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By O. Hyvärinen

Outokumpu has developed a new chloride leaching process called HydroCopper®. In this process copper sulfide concentrates are leached at atmospheric pressure in concentrated sodium chloride solution using cupric ions as the oxidant. Copper is recovered from the purified leach solution by precipitating cuprous oxide, which is then reduced to pure copper and further melted and cast directly into copper products. The key reactants: sodium hydroxide, hydrogen and chlorine are generated from the spent sodium chloride using chlor-alkali technology. An important feature in the HydroCopperTM process is that the gold contained in copper concentrates is leached as chloride complexes and recovered from solution by activated carbon. Conditions for the leaching of gold-bearing pyrite can also be achieved in the chloride system by optimizing the redox conditions. This paper deals with the basic factors affecting gold leaching from copper sulfide concentrates. Opportunities for the recovery of refractory gold associated with arsenopyrite and pyrite will also be discussed.

Jan 1, 2005

By Edgar Peek

"The two dominant metallurgical problems in treating most base metal sulphide concentrates in both the chloride and sulphate system are iron removal and sulphur elimination, respectively [1-3]. In this paper the opportunities, limitations and fundamentals of the chloride metallurgy system are described against this background. Only the treatment of oxide and sulphide ores is considered. Both hydrometallurgical and pyrometallurgical unit process operations are highlighted, since most successful base metal processes are based on a combination of the two. Hence, an overview of the technical and also some economical aspects of chloride metallurgy are presented.This overview summarizes the successful applications of chloride metallurgy, but it will not give exhaustive process and plant descriptions. It predominantly focuses on the essential technical features of the metallurgical unit process operations, while providing numerous references separately."

Jan 1, 2011

By William D. Riley

The U. S. Bureau of Mines has developed technology to recover by-product materials from aluminum scrap using engineered scavenger compounds (ESC). ESCs are structural oxides with a channel or tunnel structure that allows them to hold ions of a specitksizes and charges. The scavenging reaction is easily reversible allowing the ESC to be recharged for continued use and the ion is recovered as an electro deposit. Key features of this novel technology are; a) ESC ,systems are designed to have a high degree of selectivity for a desired ionic species. b) The recovered mater.ial requires little or no additional reprocessing prior to reuse. Two current uses for the ESC technology that are described in this paper are the removal and recycle oflithium (Li) from lithium aluminum (Li-AL) alloys; and, using ESCs as a replacement for the conventional demaging (magnesium removal) technology used by the secondary casting industry. Research indicates that the ESC technology proposed for both these applications has either 'distinct economic and/or environmental advantages over previously employed methods of recovering. metal values from aluminum scrap.

Jan 1, 1995

By Van Weert G

Ferrous chloride solution pyrohydrolysis is a successful IICI regeneration process which is carried out in either a spray roaster or a fluid bed. Pyrohydrolysis is not restricted to ferrous chloride only, every metal chloride can theoretically be hydrolysed. In practice, however, the conditions for ferrous chloride pyrohydrolysis differ from those of many non-ferrous metal chlorides eg zinc, lead and cadmium. Hence, a pyrohydrolyser could function as a ferrous/non-ferrous separator, eg an iron oxide and zinc chloride separator. Most pyrohydrolysers operate on solutions, in which case more than 50 per cent of the process energy is needed for heating and evaporating water. However, this disadvantage would be diminished by feeding a slurry to the reactor. Against this background it was decided to investigate the chloridisation of zinc ferrite which is an iron bearing solid with a non-ferrous impurity, and subject to processing and disposal problems. Synthetic zinc ferrite was chloridised in a tube furnace. The gas atmosphere in the tube furnace resembled that of a fluid bed pyrohydrolyser and consisted of HCI, 1120 and nitrogen. Gas composition, temperature, residence time and sample mass were varied. The experiments showed that complete (99.6 per cent) zinc extraction from synthetic zinc ferrite was possible in a simplified pyrohydrolyser atmosphere at 1100 K in four hours. The zinc ferrite was converted into gaseous zinc chloride, and hematite and magnetite solids. It was also found that zinc removal from the zinc ferrite lattice caused not only rccrystallisation of the zinc ferrite into hematite, but also hematite crystal

