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By M. S. Moats

As smelter feeds become more complex, the impurities that ultimately report to electrorefineries are changing. Many of these impurities collect in the electrolyte and must be controlled. As commercial copper electrorefineries look to expand their capacities by increasing their operating current density, the interaction between electrolyte composition and anode passivation can cause operational difficulties. To improve the industry's understanding of the passivation phenomena, the role of electrolyte composition was evaluated using commercial electrorefining anode samples. The electrolyte composition variables examined were copper, acid, nickel, arsenic, chloride, thiourea and glue. Finally, the role of non-commercially utilized additives, such as bromide and polyethylene glycol were studied. From this investigation, electrolyte parameters that inhibit copper dissolution through copper sulfate precipitation, copper oxide stabilization and/or absorbed species lead to accelerated passivation. Species that prevent passivation have the ability to complex cuprous ions. Possible mechanisms for the role of each.

Jan 1, 2007

By David Deglon, Martin Harris, Mehdi Safari

"This study investigates the effect of energy/power input on the flotation of a platinum ore in a pilot-scale oscillating grid flotation cell. Oscillating grids generate near ideal hydrodynamic environments, characterised by turbulence which is relatively homogeneous and isotropic. Secondary rougher feed and primary cleaner tail streams were floated in a pilot-scale oscillating grid flotation cell at power inputs from 0 to 2.5 W/kg, using 0.50 and 1.50 mm bubbles. From this study one may conclude that the effect of energy/power input on the flotation rate is strongly dependent on the particle and bubble size. Flotation rates generally increase with increasing particle size and decreasing bubble size. Increasing energy input generally leads to an increase in the flotation rate for finer particles and an optimum flotation rate for intermediate & coarse particles. Optimum conditions for intermediate/coarse platinum ore particles are using small bubbles at low energy inputs, or large bubbles at higher energy inputs. These results suggest that higher energy inputs are generally beneficial for the flotation of platinum ores as these consist predominantly of finer particles (-75 µm) and most industrial flotation applications operate with larger bubbles (1.0-2.0 mm).INTRODUCTIONThere is a considerable body of experimental and theoretical evidence to suggest that energy/power input plays an important role in flotation kinetics, particularly in the finer particle sizes where flotation efficiency is poor (Deglon, 2005). A number of excellent studies into the effect of energy on flotation kinetics have been carried out in impeller stirred cells (Ahmed and Jameson, 1985; Deglon, 2002; Pyke et al., 2003; Newell and Grano, 2006). Investigations using this type of cell have a number of limitations. The impeller influences particle suspension, bubble break-up and agitation in the cell. In addition, turbulence is highly inhomogeneous and anisotropic, with power intensities near the impeller orders of magnitude higher than those found elsewhere in the cell (Deglon, 1998; Koh and Schwarz, 2003; Schubert, 1999, 2008). These limitations have resulted in the development of novel cells for the investigation of energy input."

Jan 1, 2016

By Nicholas J. Welham

"The effect of extended ball milling has been studied for two industrial processes: (1) the processing of metal sulphide concentrates; and (2) the carbothermic reduction of ilmenite to synthetic rutile. The plant concentrates were a gold ore with 128 g/t Au of which 8% was non-refractory and a zinc concentrate with copper impurities.The gold concentrate showed complete dissolution of the gold bearing minerals after milling giving almost complete gold liberation. The zinc concentrate showed almost all of the sulphides had solubilized leaving the gangue and elemental sulphur. A study of pure pyrite and arsenopyrite shows that selective dissolution should be possible by a simple leaching process directly after milling.The milling of ilmenite and coal together results in a temperature decrease of 150°C in the subsequent carbothermic reduction of ilmenite, in addition to improving both the rate of reaction and the selectivity of reduction to rutile. IntroductionAlmost all mineral processing plants use mills for the size reduction of particles prior to physical and chemical processing. When continuing beyond size reduction the energy transferred causes deformation, breakage and realloying of particles giving a steady-state particle size. The breaking also causes increasing damage to the crystal structure by formation of smaller grains giving an increasing number of grain boundaries and increasing crystalline defects due to deformation. The increasing number of grain boundaries and defects leads to increased reactivity, which can be seen directly by a reaction within the mill, or indirectly, via a decrease in the subsequent annealing temperature required for reaction."

