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By Alvin F. Kroll

BUNKER Hill & Sullivan Mining & Concentrating Co. has three major sources of ore at present: high grade square-set ore, low grade caving ore, and low grade reclaimed washed jig tailing. These three ores constitute the feed to the 3000-tpd concentrator that produces lead-silver, zinc, and pyrite concentrates. Each ore has different characteristics of grindability, mineral particle size, composition, and content, but all are amenable to the same mill circuit and reagents. The gangue is primarily quartzite. The Bunker Hill concentrator is now using a coarse primary grind followed by a regrind of sand fractions of lead rougher tailing, lead middling, and lead cleaner tailing, and further regrind of sand fractions of zinc middling and zinc cleaner tailing. This is done in order to recover all valuable minerals at the largest size possible. The tendency to overgrind the ore particles, which are more easily ground than the gangue particles, is thereby reduced-lessening their loss in the flotation tailings. The coarse primary grind and regrind method has the advantages of: providing more selective grinding, minimizing sliming, giving a cleaner lead-zinc separation, and reducing mill requirements for reagents and water. Improved lead-zinc separation raises overall recovery and also increases lead and zinc recovery in the respective concentrates by keeping middling particles out of concentrate and tailing.

Jan 6, 1954

By M. R. Khalesi, Y. Kianinia, A. Seyedhakimi, F. Soltani

"The feed ore to Agh-Darreh gold processing plant (Takab City, Iran) contains 2 ppm Au, 7 ppm Ag, and 120 ppm Hg. After cyanide leaching of the feed at pH 10.4, 80% of the total mercury (in the form of cinnabar and metacinnabar) goes to the tailings dam. In this research, separation of mercury from the tailings by flotation was investigated. The effects of collector type and dosage, pH, and the number of cleaner stages were studied. The results showed that after two stages of cleaning a 40–62% recovery of Hg at a grade of 14.3% Hg is attainable. Recycle water from a test in which 50 g/t amyl xanthate and 30 g/t pine oil was used in the flotation stage showed no adverse effect on the leaching and adsorption of gold onto activated carbon.IntroductionMercury, most commonly found in the nature as the mineral cinnabar (HgS), is a hazardous substance because of its mobility and toxicity (Kyle et al., 2012). Roasting is the wellestablished treatment method for mercury ores, in which the sulphur is burned and volatilized and mercury recovered subsequently by cooling and condensation. However, for low-grade ores, flotation is mainly applied for primary concentration due to its comparatively low costs, low environmental hazards, and flexibility regarding feed variations (Bulatovic, 2007). Particularly, cinnabar can be easily recovered by flotation using pine oil and a collector such as xanthate (Bulatovic, 2007; Crozier, 1991). Floatability of cinnabar is reduced due to surface oxidation when it is exposed to the atmosphere after crushing. This problem can be eliminated by adding copper sulphate and lead nitrate or acetate. The best pH for cinnabar flotation is in the range of 6.5 to 8.5 according to Erspamer and Wells, (1954), who reported that cinnabar and stibnite can be floated with long-chain xanthate collectors and an alcohol frother, with cumulative recoveries of 97.8% and 94.7%, respectively. Lead acetate was used as an activator."

Jan 1, 2017

By N. Nišić, A. Cvjetić, V. Jovanović, D. Knežević, D. Nišić

The risks associated with operating water reservoirs and tailings storage facilities (TSFs) are different because of their different purposes, methods of construction and operation, and characteristics of the materials impounded and their flow behaviour. Regardless of the differences, these two types of structures are often put in the same category when it comes to risk assessment and the application of relevant methodologies, which may result in unrealistic outcomes. In this paper we investigate whether it is possible to apply the risk ranking methodology designed for water reservoirs to TDFs and overview the key differences between these two types of structures. We also provide a comparative analysis of the results obtained by conducting a risk assessment of the operation of a TSF, applying the method recommended by the International Commission on Large Dams (ICOLD), and analyse the results using the failure mode and effects analysis (FMEA) technique. It is concluded that a more realistic evaluation of risk levels can be obtained by adopting a systematic approach to risk assessment in accordance with the traditional risk definition on which the FMEA technique is based.