Jan 1, 1993

Discussion of the paper of THEODORE P. HOLT, presented at the Salt Lake meeting August, 1914, and printed in Bulletin No. 91, July, 1914, pp. 1699 to 1708. F. S. SCHMIDT, Salt Lake City, Utah.-Any furnace that can make a chloridizing roast to yield an extraction of 92 to 93 per cent. and do this at a cost of less than 24c. per ton in a 10-ton unit opens great possibilities for the treatment of certain classes of ores. This cost does not include the salt and will be reduced to about 16 1/2c. per ton in the 35-ton furnace which are now being designed. On a trial run the roaster treated 9.6 ton: of Ontario stope fillings crushed to pass a ¼ -in. opening. The ore is well adapted to chlorination with the exception of insufficient sulphur and occasional days of high lime, but is not adapted to any other process The following is an analysis of the crude ore: Ag, 9.52; Au, 0.029 oz. Pb, 0.71; Cu, 0.11; SiO, 79.4; Zn, 0.57; S, 1.00; Fe, 7.9; Mn, 0.89 CaO, 1.20 per cent. The composition of the mix was: Per Cent. Dry crude ore : 83.4 Moisture 4.4 Salt 8.3 Coal dust 2.3 Pyrite, 70 per cent. pure 1.6

Jan 11, 1914

By H. P. Allen

THE chloridizing mill of the Standard Reduction Co. is located about 75 miles south of Salt Lake City on the Tintic branch of the Denver & Rio Grande Western R. R. and 12 miles from the Tintic Standard mine. The daily capacity is 200 tons of a siliceous, low-grade, silver-lead ore from this property. It has operated continuously since it was started in January, 1921. The process consists essentially of a chloridizing roast followed by a percolating leach with a nearly saturated solution of common salt, acidified with sulfuric acid, the precipitation of silver on sponge copper and of copper and lead on tin-plate cuttings. The precipitates are shipped to a smelter. Some of the general ideas involved are said to have been used by Augustin in England, in 1840. A number of textbooks treat of the subject, especially the chloridizing roast followed by a leach with sodium hyposulfite or amalgamation. The process was revived in this district by Theo. P. Holt, N. C. Christensen, the Bureau of Mines, and others.

Jan 8, 1925

By R. Padilla, M. C. Ruiz, L. Salinas

Chloridizing roasting of bismuthinite (Bi2S3) with NaCl–O2 has been studied to remove the impurity Bi by volatilization from copper concentrates that contain bismuthinite. The study was conducted using mixtures of Bi2S3 and NaCl in a horizontal furnace and TGA apparatus. The variables were temperature, oxygen partial pressure, and NaCl concentration. The chlorination was analyzed by weight loss measurement method. XRD results of calcines reacted for short times in 21% O2, and 850–1000 °C showed the presence of Bi, BiOCl, and Na2SO4, while at longer times in the same conditions, Bi2O3 was identified as a stable phase. Thus, the chloridizing-volatilization of Bi2S3 proceeds through the formation of Bi and BiOCl. Temperature affected significantly the fraction of weight loss of samples. At 600 and 900 °C in 10.3% oxygen, the maximum weight loss was 4 and 20%, respectively. An increment in the partial pressure of oxygen in the range 1–21% affected negatively the Bi2S3 conversion. At normal roasting temperatures (550–700 °C), bismuth volatilization by NaCl roasting was found to be marginal.

Mar 1, 2017

By R. J. Nankee

Coal is our most abundant, low cost energy resource. Expansion of its use is being challenged due to "SO2 emission" regulation with even more restrictive legislation pending. High quality coal reserves are becoming more limited due to greater use in recent years. Aids to allow more use of coal in today's environment include beneficiation of coal before combustion and scrubbing the emissions from coal during combustion. Dow's efforts to help meet these challenges by beneficiating coal include true heavy media gravity separations with chlorinated solvents and liquid/liquid partitioning processes and "Dow Solvent Recovery Technology" to recover the solvent from product (or floats) and rejects (sinks), The objectives of the efforts to clean coal are to reduce sulfur and ash. In this paper, the solvents will be described; a general flow diagram presented; some of the results on upgrading various coals will be described, and projected costs of full-scale plants using chlorinated solvents will be compared to conventional cleaning plants.