Jan 1, 1997

By D. C. R. Espinosa, P. Aliprandini, J. A. S. Tenório, M. M. Jiménez Correa

"Nickel and cobalt are the main value metals in nickel laterite ore. Cyanex 272 is an extractant used to separate nickel from cobalt present in leach solutions. Previous studies demonstrated the extraction of cobalt, magnesium and manganese at pH 5 using Cyanex 272 as extractant, while nickel remained in aqueous solution. However, it is difficult to separate cobalt from a solution containing magnesium and manganese. One possibility to change the metals’ behavior during the solvent extraction process is to change the mixture of extractants used. In order to study the solvent extraction of cobalt from a solution containing manganese, magnesium and nickel, a mixture of two extractants was investigated. Cyanex 272 (CY) and Versatic 10 (VC) were selected as extractants in the organic phase. The extractant concentrations used were: 20% CY; 15% CY, 5% VC; 10% CY, 10% VC; 5% CY, 15% VC; and 20% VC. In all cases, kerosene was used as a diluent, at 80 vol%. The synthetic leach solution used in this study was prepared according to the composition of a typical nickel laterite leach solution. The pH value of the solution used was 5, and the aqueous/organic ratio used in solvent extraction testwork was 1/1. The study has shown that the composition of the organic extractant mixture did not increase cobalt separation.INTRODUCTION Sulfide and laterite deposits are the main sources of nickel. However, production from nickel sulfide ores has been decreasing over the years, relatively speaking, increasing the focus on the exploitation of laterite nickel ores (McDonald & Whittington, 2008). Laterite ore deposits are found in tropical and subtropical regions. A characteristic of laterite ores is the presence of a complex and heterogeneous mixture of metals (Crundwell, Moats, Ramachandran, Robinson, & Davenport, 2011; Georgiou & Papangelakis, 1998; Gupta, 2003; Kesler & Simon, 2015; Moskalyk & Alfantazi, 2002; Oliveira, 1990). Nickel concentrations in laterite ores are typically 1–2% (Dalvi, Bacon, & Osborne, 2004)."

Jan 1, 2017

By C. B. Webber, W. A. Grell

"Sherritt Gordon's Ruttan Mine is a copper-zinc operation located approximately 600 miles north-west of Winnipeg. Open pit mining began in 1973 at 10,000 Tpd., with a change over to underground mining in 1981 producing 7,500 Tpd. Ore is crushed to -3/4"" followed by rod and ball mill grinding (two ball mills in series), differential copper-zinc flotation, and dewatering. The copper and zinc concentrates are trucked 75 miles to Lynn Lake, where they are shipped by rail to the HBM&S smelter at Flin Flon.This paper describes a method used to predict the effect of grind particle size on metallurgical performance, as well as details on the copper and zinc metallurgical performance due to fine grinding in actual plant operation.Initial statistical studies on historical data (including ore source, head gr3de, grind, etc.) indicated that copper and zinc recoveries increased with fineness of grind, in the range of 50 to 65% -200 mesh.In order to extrapolate beyond 65% -200 mesh, and to develop expected nonlinear relations between grind, Cu concentrate grade, and recovery, the assays of screen fractions of copper concentrate and tailing were used to determin8 the relation between net smelter return (a function of grade and recovery) and grinding cost. The optimum economic grind would be that grind at which the net smelter return less grinding cost reached a maximum."