Nov 30, 2022

By L. X. Yu

As is known, the oxidized antimony and antimonate are hard to treat. A new combined mineral processing flowsheet which is controlled in grinding and classification ?bulk floatation?gravity separation is adopted to dispose of gravity-separated tailings (antimony content: 1.48%; oxidization rate: 68.92%) in Muli Antimony Mine, the grade of floatation antimony concentrates is 32.64%; the grade of gravity-separated antimony concentrates is 26.99%; the combined grade of floatation and gravity-separated antimony concentrates is 30.85%, and the total recovery of antimony is 70.64%. Adopting the new mineral processing technique have solved the problem of over grinding of antimony ore by controlling grinding and classification, recovering effectively the sulfide antimony and the easy floatable oxidized antimony of fine size by bulk floatation, recovering effectively the oxidized antimony which is hard to float by classified gravity separation. This is a great breakthrough in the mineral processing technique of the recovery of the oxidized antimony. At present, favorable economic benefit has been obtained from the application of this processing technique in Muli Antimony Mine.

Jan 1, 2014

By Ande&apos, Bengt n

The first Model 480 SALA-HGMS R carousel high gradient magnetic separator was recently manufactured (Figure 1). The carousel's design was based upon the high gradient magnetic separation technology developed and owned by Sala Magnetics Incorporated, Cambridge, Massachusetts. After a short (approximately 300 hours) in-house mechanical and electrical test and break-in period, the carousel was ready for a long term demonstration test. This paper describes the demonstration plant and operational performance of the Model 480 HGMS treating a low grade iron ore tailings. A location convenient to Sala International's main office in Sala, Sweden was desirable because Sala engineers and technicians were to operate the demonstration plant. The Strassa iron ore mine in central Sweden was chosen because it fit most of the criteria that was important for a good demonstration site. The Strassa mine is part of the state owned Svenskt Stal AB. The mine produces about 320,000 tons of iron concentrate per year consisting of 200,000 tons of magnitite concentrate and 120,000 tons of hematite concentrate.

Jan 1, 1980

By C. Ptacek, D. Hiller, J. L. Janbor, H. Mccreadie

Between 1991 and 1994, metallurgical residues from a pressure-oxidation autoclave at the Campbell Au-quartz mine at Red Lake, northern Ontario, were neutralized with dolomite-rich flotation tailings and were deposited in a polyethylene-lined disposal area. Mineralogical analysis indicates that most of the As occurs as a crystalline ferric sulfarsenate, and some is associated with Fe oxides and with K-H3O jarosite. Field pore-water geochemical data from 1994 and 1995 showed increases in Fe(II) and As(III) concentrations, suggesting that the sulfarsenate and jarosite are dissolving; none of the discrete As phases included in a thermodynamic database limit dissolved-As concentrations, and these concentrations may continue to increase. These results demonstrate the potential for Fe and As loading through groundwater systems.

Jan 1, 1998

By M. B. Haggerty, R. J. Sheets, F. Abbasy

To address coal refuse tailings storage needs for a mine operation in western Canada, a feasible solution was to convert excavated open pits into tailings cells. A slope stability investigation and design study were completed to determine the necessary width and slope geometry for in-situ native plugs as the impoundment embankment. The native plugs separate the previous open pit from the current active open pit. This configuration allows for the deposition of refuse tailings in the previous pit while allowing mining to safely continue “downstream” of the temporary impoundment. As mining progresses, each subsequent pit will be filled with refuse tailings; thereby, buttressing the upstream slope of the previous native plug. Although these cells are excavated, and the timeframe for downstream exposure is relatively limited, the native plugs are classified as dams and must be designed to meet applicable safety of dam standards since they impound water and tails above mining personnel in the subsequent open pit. The presentation and paper will discuss the investigation and analysis conducted to develop the geotechnical and hydrogeological design recommendations.

Mar 2, 2022

By R. Kingman

At Tahawus, New York, National Lead Company operates the MacIntyre Development, This operation produces 10,600 long tons of ore per day for crushing, 8,000 tons for grinding and processing, of which 2,000 tons per day reports as tailing for disposal in an impounded tailing storage area. With the change in location of mining operations at the mine of National Lead Company, a change in the area for tailing disposal by gravity became necessary, This called for the diversion of tailing from near the proposed mining area to areas more favorable for long term storage by gravity. In order to avail the operation of maximum storage for the future, a uniform minimum grade launder was proposed with a minimum velocity and maximum outfall at the discharge. Disposal of the gravity and flotation tailing was formerly accomplished in open ditch from the concentrator at a grade of 1.5 percent. It was desirable to design the new launder at grades less than this for the greater storage elevation.

