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he leaching behavior of palladium in ammoniacal solutions has been investigated by using the rotating disk technique in an autoclave. The study included the effect of rotating speed, temperature, oxygen pressure, composition of ammoniacal solutions, and catalyst on the overall leaching rate. The dissolution of palladium in ammoniacal solutions was discussed in view of the thermodynamics and a kinetic model. Even though the potential-pH diagram suggests that palladium metal can dissolve in ammoniacal solutions at room temperature, it was found that the dissolution was too slow at temperatures below 100°C but was significantly improved at higher temperatures. It was also found that the rotating speed of palladium disk had no effect on the leaching rate. The apparent activation energy was found to be 69.7 kJ/mole. Based on the results obtained, it is concluded that the dissolution of palladium is controlled by chemical reaction.

Jan 1, 1997

By X. Xie

The dissolution behavior of palladium powder in thiourea solutions has been studied. The variables studied included temperature, stirring speed, and concentrations of thiourea, acid and ferric ion. Palladium was found to be soluble in thiourea solutions in the presence of an oxidant such as ferric ion. The dissolution of palladium in thiourea solutions was found to be chemical reaction controlled with an apparent activation energy of about 2S.3 kJ/mol.

Jan 1, 1995

By Xinghui Meng

The leaching behavior of platinum powder in .iodide solutions under oxygen pressure in an autoclave was investigated. Ammonium salts were added to the leaching solution at neutral pH as a key lixiviant for platinum. The effect of temperature, partial pressure of oxygen, concentration of iodide and of arttmonium salts on the dissolution rate of platinum was examined. Under experimental conditions, the leaching behavior of platinum powder exhibited a complex reaction mechanism. A precipitate was found during the leaching process and was identified to be platinum ammine iodide, Pt(NH3hI4. The dissolution rate was controlled by a series of chemical reactions, the equilibrium concentration of the platinum ammine iodide being responsible for the ultimate limit to the recovery.

Jan 1, 1996

By D. Hong

The leaching behavior of rhenium in ammonium iodide solutions with iodine as oxidant has been studied. Experiments were carried out in an autoclave. The effect of stirring speed, temperature, pH of the solution, iodine and iodide concentrations on the dissolution of rhenium was investigated. It was found that high temperature and high pH of the solutions were essential for the efficient leaching of rhenium. The dissolution rate was found to be dependent upon the concentration of iodine and iodide. It was concluded that the dissolution of rhenium under the conditions studied was controlled by mass transfer of iodine with an apparent activation energy of 15.77 kJ/mole.

Jan 1, 1997

By S. J. Ma

To look into the ionic release laws of the sulphide tailings in some copper mine, the effects of the soaking period, temperature, oscillating speed, and pH of the solution on the heavy metals ionic re-lease were investigated through static soaking leaching experiments. The results show that Cu2+, Zn2+, Pb2+, Cr3+ are of different leaching behaviors with soaking period. The ionic concentration in soaking solution for four ions has a great gap. Zn2+ is more easily released out of sulfide tailings than Cu2+, Pb2+, Cr3+ , and the ion release concentrations are not proportion to their contents in the tailings. The time for each ion to reach its maximum concentration are different. The minimum concentration do not occur at the first day. Each ion concentration shows different fluctuation characteristic during soaking process. Solution pH, oscillating speed, and temperature affect the releases of Cu2+, Zn2+, Pb2+, Cr3+ out of the tailings to different extent. The effect of pH value on ion re-lease for the four ions ranks as Zn2+>Cu2+>Pb2+>Cr3+, the oscillating speed ranks as Pb2+>Zn2+>Cu2+>Cr3+, and the temperature ranks as Cr3+>Pb2+>Zn2+>Cu2+.

