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By M. Urda-Kiel

Stainless steel is increasingly used as a cathode starting sheet. It was shown earlier that even when depositing copper on copper, a substrate-influenced zone appears in the deposit structure, followed by a transition zone and finally by the electrolysis-influenced structure. On copper, the substrate-influenced zone has a thickness of twenty micrometers and on titanium, eventually more than fifty micrometers. Initial nucleation and growth have a considerable influence. In this paper, copper was deposited in a channel cell on four different 304 stainless steel substrates: mechanically polished and anodized, respectively, at 0, +400 and +800 mV/SHE in 1M H2SO4 at 20°C. Copper was deposited under electrowinning conditions after solvent extraction (SX) (40 g/l Cu++, 180 gh/l, H2SO4, 20 or 60°C). The results showed that, although Auger surface analyses of the substrates were only slightly different one from another including the oxide thickness, nucleation and initial stages of growth were very sensitive to the surface treatment of the substrate. Potential step and current step methods gave similar results : progressive nucleation followed by 3D diffusion controlled growth. The highest nucleation frequency was observed on stainless steel anodized at 0 mV/SHE, under various electrolysis conditions. The potential influence of these findings on industrial practice is discussed.

Jan 1, 1999

By Chuang Y-Y

A summary is made on the thermodynamic analysis for two of the four ternary systems of the quaternary Cu-Fe-Ni-S. The analysis is being extended to ternary Cu-Fe-S and then to the quaternary. An analysis is already made for the metal ternary but is not reported here. Experiment is being carried out to measure the liquidus surface. of Fe-Ni-S equilbrium with the fcc solid solutions of (Fe, Ni). The presentation at the Symposium will include all the results obtained at the time of presentation.

Jan 1, 1983

By L. M. . Chia

The successful treatment of waste materials to recover metals with acceptable purity is often hampered by the presence of undesirable impurities. The problem is further complicated by the need to insure that any residues or solutions generated are environmentally acceptable. Hydrometallurgical processes involving leaching- solvent extraction-electrowinning offer some attractive features which address many of these problems. However, in certain instances trace impurities that contaminate the organics are difficult to chemically strip and interfere with the overall process efficiency. A new processing scheme has been developed which appears to provide some attractive alternatives for impurity removal from organics. It has been shown that many ionic impurities can be electrochemically removed from the organic by spontaneous direct reduction using solid metals. A variety of metal ion impurities have been removed from selected common organic extractants such as D2EHPA and TBP using different metals as the reducing agents. The results and probable mechanisms involved in the reactions are discussed.

Jan 1, 1994

By Kezhuan Gu

Data on the kinetics and thermodynamics of dephosphorization are analyzed in terms of the dynamic partition ratio, determined by a combination of slag thermodynamic properties and the balance of oxygen supply and consumption. A recently developed phosphate capacity correlation is combined with data on kinetics to quantify the most significant parameters in controlling the interfacial oxygen potential. The role of slag composition, entrained gas fraction and temperature are discussed in addition to metal carbon and sulfur content. It is demonstrated that the mass transport of oxygen in the slag is enhanced by factors increasing conductivity and inhibited by entrained gas. The mass transfer of phosphorus in the metal is described considering surface renewal induced by CO bubbles.

By V. Ramachandran

Over the past 20 years, water treatment plants have been built at various Asarco facilities to effectively handle the process waste streams that need to be bled from the system. The current paper review the several operations with reference to (a) waste streams processed, (b) process chemistry, (c) discharge options and recovery of treated water, and (d) sludge disposal.

