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By G. Agar

"E. Joe We have been working for some years on the elimination ofCANMET thiosulphates from solutions as a means of protecting ground water and fish streams. Alot of these thiosulphates result from the mill use of S02. Is there any possibility that thiosulphates will be generated by this process?Reply I don't know the answer to your question. I can speculate that there is no question that when one adds SO2 and air to a sulphide pulp one generates a host of unsaturated poly¬cyanates. I unfortunately cannot remember the reactions which arise in the cyanide destruction scheme. I think that if thiosulphate was generated the process wouldn't work. It is not oxidizing enough to destroy the cyanide.R. Wagner Would you comment on the actual cost/ton at the Scottie goldPamour mine concentrator? Would $2.50/ton be typical or is thisPorcupine just a ball park figure.Reply We wouldn't disclose Scottie's costs even if we knew them. I depends on the amount of cyanide and thiocyanate in the effluent. I have given you a hypothetical case. With the conditions stipulated the figures I Have given are realistic.K. Konigsmann I do not believe the use of S02 adds significantly to the thio Noranda Mines salt problem Certainly we use SO2 in many of our mills and thio-salt levels in a number of them haven't changed whether we use or don't use S02. For instance, Orchan and Matagami mills treating the same ore side by side in which one was using S02 and the other wasn't, the thio-salt level was virtually the same in both pulp streams. In the case of Brunswick where alot of SO2 is used (20 tons/day) there is a thio-salt problem but it is relatively insignificant.Reply Thank you for that comment. It serves to reassure us that we are not polluting the environment."

Jan 1, 1984

"The ore currently being milled at Delnite is a high pyrrhotite ore with the gold intimately associated with the pyrrhotite. To get maximum recovery of the gold with the lowest reagent cost, it is necessary to keep the pH as low as possible In the laboratory the best results were obtained with a pH of 10 but in the plant it was found that a pH of 11 was the minimum practical pH. With the n1anual control it was difficult to maintain this figure.The Mines Branch had a conductivity probe (see Fig. 1), developed by Dr. G .G. Eichholz that they wished to try out in plant operation.On June 17, 1963, the probe with all necessary control equipment was installed in the regrind circuit (see Fig. 2) at Delnite Mines. Crushed quick lime is fed at this point but it is also necessary to feed hydrated li1ne at two other points in the cyanide circuit in order to maintain the pH close to 11. 0 throughout the circuit. Only the feeding of the quick lime at the regrind mill was controlled in this test.One thing that concerned me with the operation of the probe was whether the reading on the probe chart would correspond to the pH. A look at the graph accompanying this paper will show that except for very minor variations the graphs of the chart reading and the pH reading are parallel. The upper graph shows that the pH reading and the lime titrations also correspond and that the cyanide titrations appear to have little effect on the pH reading.Table 1 gives the results obtained for the two months before and after the instal1ation of automatic control as well as the two months of control. Unfortunately there was a decided change in ore just after the automatic controls were removed. There was much less pyrrhotitc in the ore than usual thus changing the need for lime and cyanide."

Jan 1, 1964

By W. A. Grell

"The oxygen meter has been in use at MacLeod Cockshutt Gold Mines for over a year, At MacLeod about 200 tons per day of reground concentrate containing roughly 10% pyrrhotite and 20% pyrite is cyanided in four 181 x 18' Dorr agitators. A survey by the Department of Mines showed that oxygen content of some of the agitators was rather low. Experience with our oxygen meter indicates that the oxygen content must at times have been too low for good gold dis solution.We have found that in winter, when pulp temperatures drop to about 50° F, oxygen saturation in the agitators is not difficult to maintain. In summer, however, when pulp temperatures rise to 70or 75°F, the oxygen content will drop very readily. It was found that a high oxygen content can be maintained readily if the pulp is sufficiently dilute. Thus at 50° F, a pulp of specific gravity 1. 75 i.e., about 55% solids, will saturate; but at 70 or 75° F, the specific gravity must be decreased to 1 .45, equivalent to about 40% solids, in order to maintain saturation. Very thick pulps, such as thickener underflows, are noticeably difficult to saturate in the laboratory, and lose their oxygen rapidly.Clear solutions, on the other hand, absorb oxygen readily and hold it. The primary thickener overflow, before agitation, is 100% saturated; after storage of 3 to 4 hours it is still 100% saturated at re clarifier tank. Barren solution after spraying in a launder is 100% saturated.We have not yet attempted to determine what percentage saturation is necessary for good extraction, but it is known that there is a point at which extraction decreases rapidly. So far, the measures required to maintain over 80 or 90% saturation have coincided with measures required to hold the reducing power to the point believed necessary."