Jan 1, 1985

By R. J. Nankee

Recent on-going efforts to develop commercial coal cleaning beneficiation processes using chlorinated solvents-as the separating media will be described. True heavy media separations were performed on raw coal feedstocks, including Pennsylvania anthracite; Western, Eastern, and Midwestern bituminous; and lignite coals. Separation techniques included bath-type devices, cyclones, and liquid/liquid partitioning. Details on separation devices, removal of solvent, and recovery and recycling of solvent will be discussed. Projected costs of full-scale plants will be given for various coal sizes. A number of the coals evaluated are good candidates for beneficiation using chlorinated solvents. Coal beneficiation is desirable to provide needed coal quality. These needs include low ash fuels for coal water slurries and low sulfur fuels to meet emission regulations. In October of 1985, a paper, "Chlorinated Solvents for Coal Beneficiating"' was presented at the fall meeting of SME-AIME at Albuquerque, New Mexico. It reported the results obtained from various coals as feedstocks and of the separations using chlorinated solvents. These beneficiation results depended on true heavy media gravity and liquid/liquid partitioning separations. This paper will include detailed descriptions of the liquid/liquid partitioning process used in "Dow's Advanced Coal Cleaning Technology?2 and of the "Dow Solvent Recovery Technology.?3

Jan 1, 1987

By I. Gaballah

High and low grade concentrates of tantalum and niobium oxides were obtained by leaching a medium grade tin slag. Chlorination and carbochlorinalion of these concentrates were studied between 200 and 1000 "C. Chlorination or carbochlorination of the low grade concentrate led, at 1000 OC, to the extraction of about 95 % of the tantalum and niobium oxides. However, the tantalum and niobium chlorinated compounds were contaminated by iron and manganese chlorides. Complete extraction of the niobium and tantalum pentoxide was achieved during the carbochlorination of the high grade concentrate at 300 and 500 "C, respectively. While chlorination of this concentrate at 1000 "C, allowed to the volatilization of about 60 and 80 % of Ta205 and Nb205. In both cases, condensates were composed of almost pure tantalum and niobium chlorinated compounds. The results demonstrate the advantages of the lixivialion based on the "pseudo- structure" of the raw tin slag to obtain the high grade concentrate. The conclusions of this work suggest that the carbochlorination of the high grade concentrate is an efficient process for the extraction of refractory elements from tin slag.

Jan 1, 1993

By Guolin Zheng, Tiejun Chun, Jian Pan, Deqing Zhu, Xianlin Zhou, Dong Chen

"The high temperature oxidizing-chloridizing roasting is one of useful methods to separate nonferrous metals from pyrite cinder and metallurgy slag. However, as one kind of main nonferrous metals in pyrite cinder and metallurgy slag, copper’s chlorination behavior is not demonstrated. In the present study, chlorination behaviors of CuO, Cu2O and CuS at the temperature range of 1025-1175 ? in air by simulating the chloridizing roasting process of pyrite cinder and metallurgy slag have been investigated. It is shown that Cu2O and CuS is much easier to remove, CuO is refractory to remove. In addition, the presence of SiO2 and CaO increases the Cu removal rate from CuO while the presence of SiO2, Al2O3, CaO and MgO decreases the Cu removal rate from Cu2O. Fe3O4 influences the Cu removal rate from CuO and Cu2O positively.IntroductionAs two kinds of significant secondary resources, pyrite cinder and metallurgy slag contain not only abundant iron but also considerable copper, lead, zinc etc. Millions of tons of pyrite cinder and metallurgy slag were discharged from nonferrous metal and chemical industries annually [1, 2] in China. The pyrite cinder and metallurgy slag are formed by roasting of minerals at a high temperature, so the properties of them are distinctly different from the natural iron ores, and the iron in them is closely associated with copper, lead, zinc and other nonferrous metals. High-quality pyrite cinder and metallurgy slag containing high iron and low impurities can hardly be obtained, leading to their low utilization degree [3]. Chlorination is characterized by the high reactivity of chloridion, the high volatility and low melting point of metal chlorides, and the big differences in the formation of metal chloride. Therefore, chlorination is widely applied in extraction metallurgy."