Jan 1, 1966

By J. Svoboda, W. J. C. Venter, R. N. Guest

SYNOPSIS Wet high-intensity magnetic separation was examined as a method for the recovery of copper and lead from Tsumeb flotation tailings. Bench-scale and pilot-plant tests were conducted and, despite the very low magnetic susceptibilities of all the complex copper and lead minerals, copper and lead recoveries of over 60 per cent and almost 50 per cent respectively were obtained at a combined copper and lead grade of up to 13,5 per cent. The recoveries were found to be practically independent of magnetic induction in the range 0,5 to 2,2 T, and the grade of the magnetic concentrate decreased with increasing magnetic induction. The grade of the magnetic concen¬trate improved when the flowrate of the slurry was increased, and desliming was found to have a profound effect on the. metallurgical process. Particles smaller than 10 µm were not recovered, and their presence in inefficiently deslimed samples impaired the overall performance of the magnetic separation

Jan 1, 1988

By J Siliézar

The Ernest Henry Mine (EHM) is situated 38 km north-east of Cloncurry in the Mount Isa – Cloncurry mineral district of northwest Queensland. The EHM orebody is an iron oxide copper gold deposit with an average grade of one per cent copper, 0.5 g/t gold and 23 per cent magnetite, with current reserves of approximately 88 Mt.The copper concentrator is a single line plant with a nominal throughput rate of 1300 tonnes per hour. An average of 350 000 tonnes of concentrate is produced each year containing around 100 000 tonnes of copper metal and 120 000 troy ounces of gold.In December 2009, Xstrata Copper announced approval of a $589 M investment to extend the life of EHM to at least 2024, through the transformation of open pit mining operations to a major underground mine together with the construction of an associated magnetite extraction plant.The EHM magnetite plant extracts the magnetite from the copper concentrator’s tailings stream. It will produce approximately 1.2 Mt of magnetite concentrate per annum at full capacity for export to Asia, making EHM Queensland’s first iron ore concentrate exporter. Construction of the magnetite extraction plant commenced in July 2010 and was completed in January 2011.This paper gives an overview of the development of the EHM magnetite extraction plant flow sheet from the early mineralogy and laboratory test work through to the plant commissioning and early operation of the base plant. It has a major focus on the commissioning learnings from the plant start-up and includes a review of the operational performance of the plant versus original design and expectations.

Jul 11, 2011

By Patrizia Pereyra, M. Guzman, C. Toledo, J. A. Rau, A. Ruiz

The use of mining tailings-based geopolymer binder as a substitute for cement can be an interesting alternative to the problem of final disposing of large volumes of them. Indeed, mining tailings have a high silica content, which makes them suitable for geopolymerization. The tailings samples contain Cu, Pb, Zn, As, Sb, Cd and Bi, mainly. X-ray diffraction results show that the tailings are composed of quartz, SiO2 (81%). Geopolymerized tailing samples (GTPS) were obtained by varying the ratio of SiO2/ Al2O3 and SiO2/NaOH. The compressive strength of the GTPS varied between 2 MPa and 8 MPa. The electrical conductivity in the leaching tests varies between 13.91– 16.11 mS/cm2 indicating that ions are present in the solution. However, the pH = 10.5 indicates that acidity is not being generated. The element that mostly leaches is iron. The highest percentage of leaching is observed for cadmium (69%) and chromium (93%). The concentrations of elements are under the maximum permissible limits for water according to the Peruvian legislation. However, mass transfer leach tests and immersion leach tests must be realized.

Feb 1, 2024

By Edward Kline

A review and synthesis of literature pertaining to biological consequences of submarine mine-tailings disposal (STD) was conducted. STD can result in massive sea floor sediment deposition. STD may also increase suspended sediment, trace metals, and residual milling reagents in receiving waters. These perturbations, which are highly site dependent, invariably smother benthic organisms, and could potentially affect or alter fish, plankton, and benthos through acute and chronic toxicity, bioaccumulation, behavioral changes, smothering, derived secondary effects, and habitat alteration. Much of the available information concerning STD is unpublished and limited to a restricted number of sites. From this information and related research, it appears that benthic smothering is the only major consequence of a properly designed STD system. There have been cases of metal bioaccumulation due to inadequate preliminary evaluation, Also, shallow water habitat alteration has resulted from STD. Reductions in biological production are likely due to benthic smothering. The rate of ecological recovery after termination of an STD operation varies and is difficult to assess, The consequences of STD are site specific. Additional research is needed concerning possible consequences of milling reagents and the broad implications of benthic smothering. Also, methodologies are needed to allow better prediction of benthic recolonization and ecological implications of metal bioaccumulation.

