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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 Zhou Aiping, Lv Guozhi, Niu Liping, Wang Wenbo, Zhang Ting’an, Jiang Xiaoli

The current method used to produce sponge titanium utilizes a process cycle of magnesium metal and chlorine gas via the electrolysis of magnesium chloride. The primary issues are the high capital cost of the electrolysis process and the environmental regulations limiting high chloride concentrations. Thus, the authors have developed a conceptual method involving theelectrolysis of magnesium chloride as well as a reduction process which allows for a cycle of magnesium and chloride. In the new method, MgCl2 which was distilled during the titanium extracting process as raw material was directly pyrolyzed at elevated temperatures yielding high purity magnesium oxide and chlorine. The resulting magnesium oxide can be used in the production of magnesium metal. This work examines and compares the energy consumption and economic benefits of the Kroll method as well as the new proposed method. The results from the study demonstrated that in comparison to the traditional Kroll method, the new method provided considerable economic and environmental benefits.

Jan 1, 2014

By T. T. Chen, J. E. Dutrizac

Noranda Inc., CEZinc presently employs a cement-stabilization (jarofix) process to stabilize its jarosite residues for direct disposal. In this process, Portland cement partly decomposes Na-jarosite to form Ca6Fe2(SC>4)3(OH)i2.nH20, ferric hydroxide and sodium sulphate which further reacts to form gypsum; approximately one-third of the Na-jarosite reacts in the current commercial product. The cured jarofix products also contain minor amounts of various Ca-Al-Fe silicate-sulphate-hydrate phases and traces of calcite. Together with gypsum and a Ca-Fe sulphate phase, these species bond the various particles together, imparting mechanical strength and reducing the permeability. Water-soluble Zn and Mg are stabilized in the Ca-Al-Fe silicate-sulphate-hydrate phases. The freshly cured products usually contain small amounts of moisture and traces of residual Portland cement, which provides additional alkalinity for the neutralization of any residual acid species. Jarofix products stored for more than six years in the laboratory, after curing, are usually dry and exhibit the same micro-textures as those of the freshly cured products. However, the stored products are depleted in the Ca6Fe2(S04)3(OH)i2.nH20 and Portland cement phases, and are enriched in gypsum and calcite which reflects the highly alkaline environment of the jarofix products. The aged products also have increased amounts of the Ca-Al-Fe silicate-sulphate-hydrate phases which have elevated Zn and Mg contents. This implies a further immobilization of residual water-soluble Zn and Mg during stora

Jan 1, 2000

By W. G. Davenport

The Cu content of industrial flash furnace concentrates varies considerably. It is usually around 30% Cu (Tables 2.1, 3.1) but it may be as high as 55% Cu and as low as 20% Cu. The mineralogy of flash furnace concentrates also varies. The most common Cu mineral is chalcopyrite (CuFeSz), but many concentrates contain bornite (CuSFeS4), chalcocite (CU2S), covellite (CuS) and even native copper (CUO). Pyrite (FeSz) :is also present in many concentrates. This chapter examines how concentrate composition affects flash smelting -with specific reference to 'low-Cu-grade' FeS2-CuFeSz concentrates. Chapter 14 does the same with 'high-Cu-grade' CuzS-CuFeSz?concentrates.

