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The Copper Concentrator at Mount Isa Mines Limited treats approximately 6 Mt of chalcopyrite ore annually. The ore is currently sourced from the 1100 and 3000 orebodies at Mount Isa. Chalcopyrite is the only significant copper bearing mineral, with pyrite, quartz and dolomite being major gangue minerals. Ore is typically 10% chalcopyrite, 11% iron sulphides and 79% non-sulphide gangue. Copper ore is treated through two autogenous mills, two ball mills and a flotation circuit. The autogenous mills replaced a crushing plant, two rod mills and two primary ball mills during 1991. For routine analysis of concentrator performance, composite samples of mill feed, final concentrate and tailings are generated every 4 week period. All samples are sized and size fractions are assayed. Recovery of mineral species to the concentrate are then calculated for each size fraction. Sometimes liberation analysis was also performed by QEM*SEM (Quantitative Evaluation of Material by Scanning Electron Microscope).

Jan 1, 1990

By J. Hanna, J. Jung, I. J. Anazia

Electrothermal production of elemental phosphorous (P4) generates substantial amounts of highly toxic phossy water and sludges. Because of their high phosphorous contents and lack of reliable processing technology, large tonnages of these hazardous wastes have accumulated from current and past operations. To date, the sludges are stored in ponds or above ground in special containment vessels at various locations around the country. Two samples of electric furnace phosphorous sludge (obtained from the Tennessee Valley Authority (TVA) in Muscle Shoals, AL and FMC Corporation in Pocatello, ID) were investigated using physical separation techniques to recover the elemental phosphorous and to render the residual material nonhazardous. The techniques tested include sizing, froth flotation and dispersed air oxidation. Flotation and sizing produced concentrates containing 69% - 90%P4, with recovery in the range of 60% - 87%. Dispersed air oxidation of the fines and tailings products was effective in the removal of an additional 10% - 15% of the P4 in the sludge.

Jan 1, 1993

A new $18-million hydrometallurgical ore treatment facility for the recovery of molybdenum oxide was dedicated at Climax, Colo., on November 19 by Climax Molybdenum Co., division of AMAX. The Company has been producing molybdenum from molybdenite (MoS2) ores since 1918. Now oxide values, formerly lost in the sulfide flotation plant tailing are to be processed to yield about 3 million Ib per year of metal. In 1965, Climax produced about 50% of world molybdenum from sulfide ore. This ore is mined at the rate of 42,000 tpd. The annual production of concentrate is 100 million lb containing 90% to 92% MoS2, or about 50 million lb of metal. As only a relatively small portion of the total Climax production is sufficiently enriched in oxide molybdenum to process through the new plant, 5700 tpd of selected ore containing 1.5 lb of Mo per ton is now treated by the new process.

Jan 12, 1966

By Murray G

This paper outlines the processes involved in acid generation from sulphides and the key factors in identifying and predicting time dependent changes in the nature of acid mine drainage (AMD). The initial characteristics of mine water; whether drainage from exposed mining faces, runoff and seepage from waste dumps or seepage from process tailings; do not necessarily reflect the long term potential for AMD. The ultimate quality of mine drainage is a function of the interaction of many chemical, physical, biological and engineering (operational) factors. Techniques are available that can predict the qualitative nature of acid generation and mine drainage quality but accurate quantitative predictions are essentially impossible at the present time. As more and more mining operations exploit sulphide ore, the AMD and acid waste issue will become a. far greater concern. The significance of time dependent factors in the production of acid mine drainage-.is not just restricted to particuL.ar mine sites but needs to be recognised by the mining industry as a whole.

Jan 1, 1988

By R. W. Mancantelli, J. R. Woodward

In 1947 a large low grade deposit of uranium was located in the northwest corner of Saskatchewan, in the Beaverlodge property of Eldorado Mining & Refining Ltd. Most of the values occur as thin seams and as coatings on other minerals, and the ore, which is light and friable, is not amenable to the usual gravity methods of concentration. The acid leaching process developed to retreat the company's Port Radium tailings was also impractical, as the percentage of carbonates is high and the percentage of sulphides relatively low. Construction of the mill building and installation of equipment, both scheduled for 1952, awaited selection of a satisfactory ore dressing process on or before Oct. 1, 1951. An extensive research program on dressing of ores from the Ace mine therefore was undertaken by the Mines Branch in Ottawa, a research team at the University of British Columbia and an ore dressing group at the company's Port Hope refinery.