Jan 1, 1964

By A. Fyson, M. P. Smith, Margarete Kalin

"Drums containing 0.07 m3 of pyritic waste rock from a zinc/copper mine, with and without Natural Phosphate Rock (NPR) additions (40 kg/t waste rock) were set up outdoors. The effluent chemistry has been monitored over a period of 709 days. In the presence of NPR, the cumulative acidity discharged from 7 out of 8 drums was 80 % less than that discharged from drums without NPR. Mixing NPR throughout the drums was more effective in reducing acidity than placement of the NPR in a layer in the drum. NPR treatment was more effective for 'fresh' waste rock than for material which had weathered for more than 4 years. It is concluded that application of NPR to pyritic waste rock could greatly reduce acidity generated.IntroductionWaste rock dumps are significant sources of acid, generated by the oxidation of sulphides. Oxygen and water supply are oxidation rate-controlling factors, and oxygen supply is the reaction-limiting component. If secondary minerals can be formed on the surface of the rocks, pyrite oxidation and therefore acid generation may be reduced or inhibited. Neutralizing agents such as limestone do not prevent the oxidation of pyrite but treat the resultant acid mine drainage (AMD) or acid rock drainage (ARD).Meek et al. (1991) determined in pilot scale experiments that apatite mixed with tailings was more cost effective than liming or use of PVC liners to reduce oxygen and water penetration into the tailings. Phosphate is not acting as a neutralising agent but it creates a chemical condition on pyritic surfaces which prevents reactions of the oxidising agents. Precipitation of principally iron phosphates physically isolates the pyrite surfaces from further oxidation (Ziemkiewicz 1990; Evangelou 1994). The chemistry of the phosphate effects has been described by Spotts and Dollhopf (1990). Secondary mineral formation was evident in the leach pile at Gibraltar when leaching was impaired (Scott, 1991) indicating a link between oxidation and the mineral-formation process. This paper reports on effluent characteristics when waste rock is generating acid in the presence of NPR."

Jan 1, 1995

By E. G. Joe

"The continuing increase in costs for power and fuel are affecting significantly the cost for milling of Canadian ores. In order to ascertain the scope and source of energy consumption in Canadian mills for further energy conservation opportunities, CANMET, in cooperation with the Canadian Mineral Processors carried out a detailed survey on the metallic mineral processing plants. Sixty-seven mills representing almost 300 million tons of annual capacity participated in the survey. The results indicated that in general, copper and iron ore mills, representing the bulk of the tonnage, consumed relatively low levels of energy per ton. Nickel-copper, zinc-copper and lead-zinc ore mills were moderate in energy consumption, while uranium and some copper-lead-zinc mills reported much higher energy consumptions due primarily to the increased heat requirements of the ore pulp. While the plant heating costs varied somewhat with the location in Canada, it was apparent that the lowest heating costs were achieved by those built with proper insulation and energy conservation design. If the cost per million British Thermal Units is taken as $2.00, the mill energy costs generally ranged from 40¢/ton up to $1.20/ton.IntroductionEnergy consumption in milling has always been of significant interest to Canadian mineral processors who traditionally calculate their milling costs down to pennies per ton. Even the measure of an ore's grindability is expressed empirically by Bond in kilowatts hours per ton milled. Some of the mills located in isolated areas were obliged to generate their own electrical energy from fossil or petroleum-based fuel at a cost substantially higher than that obtainable from captive or public hydro-electric installations. These mills practiced energy conservation long before the advent of O.P.E.C. energy cost inflation. One example of this was the use of electric boilers to produce process and heating steam during off-peak electrical load periods."