Jan 1, 2001

By John E. Dutrizac, Tzong T. Chen

"Laboratory experiments were undertaken to ascertain the behaviour of various ""impurity"" species during the precipitation of hematite directly from ferric sulphate solutions at 225°C and in the presence of 40 or 50 g/L of hematite seed. Copper, nickel and zinc are not significantly incorporated in hematite precipitates. Although the nickel contents increase to 0.22% as the Ni concentration increases to 60 g/L, the solid solution Ni contents are low relative to the bulk Ni contents. Thus, part of the Ni may be present as tiny particles of poorly soluble NiS04.H20. Increasing germanium concentrations to 2,000 mg/L result in products containing up to 0.59% Ge. The Ge occurs either in solid solution in the hematite or is adsorbed on the nanometer-size hematite crystallites. Increasing As(V) concentrations to 10 g/L increase the AsO4 contents of the products up to 16%. However, the products consist of hematite together with As04-containing Fe(SO4)(OH) and an unknown Fe-AsO4 species. All the hematite products contain - 1% SO4 which appears to be adsorbed on the nanometer-size hematite crystallites.IntroductionDissolved iron is a common impurity in most hydrometallurgical processing circuits and, generally, the iron must be eliminated prior to the subsequent recovery of the sought after metals. Simple neutralization of iron-rich solutions yields slurries which are difficult to thicken and filter. To address these concerns, the jarosite, goethite and hematite processes have been developed, and these technologies are able to reject the dissolved iron in a readily filterable form [1]. Despite its higher cost, the hematite process offers several potential advantages. Because hematite is an iron-rich, high-density compound, its formation would reduce the mass of residue and the size of a residue storage pond. Hematite is the thermodynamically stable form of iron under ambient conditions, and this stability would facilitate the impoundment of hematite precipitates. Furthermore, hematite has some market potential for cement manufacture and even for iron making [2]. In all the commercial hematite processes, the dissolved ferric ions are first reduced to the ferrous state. Subsequently, the pH of the solution is adjusted and the solution is then oxidized. The ferric ions gradually generated during solution oxidation hydrolyze and precipitate mostly as hematite, although commercial hematite products also contain trace or minor amounts of jarosite-type compounds. However, there are potential advantages in precipitating hematite directly from ferric ion solutions, as the reduction and subsequent oxidation of the iron are not required. Although the precipitation of hematite directly from relatively concentrated ferric sulphate solutions (up to 0.5 M Fe(SO4)is) was recently demonstrated [3], the behaviour of impurity species during hematite precipitation is largely unknown. To address this issue, systematic laboratory experiments were undertaken to ascertain the behaviour of a number of ""impurity"" species during the precipitation of hematite directly from ferric sulphate solutions at 225°C, and the results are presented in this paper."

Jan 1, 2012

By Kelly J. Driscoll

Impure zinc (98.5%) was electrorefined to high purity zinc (99.995%) using a fused salt electrolyte of molar composition 1: 1 : 1 :2 ZnCI2:KCl:NaCl:LiCl with a current efficiency of about 90% at 723 K. A recessed channel electrorefining cell was developed which achieved excellent mixing of both the metal and the electrolyte and by ensuring a very small anode to cathode spacing the voltage drop across the cell was 0.16 V and at a current density of 7kA m-2. The physical characteristics of the cell were observed using a low temperature model which showed the formation of droplets and rivulets which provided mixing. The application of this approach to the refining of other non-ferrous metals is discussed.

Jan 1, 1992

By Y. C. Avniel, T. E. Warner

"It is important to be able to measure sulphur and nitrogen in molten iron and steel on-line. Unfortunately, there are no solid electrolytes which are stable under oxidising conditions and in which the ions of these species are mobile. In this paper, novel applications of beta alumina electrolytes are described and the results compared with data in the literature. Excellent agreement was obtained confirming this approach as a viable method of detecting elements whose ions are not mobile in solid electrolytes.IntroductionIn the past two decades, the rate of many processes in steelmaking have increased substantially. The sophistication of chemical analysis has also increased but analytical equipment is frequently positioned a significant distance from the operating plant which dictates that time is spent taking a sample, sending it to the analytical facility, preparation of the analytical sample prior to insertion into the equipment. In many cases, the time taken for analysis can exceed the time taken for the actual process and is now the rate defining step. Furthermore, most analytical methods are unable to distinguish between the amount of element which is dissolved in the steel and that contained in precipitates. Even if the analytical equipment could separate these two, the distribution would have changed between the molten steel and the cast sample. It would be an advantage to analyse on-line for impurities in molten metals."

Jan 1, 1997

By Dalene Joalet Steenkamp

Excavation of industrial-scale furnaces allows for the systematic study of reaction sequences by identifying the different reaction zones within the furnace. In 2013, Transalloys excavated a 48 MVA submerged arc furnace that was used for silicomanganese production using the ore-based route. The excavation method was reported elsewhere as was observations made in terms of refractory wear and modes of electrical energy dissipation prior to excavation. The paper presented here, reports on the process reaction zones observed during the excavation and subsequent phase chemical analyses of a number of process samples obtained during the excavation. The zones identified were a loose burden zone, a dry coke-bed zone, a wet coke-bed zone, a hard build-up zone, and an alloy zone. The results are compared to observations made in the excavation of an industrial-scale SAF and a pilot-scale SAF in Norway. The presence of the hard build-up zone below one of the electrodes and absence of a slag zone below all three electrodes are unique features of the SAF excavated at Transalloys.

