate. As the major silver-bearing minerals also contain Sb and Bi, recovering silver by ore dressing would involve recovering elements that are troublesome in extraction metallurgy. Cadmium and most of the indium and mercury occur in solid solution in sphalerite and consequently are recovered in the zinc concentrate. Some indium is also recovered in the copper concentrate in a second unidentified indium-bearing mineral. Cobalt and arsenic occur principally in arsenopyrite, cobaltite, a Co-Fe sulfarsenide, pyrite and pyrrhotite, and follow these minerals into the tailings. Recovery of the valuable element cobalt, therefore, would involve recovering the environmentally hazardous element arsenic. Most of the tin and gold report to tailings. The ore was ground to about 80% minus 270 mesh. At this grind, 71 wt % of the sphalerite, 66 wt % of the chalcopyrite and 59 wt % of the galena were liberated. Recoveries of free sphalerite, free chalcopyrite and free galena in the Zn, Cu and Pb concentrates were 85 wt %, 76 wt % and 80 wt %, respectively; recoveries of unliberated sphalerite, unliberated chalcopyrite and unliberated galena in the respective Zn, Cu and Pb concentrates were 48 wt %, 5 wt % and 7 wt %. Some of the unliberated sphalerite and chalcopyrite grains in the tailings are large and would have been liberated at a finer grind.

Jan 1, 1980

By N. Kuyucak

Uranium ores leached with sulphuric acid typically have poor dewatering properties. Applicability of controlled neutralization where gypsum precipitates as crystalline particles, improving dewatering properties of material was examined. Neutralized material was then dewatered using paste thickener to increase underflow density. The objectives are to maximize water recovery due to water scarcity and to produce suitable cemented paste fill material for optimizing underground ore recovery. Work was performed on acidic slurry samples collected after leaching but before the resin-in-pulp uranium recovery. Controlled neutralization greatly reduced volume of neutralization precipitates giving higher densities and improving the dewatering (filtration) properties. As a result it would be possible to increase the water recycled to the mill. Another benefit is that sulphate in the recycled water was reduced to saturation limit minimizing scaling of process equipment. Additional test work is needed; however, Unconfined Compressive Strength results imply that controlled neutralization increases the strength of paste fill and possibly offers opportunities towards a viable backfill option.

Jan 1, 2010

By R. D. Williams

INTRODUCTION In a very real sense, Butte, Montana is where the copper came from that won two world wars. The price for that unrestricted mining and smelting of copper came due in 1983 when Butte was declared a Superfund site. One of the most complex remnants of this mining history is the Parrot Tailings site. The Parrot Smelter operated from 1881 to 1910 moving Upper Silver Bow Creek to the south and depositing approximately 370,000 cubic yards of waste material adjacent to Upper Silver Bow Creek. Additional mixed waste from other processing sites deposited below the Parrot Tailings at the Northside Tailings, Diggings East and Blacktail Creek Berm total approximately 224,000 cubic yards of contaminated material. As part of the Superfund remedy, EPA proposed leaving all these wastes in place. This proposal prompted a detailed criticism by several geoscientists familiar with the site. The EPA’s 2006 Record of Decision (ROD) did in fact leave these wastes in place which set the stage for a series of studies which continue to refine how best to protect extensive completed Superfund Remediation downstream of Upper Silver Bow Creek. HISTORY OF THE PARROTT MINE AND TAILINGS Placer gold was discovered in Silver Bow Creek, the headwaters of the Clark Fork River only a few miles from the Continental Divide in 1864. This discovery lead to a stampede of placer and lode claims. One of the first lode claims located was the “Parrot” lode, located in the fall of 1864. Progress for developing a mine at the Parrott was slow and it wasn’t until 1876 that one of the original locators encountered a rich vein of chalcocite. Butte had its first copper mine! The Parrott mine was patented in 1879 and by 1881 the Parrot, having lost a “t”, concentrator and smelter were built. The smelter and concentrator were an immediate success with a considerable market for processing the District’s rich copper ore. In 1885 the employees of the Parrot Smelter bought the Butte Reduction Works, another smelter. In 1885 dams were built on Silver Bow Creek to contain various tailings and related processing byproducts (Figure 1).