Jun 1, 1955

By D. Dimitriadis, T. Polglase, M. M. Botz, W. Phillips, R. Jenny

Cyanidation processes for the extraction of gold or silver frequently involve the processing of sulfurcontaining solids that react with cyanide to form thiocyanate. The formation of thiocyanate is an undesired side reaction, and in many cases thiocyanate formation results in a prohibitively high cyanide consumption. With the increasing number of gold and silver mining projects that involve the processing of sulfide ores or concentrates or that utilize a biological sulfur oxidation process, more attention is being focused on minimizing thiocyanate formation and in treating solutions to eliminate toxicity associated with thiocyanate. Though not yet commercially applied, there are several processes available to recover cyanide from thiocyanate to reduce tailings toxicity and to allow the reuse of cyanide otherwise lost to thiocyanate formation. This paper presents an overview of processes available for recovering cyanide from thiocyanate along with estimated operating costs for the processes. Pilot testing data suggests that in some cases cyanide can be recovered from thiocyanate for approximately the same cost as newly purchased cyanide.

Jan 1, 2002

By James L. Hendrix, R. Y. Wan, David W. Kappes, Jinrong Zhang, M. C. Fuerstenau, Terence E. Albert

A sulfidic flotation tailings ore, containing 1.14 gpt Au and 242.53 gpt Ag, was treated by low temperature chlorination roasting followed by cyanide or chloride leaching. Eighty per cent of Ag was recovered by cyanide leach when the ore was roasted at 350°C with 50 kg/ton calcium chloride. About seventy per cent of Au was made cyanide amenable by chlorination roasting at 400°C. A better Ag recovery, ninety per cent or over, was achieved by chloride (NaCl+HCl) leach. However, Au recovery was very poor by chloride leach. Ferric chloride, calcium chloride, and cupric chloride were compared as chlorinating agents for their effect on Au and Ag recovery. The study found that low temperature chlorination roasting is very effective for recovering Ag associated with stibnite and stannite. The best chlorinating agent for the test ore was found to be calcium chloride. Parallel roasting-leaching tests indicated that chlorination roasting was far superior over oxidation roasting for both gold and silver recovery.

Jan 1, 1993

By J. E. Forciea, O. E. Palasvirta, L. G. Hendrickson

All pilot and commercial plants working with Mesabi Range taconite employ wet magnetic separation. Progress is being made with a dry magnetic process,9-11 but this has not yet been applied to taconite except in the laboratory and pilot plants. Flowsheets in all five plants are similar, including multiple stages of grinding, each followed by magnetic separation, which rejects a finished tailing. The flowsheet in Fig. 1 is used in the preliminary and commercial plants of the Erie and Reserve mining companies and in Unit C of Oliver's Pilotac plant. At Erie the desliming operation is omitted. In the Reserve plants two ball mills are used with each rod mill. Pilotac uses flowsheet 2 in Unit A-B and has also tested flowsheet 3. Principal differences in the three flowsheets are the number of grinding stages and the type of roughing. The terms open- circuit roughing and closed-circuit roughing are best defined by the diagrams.

Jan 12, 1958

By M. Faure

The experiment with backfill at Harmony was mainly an attempt to improve the ground conditions in a particularly tricky section of the mine. A deslimed tailings system was selected for financial reasons, ease of installation, ease of operation and its low level of mechanization. The design of the system has three aspects: firstly, to define the parameters of the backfill material itself; secondly, to design a reliable plant which could consistently produce the desired material in sufficient quantities; and, thirdly, to design a fill distribution system for maximum reliability in the shaft, minimum pipe line velocity, pressure and cost in the horizontal sections and consistency and reliability at the delivery end. The system was successfully constructed and commissioned in a short space of time at a relatively low cost. It has now reached its full potential, and very encouraging results have been achieved with regard to cost savings, ground conditions, ventilation and accident statistics.

Jan 1, 1988

By G. M. Wilemon, Swanton. R. G.

The Bureau of Mines is investigating the leaching of phosphate values from Florida phosphatic clay wastes using sulfuric acid (H2SO4) as the leaching agent and methanol as the solvent. Acid-alcohol leaching studies were done using research-grade apatite to investigate potential causes of flow extractions. It was found that a barrier layer formed on the surface of the apatite. Initial discovery of the barrier layer was made by analyzing pure apatite samples before and after leaching using an optical microscope. Subsequent Fourier Transform Infrared (FT-IR) analyses proved the coating was calcium sulfate hemihydrate (CaSO4•1/2H2O) and that unreacted apatite existed beneath the coating. Scanning electron microscope (SEM) photographs depict the CaSO4• 1 /2H20 morphology as long slender crystals adhering to the apatite surface. It was also shown that grinding the leached tailings and leaching again in the H2SO4-methanol system resulted in additional phosphate extraction.