Jan 1, 1982

By Douglas W. Reeve, Hoc Nghia Tran, David Barham

"In kraft pulp mills, the spent pulping liquor is concentrated and used as fuel in the recovery boiler. Extensive deposits are formed in the upper section of the boiler due to the very high ash content of the fuel and the volatite nature of the ash. Deposits are composed of Na2S04, Na2C03 NaCI, some potassium salts and reduced sulphur compounds. Chloride and potassium salts are enriched in the deposits compared to the smelt due to their greater volatility.Severe corrosion of the superheater tubes in a recovery boiler has been traced to three factors: a non-uniform temperature profile in the flue gas, relatively high operating steam temperatures and high chIoride and potassium concentrations in the fireside deposits.The role of chIoride and potassium in superheater corrosion has been explained by increased Iiquid phase in deposits. Together with corrosion studies, this information has shown that kraft recovery boiler superheater corrosion due to the chloride and potassium enriched deposits can be avoided by proper boiler operation and by judicious superheater and furnace design.IntroductionThis investigation of recovery boiler superheater corrosion was initiated by the discovery of severe corrosion of the final stage superheater in a Canadian closed cycle kraft pulp mill(1,2). Metal wastage leading to tube failure was believed to be due to three major causes : the relatively high operating steam temperatures, the uneven distribution of flue gas temperature across the boiler which created a "" hot spot"" and the high chloride and potassium concentrations in the superheater fire side deposit. The mill responded to the problem by i) lowering the superheater steam temperature from 475°C to 380°C, ii) minimizing the chloride content in the spent pulping liquor, by controlling the NaCI input with recovered bleach plant effluent, and iii) changing metallurgy of a portion of the bottom section of the final stage superheater platens from ferritic steel T-22 (2.25% Cr, 1% Mo) to austenitic steel Incoloy 800H (21% Cr, 32% Ni, 46% Fe, 1.5% Mn, 1% Si, 0.6% Al) (Fig. 1). Details of the superheater corrosion experienced in the closed cycle mil! have been reported in previous papers(l ·2,) ."

Jan 1, 1984

By Douglas A. Sloan

AN INCENTIVE or bonus system of payment is commonly used for mine labour. The incentive rates are usually based on the variable units of production, such as dollars per foot of drift advance or cents per ton of ore broken. Portions of the work that must be done to achieve the units of ?production are independent of the production unit and tend to be fixed work for the shift or for the job, irrespective of the production achieved. This fixed work has considerable influence on the equity of incentive rates. The effect of fixed work is to accelerate the loosening or tightening of rates if the production task is over-achieved or under-achieved, and of increasing the loosening of rates loosely set and the tightening of rates tightly set. This gearing of mine contract rates is the reason why changes in conditions, materials, equipment, or methods will change the incentive earnings to a greater degree than would normally be expected. Obviously, therefore, care should be taken in the setting of the task for incentive rates as the gearing makes an inequitable rate even more inequitable.

Jan 1, 1958

By Mohammad Mehdi Salarirad

Fouling of activated carbon by different organic species is a great concern in carbon in pulp and carbon in leach circuits. The rate of gold adsorption onto activated carbon from synthetic gold solutions which contained different amounts of potassium ethyl xanthate (PEX), potassium isobutyl xanthate (PIBX), and methyl isobutyl carbinol (MIBC) were evaluated through batch adsorption tests. The results of these experiments showed that the gold sorption rate rapidly decreases by increasing the concentration of collectors including PIBX and PEX up to 20 mg/L and at higher concentrations, the kinetic rate constant (i.e., k value) gradually leveled off to 53 h-1 for PIBX, which is 46% lower than the 99 h-1 for PEX. At low concentrations of up to 15 mg/L of MIBC, a gradual decrease of the rate constant was observed, while at higher concentrations, the decrease was found to be more pronounced. The gold loading capacity and the rate of gold adsorption onto activated carbon from a gold cyanidation pulp in the presence of flotation reagents were also evaluated. The results of these experiments showed that the flotation reagents have no significant effect on the gold loading capacity of activated carbon at prolong adsorption period, meanwhile the rate of gold adsorption continuously decreases by increasing the concentration of flotation reagents in the leaching solution.