Jan 1, 2019

By Euler M. De Souza

"This paper presents a study developed to assess the potential use of salt tailings as backfill to control surface subsidence and as a mine bulkhead used to isolate areas of potential brine inflows. A series of index tests, centrifuge model tests and quasi-static creep consolidation tests were performed to evaluate the creep and consolidation properties of salt backfl1l and to evaluate the effect of brine inflow and brine infiltration on backfill behaviour. The centrifuge models were specifically designed to study the mechanical behaviour of salt backfill under the influence of brine inflow and long-term creep. Test results indicated that salt tailings could be applied underground as a bulkhead to seal off areas of potential inflow with the additional advantage of controlling room closure and local back instability. It is hoped that this study will assist mine planners to develop contingency plans to control brine inflow occurrences. IntroductionThe use of salt tailings as backfill is a technique that would assist in the control of potential brine inflows and surface subsidence, and would have a very positive impact on the long term adverse environmental effects associated with surface salt tailings disposal. A centrifuge testing program was designed to evaluate the behaviour of salt tailings as backfill; the effectiveness of salt backfill as a bulkhead used to isolate areas of potential brine inflow and the potential use of salt backfill to control surface subsidence. The testing program was also used to evaluate the creep-consolidation properties of salt backfill and to evaluate the effect of brine inflow and brine infiltration on backfill behaviour. This paper presents the results of the experimental centrifuge testing program and a comparison of results with models tested under I g using an electrohydraulic servo-controlled loading frame (MTS) and a creep consolidation cell. Applicability of the results to practical underground soft rock mining is also discussed.The centrifuge model tests indicated that salt tailings could be applied underground as bulkheads to seal off areas of potential inflow with the additional advantage of controlling room closure and local back instability. Because the applied stresses (1.2 MPa) and fill consolidation were relatively low, the centrifuge simulated bulkhead did not retain all of the inflow brine. At typical field stresses it is expected that the fill would be able to seal off panels with brine inflows. Further testing at higher stresses are suggested to determine the consolidation and stress levels at which the fill would completely retain any potential brine inflow head and positively act as mine bulkhead."

Jan 1, 1994

By Peter Van-De-Velde

"The results of applying Japanese pumping technology to North American reclamation projects will be discussed.IntroductionEconomic high-density mineral reclamation and subsequent transportation of solids through a pipeline presents many formidable engineering problems.In North America where space has rarely been a facto r, industry has generally relied upon the traditional dumping of material into various waste ponds, etc . In Japan, where confined geographies and stringent environmental regulations are enforced, alternative solutions for disposal or reprocessing of waste material had to be found.High-density mineral reclamation system designed by theoretical formulas is almost impossible because solids composition, i.e. concentration, velocity, moisture content, solid size, etc. can only be controlled within a broad pumping range.Traditionally this has restricted dredging of material to relatively dilute slurries whereas it is usually preferred to reduce water usage with its associated power and reclamation costs to a minimum and reclaim as high a concentration as possible.At Toyo Pumps North America we recognized these pumping limitations and began to develop and manufacture a range of standard and auxiliary equipment that would exceed these technical restrictions."

Jan 1, 1985

By B. C. Williams

The Bureau of Mines monitored the water quality in and near a copper/gold tailings impoundment. Samples were taken from the unsaturated %one. and dissolved concentrations of over ZO constituents were measured with respect to depth and moisture content, factors which influence oxygen content and sulfide oxidation. Samples from the unsaturated %one were drawn through a hypodermic needle sampler into individually evacuated vials fitted with septa, thus preventing oxygen contamination. Water quality was also monitored in the saturated tailings, underlying colluvial aquifer, and weathered andesite bedrock.