Jan 1, 1992

This paper discusses a number of examples of how environmental factors have been incorporated into the water management aspects of mining projects. One example is the Kingston Lignite Project in the South-East of South Australia where the most significant environmental issue was considered to be the effect of drawdown from mine dewatering and depressurization upon the use of groundwater. By proposing an alternative mining method - the use of floating dredges rather than a two-pass dragline method - drawdown effects would no longer be a concern. A second example is the Pine Creek Gold Mine in the Northern Territory. In this case the principal area of interest from an environmental viewpoint was water management - in particular effluent disposal, contaminated seepage from overburden and tailings, and project water supply. Through an integrated approach to the management of wastes and water the potential for water quality impairment was significantly reduced and external water supply requirements were also significantly reduced.

Jan 1, 1987

By B. Dubreuil, J. -M. Lalancette, D. Lemieux

Arsenical bearing deposits are posing an increasing challenge for the mining industry. In particular, mineral processing of arsenical sulfidic ores for base and precious metals recovery tend to produce problematic arsenical tailings. This arsenic being present as sulphide or oxide, tends to dissolve, contaminating surface and underground waters. To meet the environmental regulations, it is often required to stabilize the arsenic in a more insoluble form. Dundee Sustainable Technologies has developed a new patented technology using vitrification of the arsenic oxide in an economical way, giving a stable solid, well below the environmental regulations upon leaching. This technology involves oxidation of As (III) to As (V), by using a low cost natural oxidizer and transformation of the produced arsenic oxide into calcium arsenate, which is not volatile at the melting temperature of the glass. This technology would be advantageously applied on most cases around the world where sulphide concentrates are oxidized in a fluidized bed or smelted.

Jan 1, 2014

Prior to 1972, Peko Mines Limited at Tennant Creek used a central mill at Peko, which consisted of a gravity circuit, flotation and cyanidation of tailings, to treat their gold ore from surrounding mines. The Warrego concentrator, 60km to the west, was built to handle the Warrego and Gecko copper orebodies, which were being developed concurrently with the concentrator construction The Warrego concentrator consisted of a flotation circuit, using autogenous grinding at 100 - 120 tph, compared to the more elaborate Peko circuit employing rod mill - ball mill grinding capable of treating only 18 - 26 tph. During development of the Warrego copper mine a gold rich ore pod was located, because of which the mining plans were adjusted to allow simultaneous, separate extraction of the copper and gold ores. However, the plans for the concentrator were not affected as the gold ore could be carted 60km to the Peko central mill for processing.

Jan 1, 1978

A model of dense medium cyclones is developed based on balancing the terminal velocities of the discrete particles due to the centrifugal force against a random turbulent eddy velocity. An expression of the form P - or Q _ 1 100 e."A where P is the partition number., is a function of cyclone efficiency, particle size and density differential between the particle and medium, and is the slurry spilt, is developed and may be used for regressing plant partition data. The above model provides an understanding of peculiarities of the partition curve such as asymmetrical tails, elevation of the cutpoint and elevation of the d50 with increasingly fine feed. The effect of feed variables such as particle size, medium viscosity and changes in spigot or vortex finder diameter is shown. A means of convoluting the partition curve with various feed washabilities is discussed together with how the above may be used to evaluate plant performance for a variety of feeds.

Jan 1, 1987

By Alex Probst

With low metal prices and increasing value placed on plant water and tailings disposal areas, it is becoming essential to maximize the efficiency of solid-liquid separation processes. Since its inception over 25 years ago, the use of flocculants has grown exponentially. While flocculants play a crucial role in the sedimentation process, they are also a major operating cost. Recent innovations are revolutionizing the industry once again. New thickener designs are doubling thickener-loading rates while reducing flocculant consumption. The feedwell, as the heart of the thickener, is undergoing many significant metamorphoses. With CFD and other advanced computational analysis, feedwell technology is being examined more thoroughly than ever before, The result is that the thickener previously thought of, as a "Black Box" is now being mapped and controlled with state-of-the-art technology. This paper will examine recent design improvements in all aspects of thickeners in order to project the look of the thickener of tomorrow.