Jan 1, 1978

By R. J. McLeod

"Grinding at Giant Yellowknife is done in two stages. The first stage consists of two ball mills in closed circuit will: mechanical classifiers, the classifier overflow goes to flotation. The tailings from this flotation circuit are separated into slime and sand portions, the slimes go to an additional flotation circuit while the sand portion goes to the secondary grinding mills.Milltronic automatic sound control units were installed on each of the two primary ball mills and put into operation September 3, 1966. After about two months adjustments to the control units the operators were taken off the grinding circuit on November 1, 1966. The operators duties were taken over by the automatic control units while miscellaneous work was taken over by the helpers in the mill. The labour saving by the installation is approximately $2250.00 per month.After a short period of operation it was apparent there was a slight change in grind. The screen analysis of the primary classifier overflows for the period May to August 1966 averaged 62.23% minus 200 mesh while for the period September to December 1966 averaged 62.0% minus 200 mesh. This is a slight coarsening of the grind. The change in the primary grind has manifested itself also in the screen analysis of the final grind. Before automation the final grind was 80% minus 200 mesh, while after automation it was 7 5% minus 200 mesh. A change of grind was also noted in the flotation concentrates. Before automation the screen analysis of the concentrates was 35% plus 200 mesh and 50% minus 3 25 mesh. A ft er automation it was 35% plus 200 mesh and 47 .5% minus 325 mesh. In other words the grind got coarser or perhaps we should say there was less overgrinding. Another significant change took place in the assays of the slime and sand portion of the primary flotation tailings. It was usual under manual control for the slime portion to assay higher in gold than the sand portion. The reverse is now true."

Jan 1, 1967

By Samantha A. Beeson, Colin Poole

"A selective flocculation-floe flotation technique is already in use in the Canadian potash industry for the removal of water-insoluble slimes, but the process is currently limited in its application and reagent consumption is high. The work described in this paper details investigations into the separation of sylvite and halite from water-insoluble slimes by selective flocculation utilising other commercially available polymers.Preliminary flocculation performance tests were undertaken on pure sylvite, halite and water-insoluble minerals to establish the importance of a number of experimental parameters. The data from these single mineral tests was then used to establish optimum parameters for the separation of sylvite and halite from a synthetic mixture similar in composition to a typical potash ore. Under optimum conditions, separation was achieved with high recoveries using dispersant and flocculant dosages significantly lower than those currently used in plant practice.IntroductionAlthough solar evaporation and solution mining are used in some operations, most potash is extracted by conventional underground techniques and is then processed by froth flotation, in order to separate the valuable sylvite (KCI) from halite (NaCl) and water-insoluble gangue minerals. Flotation is usually preceded by a desliming stage to remove water-insoluble fines, but this also results in high potash losses and affects process brine requirements. To overcome these problems, one plant operated by Cominco in Saskatchewan uses a selective flocculation -floe flotation process (Banks, 1979/1980), in which a polyacrylamide flocculant is added to produce floes of water-insoluble material with minimal incorporation of halides. A collector is then added to render the floes hydrophobic and suitable for removal by froth flotation.A number of workers have also investigated the use of selective flocculation in potash processing, but these too have been concerned with developments in the removal of water-insoluble slimes (Williams, 1960; Brogiotti et al, 1974; Chan et al, 1986; Thompson et al, 1979; etc.). This paper provides results of novel investigations into the selective flocculation of sylvite and halite from water-insoluble minerals and discusses the possible mechanisms involved."