Jan 1, 1989

By F. Qunfei, W. Laijun, X. Lyu, A. Cosmos, L Huijuan, Z. Jinggang, G. Yao

In this paper, four cyanide-utilising bacteria strains were isolated from a gold mine tailings pond, in the Hexi gold mine (Shandong), China. Based on the analysis of 16S rRNA gene sequence, four strains were identified as Aeromonas hydrophila strain C1, Bacillus mojavensis strain C3, Shewanella oneidensis strain C2, and Aeromonas media strain C4. Based on the cyanide detoxification efficiencies results of the four strains, Bacillus mojavensis C3 had the highest detoxification rate and was selected to assess its efficiency to detoxify cyanide in gold mine tailings at different pH (9.5, 10.5, and 11.6) using cyanide as the sole carbon and nitrogen sources. Strain C3 was capable of detoxifying 46.5mg/L of cyanide-containing gold mine tailings at all tested pH ranges at 30oC, 196 rpm after 15 d. The amounts of cyanide detoxified were 87.5%, 84.1%, and 73.5% at initial pH 9.5, 10.5, and 11.6, respectively. The highest cyanide detoxification was observed at pH 9.5. The viable cells growths in the samples were 5.7E+09CFU/mL, 4.9E+09CFU/mL, and 2.7E+09CFU/mL at pH 9.5, 10.5, and 11.6, respectively. Free ammonia concentrations were observed as a byproduct of the biodetoxification process. The residual ammonia concentration in the sample was observed at 0.35 mg/L, 0.33 mg/L, and 0.62 mg/L at pH 9.5, 10.5, and 11.6, respectively. The highest residual ammonia concentration was noticed at pH 11.6. The amount of cyanide removed by mixed culture strains C3 and C4 was 94.6% and showed a residual free ammonia concentration of 2.6mg/L, as biodetoxification byproduct after 120 h. The pure culture Bacillus mojavensis strain C3 and its mixed culture strains exhibited higher potential to detoxify cyanide-containing gold mine tailings at alkaline pH values and could be cultivated to detoxify gold mine effluents at large-scale. Keywords: Bacteria strain, cyanide, detoxification efficiency, gold mine tailings

Jan 1, 2020

By J. A. Folinsbee, Andre Robitaille, Tony Y. Chong

"Conventional crushing and grinding is followed by gravity concentration to recover free milling gold. Refractory gold, finely disseminated within the pyrite and arsenopyrite matrix, is recovered by flotation followed by pressure oxidation, cyanidation, and cementation. Non-refractory gold is recovered by cyanidation/CIP processing.A comprehensive test program aimed at improving flotation recovery was initiated in June 1995. Plant testing indicated that flotation tails value could be significantly reduced using mechanical flotation cells. Based on these test results, a flotation circuit that incorporates one of the three existing flotation columns followed by mechanical scavenging and cleaning was developed. Detailed engineering was completed in September 1996. Construction started in February 1997 and was completed in 3 months at a cost of $3.2 million. An on-stream X-ray analyzer was installed to provide real-time arsenic and iron analyses and inferred sulphur value on some of the flotation streams.The new flotation circuit proved to be an instant success. Flotation tails values were reduced by 27%. Final tails solids and solution was reduced from 0.035 to 0 023, and 0.0014 to 0.0006 oz/ton Au, respectively. Based on a mill throughput of 1,620 stpd this decrease in tails value will result in an increased gold production of 7,500 ounces per year. The payback for the project is estimated to be around one year.This paper describes the evolution of the project from bench scale testing to commissioning. Also, the operational and economic impacts on down stream processing are highlighted"

Jan 1, 1998

By Donald R. East

In the planning and design of mill tailing disposal facilities, the chemical and physical nature of the fine­grained material to be disposed of has to be fully understood in order to create an efficient and safe long­term storage facility. Disposal of tailing has become the single most important source of environmental impact for mining projects and the development of many promising ore bodies has been delayed or even stopped due to problems of facility siting and design under pressure of increasingly stringent regulatory require­ments.

Jan 1, 1992

By Raymond E. Healy

Circuit samples from the HBMS concentrator were assayed for l5elements, and image analysis was done to determine the modal quantities, grain size distributions and liberations of the minerals. Material balancing was done on these data to determine distributions of the minerals in each concentrator stream. The ore was mined during the waning phase of mining and is diluted with backfilled copper smelter slag. The mill feed grades 1.0 wt% Cu and 1.8 wt% Zn. The slag silicate phases contain approximately 1.5 wt% Cu and 6.5 wt% Zn; therefore the Cu and Zn head grades due to chalcopyrite and sphalerite, are only 0.9 wt% Cu and 1.2 wt% Zn, respectively. The recovery of Cu in the Cu-concentrate (Cu-conc.) is only 77.5 %, whereas that of chalcopyrite is 89.2 %. Similarly, the recovery of Zn in the Zn-concentrate (Zn-conc.) is only 40.2 %, whereas that of sphalerite is 71.0 %. The slag accounts for approximately 60 % of the Cu, and 85 % of the Zn in the tailings. The main loss of chalcopyrite is as unliberated grains to the tailings, and the main loss of sphalerite is as free grains to the Cu-concentrate and as unliberated grains to the Cu-concentrate and tailings. Approximately 85 % of the chalcopyrite and sphalerite lost to the tailings. accounting for 8 % and 10 % of the respective minerals in the concentrator feed, occur as middling and sub-middling grains. Extremely fine regrinding would be required to liberate these minerals.