Jan 1, 2012

By M. A. Gimenes

The utilization of chlorination in extractive metallurgy and advanced ceramics areas has been widely investigated as a preparative route in order to obtain intermetallic compounds, which are used as precursors in the development of new materials and process. The major processes of decomposition of rare-earth ores use sulfuric acid and sodium hydroxide at high temperatures, however there is not much work concerning chlorination. In this work, a systematic study of the reaction between xenotime (REPO4), chlorine and carbon has been performed. Particular emphasis was on kinetic studies to establish optimized conditions for the reaction. The kinetics of chlorination of xenotime raw material by rare-earth elements/compounds has been studied over a temperature range from 600°C to 950°C. The influence on the rate of conversion of xenotime to RECl3 on temperature, partial pressure of chlorine, carbon content, and particle size were investigated. A global rate equation that includes these parameters has been developed. The results show that the process follows the unreacted core-shrinking model with a formation of a porous product layer. The powder X ray-diffraction technique corroborated this model showing clearly the patterns related to the formation of yttrium oxychloride (YOCl), indicating that the reaction mechanism involves the presence ofan intermediate step before the formation of lanthanide chloride.

Jan 1, 2000

By A. L. Sweetser

EXPERIENCE has shown how difficult it is to obtain information regarding laboratory-tests in connection with the chlorination-process for the extraction of gold from its ores, and I therefore present the following method, somewhat in detail, for the benefit of those who may desire to pursue research work in this field. The ore chosen was a partly-decomposed porphyry, extremely siliceous and comparatively soft, and of an average value per ton of from $25 to $26 in gold and $0.68 in silver. The chemical analysis gave : Fe2O3, 3.35; FeS, 1.5; MnO2, 0.75; ZnCO3, 4; Al2O3, 3.20; H2O, 0.7; SiO2 and insol., 83.50 per cent. The ore, first crushed to pass a 5-mesh screen, was coned and quartered until a 45-lb. sample was obtained. Samples for assay and analysis, taken from this lot by means of riffles, were crushed to pass a 100-mesh screen. The original sample was divided into 10-lb. lots, of which one was crushed to pass a 10-mesh screen; another through a 12-; the third through a 20-, .and the last through a 30-mesh screen. Because of the production of an excessive quantity of fine material it was deemed inexpedient to crush finer than 30-mesh size-a decision the wisdom of which is proved by the data in Table I., showing the increasing ratio of slimes, in proportion as the ore is crushed finer. In order to determine the percentage of fines in the sizes of mesh selected, a weighed sample of ore was taken from each mesh, and a sample from the 10-mesh lot was passed successively through 20-, 40- and 100-mesh screens, and the percentage remaining on each screen determined. This process was repeated with the 12-, 20- and 30-mesh lots, with the results given in Table I.

Sep 1, 1907

By Charles H. White

Discussion of the paper of A. L. Sweetser, Tress., xxxviii., 236. CHARLES H. WHITE, Harvard University, Cambridge, Mass. (communication to the Secretary*) :-Those interested in leaching-processes gladly welcome contributions on either laboratory-tests, such as are reported by Mr. Sweetser, or mill-tests made on larger quantities of ore; and especially valuable are such reports, even with negative results, if the methods adopted are the outgrowth of an intimate knowledge of the conditions to be met. It seems very desirable, therefore, that as complete details as possible be given in reporting on such tests, particularly with respect to the character of the ore under examination. The questions of prime importance in the sampling of ore and in the making up of charges for laboratory-tests are those respecting its grade and uniformity. In Mr. Sweetser's paper the first of these questions is answered definitely, but the answer to the second, although apparent on inspection of the results, might be more strongly emphasized with advantage, since it is only on account of the extreme uniformity of the ore that the satisfactory results obtained could have been expected. It seems advisable to call attention to the fact that for average ores, and more particularly for uneven, or 11 spotted ores, the quantities used by Mr. Sweetser are not usually regarded as sufficient to insure satisfactory results. Let us examine the method somewhat in detail. The ore was crushed to pass a 5-mesh screen, and a sample of 45 lb. was taken out for assay and tests. The diameter of the largest grain of this ore, according to Richards,1 was 2.67 mm.-assuming here, as in the following cases, that the wire of the screen was of the largest size used for the purpose, thereby giving any possible advantage in this respect in favor of the method. The table in Richards's book, p. 852, shows that 45 lb. is a sufficient quantity for a fair sample with uniform ores, but not sufficient