Jan 1, 1995

By M. Remillard, M. Garon, M. Thérien, D. Hodouin

"Nineteen streams in the copper and zinc flotation units of the concentrator of the Matagami division of Noranda Inc. have been sampled for particle size distributions chemical assays, and gold and silver contents in each particle size interval. The data are upgraded by the mass balance package BILMAT. These results, in conjunction with microprobe analyses, are used to quantify the silver losses in the copper clean¬ing flotation stages and to determine how overall silver and three silver-bearing minerals are recovered in the plant final products.RESUME Dix-neuf produits des circuits de flottation du cuivre et du zinc de la division Matagami de Noranda Inc. ont été échantillonnés pour leur distribution granulométrique, analyses chimiques et le contenu d'argent et d'or des fractions granulométriques. Le bilan métallurgique a été obtenu à l'aide du programme BILMAT. Ces résultats couples avec une etude minéralogique à la microsonde sont utilises pour quantifier les pertes argentifères aux diverses étapes de nettoyage du circuit de cuivre et pour de¬terminer la repartition de l'argent et des trois minéraux argentifères dans les produitsfinals de l'usine.INTRODUCTION The Matagami division of NORANDA INC. produces copper and zinc concentrates from a sulphide ore which contains a small amount of gold (0.34 g/t) and silver (20 g/t). Only a small part of the silver (~-30%) is presently recovered in the copper concentrate. The objective of this study is to evaluate the metallurgical behaviour of silver, and secondarily of gold, in the flotation circuit for a possible increase of its recovery in the copper concentrate to improve its economic value. For that purpose it is worthwhile to study the minerals which bear silver and to determine the critical flotation stages where they are lost in the tails. The methodology proposed here is based on a coupling of a mineralogical study and a detailed mass balance calculation in the flotation circuit."

Jan 1, 1986

By C. R. Edwards

"Equity Silver Mines Limited owns and operates an open pit silver 111ine located 37 km southeast of Houston in the Omineca District of British Columbia, Figure l. The property is approximately 300 km west of Prince George and 417 km east of the Port of Prince Rupert. This is the largest silver mine in Canada and one of the five largest silver producers in the western world.Equity is a subsidiary of Placer Development Limited, which owns 70 of Equity's issued and outstanding Common Shares. Placer was responsible for financing and developing the property and now provides general management and administrative services.All mining is by open pit methods, using three 1600 P&H shovels, eight 45 t Caterpillar and three 77 t Wabco trucks, three 40R rotary drills, two D8 dozers, one 9888 front-end loader and other ancillary equipment. We are now phasing out the Southern Tail pit and bringing the Main Zone into productior; see Figure 2.Comminution in the concentrator is achieved via primary, secondary, and tertiary crushers followed by rod and ball mills. The flotation circuit, with Southern Tail ore, produces a concentrate containing between 3 000 and 6 500 g of silver per tonne, 201 copper, and 25 g of gold per tonne. Running Main Zone ore, concentrate grades will decrease to about 4 500 g of silver per tonne, 17% copper, and 18 g of gold per tonne. Design milling capacity is 5 300 tpd, which is often exceeded.The flotation concentrate contains relatively high levels of antimony and arsenic. Because of these impurities Equity was unable to obtain economically acceptable long-term sales contracts for flotation concentrate. For this reason a leaching process was developed, and a plant designed, built and placed in operation. Due to the complex and unique technology involved, many start-up problems were encountered but by March 1983 the leach plant was considered ""fully operational."" This paper describes some of our progress and problems in bringing the Equity leach plant through its start-up period."