Jan 1, 1989

By N. J. Bolin

The Aitik copper mine is located in the north of Sweden and the concentrator at the mine started its operation in 1968 with a capacity of 2 Mton ore per year. Since then the operation has expanded several times, and today the capacity is 18 Mton ore per year. A new expansion plan was investigated in 2005, however the building of a new concentrator was found to be a better solution, giving a lower cost, a longer lifetime and more concentrates to the smelter at the coast. Several choices have been evaluated regarding the location of the concentrator?s crushing, grinding, dewatering, process control and water systems. The energy cost comparison has been an important factor when evaluating the various choices. The single most important decision has been in choosing the grinding system. Once again the fully autogenous grinding system, favoured by Boliden, has proven to be the best solution. The location of the mill, high and near the tailing pond, gives lower transportation costs for the ore and tailings, compared to other locations. A new water system will lower the pumping cost, give better control of water circulation to the plant and create fewer problems with permafrost in the tailings disposal area. Molybdenum recovery will be installed in the new plant using a separate circuit. A gold leaching plant was also investigated. Gold leaching is planned for the pyrite concentrate from the chalcopyrite/pyrite separation and the concentrate from the desulphurisation circuit that will be installed on the bulk flotation tailings.

Jan 1, 2014

By Kanaya Koichiro

Doyashiki Concentrator in Hanaoka mine of Dowa Mining Co., Ltd. started the operation in 1943, treating the ore mainly from Doyashiki ore deposits. After the cessation of mining of Doyashiki ore body in 1973, Doyashiki concentrator began to treat the ore from Fukazawa mine instead in the same year, and furthermore the ore from Ezuri mine In 1979. Currently Doyashiki concentrator treats 30,000 tons of ore per month from both mines. The ore treated is complex sulfide ore, called "Kuroko" (black ore), and the principal minerals contained in the ore are chalcopyrite, galena, sphalerite, pyrite, and barite, accompanied by bornite, tetrahedrite, tennantite, covellite, stromeyerite, argentite, polybasite, pearcite, and electrum. The Principal gangue minerals are quartz, gypsum and clay. The grade of the feed to the concentrator averages 0.9 p/t Au, 110 g/t Ag, 1.1 %Cu, 2.8 %Pb, 9.2 %Zn, 5.0 % Pyrite-S and 13% BaS04. The feed is, after washing and two-stage crushing, ground to 88% minus 200 mesh in a rod mill and ball mills with cyclone classifiers, and five concentrates are recovered by flotation. Primarily copper-lead bulk concentrate is recovered by sulfurous acid-lime flotation method, and then from the bulk concentrate copper concentrate is floated by hot pulp flotation method, sinks being lead concentrate. Zinc concentrate is recovered from the tailing of copper-lead bulk flotation, pyrite concentrate from zinc flotation tailing and barite concentrate from pyrite flotation tailing. The final tailing is sent to the dam.

Jan 1, 1982

By K. Shame, H. H. Thiele

South Africa holds approximately 70% of the world’s total chrome reserves. Chromium is an important raw material for the production of stainless steel and ferrochrome alloys. Over the years, China’s ferrochrome market has been dependent on chrome ore imports from South Africa. In 2018, China imported about 75% of its chrome ore from South Africa. With the low platinum price and the rising demand for stainless steel, a growing demand to evaluate new beneficiation technologies for fine chromite was identified. Gravity separation of fine chromite (<75 μm) is inefficient, due to the limitations of the technology to upgrade fine particle size distributions. Potential for the recovery of chromite, specifically fines (<75μm), lost to UG2 PGM flotation tails, but also from MG and LG6 spiral circuit tailings exist. Flotation of podiform chromite ores from India and Kazhkhstan is successful in generating saleable product grade, following acid or alkaline flotation routes. Application to South African stratiform chromite is being evaluated, with the emphasis on evaluating reagent type, optimising dosages and the effect of pH in optimising flotation performance. Flotation can unlock additional value from chromite tailings material. The effect of ore characteristics such as gangue mineralogy, particle size distribution and headgrade are reported. The performance in terms of product quality and yield are evaluated and compared. The impact of this paper will not only prolong life of mine and increase revenue but will also alleviate environmental burdens associated with tailings dams. Keywords: Chromite, flotation, fine material, performance optimisation