By Gary A. Campbell

This article is an analysis of world consumption of copper for the years 1976-1996. It is done to provide a comparison of what has happened over this time period with the trends of previous years in order to learn more about world copper consumption and to forecast world copper consumption trends into the new millennium. The analysis begins by documenting the geographic trends of consumption of copper for the years 1976-1996 by both region and key individual countries. Then, variables based on the intensity-of-use hypothesis and the intensity-of-use technique are selected and tested against the observed consumption behavior by using the correlation coefficient technique. The statistical results and their implications for the consumption of copper into the new millennium are discussed.

Jan 1, 1999

By John C. Carpenter

Solvent extraction is used to selectively recover copper from iron containing sulphate solutions. As chloride ions, which form series of complexes with transition metals, are now being used for leaching. Little is known of the effects of these ions, on solvent extraction behavior. It has been shown that commercial extractants of the L1X and Kelex series extract copper from chloride media: (1) at a lower pH than from sulphate solutions, (2) to similar loadings, and (3) except for L1X 63, without transferring chloride to the organic phase. A process for selective extraction of copper from chloride solutions should thus be compatible with standard stripping and electrowinning practice. Of several diluents, differing in aromaticity, kerosene gave enhanced extraction only with L1X reagents. A change in operating temperature had little effect on equilibrium metal distribution. All the extractants examined gave adequate separation of copper from iron. L1X 63 was found to extract chloro-complexes of copper and of iron and thus transfer chloride to the organic solution. Consequently this extractant and its mixtures are not preferred for use with chloride leaching systems.

Jan 1, 1976

By Fathi Habashi

Laboratory autoclaves for hydrometallurgical investigations are available in a variety of sizes, models, and materials of constructions. They vary in sizes from 25 mL to 2 liters for laboratory studies and 5 to 50 gallons for pilot plant work. They are essential tools for studying aqueous oxidation of sulfide concentrates, dissolution of oxide minerals at high temperature and pressure and hydrothermal precipitation reactions. The maximum pressure and temperature at which any pressure vessel can be used will depend upon the design of the vessel and the materials used in its construction. Since all materials lose strength at elevated temperatures, any pressure rating must be stated in terms of the temperature at which it applies. A review of existing models and their accessories will be given.

Jan 1, 2000

By Kevin L. Purdy

The Process Technology Center of Phelps Dodge Mining Company (PDMC) .operates several test facilities in support of PDMC's hydrometallurgical operations. Installations for small leach columns, 15 cm. to 60 cm. (6 in. to 24 in.) in diameter, and large leach columns, 1.8 m. (6 ft.) in diameter, are complemented by a test facility for SX/EW operations. The history of and reasons for the development of these facilities: are described and original standard equipment is discussed. A primary goal in constructing these facilities was achieving a high degree of flexibility to allow for testing of new equipment and new leaching and SX/EW technologies as they are developed. Along with this challenge, construction of a filly integrated SX/EW test facility on a small scale brought its own set of challenges.

Jan 1, 1999

By Asmo Vartiainen

The viscosity of liquid Fe-O-SiO2 -slag at copper smelting conditions was investigated using a rotational viscometer. The studied parameters were oxygen partial pressure (10" °-10"6 bar ); composition of slag (Fe/SiO2 ratio), and temperature (1250-1400°C). The viscosity values are presented as a function of Fe/SiO2 ratio and as isoviscosity curves in ternary FeO-Fe2O3-SiO2 diagrams. The increase in viscosity at high silica contents can be accounted for the polymerisation effect of SiO4/4- -anions. At high ferric iron contents, higher oxygen partial pressure, the viscosity values increase especially near the magnetite saturation. The viscosity values are highest at the double saturation of silica and magnetite. At lower silica and ferric iron contents the viscosity values are more uniform in the homogenous liquid slag area. The viscosity values were lowest at high temperatures and highest at low temperatures.