Jan 1, 1984

"An atmospheric process for the synthesis of scorodite(DMSP®)has beendeveloped,in which ferrous ions were oxidized by oxygen gas in the presence of pentavalent arsenic ions. Under optimum conditions, the synthesized scorodite (FeAsO4?2H2O) was well crystallizedand the crystals were faceted; the particle sizes were as large as 15 to 20 µm. In thisstudy, iron (II) sulfate heptahydrate (FeSO4·7H2O), arsenic (V) acid (60mass%H3AsO4) solution, and distilled water were used to prepare the original experimental solution (700 ml), in which the contents of Fe(II) and As(V) were 55.9 g/L and 50.0 g/L, respectively. The effects of various synthesis conditions, such as the type of oxidizing gas (oxygen or air), gas injectionrate, temperature, and the initial pH of the solution, on the size and shape of the scorodite particles wereinvestigated. From the experimental results, the changes in the shape of the scorodite particles wereclarified for each synthesis condition.INTRODUCTIONDOWA has developed a processfor synthesizing well-crystallized scorodite (FeAsO4•2H2O). Theprocess is named andtrademarked as DMSP® (Fujita et al., 2008; Fujita et al., 2008; Fujita et al., 2009; Fujita et al., 2009; Fujita et al., 2009, Kubo et al., 2010). In this process, oxygen gas is injectedinto a sulfuric acid solution containing large amountsof As(V) and Fe(II). Facetedscorodite crystals as large as 15–20 µm weregeneratedthrough the pre-formation of a gel-like precursor. However, it is known that the synthesized scorodite particles become smaller and more aggregatedwhen the synthesis conditionsdeviate from the optimumconditions.In this study, the effects of various synthesis conditions, such as the type of oxidizing gas (oxygen or air), gas flowrate, temperature, and initial pH of the solution, on the size and shape of the scorodite particles were investigated. The experimental resultsclarified the changes in the shape of the scorodite particles for each synthesis condition."

Jan 1, 2016

By Brian Penney

"The Iron Ore Company of Canada (IOCC} has mined and processed an iron deposit in Western Labrador along the Labrador-Quebec border since the early 1960' s. Current capacity is approximately 19 million tonnes of concentrate, 70% of which is converted to pellets at these facilities.The ore deposits are located in hills rising 150 to 300 meters above the adjacent lakes, and most of them are covered by very little overburden. The iron minerals are mainly specular hematite and magnetite hosted in a quartz formation. Crude ore averages 40% Fe that is concentrated to 65.5% Fe. IOCC is an extensive operation extending from Labrador City NF to Sept-Iles Quebec. It consists of open pit mining, concentrating, pelletizing, railway, and dock loading facilities.The Concentrator liberates iron from the silica by crushing and grinding the ore to -14 mesh (-1190 microns}. Ground ore is concentrated in spirals and filtered for transportation. Spiral tailings are treated in a Magnetic Recovery Plant (Mag Plant} that recovers lost magnetite. Mag Plant tailings are treated in a Hematite Spiral Plant (Hem Plant} that scavenges fine iron remaining in this stream.This report deals with ongoing iron recovery initiatives within the ConcentratorIron Losses in the ConcentratorIron recovery at IOCC is approximately 70.5%. This compares with recoveries that range from 65% to 86% for other iron mines in our area. Other world large iron ore producers dig and ship concentrate quality ore directly out of the ground and have an effective recovery of 100%. With increasing costs for labour, equipment, and supplies an increase in iron recovery is required to keep the operation viable.More than 90% of our iron concentrate is produced by gravity separation using spirals. The physical characteristics of the iron minerals that most effect recovery are mineral grain size and degree of liberation. We divide ground ore into three size categories -coarse, intermediate, and fines. A brief description of each size fraction follows:"

Jan 1, 1996

By T. Gonzales, M. Maleki, M. Caçador, B. Climenhaga, H. Ghorbani

The campaign life of furnaces in the copper making industry is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper monitoring and assessment of the existing conditions and safe remaining life of the vessels and their refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a typical fitness-for-service assessment of a rotary vessel shell in order to extend the campaign life of copper making furnaces and enhance safety. The furnace refractory management system involves a comprehensive audit and monitoring program which has been successfully applied to several copper flash smelting furnaces over the past 15 years. The fitness for service assessment provides a methodology to estimate the remaining life of the steel components by evaluating the existing damage and future damage rate based on the historical survey data. The above practices can provide the operator with cost savings by extending the vessel campaign life in lieu of implementing more rigorous refractory lining and shell audits. INTRODUCTION Primary copper smelting is generally divided into two categories of flash smelting and open bath smelting. There are many different types of vessels that can be used to convert the copper concentrate to blister copper including, flash, Mitsubishi process, TSL (ISA/Ausmelt), SKS Bottom Blowing, Anode and Teniente/Noranda furnaces, reactors, and Peirce Smith converters. Each vessel has its own complex and intense environment which requires careful refractory and steel design to withstand the severe thermomechanical and thermochemical attacks within its campaign life. The campaign life of furnaces is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper assessment of the existing conditions and safe remaining life of the vessels and/or refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a fitness-for- service assessment methodology to estimate the remaining campaign life of copper making furnaces resulting in cost savings and enhanced safety of the operating vessels.