Jan 1, 1994

By KNona C. Liddell

"Many metal recovery processes involve heterogeneous reactions in porous media accompanied by fluid flow and mass transfer. Modeling such systems is difficult because of the number of rate processes involved and is complicated even further by effects due to the structure of the porous medium. Data on even bulk properties of the medium are often severely limited or unavailable. For many such systems, a cell model may provide an attractive alternative to models based on transport mechanisms in a differential volume. A cell model requires less input data and is computationally efficient. Identical residence time distributions can be obtained with models of both types, and both allow conversion and concentration profiles and histories to be calculated. The formulation and properties of cell models for metals recovery are described in detail.INTRODUCTIONSubsurface mineral leaching, metal ion recovery and environmental remediation are significant modeling challenges. They involve one-, two-, or three-dimensional reactive flow in multi reaction open systems. Dissolution, adsorption, and ion exchange occur at reactive mineral surfaces, and these heterogeneous reactions are accompanied by acid-base, complexation, and redox reactions in the solution phase.Two approaches have been taken to reactive flow problems in geologic media: mixing cell and transport models. Mixing cell models, the subject of this review, are based on a network of connected cells, each with a specified number of input and output streams. Mixing cells are assumed to have no internal solute concentration gradients, that is, each cell behaves like a well-stirred tank. Gradients are represented by step changes between connected cells; this results in a model that is spatially discretized at the outset. Different methods of discretizing the model equations are considered more fully below. Although this is not strictly necessary, it has usually been assumed in subsurface flow applications that all cells in the network have the same volume and identical initial properties."

Jan 1, 1996

By J. E. Dutrizac

The behaviour of the lanthanide elements during the precipitation of sodium and potassium jarosites at 98oC from sulphate media was systematically investigated. The lower atomic weight lanthanide elements (La to Gd) are characterized by the limited solubility of their alkali-lanthanide sulphates, such as NaNd(SO4)2?H2O, which commonly contaminate the jarosite precipitates. Hence, single phase jarosite-type compounds were precipitated only from solutions containing <0.4 g/L of the La to Sm lanthanides. The sulphates of the heavier atomic weight lanthanide elements (Tb to Lu) have significantly higher solubilities such that solutions containing up to 10 g/L of the individual lanthanides could be investigated without the precipitation of alkali-lanthanide sulphate species. Even at the highest lanthanide concentrations, however, none of the jarosite precipitates incorporated more than 0.3 wt% of any of the lanthanide elements. Comparable levels of lanthanide incorporation were noted in both Na-jarosite and K-jarosite. Although the mass of the jarosite precipitate increased with increasing ferric sulphate concentration or with increasing pH, the extent of lanthanide incorporation was not significantly affected. Clearly, the trivalent lanthanide elements do not substitute extensively for trivalent iron in the jarosite structure. This behaviour likely reflects the significantly lower hydrolyzability and larger size of the trivalent lanthanide ions relative to ferric ions.

Jan 1, 2003

By V. M. Sokolov, Jr. DuPlessis

The technology for recycling fine wastes, grinding sludges or swarf, and solids such as slag that contain a high content of metallic material needed improving. The objective was to produce a master alloy that met the tight requirements of the Western manufacturers and achieve improved yields of the valuable elements, nickel and cobalt, to over 90%. The alloys from the production of Alnico. magnets were chosen for this work. A process is proposed .that is able to recycle all kinds of Alnico wastes. This process is able to reduce all detrimental elements, which are contained in the wastes, to within acceptable levels. Molybdenum, if present, is an exception. Depending on the levels of contamination of undesirable elements different recycling strategies are proposed. The majority of the wastes can be recycled by melting and refining with a liquid slag layer using unique melting furnaces that are relatively unknown in Western nations. The proposed technologies allow the reelamation of over 95% of the cobalt and nickel from any kind of Alnico waste, and there are no limitations on the proportion of different kinds of wastes in the charge. The proposed processes and technologies could also be applied to the recycling of grindings, turnin

Jan 1, 2000

By Wayne Bell, Shivanand A. Borkar, Anish Upadhyaya, Matthew D. Andriese, Jiann-Yang Hwang, Zhiwei Peng