Jan 1, 1994

By Leandro Seixas Bicalho, Wagner Cláudio da Silva, Kelly Laranjeira Nunes, Lívia Marques Faustino

Seeking to implement the Global Industry Standard on Tailings Management to improve the safety of their tailings facilities across the globe, mining companies are motivated to invest in new technologies to recover valuable minerals from tailings. One solution being evaluated is the use of magnetic separators to concentrate ultrafine tailings. However, this approach is energy intensive as it requires high magnetic fields to recover ultrafines, and usually results in low mass yields. Additives that modify pulp rheology and surface properties, or selective flocculate ultrafine particles, were tested to enhance the performance of magnetic separation. Promising results in laboratory scale testing different types of additives confirmed that it is possible to increase the mass yield, keeping the selectivity of this stage.

Feb 1, 2024

By Robert C. Bates

Variations in water percolation rates through saturated cohesionless hydraulic backfill (mill tailings) have been studied in relation to the many physical aspects of the system to determine the most important factors in causing these variations. A multivariable least-squares regression analysis of 135 separate tests shows that a reliable estimate of the water percolation rate through typical hydraulic backfill can be made if the bulk density (unit weight), specific gravity, and grain-size distribution are known or can be estimated. Equations for predicting percolation rate, coefficient of permeability, and seepage velocity are developed both for the range of data analyzed and for extrapolation. The prediction equations for use within the range of data studied are quite accurate, while the ones developed for extrapolation are less accurate but still give reasonable approximations of flow rates. The equations given are valid for cohesionless materials that are free of clays and micas, exemplified by classified mill tailings. The equation developed for predicting percolation rates is as follows: 1nP20 = 11.39147 + 2.853422 1n(VR X D10) + 0.1747436 VR X Cu -178.8039 D10 X D50 + 311.7034 (DSO )3, (1) where P20 = percolation rate, inches per hour at 20° C; correction factors for other temperatures are given in the text, VR = void ratio, dimensionless, D10 = effective size (diameter in millimeters of the 10-percent-finer grain size), Cu = coefficient of uniformity (ratio of the diameter of the 60-percent-finer grain size over D10), D50 = average size (diameter in millimeters of the 50-percent-finer grain size), and In = natural logarithm (log to the base e). The prediction equation for percolation rate is rather complicated, so for practical use a nomograph that simplifies the solution of the prediction equation was developed.

Jan 1, 1967

By J. Stowe

While there has been continual improvement in the models for predicating deposition velocity and pipeline friction losses, the fundamental design philosophy for tailings pipelines has remained the same over the previous three or four decades: design the tailings pipeline to operate in turbulent flow without a deposit on the pipe invert for all expected process conditions. Recent improvements in the understanding of slurry pipeline flows allows for the design of pipelines operating with a deposit on the pipe invert, allowing the system to operate at moderate velocities for high flow rates, optimizing system energy requirements and maximizing pipeline wear life. Tailings pipelines are increasingly operated at higher solids concentrations where there is a greater likelihood the pipeline will operate in laminar flow due to the higher tailings slurry rheology. Historically, slurry pipelines have been operated in turbulent flow. We now know that laminar flow pipeline operation is possible as there are several operational laminar flow tailings pipelines. While our understanding of laminar flow slurry pipelines is not complete, recent research provides guidance on how to approach the design of these pipelines. New sensor technology has led to the development of an instrument that provides reliable indication of bed deposit height within a pipeline. This technology is particularly helpful in operating pipelines with a deposit and has application in better understanding the operation of laminar flow pipelines. Failure of tailings pipelines can result in consequent damage to surrounding structures and unacceptable spillage into the environment. Modern pipeline leak detection technology allows operators to better manage this risk by ensuring safe and reliable operation of tailings pipelines. This paper provides an overview of the each of the above advances in the design and operation of tailings pipelines. OPERATING TAILINGS PIPELINES WITH A DEPOSIT Tailings pipelines, and slurry pipelines in general, are typically designed to operate without a deposit on the pipeline invert. Meaning, the minimum operating velocity exceeds the turbulent deposition velocity. This approach has been adopted due to concerns regarding potential blockage of a pipeline operating with a deposit. Figure 1 shows a typical design duty envelope for a tailings delivery system as a function of solids throughput and solids concentration by mass. If the pipeline is designed to operate without a deposit at the lowest duty flow rate (4,480 m3/h), the operating velocity at the highest duty flow rate (6,590 m3/h) is about 1.5 times the deposition velocity. Considering that the pipeline wear and energy requirements are roughly proportional to the square of operating velocity, this results in an operationally inefficient system when the pipeline is operated close to the maximum design flow rate. Most systems tend to operate at maximum throughput after post-startup debottlenecking optimizations. Alternative Approach An alternate design approach is to ensure the pipeline operates without a deposit for the nominal flow rate and accept that a deposit will form when operating at lower flow rates. This results in the selection of a larger diameter pipeline, but there are significant benefits in terms of reduced energy requirements and lower pipeline wear.