Jan 7, 1908

By Ana E. Bohé, Georgina De Micco, Gastón G. Fouga

"In this work the reaction between several manganese oxides and chlorine is investigated in the range of temperatures between 650 and 950ºC. The reaction path for the chlorination of MnO is established, which involves recrystallization of high valence manganese oxides (Mn3O4 and Mn2O3) followed by its chlorination. For this reason, to understand the reaction of MnO, the chlorination of Mn3O4 and Mn2O3 has also been studied. The starting temperature for the reaction of these oxides is determined by non isothermal thermogravimetric measurements. The effect of temperature is analyzed, under our experimental conditions the rate of reaction is mass transfer controlled. The activation energy for the volatilization of manganese chloride is obtained which is in good agreement with volatilization enthalpy of MnCl2.IntroductionMany oxide ores which have manganese in their composition undergo chlorination processes during the extraction of refractory metals such as Ti, Hf and Nb. Although several works related with chlorination of manganese oxides were reported, there is no information about the kinetics or mechanism of the manganese oxide chlorination reactions. Okahara et al. studied the chlorination of manganese oxides and established that they react easily above 250ºC. Furthermore, they observed that chlorination of high order oxides is more difficult when there is not a reducing agent [1]. Ivashentsev studied the chlorination of manganese oxides using Cl2(g), CO(g) + Cl2(g) and COCl2(g). He established that the reaction of MnO started at 130ºC with formation of MnCl2(s), he also observed recrystallization of manganese sexquioxide [2]. Finally, the chlorination of manganese nodules from the ocean was studied by Murthy and Reddy [3] with the aim of obtaining valuable metals such as Mn, Co, Ni and Cu. They studied the influence of several parameters like temperature, carbon percentage, particle size and chlorine flow rate in the extraction process."

Jan 1, 2008

By Samantha Moodley, Rauf Hurman Eric, Cevat Kucukkaragoz, Aditya Kale

"Two titania slags, rutile and synthetic rutile were chlorinated with petroleum coke and CO in a small bubbling fluidized bed reactor. The study aims to identify differences in chlorination mechanism, compare conversion rates, blowover and the chlorination of impurities for the various titania feedstocks at different temperatures.Chlorination rates were highest at 1000ºC; rutile chlorination significantly increases as temperature increases from 800ºC to 1000ºC. At 1000ºC, synthetic rutile had the highest chlorination conversion rate; this was followed by Slag B which in turn was more reactive than Slag A and rutile. The mechanism for slag, synthetic rutile (SR) and rutile chlorination differs. Synthetic rutile feed is porous, providing a larger surface area for the chlorination reaction, hence the highest conversion rates was attained. Titania slag becomes porous with the initial chlorination of FeO and MnO whilst rutile remains solid. As the porosity of slag particles increases so does its tendency to be elutriated. Ti2O3 is oxidized within the early stages of chlorination during the chlorination of FeO and MnO. Ti2O3 not oxidized is then rapidly chlorinated.IntroductionTitanium dioxide (TiO2) is the most common compound of titanium, about 95 - 98% of extracted titanium minerals is processed into TiO2 white pigment [1]. Two processing routes exist for the production of TiO2 pigments, i.e., sulphate and chloride based processes. In the chloride process, titania feedstock is fed to a fluidized bed reactor together with reductant and chlorine gas. A temperature of 1000 - 1300ºC is maintained in the reaction zone and apart from silica, zirconia, uranium and thorium, virtually all the metal oxides in the feed are converted to their respective chlorides [2]. Chlorine is recovered and recycled back to the reactor."