"A refractory ore body located in Michigan’s Upper Peninsula contains high concentrations of nickel and copper chiefly occurring in the minerals pyrrhotite, chalcopyrite, and pentlandite. Refractory ore bodies are difficult to treat by conventional mineral processing methods so microwave pre-treatment of the ore is employed to increase metallic-particle/host rock liberation by making use the differential thermal expansion properties of the mineral phases that absorb microwave energy. The Curie temperature measurement of metallic-bearing particles is in agreement with the known value for pyrrhotite occurring at 325°C. A nickel-rich iron sulfide mineral is found to be in occurrence and also appears to be magnetic under BSE imaging. It is shown the ore particles heat rapidly when exposed to microwave radiation for short durations of time mainly due to the high concentration of ferromagnetic mineral phases. Rapid heating causes thermal expansion of constituent mineral phases that produce cracks within ore particles. SEM imaging shows fracture occurring along grain boundaries and throughout host rock matrix. Ball milling experiments show an increased grindability of the ore resulting in a decrease in work index values.IntroductionCrushing and grinding operations consume 50-70% amount of energy used in mineral processing operations [1]. All these processes involve high energy impact of surfaces to develop compressive stresses that produce cracks to initiate breakage of particles. Microwave pre-treatment of certain types of ore might provide an energy savings to these type grinding processes by taking advantage of the tensional forces produced by selectively heating microwave absorbing type mineral phases. The heating of ore particles produces cracking by thermal expansion of the constituent mineral phases. Conventional heating has been investigated and shown to positively aid in grinding processes but determined energy inefficient to be used commercially."

Jan 1, 2011

By S. Kang, D. G. C. Robertson

Slag cleaning furnaces require extensive wall cooling because they involve the use of a superheated slag bath. This is in direct contrast to smelting furnaces, where the wall is usually (but not always) protected by "banks" of solid charge. In order to understand the heat transfer and flow in slag cleaning furnaces a number of model studies have been carried out using molten wax to simulate the slag. Wax is a good model fluid because, like slag, it has a high Prandtl number. Wax models cannot simulate the behavior close to the electrodes (DC arc or AC submerged), but they can and do give a good simulation of the flow and heat transfer at the wall. An understanding of these phenomena is of great importance for the design of cooling panels in the walls of the furnace. Since superheated molten slag will eventually wear away any refractory bricks over time, the wall cooling must be designed to form a frozen layer of slag, referred to as a "freeze lining". The heat flux density (Q/A) required to do this is given by Q/A = h* ?T, where h is the heat transfer coefficient in the slag phase at the wall, and Li T is the superheat (bulk slag temperature - liquidus temperature). Both Q/A and h vary greatly with position and both are usually a maximum at the slag line. In the experimental work, molten wax baths were heated by block heaters to provide similar energy densities to those encountered in commercial slag cleaning furnaces. This procedure was considered to be valid since most of the electrical energy into a furnace bath is dissipated as heat. Gas stirring was used to simulate the gas evolved at the electrode in real furnaces - the quantity of this gas could be estimated from a knowledge of electrode consumption. The similarity criterion was equal gas flow per unit area of surface, in model and prototype. Heat transfer coefficients were measured in the models as a function of power density, gas stirring, superheat, etc., and could be related to the full-scale operations by the use of dimensional analysis.

Jan 1, 1999

By J. Brent Hiskey

The global copper industry faces a technological challenge to develop processing schemes for more complex and impure ores and concentrates. Group 15 elements (As, Sb, and Bi) represent impurities that are troublesome both in terms of environmental impact and metallurgical processing. The so-call "dirty" copper concentrates are either unacceptable for smelting and refining or incur high penalties. The behavior of the Group 15 elements is of critical importance in maintaining product quality and process performance. During copper electrorefining, these impurities are released to the electrolyte by electrochemical dissolution of the anode. Depending mainly on temperature, concentration, and oxidation state these elements can spontaneously coprecipitate from the electrolyte. Some of these precipitates are referred to as floating slimes and their formation can have extremely detrimental effects on cathode purity. In this paper, the mechanism of floating slimes formation will be discussed and the fundamental thermochemistry will be reviewed and analyzed in relation to solubility properties.