Oct 1, 2011

By M. S. Greenfield, E. Cocchiarella, S. Chubbs

"At Dofasco sNo.4furnace a variation of the basic oxygen steelmaking process, Klockner Oxygen Bottom Maxhutte (K-OBM), is used to produce steel. Large amounts of hot gases and particulate are evolved during the oxygen blowing phase. A quencher and a venturi scrubber are used to collect over 99.5% of the particulate. In 1990, the authors investigated the effect of foamy slag on the particulate emissions from the furnace. The amount of lime added early in the oxygen blowing cycle was reduced, making the slag more viscous and foamy for a six minute period. Instantaneous particulate emissions were monitored throughout the blow by analyzing quencher and venturi scrubber water samples. Particulate emissions were reduced up to 59% during the foamy slag period.IntroductionThe reduction of the amount of waste generated or control of emissions at source is at the top of the ""3R"" hierarchy of waste management (Reduce, Reuse, Recycle). Process and operating practice changes are often the most economical ways to control emissions. At Dofasco, they investigated the effect of operating practice changes, specifically the formation of a foamy slag, on particulate emissions from steelmaking furnaces. They wanted to determine if a foamy slag changed the quantity and nature of the fume evolved from the K-OBM furnace over the oxygen blowing phase.The Basic Oxygen Steelmaking ProcessesThe basic oxygen steelmaking process (BOP) generates large quantities of very fine metal oxide fume during the production of steel. In this process, scrap steel and molten iron (hot metal) from blast furnaces are charged into a steelmaking furnace. Then pure oxygen is injected and a lime flux added . The oxygen reacts with carbon dissolved in the molten iron to form carbon monoxide. Oxygen lances can have secondary nozzles which introduce oxygen above the bath. This secondary oxygen, known as post-combustion, bums some of the carbon monoxide radiating the heat back to the bath. Particulate and gases rich in carbon monoxide are ejected from the furnace during the oxygen blowing phase and are collected by a gas cleaning system. Typically, more than 99.5010 of the particulate is scrubbed from the gases by high energy venturi scrubbers or electrostatic precipitators. The oxide sludge or dust collected may have elevated zinc and lead which is unsuitable for recycle to blast furnaces. Due to the presence of these undesirable elements, treatment and disposal of steelmaking oxide dust is difficult and costly."

Jan 1, 1993

By M. H. Moys, C. Bhondayi, Eric Tshibwabwa

"The relationship between froth bubble size as represented by intra-bubble impact distances (IID) and flotation performance is investigated. Experiments using an artificial ore comprising of limestone as the floatable component and silica as the gangue material were carried out in pseudo-steady-state circuit. A method to measure IIDs developed by Bhondayi and Moys (2014) was then used to measure froth bubble sizes as a function of froth height, superficial gas velocity and feed rate (pulp residence time). Samples for flotation performance evaluation were also taken. Results indicate that as froth height increases, the average IID values increase resulting in a decrease in limestone recovery while the grade increases. Increasing gas rate results in lower IID values that are accompanied by high recovery and low limestone grade. The effect of feed rate on both flotation performance and average IID values depended on froth depth and gas rate. Average IIDs largely remained the same at the two feed rates investigated.INTRODUCTIONIt is regarded as obvious that froth bubble size distribution (FBSD) will have an effect on plant grade and recovery. As the average FBSD decreases, the bubble surface area flux will increase making possible an increase in recovery and a decrease in grade. The opposite trend will be involved if average FBSD is increased. However: no method for measuring FBSD has been available, so these trends have not been studied in any detail. Bhondayi and Moys (2014) developed a novel technique to measure or estimate froth bubble sizes. Although they established a potential to use this method in understanding froth performance, no data have been presented to assess the correlation between the estimates of bubble sizes and actual flotation performance. In this work experiments are done to establish how the change in froth phase bubble sizes affect flotation performance. Flotation performance will be evaluated using recovery and grade. How water recovery to the concentrate changes with changes in bubble sizes will also be evaluated. It is known that water recovery is approximately inversely proportional to the square of the overflowing bubbles sizes. It will be interesting to establish how the rate of change of bubble sizes changes with water recovery."

Jan 1, 2016

By E. W. Thorton

"This paper is intended to show that grinding media can effect flotation results, and in some instances the results are vastly different. I do not suggest that autogenous grinding is beneficial in the separation of all complex base metal ores, in fact I know of a case where autogenous grinding would be disastrous to any selective flotation. I do however contend that prospective flow sheets and reagent balances of any complex base metal ores should be tried with both types of grinding. I feel fortunate that the paper to be presented later in this meeting by Rene Klymouski will, I think, compliment and perhaps explain some of the results of my research.This paper will be in four parts. The first part deals with a literature survey on the subject. Surprisingly enough, very little information has been printed on the subject. The second part deals with results obtained in bench scale tests on an ore from Chester Mines Limited. It was in these tests that we realized that steel grinding was giving us serious problems. So bench scale tests are shown giving results with a stainless steel grind, a mild steel grind and a grind in a ceramic mill with ceramic billets.The third part of the paper deals with results obtained in the Mines Branch Pilot Plant, when again we ground the ore with three different types of grinding media. Incidentally, this has been the only pilot plant test made at the Mines Branch in which flotation products were made on both steel grinding and autogenous grinding on the same ore. The fourth part deals with two other New Brunswick ores and is intended to show that changing the grinding media effects ores other than that of Chester Mines Limited."