Jan 1, 2019

By Dr. Rory Cameron, Dr. Terry Cheng, Dr. Nadia Mykytczuk, Ian Bedard, Kristie Tarr, Ruth Joseph

Pyrrhotite-rich tailings from Sudbury, Ontario mining activities are known to contain significant amounts of EV metals, such as nickel and cobalt.  CanmetMINING has developed a novel bioleaching process for recovering the nickel and cobalt (both critical for vehicle electrification) from tailings that incorporates partial neutralization and iron removal in the bioleaching circuit; bench- and pilot-scale tests have resulted in excellent nickel and cobalt recovery over a range of temperatures.  However, this process generates a new bioleach residue, which is heavier and more voluminous than the original tailings due to pyrrhotite oxidation and jarosite formation during the bioleaching process. This residue can be problematic for long-term disposal. A possible solution being studied for this issue is using mine backfill as a disposal mechanism.  Underground backfilling allows for effective re-purposing of mine waste, however the backfill must also be able to withstand a certain load, since in most cases it is used as an underground construction material, such as in cut-and-fill mining methods.  Therefore, the composition of the mill tailings must be such that when combined with a binding agent and pumped underground, the slurry or paste is then able to withstand a certain load within a certain period of time.  Unfortunately, high-pyrrhotite content tailings, commonly found on mine sites in Ontario and Quebec, are reactive and can easily oxidize and produce acidity and/or metal-laden drainage under some disposal conditions. This means that the pyrrhotite in the tailings can lead to acid mine drainage if stored on surface, however, it can also cause strength degradation if re-purposed in too high a concentration in underground mine backfill.  The research being done at CanmetMINING related to the bioleaching process is aimed not only at recovering nickel and cobalt from mine tailings, it is also being used in an attempt to neutralize the tailings to reduce acid mine drainage and backfill strength degradation. This paper reports on research designed to evaluate different options for transforming the bioleaching process residue into iron oxide phases, for better environmental footprints if stored on surface, as well as increased disposal possible, if re-purposed as underground mine backfill material.  

May 4, 2025

By R. S. Shoemaker, F. W. McQuiston

ECOLOGICAL CONSIDERATIONS IN CYANIDATION PLANT PRACTICES Ecological controls of tailing storage pond effluents are mainly concerned with mercury and residual cyanide and are critical problems for gold and silver recovery plant operators. For example, mercury for direct amalgamation of gold ore was used at Homestake until 1970 when the Federal Government acted to prevent further use thereby forcing the abandonment of this practice. However, mercury is still used at Homestake, in South Africa, and also in some Philippine and Canadian cyanide plants for barrel amalgamation of free gold gravity concentrates. The amount of mercury involved in South African practice is about 150 ounces for every 100,000 tons of ore milled and which reaches the tailings ponds. In South Africa, it is obligatory that no pond effluent be discharged beyond the mining property boundaries and waters containing mercury and cyanide must be evaporated either on the tailings pond or in specially constructed evaporation dams. In some cyanide plants such as Carlin and Cortez in the United States, all supernatant solution from the tailings ponds are returned to the plant for reuse. When all solution cannot be reused, the excess may be evaporated in shallow ponds, or by spray systems inside of the tailings dam. Such plant reuse of tailings effluent depends strictly upon the degree of contamination or "fouling" of the barren solution and tailing water with cyanicides. Foul solutions that cannot be returned from the tailings pond may be effectively cleaned up by cyanide regeneration. Cyanogens in plant solutions exist chiefly as free alkali cyanides, as zinc and copper double cyanides, as sulphocyanides, and as ferrocyanides. From the free cyanides and zinc double cyanides, substantially complete recovery of the cyanogen is easily effected. Part of the cyanogen combined with copper is also recovered; regeneration of the remainder with the cyanogen existing as sulphocyanide and ferrocyanides requires special treatment. At this stage, it would be generally more practical to eliminate the remaining cyanogen by chlorination. Ecologists have also been concerned with pollution of ground waters from cyanidation plant tailings. An accute problem is solution seepage down stream from tailings dams and strict attention to prevent or collect any seepage is a real necessity. The tolerance levels of the total cyanide content in potable water set by several U. S. and State health agencies are 10 ppb, and 20 ppb discharged into ground waters. The cyanide content so established is, strangely enough, based on the survival tolerance of fingerling trout. Decanted waste water from tailings dams may be effectively treated by the use of alkaline chlorination for the purpose of destroying all cyanide compounds. The alkaline hypochlorite used to oxidize the cyanide to cyanate is applied as such or generated within the waste solution by adding caustic (NaOH) and chlorine gas. A second chemical reaction involves complete destruction of the sodium cyanate.