Jan 1, 2012

By D. S. Barr

The Freeport McMoRan chlorination circuit is comprised of four related processes: ore chlorination, gas scrubbing, chlorine supply and excess hypochlorite removal ("hypo kill"). Each of these processes are described, and related to the pertinent mineralogic components in the carbonaceous ore feed to the plant. The chlorination system in use at Jerritt Canyon to passivate forms of preg-robbing organic carbon has proven to be a cost effective and efficient means to recover gold from low sulfide carbonaceous refractory ores, where gold may not be specifically associated with either the sulfides or the carbon.

Jan 1, 1987

By G. L. Hundley, R. E. Mussler, D. H. Yee, F. E. Block, R. S. Olsen

An anhydrous chlorination process for the recovery of copper from chalcopyrite was investigated. Pelletized concentrate was reacted continuously with gaseous chlorine in a vertical shaft reactor at 550° and 650°C. While the copper was withdrawn with the rest of the metals as molten chlorides from the bottom, 99-x% sulfur was distilled from the top of the reactor. The molten metal chlorides were pulverized and dropped continuously into an upward flowing stream of oxygen. Over 80% of the iron chlorides were converted to oxides in a vertical free-fall reactor at 800° and 900°C while the copper chlorides passed through unoxidized. Copper was recovered in powder form by electrowinning continuously from chloride solutions using a diaphragm cell. The product sloughed easily from a titanium cathode and showed no iron contamination. Current efficiencies and power consumptions averaged 77% and 1.24 kw-hr per lb.

Jan 1, 1974

By Larry G. Twidwell, Jannette L. Chorney, Bryce D. Ruffier, Katelyn M. Lyons, Daniel W. Gaede, Ryan J. Foy, Jerome P. Downey

The Army Research Laboratory (ARL) has an interest in the recovery and processing of rare earth elements. At Montana Tech, elevated temperature chlorination of rare earth oxides is presently being investigated as a potential methods leading to the selective separation and recovery of rare earth elements from mineral ores and concentrates. After preliminary evaluation, ytterbium oxide was selected as a representative surrogate material for more extensive chlorination studies. Methods used to determine the conversion and recovery include X-ray diffraction (XRD) and inductively coupled plasma (ICP) analysis. These data have been used to establish and confirm process conditions that consistently yield high conversion efficiencies. Based on the promising results, further experimentation is underway to ascertain the metallurgical response of other rare earth oxides and carbonates to the roasting process. The end goal is to create a robust process for industrial application that has greater efficiency and reduced costs as compared to current practice.

Jan 1, 2015

By John B. Wright

This paper describes laboratory techniques and subsequent results of Bureau of Mines research to produce chlorination-grade feedstock from an abundant, low-grade, domestic, rock ilmenite ore. The research is part of the Bureau-effort to devise technology that may help decrease U.S. dependence on imported raw materials. A rock ilmenite containing about 46 wt pct Ti02 was smelted in an electric arc furnace with coke, woodchips, and Na2C03 to separate most of the iron as pig iron and to form a low-iron, titanium-enriched slag. The slag was ground, pelletized, oxidized, and then sulfated with mixtures of SO2 and air in a continuous 4.5- to 9.1- kg/h apparatus at 1023 to 1223 K. The sulfated slag, containing 60 to 65 pct TiO2, was reground and leached in water at ambient temperature to decrease the combined levels of the troublesome Ca, Mg, and Mn impurities from 7.0 to 1.7 wt pct. At levels of more than 3 wt pct, these impurities contribute to operating and disposal problems, The final product was upgraded to about 84 wt pct TiO2. This material potentially is suitable for chlorination in a fluidized-bed reactor. The most favorable results were achieved with leached slag sulfated at a feedrate of 4.5 kg/h and at temperatures of 1073 to 1123 K.

Jan 1, 1984