Jan 1, 2012

By Glenn C. Miller

Cyanide is the most common lixiviant for extraction of precious metals, and its efficient use in open pit mines has contributed to the increase in world production of gold during the previous two decades. With this expansion of precious metals mining using cyanide, public concerns have developed in response to the inherent public perception of cyanide as a tox-in as well as a series of environmental problems that are directly or indirectly related to cyanide use in open pit mining. Some of these have technical or regulatory solutions, while others will require additional investigation before they can be resolved. The following discussion will summarize some of the public concerns regarding the use of cyanide in precious metals mining. The issue of cyanide use in large scale precious mining is controversial, and recent banning of cyanide use in open pit mines in Montana, or complete banning of cyanide in Hungary has reflected those international concerns. The underlying basis for some of the bans (e.g. Montana) reflect a general concern of the environmental impacts of mining, rather than specific issues related to risk of cyanide use. The human safety record for cyanide use, particularly in the developed countries is good, and cyanide is clearly the most efficient lixiviant for precious metals. Indeed, the relatively short lifetime for cyanide in the environment renders it one of the less problematic chemicals used for extraction of gold from ore. Mercury amalgamation or use of bromide as a lixiviant are certainly much more problematic. What, then, is the basis f~r the international concern about precious metals mines that use cyanide. Ifuse of cyanide for gold recovery is to continue, what additional infom1ation needs to be developed that will allow a better management of the risks of cyanide? Which of the concerns are directly related to cyanide, and which are indirect concerns that more related to effects of open pit mining. Under what conditions should use of cyanide be discouraged? DUl::ing June of 2000 the United Nations Environmental Program conducted a two-day workshop on cyanide use in mining that was attended by persons from the mining industry, regulators and environmental community. Support and attendance at this workshop reflected concerns by all three groups on further restrictions for cyanide use and the emerging potential for guidelines for better management of the catastrophic release of cyanide-containing mining waste. The following discussion is meant to focus-on those risks and explore the basis for an often negative reaction to precious metals mining using cyanide. It is presented in the format of a statement, followed by a discussion. While admittedly provocative, the statements are meant to express the concerns? of the public regarding cyanide and the mining industry.

Jan 1, 1998

By John P. Hager

One possible approach for the recovery of mercury from acid plant blow-down sludge in smelters is the use of a vacuum retort reactor for the vaporization of the mercury with subsequent collection of metallic mercury in an attached condenser. An experimental study was undertaken to investigate the vaporization of mercury when present as a compound in the Hg-S-O and Hg-Se systems. A transpiration reactor was modified to operate under reduced pressure and with a continuous flow of inert sweep gas. A water-cooled cold finger was inserted into the exit end of the reactor to provide for the rapid quenching of the exit gases and the collection of the mercury. Rates of vaporization were measured at several temperatures for each compound and were generally found to be constant at a given temperature and reactor pressure. The results were used to determine the minimum operating temperature for 100% vaporization of the compound in 60 minutes. For Hg2SO4, HgSO4, and HgSO/2HgO the required temperatures. for a reactor pressure of 50 Torr. were determined to be 634°, 622°. and 603°C. respectively. For HgO, HgSe, and HgS the required temperatures were 557°. 451°. And 382°C. respectively. For HgS the rate of vaporization was also measured at a reactor pressure of 1 00 Torr. The rate of vaporization was found to be slower resulting in an increased vaporization temperature of 423°C. It was also determined that the mercury vapor present in the exit gas stream can back-react with the other gas species to a significant extent and can result in very little condensed metallic mercury. For the vaporization of both HgS and HgSe. the mercury vapor was found to have back-reacted to the extent that virtually none of the vaporized mercury was collected in metallic form. For the vaporization of Hg2S04? HgS04, and HgSO4*2HgO the amount of metallic mercury recovered varied from approximately 50% for HgS04*2HgO to approximately 10% for Hg2S04 and HgS04. Only for HgO was the recovered mercury in primarily elemental form. The application of the results to the operation of a vacuum retort and condenser system is discussed.