Jan 1, 2020

By Jianmin H, Shuqin L

This paper analyses problems associated with cement based technique in metal mines, and provides a new mining technique of packing the tailing sand and water into solidified material. In the new technique, the cementing agent is a solidifying material containing a high percentage of water and tailing sand. The sand and water solidify into the packed body at the same time. The slurry of the packed body varies widely in concentration and suitability for different mining conditions. It consists of two liquid materials, each composed of a cementing agent, tailing sand and water. The thick liquids are transported to the packing place at the same time. Their pumping property is good; they do not solidify or block up the pipe during transportation. At the packing place, the two liquids are evenly mixed.The mixed liquid flattens out well and its solidifying speed is so fast that it solidifies into a packed body within a very short period of time. The compressed strength of the packed body can reach 1.2-5MPa in one day, which satisfies the requirements of different mining conditions. By using the new mining technique, all the tailing sand can be used again in the packed body and it is low in cost and investment. The mechanisation is easy and thus the working efficiency is increased and the use of labour decreased. Contamination problems caused by i. Dr & Prof. Beijing Graduate School, China University of Mining Science and Technology,Beijing,China 2 Northern .Jiaotong University, Beijing,China 3. Beijing Graduate School, China University of Mining, Science and Technology

Jan 1, 1992

By G. W. Poling, J. R. Hoffert

"A common practice in preventing gravity-fed tailing pipeline from plugging due to sanding is to add lime at relatively high dosage rates. The lime affects the rheology of the slurry ""vehicle"" which transports the coarser particles in a heterogeneous flow-regime. This paper describes a study of the effect of lime on the critical deposition velocities and friction head losses of Highmont Mining Company&apos;s mill tailing. Testwork using an extrusion viscometer was related to the surface chemistry changes brought about by the high lime additions. Lime additions which typically increased the tailing pH from 8.4 to 11.5 reduced the critical deposition velocities from approximately 3.2 to 1.7 m/sec.IntroductionCommon practice, at least in base metal concentrators in British Columbia, is to use high lime additions in tailing slurries when the pipeline becomes partially plugged with sands. Some concentrators, such as the Highmont Mining Co. Ltd. plant, add lime on a continuous basis during low tonnage periods and during startup/shutdown transition s. Although such high lime additions often solve temporary plugging problems, the effect of lime on slurry rheology has not been well understood.Slurry pipelines are designed to transport solids in suspension with a minimum of wear and without particle deposition. These two objectives are in opposition because the velocities required to maintain solid suspension can sometimes cause high head losses in the pipeline, as well as wear problems. Thus, the optimum velocity for a long distance slurry pipeline is in the range that provides just enough turbulence to keep coarse particles in suspension."

Jan 1, 1985

By Gordon. I. M., J. Blackshaw

"The wet part of the Hollinger mill is housed in two buildings, the grinding plant and the treatment plant. In the grinding plant there are four 7xl5 ball mills which grind the ore to 60 per cent through 200 mesh. There is also one 7xl5 ball mill which grinds table concentrate to 9 5 per cent through 200 mesh.The primary classifier overflow is pumped to the treatment plant where it passes over shaking tables, producing a concentrate and a tail. The concentrate pulp is pumped to the grinding plant for secondary grinding. The secondary classifier overflow is pumped back to the treatment plant where it is thickened in two 40xl 5 tray thickeners and agitated in four 20x24 mechanical agitators arranged in series of two (with 2 spares).The table tail pulp is thickened in nine primary 40xl5 tray thickeners (with 2 spares). The thickened table tail pulp and the concentrate agitator discharge pulp are mixed and pumped to four rows of 15x45 pachuca agitators, arranged in series of five tanks.The pachuca discharge pulp is divided. The smaller part goes to two rows of 3-step decantation using 40xl5 tray thickeners (with 2 spare rows). The larger part goes to six 14xl6 drum filters. The repulped primary filter cake and the thickened pulp discharged from the decantation rows are fed to eight 14xl6 drum filters. The secondary filter cake is mixed with water and pumped to the tailing line."