Jan 1, 2019

By Jorge Lema Patino

"The purpose of this paper is to determine the best techno-economic alternative for the treatment of tin ores, when comparing the various mining and metallurgical methods of preconcentration, concentration, volatilization (fuming) and up to the final stage of selling to the smelter. The study of the influence of the different parameters of cost, recovery and quoted prices has motivated the paper, which aims at determining the best process of ore treatment in terms of its techno-economic feasibility.Method of WorkOn the basis of head grades of 0.5, 0.75 and 1.0% tin, the following alternative methods of treatment are possible:1. Mining-Concentration-Smelting.2. Mining-Preconcentration-Fuming-Smelting.3. Mining-Concentration-Separation, with direct smelting and fuming/smelting of the two fractions produced.The calculations were carried out on the basis of existing information and publications concerning the various experimental recoveries in the operating mills of COMIBOL. In the case of concentrates containing less than 10% or more than 40% tin, extrapolation was required.The final grade of the tin concentrates and the gross weight of the concentrate were calculated on the basis of the estimated recovery curves. The costs of the preconcentration and the concentration stage s were determined per gross metric ton of material treated.The calculation of the cost of the smelting stage is similar for all the alternatives, the royalties and smelting charge varying with the grade of the final concentrate. In the fuming stage, a uniform recovery of 90% was taken in all cases so as to obtain a final concentrate grade of 65% tin, regardless of the head grade of the feed material.For the separation stage, the following considerations apply:a) Separation can be carried out by gravity separation or flotation methods.b) The cost of separation varies depending on the system used.c) The distribution of tin in the two fractions is 100% at this stage. In other words, there is no loss of tin during separation."

Jan 1, 1982

By H. W. Smith, D. Guerin

"This paper describes the development of a dynamic model for a full- scale (64 tonnes/hr) grinding circuit consisting of an open-circuit rod mill feeding a ball mill operating in closed circuit. The object of constructing such a model is to investigate the static and dynamic behaviour of potential control schemes. Model selection, data processing and parameter estimation are discussed, and static and dynamic results of simulations are presented. Conclusions are drawn with regard to the adequacy of the model for such studies, and likely limitations on control of such circuits.IntroductionThe model developed here is intended to represent the grinding behaviour of the grinding circuit at the Lake Dufault division of Falconbridge Copper Mines Ltd. This circuit has been fully described elsewhere. At the time of the work described, the ball mill discharge was fed to a four-cell flotation bank , the tailings of which are combined with the rod mill discharge in the cyclone feed pump sump. As the emphasis in this study is on grinding behaviour, and as only I per cent of cyclone feed material is removed in these cells, no attempt has been made to include flotation behaviour in the simulation. One effect of the presence of flotation cells in the grinding circuit is, however, to necessitate somewhat finer grinding than might otherwise be needed: this will have effects on mill parameters.The paper is organized as follows. First the selection of models is discussed, paying particular attention to the problem of selecting simple, yet adequate, flow models for the mills. Then data gathering is described. The method of deriving mill model parameters is explained, and the results presented. Mass balancing of cyclone data and the determination of cyclone model parameters are described. The linking of the models into a full simulation is outlined, and sta tic and dynamic results presented. Finally, the results are discussed and conclusions drawn ."