Jan 1, 2001

By Tam Tran

Most new gold mines in Australia are required to treat tailings to reduce cyanide levels before they are discharged to storage ponds to meet the stringent environmental management guidelines set by government statutory authorities. In New South Wales, the average level of weak acid dissociable cyanide allowed in open tailings dams is now limited to 20 ppm, As stringent environmental guidelines continue to be enforced in Australia as well as in other parts of the world, the management of cyanide wastes has emerged as a key issue for the gold mining industry. This paper provides a critical evaluation on technical aspects related to the use of ion exchange resins and affiliated processes in cyanide management and tailing treatment. Particular attention is given to the use of ion exchange resin technologies in the processing of mixed gold-copper ores, where cyanide usage is the highest. Information presented in this paper is heavily drawn from the experience on the application of ion exchange resin processes at May Day mine, in New South Wales, Australia and associated pilot plant test programs. Issues related to the management of thiocyanate, especially with respect to the recovery of cyanide from thiocyanate are also discussed.

Jan 1, 1998

By Chu Yong Cheng

The solvent extraction (SX) processes of the three pressure acid leach (PAL) nickel plants in Western Australia have been reviewed. The use of intermediate precipitation, solids/liquid separation and re-leach in the three W A nickel plants make these processes complicated and costly in capital and ?operation. The research work carried out by the SX group at the AJ Parker Cooperative Research Centre for Hydrometallurgy I CSIRO Minerals has led to the development of direct solvent extraction (DSX) processes to recover and purify nickel and cobalt from leach solutions. The simplicity of the process flowsheets and the expected savings in capital and operating costs are the major advantages of the new DSX processes over the reviewed processes. By using a new synergistic organic system in semi-continuous tests with a pilot plant leach solution (after iron precipitation) from BHP-BiIliton Stainless Steel Materials, the metal values (Ni and Co) together with zinc and copper were separated from the major impurities (Mn, Mg and Ca, together with Cl) in the first SX circuit. The co-extracted manganese, magnesium and&apos; calcium were easily scrubbed out. After stripping, the metal values (Ni and Co) together with zinc and copper were concentrated, resulting in a much smaller second SX circuit and equipment in the down stream processes. The extraction and stripping kinetics of the metals with the new synergistic organic solution were very fast. The extraction and stripping almost reached steady state within 0.5 minutes. Semi-continuous test work with a synthetic leach solution, a cobalt pilot plant leach solution from Peko Rehabilitation Project Pty Ltd and a synthetic solution to simulate a concentrated laterite leach solution from BHP-Billiton Stainless Steel Materials showed that manganese, calcium, copper and zinc can be effectively and efficiently separated from nickel, cobalt and magnesium by extraction, scrubbing and stripping. This led to the development of another DSX process using D2EHPA to recover nickel and cobalt from leach solutions.

Jan 1, 2003

By C. Dresen

This paper discussed the reaction kinetics of the reduction of industrial phosphate pellets for P4 production. In the reduction process of various phosphate ores, the following parameters were investigated: Diffusion of the phosphorus oxide and other gaseous products from the reaction mixture after having been reduced on the reductant (coke) surface. Influence of silica on reaction kinetics. Temperature influence. Influence of mineralogy and morphology (different ores). Particle size influence of silica and coke. Experiments were carried out in a vertical tube furnace with 2 ores, 2 types of silica and one type of coke. Selected experimental conditions simulated furnace conditions. The following conclusions could be made: Reaction rate increases with the addition of SiO2 into the pellet. The reaction rate increases above 1350°C (melting temperature of the slag). Mixing different ore types in the pellets leads to differences in reduction kinetics. Particle size of the coke particles influence reduction kinetics. The amount of slag formation is influenced by silica addition, coke size and temperature.

Jan 1, 2002

By B. I. Hop

The f10wfield set up by the standard Hydro gas purging unit has been predicted by a numerical model. In addition turbulent trajectories of gas bubbles released from the rotor are predicted. Here the deposition rate of inclusions on the bubble surface is allowed to depend on local conditions. Bubbles passing regions with strong pressure gradients, such as close to the rotor, will experience large slip velocities and large deposition rates. This model can deal with any distribution of bubble sizes and inclusions. The results found from the multidimensional flow predictions are discussed in relation to what can be obtained from a much simpler metallurgical engineering approach. The differences between these two approaches to predict the same process will be discussed. The model predictions are compared to experimental results for removal of SiC-particles from aluminium.