Jan 1, 1968

Sulfuric Acid and Phosphate Industries at Anaconda Reduction Works. BY E L LARISON (Contribution 70-Preprint 4400 words) In 1915 a 135-ton chamber sulfuric acid plant was built at the Anaconda Reduction Works to supply acid for leaching sand tailing Sulfur dioxide was obtained by roasting copper table concentrate Soon after, the flotation process, using acid as a reagent, together with development of a leaching and electrolytic process for zinc recovery, necessitated building a second acid unit of 50 tons&apos; per day capacity, completed in 1917 Within a few months the zinc process was so improved as to consume no added acid Mean¬while research work was in progress on production of phosphate fertilizers using sulfuric acid and Idaho phosphate rock In 1920 a pilot plant was put into operation which treated 50 tons per day of phosphate rock to produce treble superphosphate containing 45 per cent available P205 In 1922 the original plant and process for leaching of sand tailing were abandoned in favor of a combination leaching and flota¬tion process In 1923 the sand leaching plant was remodeled into a treble super¬phosphate plant designed to treat 160 tons per day of phosphate rock. In 1924 the copper flotation process was put on alkaline circuit and acid, has not since been used as a reagent Practically all the acid produced is now used for making phos¬phate products In 1931 the use of copper concentrate as a source of sulfur dioxide was abandoned and an iron concentrate, floated from the copper concentrator tail¬ing, substituted

Jan 1, 1935

By V. G. Lofftus

Since plant start-up in January 1983, Golden Sunlight Mines, Inc. has implemented several process changes and improvements. Among these are a pulp preaeration circuit and a sand tailings retreatment circuit. The purpose of the preaeration circuit is to reduce reagent consumption during gold dissolution. The sand tailings retreatment circuit was added to recover and process separately pyrite concentrate to maintain acceptable overall gold recoveries with the more refractory sulfide ores. Research and plant operating data along with process and improvement descriptions are presented.

Jan 1, 1987

By Paul E. Bennett, James S. Browning, Thomas L. McVay

Continuous 1.0- to 1.5 - ton - per - hour pilot plant flotation tests were made to determine the technical and economic feasibility of recovering beryl from flotation tailing being produced at the spodumene concentrator of the Foote Mineral Company , Kings Mountain , N.C. Best results were obtained when using an acid circuit and amine - type cationic reagents to float a bulk beryl - feldspar concentrate and then selectively floating the beryl away from the feldspar , using a petroleum sulfonate collector . About 77 percent of the beryllium in the pilot plant feed was recovered in a concentrate containing 6.42 percent BeO . The capital cost of an integrated flotation plant capable of treating 500 dry tons of spodumene mill tailing per day for recovery of beryl and feldspar is estimated at $ 210,000 . Indicated direct operating costs for the plant were $ 1.52 per ton of dry spodumene flotation tailing treated . This approximated the value of beryl recovered , but potential profits of about the same magnitude lie in concurrent recovery of marketable spodumene and feldspar byproducts .

Jan 1, 1964

By Cai S, Zhou C

The requirement of mineral resources is increasing as the economy is continually increasing in China. The mining practices of various complex sulphide deposits and precious metal deposits, and the treatment of mine tailings and other industrial solid wastes, have greatly improved fill quality in the metal mines. From early-1980s, high density slurry fill, paste fill of total tailings, low density slurry fill with addition of silicate gelling agent, etc has been studied and applied, and some progress has been achieved. Today, mine tailings, slag and power plant ash are widely used as fill materials in mining. A number of cemented fill mines equipped with large capacity trackless mining machines are in production. Recently, The 3rd Chinese National Symposium on Mining with Backfill, 1995 has been held, the progress of fill technology in mines has been exchanged and summarised. In the paper, some aspects of fill technology used in the metal mines of China are briefly introduced and some applications of fill research work are presented.

Jan 1, 1998