Jan 1, 1973

By Don Langlois, Udo Weber

"The effect of grinding media performance on milling and operational behaviour was demonstrated under different selected milling conditions of calcium carbonate. A variety of grinding media material and bead sizes, along with 2 different stirrer tip speeds, were used in the grinding process to generate a particle size reduction of the calcium carbonate (CaCO3). To determine the optimum milling parameters the collected test data was used to calculate and evaluate specific energy as well as stress intensity under different milling conditions.INTRODUCTIONAs a result of increasing energy costs and raw material prices it is important to look at the efficiency of the manufacturing process. Wet grinding and ultrafine wet grinding in stirred media mills are now a possible cost-effective production step for the processing of industrial minerals and precious metals. The generation of smaller particle sizes (down to micrometer scale) can be advantageous for further production steps – i.e. better liberation of precious metals or improved industrial mineral qualities.In the following, the wet grinding process is investigated under different milling conditions of calcium carbonate. Different grinding media materials varying in specific weight, bead size and stirrer tip speeds are used to determine optimal energy consumption for the milling process.DESCRIPTION OF TEST CONDITIONS AND PARAMETERSMill ParametersTests were performed in a Drais Laboratory Perl Mill PML H/V with an installed electrical motor power of 3 kW. The mill chamber is silicon carbide where as the stirrer and discs are made of yttria stabilised zirconia. The mill working volume is 0.9 l. A screen cartridge with a screen size of 0.2 mm is used to separate the beads and slurry at the outlet of the mill. All tests were performed with a grinding media filling grade of 80% (Mill working volume). The stirrer tip speed was set at 6 m/s and 10 m/s."

Jan 1, 2009

By A. M. Elmahdy, T. Hirajima, M. Farahat

"Separation of molybdenite from chalcopyrite is mainly achieved through flotation technique to raise their grades up to 25% to be extracted in smelters. However, molybdenite recovery and grade obtained from this process is not satisfactory in most cases due to the natural hydrophobicity of both minerals. The influence of conventional heat treatment on pure chalcopyrite and molybdenite minerals was investigated. X-ray diffraction, TG-DTA and XPS techniques were applied before and after heat treatment. The results showed phase changes for both minerals. Magnetic separation experiments of single and mixture of both minerals at several magnetic intensities were made. For single mineral experiments, 100% molybdenite recovery was obtained at all experimental conditions as non-magnetic fraction whereas ca. 100% chalcopyrite recovery was achieved under certain conditions. In case of 1:1 binary mixture, 87% and 83% molybdenum and copper recoveries, respectively, were obtained. The increase in chalcopyrite magnetic recovery is related to the formation of more magnetic phases as confirmed by characterization techniques.The objective of this study is to increase the magnetism of chalcopyrite to make its separation from molybdenite by magnetic separation possible and to understand the mechanism behind it. These results indicate that it may be possible to upgrade molybdenite ores using heat treatment followed by magnetic separation to the required grade by smelters which is difficult to attain by froth flotation and the other techniques."