Jan 1, 1998

By E. M. Timoshenko

Autoclave technology can be effectively used for liberation of complex mineral raw materials to transfer valuable constituents into solution for their subsequent recovery or to refine feed material in order to remove deleterious impurities, e.g. arsenic. The high efficiency of autoclave processes is attributed to the fact that they are suitable for processing of refractory raw materials, have low capital costs, high productivity, ensure integrated utilization of the feed materials, produce innovative saleable products, completely eliminate any air emissions, provide low-waste or waste-less technology, mechanization and automation of production processes. In the immediate future, it might be expected that autoclave processes will find use on commercial scale for production of heavy non-ferrous metals in combination with other processes (hydrometallurgical, upgrading, sorption and solvent extraction, electrolytic, etc.) to ensure: - Integrated processing of refractory ores and concentrates (pyrrhotite, zinc, copper-zinc, lead-zinc, nickel-cobalt, copper, pyrite); - Production of metallic powders of base metals as commercial products (copper, cobalt, nickel, etc.), composites on their bases, final products and saleable salts. Processing technologies incorporating autoclave processes, especially at the stage of liberation (chemical upgrading) of raw materials are characterized generally by high overall recoveries of valuable constituents significantly exceeding recoveries common for conventional pyrometallurgical processes (for some elements by several tens of percent).

Jan 1, 2006

By Edilson Marques Magalhães, Danielly Da Silva Quaresma, José Antônio da Silva Souza, João Nazareno Nonato Quaresma, Emanuel Negrão Macêdo

"The production of alumina from bauxite using the Bayer process generates the residue known by red mud. This material can be used as a source for concentration and recovery of valuable metals such as titanium. The greatest difficulty of concentrating titanium compounds is the large amount of iron oxide present in red mud > 30% by weight. This paper shows the extraction of the iron oxide, increasing the concentration of the titanium compounds. The red mud is leached with H2SO4 at 30% concentration and temperature range of 60°C, 80°C and 90°C. Analysis of the leached and concentrated was made to verify the effectiveness of the iron oxide extracting process and the titanium compounds concentration viability.2- IntroductionIn the state of Para is located the Industrial Unit Alumina of North of the Brazil S/A (Alunorte), created to produce and commercialize alumina, aluminum oxide, Al2O3. The ALUNORTE with Aluminum Brasileiro S/A (Albras) and Mining Rio do Norte S/A (MRN) complete the production cycle of aluminum in the state. Although the production of alumina has a importance great for the socio-economic status, as in any area other, generates a amount large of waste, thus causing environmental impacts. Pará State has produced 25 million tons of bauxite (Pinto, 2010), of which 60% were transformed into alumina plant in Pará and Maranhão by ALUNORTE by ALCOA, generating approximately 9,000,000 tons of waste known as Red Mud (RM). Due to the high reactivity this material is considered dangerous, and currently there is no operating process capable of reuse all the material produced at any refinery. The search for industrial processes in order to reuse these wastes can contribute to reducing the overall impact of the production process of alumina. In this context, this paper estudied, within the research line of Solid Waste valorization the mining industry and metallurgy, a process that makes possible the red mud as a source of raw material, in other words, seeks to maximize the use of red mud as raw material for concentration and or recovery of titanium compounds in both the crystalline form of anatase, rutile and ilmenite, avoiding the predatory extraction of natural resources of these materials. Applying definitely the concept of self-sustainability of the mining industry and metallurgy."

Jan 1, 2012

A study of iron-removal from copper leachate by limestone precipitation followed by direct eltctrowinning of copper has been conducted. The influence of the process parameters: amount of precipitant, contact time, temperature, agitation intensity, solution acidity, and initial iron concentration on iron-removal was investigated. The effects of electrowinning time, cell voltage, solution acidity and temperature on electrowinning of copper were also investigated. The leachate contained 1.20 g/l cu2+ and 7.41 g/l total iron at pH 2.27. After contacting with 14 g/l precipitant for 2 hours, 97.5 % of the iron was attenuated and 98.4 % of the copper was recovered. With the following electrowinning conditions: 3 lead anodes, 3 stainless steel cathodes, 2.5 V, 39.41 A/m2 and 10.7 g/l H2S04, copper recovery-was 98.4 %. Purity of copper powder produced was greater than 99.0 %.