Jan 1, 1975

By N Eriksson

BolidenÆs Aitik mine, located 100 km north of the Arctic Circle in northern Sweden, is one of EuropeÆs largest copper mines. This operation produces 18 Mt of ore per year and 99 per cent of this tonnage needs to be deposited as tailings. The tailings pond currently covers 1140 ha but at closure it will hold around 760 Mt of tailings and occupy an area of 1400 ha. The tailings contain less than one per cent sulfur and only 0.03 per cent copper. The carbonate content is as low as 0.25 per cent but silicates, with substantial accumulated buffering and metal adsorption capacity, are abundant. A programme of comprehensive material characterisation and investigations has been undertaken in order to design a decommissioning strategy that meets the regulatory requirements and secures long-term physical and chemical stability of the deposited tailings. The main concern has been the potential generation of acid rock drainage (ARD). Standard tests using acid base accounting (ABA) and humidity cell tests during 15 weeks were not conclusive regarding the ARD generation potential. The accuracy of the results was considered to be unsatisfactory, therefore the humidity cell tests were gradually extended in time. In total, almost four years of humidity cell tests were conducted on various sets of samples which were selected in order to cover the variation in sulfide content existing on site and in various points of the milling process. The laboratory results were subject to detailed geochemical evaluation in order to fully understand the interaction between acid-producing reactions, driven by weathering of sulfides, and acid-consuming reactions by dissolution of carbonates and alumino-silicates. The laboratory tests were followed by both pilot tests and field measurements in order to estimate the scale dependency between the laboratory and the field results. It was found that sulfide weathering rates in the field were considerably lower than those measured in the laboratory. However, it could not be verified that the tailings top layer would maintain a neutral pH in the long term even under field conditions. Thus, a reviewed tailings management plan, including de-pyritisation of the tailings during the last years of operation, is considered.

Jan 1, 2003

By G. I. McPhail

The authors have conducted several investigations into the prediction of the beach profile of high-density tailings and have found that there are a number of important considerations that need to be taken into account over and above the reliability of the beach profile prediction methodology. These considerations relate to the chemistry of the slurry and specifically the influence of pH and flocculation, both of which have a significant influence on the rheology of the slurry at the discharge point. A number of examples are discussed and a potential testing protocol that will assist in assessing the influence of chemistry ahead of the beach profile predictions is presented. A methodology for assimilating the results of the testing into the likely field situation is discussed, and suggestions are made on application of the results in beach profile prediction. The suspected influence of residual flocculation on the stratification

Jan 1, 2012

By Roy E. Williams

1. The major sources of heavy metal ions in wastewaters of the Coeur d'Alene Mining District are effluents from metallurgical processing plants. Effluents from mine operations provide a secondary source for heavy metal ions. Based on data analyzed by the Federal Water Quality Administration, mercury and arsenic are escaping from some mill waste ponds, but they are in the suspended phase. 2. Metallurgical process effluents are combined with mine waste effluents in tailings ponds, consequently, the quality of the tailings pond decant water is highly degraded. The major effect of combining metallurgical process effluents with mill wastes is a drastic change of pH in the tailings pond. This chance may be an increase or a decrease depending on the metallurgical process. The effluent from an electrolytic antimony plant drastically increases the pH in the tailings pond permitting the removal in solution of antimony via the pond decant system. The addition of low pH effluents to the tailings ponds results in a net decrease in pH which makes lead, zinc, cadmium, and copper more soluble, thereby permitting their removal from the pond via the decant system.

Jan 1, 2011