Jan 1, 2016

By X. -G. Chen, K. Liu, N. Parson

"Precipitation-strengthening 6xxx alloys are widely used for building and construction and automobile industries due to their high strength, good weldability and corrosion resistance. However, their elevated-temperature properties (250–300°C) decrease rapidly due to the fast-coarsening of precipitates. On the other hand, dispersoid-hardening is one of the most efficient approaches to improve the elevated-temperature properties. Therefore, the evolution of dispersoids during heat treatment in Al-Mg-Si-Mn alloys was investigated in the present work. Various heat treatments with temperatures ranging from 300–500°Cfor up to 24 hours were performed. Electrical conductivity and microhardness were measured while optical and scanning electron microscopy combined with image analysis, were used to characterize the dispersoids. Special emphasis was placed on the evolution of the size and number density of the dispersoids under different heat treatment parameters. It was found that dispersoids started to precipitate at 350°Cfor 1 hour. Nucleation and growth were the main mechanisms at temperatures up to 400°Cwhile coarsening occurred at temperatures above 450°C. The quantitative results indicated that dispersoids precipitation reached the optimum condition after treatment at 400°Cfor 2–4 hours with a fine size and high number density, resulting in the highest hardness.INTRODUCTIONThe precipitation-hardening 6xxx aluminum alloys with major alloying elements of Mg and Si are one of the most widely used aluminum alloys for building and construction and automobile industries (Dinaharan & Murugan, 2012). In recent years, there have been increasing demands for aluminum structural components to be used at elevated temperatures (250–300°C) for different applications. However, the mechanical properties of 6xxx alloys at elevated temperature are typically reduced due to the rapid coarsening of precipitates at 250–300°C. It has been reported that the yield strength (YS) of AA6061 in the peak-aged condition drops from 276 MPa at room temperature to only 30 MPa at 315°C(Kaufman, 1999). Therefore, it is a key concern to improve the elevated-temperature properties of 6xxx alloys to expand their range of applications."

Jan 1, 2018

By P. Au, P. Patnaik, H. Saari, D. Y. Seo, S. Bulmer

Ti-48Al-2Cr-2Nb-xW (where x = 0, 0.5, and 1 at. %) powders were consolidated by hot isostatic pressing (HIP). The consolidated samples were solution heat treated above the a-transus followed by stepped cooling to room temperature. Selected samples were subsequently aged at 950 °C for up to 72 hours. The post-heat treatment microstructure was examined by optical and electron microscopy. It was found that increasing holding time above the a-transus generated a more homogenous microstructure. Utilizing the stepped cooling sequence, especially with a transition temperature of 1270 °C, massively transformed gamma (Ym) was diminished. Aging caused coarsening of the fully lamellar microstructure in the 0%W and 0.5%W samples while almost no coarsening occurred in the 1 % W sample. Precipitates began to form during aging in samples containing W. The precipitates at the lamellar interfaces are fine and have a needle like shape, while precipitates at the lamellar grain boundaries are coarse and circular in shape. The size and number of the precipitates depend on aging time and the level ofW content.

Jan 1, 2006

By M. Pelser, H. Gous, J. H. W. M. Brits, K. Voogt

The effect of impurities; iron and selenium, on current efficiency and deposit quality in copper electrowinning was investigated using an at-line cell at the Rustenburg Base Metals Refiners’ Copper Tankhouse. The tankhouse is a direct leach, starter sheet electrowinning operation with a unique electrolyte composition. The effect of iron on the current efficiency in the at-line cell was found to be less severe than reported in literature. However, the distinction between the chemical impact of iron on current efficiency and the physical impact of iron on a starter sheet operation needs to be realised. The effect of selenium on current efficiency and deposit quality is also reported. INTRODUCTION Anglo American Platinum’s Rustenburg Base Metals Refiners’ (RBMR) Copper Tankhouse serves a unique function in the refinery flowsheet in that its primary functions are the removal of copper and production of acid for the leach circuit. This function needs to be performed to specification in order to keep the leach circuit operating optimally. Current efficiency is seen as an indicator of tankhouse performance and efficient energy use. The tankhouse plates starter sheets onto titanium blanks which, after stripping and preparation, return into the production cells for further plating. Approximately 78% of tankhouses surveyed (Robinson et al., 2013) use permanent cathode technology whereas older, smaller operations still use starter sheet technology. Drawbacks of the starter sheet technology are that the starters are thin and bend easily leading to increased shorts, they need to be removed after two days of plating to be manually straightened and are then returned to the cells. In addition, the electrolyte levels in the cells need to be adjusted during the plating cycle to improve loop adhesion to the cathode and reduce loop corrosion. The starter sheet technology is therefore very labour intensive and the final quality of the product is dependent on the quality of the starter sheets produced. In comparison, permanent cathode or direct plating technology lends itself to automation, allows for higher current densities and better cathode quality due to more robust geometric stability in the cell. The present barrier to implementation of permanent cathode plating at the RBMR Copper Tankhouse is the capital cost required to convert the tankhouse, i.e. blanks and stripping machine. At present, alternatives to titanium blanks, such as LDX 2101, are being investigated for the unique electrolyte composition of this tankhouse.