Jan 1, 1980

By Robert B. Parsons

"Thrashing one's way through a jungle of entangled tax rules is no easy task today, as Canada's legislators continue to complicate this country's income tax laws. At some magical point, there is presumably some balance to be struck between legislation which is equitable and unambiguous on the one hand, and at the same time is efficiently complied with and administered. It is becoming increasingly clear that the legislation is not in balance, and that there is a pressing need for some simplification of the income tax rules and regulations especially in areas of critical importance to the resource industries-even though such simplification will likely involve some sacrifice of equity and certainty.Introduction""I would wish that ... the superfluous and tedious statutes were brought into one sum together, and made more plain and short, to the intent that men might better understand them."" Edward VI, King of England 1547-1553.Regal decree notwithstanding, the legislation of most developed nations continues to become more complex, lengthy, and confusing than ever before. Canada's income tax system cannot escape the accusatory finger pointing at examples of this serious problem which plagues all of us today.Although many areas of our tax law could easily warrant some degree of simplification, for those who are involved with this country's resource sector, that horrendous, almost unreadable, part of our tax legislation which deals with resource taxation must be regarded s a prime candidate for simplification. In recent decades, but particularly in the past ten years since tax reform in 1972, budget after budget, both federal and provincial, have amended, massaged, modified, and manipulated the resource tax rules to the point where not only are taxpayers totally lost in their attempts to make some sense of the entanglement of rules, but also the draftsmen themselves acknowledge that the outer limits of complexity are being reached, and that the whole set of rules is on the verge of grinding to a groaning halt under its own weight of nonsense and frustration."

Jan 1, 1982

By E. Ghali, W. Zhang, C. Su, N. Sorour, G. Houlachi

"Zinc electrowinning process is sensitive to the presence of metallic impurities such as lead ions that could substantially and negatively affect the power consumption and surface morphology. Effect of various concentrations of manganese ions and traces of lead ions up to 2.5 mg L-1 added to the electrolyte during electrowinning have been investigated by electrochemical polarization techniques such as galvanostatic, cyclic voltammetry and potentiodynamic. Scanning electron microscopy (SEM) and X-ray diffraction (XRD) were applied to examine the surface morphology and crystallographic orientation, respectively. Microwave plasma-atomic emission spectrometry has been used to determine the lead content in the deposit and in the electrolyte after electrodeposition. It was found that lead ions added to the zinc electrolyte led to the increase of cathodic overpotential and current efficiencies of zinc deposit. However, addition of Mn2+ into the zinc electrolyte resulted in the decrease of cathodic potential and current efficiency of zinc deposit. It can be stated as example that galvanostatic polarisation at 40?C and 52.5 mA cm-2 in presence of lead up to 0.1-0.2 mg L-1 in the electrolyte and 8-12 g L-1 Mn could be better considered as practical working parameters for electrowinning process.INTRODUCTIONApproximately 85% of the global production of zinc metal is produced by electrolytic process (Sinclair, 2005). The main goal of the zinc plant is to obtain pure zinc deposits with high current efficiencies. Smooth and pure deposits with high current efficiency (CE) are always obtained from pure electrolytes (Alfantazi & Dreisinger, 2001). The presence of metallic impurities in the zinc sulphate electrolyte is a major problem for the zinc electrowinning industry, since decreases in current efficiency, changes in deposit morphology and cathodic polarization are observed for electrolytes containing small concentrations of impurities (Ault & Frazer, 1988; Muresan, Maurin, Oniciu & Gaga, 1996). Low level concentrations as parts per million of these impurities are reported to influence negatively the zinc electrodeposition (Beshore, Flori, Schade & O'Keefe, 1987)."