Jan 1, 1996

By N. J. Tannyan

Laboratory bath smelting tests were conducted to determine the feasibility of smelting nickel- copper concentrate directly to a low iron matte in one step. Concentrate, air/oxygen and a silica flux were fed to a matte-slag bath kept in MgO crucibles at I 300°C. Various oxygen enrichments and smelting rates were used to maintain 0.5 to 2 wt% iron in the product matte. The slag losses are comparable to those obtained under similar industrial conditions. Due to the highly oxidizing conditions required to produce mattes with low iron contents, it is not recommended to target iron contents below about 2% in the matte. The specific bath smelting rates were equal to or higher than those in various industrial bath smelting processes.

Jan 1, 2003

By Jae-Ho. Lee

Mercury is a toxic contaminant encountered in the extraction of gold from mercury-containing ores. It is found in the leach solution of CIP circuits and in the cathodes of the electrolyte cells used to electrodeposit the gold. The research conducted has examined the potential for selective removal of mercury from synthetic Au-Hg-CN electrolytes containing only gold, only mercury, and both gold and mercury. Cyclic voltammograms were obtained for mercury- coated copper electrodes with these cyanide electrolytes. Exchange current densities were determined from the voltammograms in the low overpotential region. Laboratory scale batch- electrolysis at constant current was also conducted for the electrolytes containing only mercury and both mercury and gold, at 20 and 50°C and from which the kinetic behavior associated with the process was analyzed. The surface of cathode before and after electrolysis was also characterized and a copper mercury intermetallic (CyHg,) was identified. Gold codeposited with mercury during the later stages of electrolysis. The results of this work corroborate that the mercury-amalgam copper electrode offers a viable electrochemical technique for the selective removal of mercury, however a high-surface-area cathode-cell would have to be utilized to provide for the desired selectivity.

Jan 1, 1994

By Qixing Yang, Björn Haase, Fredrik Engström, Anita Wedholm

"MiMeR, the Minerals and Metals Recycling Research Centre at Luleå University of Technology, Sweden, started industry-related research work to stabilize an EAF slag from Höganäs AB with of CaO/SiO2 values near to 3.2. The intention was to determine the suitability of the EAF slag as construction material after stabilization. Some of the stabilizers effective for preventing disintegration of AOD slag were examined in laboratory tests for stabilization of the EAF slag. The EAF slag samples were mixed together with stabilizers for melting at 1600ºC and then cooling in a crucible system. With a laboratory granulation system, the liquid EAF slag was cooled rapidly using air. Samples from the tests were characterized by different techniques, including XRD and SEM. Some of the results from the work may be useful for preventing slag disintegration and, thus, enhancing utilization of the EAF slag as construction material, thereby saving valuable natural resources.IntroductionAt steel plants of some MiMeR member companies, steel grades with low carbon contents were produced in Electric Arc Furnaces (EAF). EAF slag was tapped out of the furnace to slag pots for transport to slag yards, where the slag was poured on the ground. Very often, the EAF slag disintegrated during cooling on the ground. Fines from the slag disintegration were often difficult to handle and use as material for road construction [1, 2]. The EAF slag disintegration was rather similar to the disintegration occurring in the slag tapped from some AOD (Argon Oxygen Decarburization) converter heats.Literature studies show that formation of dicalcium silicate (C2S) is the main reason for the disintegration or dusting of AOD slag from production of stainless steel [1-4]. The C2S can form while the content ratio of CaO to SiO2 of the slag, CaO/SiO2, exceeds 1.4 [1]. Values of CaO/SiO2 for the AOD slag are often near to 2.0, making the C2S a predominant slag component. The ß phase of C2S, ß-C2S, transforms into its low temperature ? form in the temperature range of 400-500°C, inducing a volume expansion of nearly 12% [1-4]. The volume expansion disintegrates the AOD slag."

Jan 1, 2008