Jan 1, 1999

By M. -L. Metsärinta, J. Nyberg, A. Roine

The Kokkola roasting plant uses more than 20 different zinc concentrates every year. During recent years there has been an increase in the variation in concentrate quality; in particular, the amount of fine-grained and low-grade concentrates has increased. Therefore, a great deal of development work has been carried out to improve the operational flexibility, online availability, production capacity and calcine quality of the roasting plant. This work consisted of a heat and material balance model of the roasting based on HSC Chemistry software, a new water admixture system for the concentrate feed, a new oxygen feed system for the process air line, improved cooling of the furnace and boiler and improved operational practices, dust removal and concentrate mixing. These new process control tools and other improvements, as well as some test results, will be described in this paper. The oxygen feed may be used to improve production capacity and to decrease the sulfide content of the calcine. The water feed may be used with fine concentrates to decrease both dust carry-over and the temperature in the top part of the furnace. Improved cooling of the furnace and boiler helps to increase production capacity.

Jan 1, 2000

By I. H. Bonekamp

Norddeutsche Affinerie has operated an Outokurnpu type flash smelter for more than 25 years. In recent years, the concentrate smelting capacity was increased to 750,000 tpy, double of the initial design capacity. Maintaining this high performance level requires very good metallurgical process control. It is not enough to base this control on thermodynamics and kinetics since the furnace is highly dynamic. Understanding these dynamics and being able to quantify them can have significant financial benefits. A mass and energy balance model was developed and is discussed based on measured data. At this time the model is based on the major elements copper, iron, sulfur and silica. With it is possible to quantify the differences of the various material streams such as matte, slag, flue dust, matte carry-over to settling furnace, etc. Comparing the reconciled energy balance model with the fundamental thermodynamic model makes it possible to quantify the dynamics. These dynamics are related to input parameters using methods such as statistics and neural nets. In the future this database will be used to improve the understanding of the dynamic process as well as the dynamic process control.

Jan 1, 1999

By Aliye Arabaci

In the present investigation, an attempt has been made to study the reduction behavior of composite pellets consisting two different types of iron oxide fines (high and low grade) and coke or lignite fines as reducing agents. The reduction experiments were carried out under isothermal conditions and constant flow of nitrogen gas in the temperature range of 900 ? 1100 oC. The variables investigated are the reduction temperature, Cfix/Fe3O4 ratio, reduction time and reductant type. In the experiments carried out with Sivas Divrigi A head magnetite ores (high grade) at a temperature of 1100 oC, for 40 minutes and on the condition that Cfix/Fe3O4 ratio is 0.2, in which lignite and coke fines were used as reductant, degree of reduction was obtained as 91 % and 86 %, respectively. Whereas, in the experiments realized with Malatya-Hekimhan-Hasançelebi magnetite ores (low grade) on the same conditions by using lignite as the reductant, around 60 % degree of reduction was found.

Jan 1, 2008

By James J. Komadina, Robert R. Beebe

"Heap leaching of gold ores has enjoyed wide acceptance in the arid and semi-arid regions of the western U.S., Australia and South America. There are areas, however, where cold weather or excessive precipitation can limit the usefulness of heap leaching. Even in Nevada difficulties can be encountered which call for special operating procedures.The paper discusses the general climatological limitations and some of the techniques used to extend heap leaching into colder conditions. A case history of the Cripple Creek and Victor Gold Mining Company operation in Colorado examines 'valley-fill"" technology as an ideal cold-climate solution, where topographical conditions permit.AcknowledgementsThe authors gratefully acknowledge the management of Independence Mining Company for their permission to present this paper. The contributions of Golder Associates; Pikes Peak Mining Company Environmental Affairs Department staff, and the Colorado Department of Mines and Geology were all invaluable in successfully concluding this project.BackgroundWhen heap leaching of oxidized gold ores was introduced in the rnid-1970s it was aimed at large tonnages of low-grade material which could not, in general, be developed via conventional milling. Normally, the reasons for this were economic rather than technical: capital and operating costs for milling were simply too high. At first, grade was almost the only criterion, but deposits too small to justify milling were also potential targets for heap leaching.Renewed interest in gold was centered in the western United States and Australia, where exploration for milling-grade deposits had never really died. When heap leaching came along, it was in these areas that the first developments took place. Coincidentally, these same areas were favorable for heap leaching: climates were temperate for most of the year and competing land uses were limited. Even so, in the western United States in particular, winter temperatures were low enough to complicate operations. Precipitation was less a problem, although heavy snowfalls could periodically interfere as well."

Jan 1, 1997