Jan 1, 2016

By R. Ricketts, D. G. Vanderveer, F. Kirby

"A program of mapping, sampling and testing to determine the potential of the aggregate resources throughout Newfoundland and Labrador was commenced in 1978. A 6-km-wide corridor parallel to all transportation routes was selected as the study area.The study consisted of a preliminary surficial interpretation followed by field mapping and sampling. Granular resources were sampled at a 500-m interval and non-granular materials were sampled at a 1.5- to 2-km interval. Data were gathered on landforms and stratigraphy as well as a field description of the sample and deposit types. Particle-size and lithologic analyses of the samples were also conducted.The locations of samples and deposits were mapped on 1:50,000 topographic base maps and summarized for publication at a 1:250,000 scale. Surficial geology maps were produced at a 1:50,000 scale. A computer program is being developed for the search and retrieval of field data.The first phase of the Inventory of Aggregate Resources program is nearing completion and to date has collected information on approximately 9,500 samples throughout Newfoundland and Labrador. This data will be used to provide f or the effective management of these vital non-renewable aggregate resources.The mineral aggregate industry in Newfoundland and Labrador is very competitive, but undercapitalized when compared to other provinces. The demand for aggregates is forecasted to increase to in excess of23 million tons per annum by the year 2002.In terms of over-all mineral production, the extraction of sand, gravel and stone has a significant impact when compared to other minerals in terms of value and tonnage produced.IntroductionThe province of Newfoundland has approximately 18,000 km of coastline. A large proportion of the population lives in coastal communities where, traditionally, beaches have been a major source of construction materials. Beaches have, therefore, often been subjected to overutilization, resulting in erosion of back beach areas or adjacent headlands."

Jan 1, 1982

By Kazem Oraee

Implementing stable final pit walls in open pit mines during the open pit design phase and mine exploitation has become one of the most important concerns of mining engineers in recent years. Failure of mine walls can potentially cause loss of life, roads blocks, and damage to mining machinery, temporarily or permanently halt production at the face and in the worst case scenario, close the mine. Thus, the role of creating stable mine walls is crucial to safety and general operations of mines and any effort to optimize this process has great economic value for mine operators. This paper aims to illustrate the technical aspects of use of the successful presplit blasting method in metal open pit mines. It also shows that using this approach not only increases safety in mines through wall stabilization, but also will boost economic profits by savings in stripping ratio. In this regard, appropriate practical concepts for designing the drilling and blasting patterns and the role of selecting the correct drilling rigs are elaborated in detail through conducting a case study in Chador Malu iron ore mine in Iran.

Aug 1, 2013

By Paolo Farina

Slope monitoring radar has emerged in the last ten years as a leading edge tool for monitoring movements in natural and engineered slopes, thanks to its ability to rapidly measure slope movements with millimetric accuracy over wide areas in any weather conditions, obviating the need to install artificial reflectors. In recent years, slope monitoring radar has experienced significant improvements in its performances, mainly thanks to the use of different radar technology, such as Synthetic Aperture Radar (SAR), able to overcome some of the limitations of the first generation of radar and to provide higher spatial resolution, longer working distances and faster acquisition time. One of the leading applications of this technology is today represented by slope monitoring within the surface mining industry. Prominent international mining groups currently employ slope monitoring radar, based either on Real Aperture Radar (RAR) and SAR technology, for active and background monitoring of open-pit walls. The main technical features of typical slope monitoring radar based on SAR technology along with their possible applications are presented in this paper along with recent case studies of successful slope monitoring in open pit mines.

Jun 1, 2012

By Karen M. Moffitt

Developments in mining methods and rates have led to an increase in the frequency and size of waste dump failures that have not been adequately predicted or explained in terms of conventional slope stability analysis. Monitoring data and visual observations suggest a unique mode of failure responsible for large runout events. A numerical model has been developed to capture commonly observed patterns of deformation, and investigate the formation of the ensuing failure mechanism. The analyses have yielded a good correlation with observed field behaviour and have provided significant insight into the coincident stresses and deformations within the dump. Results indicate that while creep effects in the waste dump could have a significant effect on the magnitudes of displacement, the overall dump stability is governed by the strength of the foundation soils underlying the toe region. Consideration of both the stress and velocity fields leads to factors of safety for various dump heights and foundation slopes that are considerably less than predicted from limit equilibrium analysis. The results suggest that conventional limit equilibrium analysis techniques may not be adequate when assessing the stability and design of waste dumps on steeply sloping terrain.

May 1, 2001