Jan 1, 1999

By T. G. Carnahan

The Bureau of Mines investigated a flotation-nitric acid leach procedure as part of the goal to maximize minerals and metals recovered from primary and secondary domestic resources. Studies were conducted on an ore that contained carbon-bearing and sulfide mineralization that rendered a portion of the ore refractory (resistant) to conventional leaching technology. The procedure investigated for treating the ore consisted of the following: (1) separation by flotation of the carbonaceous and sulfidic components from the ore, (2) leaching the flotation concentrate with nitric acid at 100° to 110° C, (3) leaching the flotation tailings with sulfuric acid, and (4) processing the combined leached slurries in a conventional manner to recover yellow cake. In step 2, HN03 is converted to gaseous products from which it is regenerated by reacting these products with air and water for further leaching. An overall uranium extraction of 96 pct was achieved by this procedure.

Jan 1, 1979

By Harbort G. J

Jameson cell pilot plant trials were conducted in a number of areas in the Red Dome gold/copper flotation plant during 1991. The aim of testwork was to reduce the amount of cyanide soluble copper reporting in tailing to the gold leach plant, and hence reduce cyanide consumption. A number of these tests were conducted in parallel with conventional flotation columns. Laboratory tests were conducted to simulate the performance of the existing plant when compared with the addition of the Jameson cell allowing an extra sulphidising phase. This paper reviews the comparison of the Jameson cell and conventional columns in recovering cyanide soluble copper, while maximising copper concentrate grade. Key parameters that are investigated are residence times, air/pulp ratios, superficial gas recoveries and size by size recoveries. A brief comparison of performance of the Jameson cell and column in the controlled potential sulphidisation process is also presented. An estimation of the reduction in cyanide consumption is given.

Jan 1, 1992

By A. Probst, J. A. Finch

"Solids concentration (weight % solids) profiles and the clear liquid/slurry interface level (mud-line) were determined from conductivity measurements in thickeners at Falconbridge's Strathcona mill (Ni concentrate and non-mag thickeners) and Kidd Creek Division (Cu, Zn concentrates and final tails in the concentrator and a jarosite thickener in the Zn plant).The conductivity measurement system consisted of a probe 1.5 m in length with 23 electrode cells, a 24-channel relay, a conductivity meter and a computer. A conductivity profile is collected which locates the mud-line by a sharp change in conductivity. The conductivity measurements are readily converted to solids concentration profiles using a model due to Maxwell.The results from conductivity measurements were compared with direct measurements on local samples obtained using a modified Noranda sampler.It is concluded that the conductivity technique is applicable to locate the mud-line and to estimate the solids concentration profiles in thickeners. This on-line informnation could improve the operational efficiency of industrial thickeners."

Jan 1, 1994

By R. H. McCorkell

Radon emanation rates. J. and compositions of materials from the Canadian Certified Reference Materials Project (CCRMP) indicate that the fraction of the Rn produced that escapes, E, is usually 1% to 8% and the 238U decay chain is nearly in secular equilibrium. For some Chinese ore standards, E values are higher than f or the CCRMP materials. Standard ore tailings have 2/oPb concentrations higher than expected from their 226Ra concentrations and Rn escape rates (12 - 34%) . Biological materials have E ˜100%, their ash E ˜0%. Pelletizing ore powders at 20 000 lb/in2 reduces J by only 20 - 40% but only 2% of Rn escapes from pressed Li2B4O7 pellets and 0.01 - 0.04% from fused Na2B4O7 or Li2B4O7 discs. Large, difficult-to-explain random variations in J are observed in some situations but several highly reproducible ways of measuring emanation rates have been found.

Jan 1, 1986

By Frank F. Aplan

In looking into the crystal-ball scene of minerals beneficiation for 1970 and beyond, an impartial observer becomes con- fused as to what has been done and what is possible in the way of improvements. The mining official must balance the burden which is being shouldered by the mineral industry for maintaining our environment in its pristine state on one hand, with the solid advancements made in technology and cost savings on the other hand. Then the official wonders if this is really the criterion for judging overall performance today. As a prospective engineer, Dave Bolles sees it this way: "Specialization has caused many engineers to subordinate everything to doing their job the cheapest and fastest way. In the future this won't be enough. The engineer in the mineral industry will be responsible not only for mining, concentrating and smelting, but also for the resultant wastes- the mine dumps, the tailing dams and the slag piles." (See Drift on page 63).

Jan 1, 1971