Jan 1, 2019

By I. Cifuentes

Chemical and physical characterization of copper electrodeposits obtained on mesh and plate cathodes in conventional and reactive electrodialysis lab-scale cells have been carried out. The effect of Fe, Mn, Cl and guar on electrodeposit composition and morphology were determined by means of metallographies, scanning electron microscopy (SEM), atomic force microscopy (AFM), inductively coupled plasma mass spectrometry (ICP-MS) and Auger electron spectroscopy (AES). Each impurity influenced the deposit morphology in a different way (e. g. grain size, surface roughness). AFM proved to be a very effective tool for quantifiying and classifiying surface quality.

Jan 1, 2007

By P. Stantke

Glue, thiourea and chloride are the common reagents which are used in copper electrorefining to achieve fine-grained and smooth copper deposition. While thiourea and chloride are clearly defined chemical substances, glue is a natural product consisting of a mixture of amino acids in different constituent parts. The amino acids are continually degraded in the electrolyte by hydrolysis until they finally become completely inactive. This effect is well-known, largely constant and balanced by the continuous dosing of glue. In contrast, the behaviour of glue in the batch container is not always foreseeable. The glue can become partly or completely inactive by contamination with acid or bacteria, for instance. Usually this is not noticed, and partly or fully inactive glue is added to the electrolyte over a longish period, with the result that the copper growth is significantly impaired. Examples from operational practice show that, with the help of the CollaMat system, partly or fully inactive glue batches can be quickly detected.

Jan 1, 2007

By Walter Nelson

Considerable thought and discussion had been devoted toward provision for increased milling capacity at the Allenby concentrator of The Granby Consolidated Mining, Smelting and Power Company operations as early as 1940. During the war, plans for the installation of an additional grindi.ng unit were formulated but were not carried out. It took considerable time after the war to prepare the mine again for its highest rate of production since, with the scarcity of men, development work had been sacrificed in favour of maximum ore shipments. In order to prepare for the expected competitive market in copper, the matter of increasing the capacity of the concentrator was again given serious attention. As ore became available in the mine and the output of the mill was increased, limitations for a satisfactory recovery were observed, due to the overloaded condition of the classifier units at the higher tonnages. From a metallurgical standpoint, the ore from the Copper Mountain mine had always been difficult to treat on account of the finely disseminated character of much of the mineralization and the difficulty of grinding to the desired fineness. This trouble has become aggravated with time as an increasing amount of wall-rock has diluted the ore and lower grade orebodies have been mined. It therefore has been, and is, the practice at the concentrator to maintain the maximum amount of reduction in size of material fed to the flotation units as is possible for the daily tonnage to be milled. This means that all the grinding units are operated for the maximum possible time, and that the classifiers are regulated to return as much feed ?as possible to the mills. The number of tons to be milled for ?the day is determined by the concentrator superintendent after ascertaining the availability of ore from the mine, the condition of the concentrator ore bins, and the need for repairs in the rolls plant. In order to maintain the tonnage des1red, hourly estimates are made of the rate of grinding by weighing the amount of ore contained on one foot of each of the belts conveying the feed to the various ball-mills, and adjustments are made accordingly.

Jan 1, 1950

By Herbert D. Drechsler, James B. Stephenson

"AbstractThe paper identifies the major components of price change which affect the determination of the internal rate of return of a mining project. It illustrates several price indexes which may be used to measure general price increases, identifying techniques which might incorporate them into mine valuation procedures. The ""hurdle rate"" for new investment is discussed.IntroductionDECISIONS regarding the value of an orebody and whether or when to mine an orebody are based on an estimate of the cash flows which would be associated with its exploitation. These pro-forma cash flows represent cash in or out of a possible mineable deposit and include exploration and construction costs, sales revenues, operating expenses, taxes, debt borrowings, interest and debt repayment. It has long been recognized that when combining the cash flows occurring over the life of the mine, dollars received or paid out early have greater value than those received or paid out later. This time value of money has been incorporated into mine evaluation calculations through the use of present-value or discounting techniques. Another factor affecting mine valuation, inflation, is not accounted for by valuation techniques now in use. In recent years, price changes have become more volatile. In the past, price change could easily be ignored when evaluating potential mines. Today, the combination of larger investments and greater price volatility leads to great concern about the effects of price change on mine valuation."

Jan 1, 1977