Jan 1, 2014

By Y. C. Guan

The leaching kinetics of chalcopyrite in ammonium iodide solutions with iodine has been studied using rotating disc. The variables studied J include the concentrations of lixiviants, rotation speed, pH of the solution, reaction temperature, and reaction product layer. The leaching rate was found to be independent with the disc rotating speed. The apparent activation energy was measured to be about 50 kJ/mole from 15°C to 35°C, and 30.3 kJ/mole from 35°C to 60°C. Experimental findings were described by a kinetic model, and the overall rate was found to be: (formula)

Jan 1, 1997

By Song Ning

The main mineral constituents of phase-transformed chalcopyrite are CuS and FeS2. CuS is more soluble than FeS2 in an acidic CuCl2 solution with presence of excessive NaCl, so that FeS2 could be removed as solid residue from a leached solution. The experimental results showed that the efficiency of the copper leaching is heavily determined by the following two factors: a) The degree of the chalcopyrite phase transformation to CuS and FeS2 b)The concentration of Cl during the leaching, specifically whether there is enough excessive Cl existed. I t was found that when the Cl in the solution is not excessive, the slowest step of the leaching is the dissolution of Cu2Cl2.

Jan 1, 2006

By D. L. Jones

The leaching of natural chalcopyrite was studied in both ferric sulfate and ferric chloride solutions. The chemistry of the two reactions is quite different, with substantial sulfur oxidation occurring only in sulfate solutions. Ferric chloride is a more effective reagent, for the leaching rate is greatly increased with finer particles and higher (Fe+++), and unaffected by high (Fe++); none of these advantages are found with ferric sulfate. In addition, the morphology of leaching appears to be different, with ferric sulfate leaching preferentially along fissures or grain boundaries. The kinetics were found to be linear over an extended period, rather than parabolic as reported elsewhere. An explanation of the observed phenanena is proposed using the electrochemical theory of leaching, and experimental measurement of the mixed potential of the mineral during leaching.

Jan 1, 1976

Discussion of the papers of FREDERICH LAIST and HAROLD W. ALDRICH, FREDERICH LAIST and F. F. FRICK, W. L. AUSTIN, and STUART CROASDALE, presented at the Salt Lake meeting, August, 1914, and printed in Bulletins No. 91 and 92, July, and August, 1914. R. C. CANBY, Wallingford, Conn. (communication to the Secretary*). -Apropos of the experimental reduction of copper from cuprous chloride by fusion with ground limestone and coke, as described by Messrs. Laist and Frick, and also proposed for use at Chuquicamata., it might be of interest to refer briefly to the experience at Argentine, Kan., in the smelting of the suboxide of copper (as produced in the Hunt and Douglas process) containing appreciable quantities of chlorides. In the Engineering and Mining Journal of June 10, 1911, I. gave some account of what had appeared to be crucial difficulties with the Hunt and Douglas process, and how the filter-press difficulty had been so rapidly overcome after my discovery of the effect of the sodium salts in the leaching solutions. I also described the effect of the cupola furnace gases upon the lime rock with which I had replaced the coke in the absorption tower, which apparently showed the presence of considerable-chlorine, clue to the imperfect conversion of the subchloride to suboxide, as well as that clue to the difficulties of washing thoroughly the suboxide with the means provided, and also to the presence of chlorides in the cement copper which was smelted With the suboxide. The fumes from the cupola, when issuing direct into the atmosphere, were very dense and white, except when a hot furnace top caused a reddish tinge due to dust. These, fumes had a peculiarly suffocating effect, so that when they were blown over into the town the population suffered the greatest actual distress, not the mere annoyance which would have been clue to a disagreeable odor. The absorption tower had been built at the instance of the city authorities to abate this nuisance, and not with the idea of making any additional metal savings. A serious feature in connection with the Argentine experience with the Hunt and Douglas process was, that it had been implicitly assumed that there could be no metal loss in the Hunt and Douglas process itself, and that all attempts to locate the ruinous copper losses which were accumulating in the metallurgical accounts were devoted to the roasting department. It is possibly unfortunate for metallurgical advance along these particular lines that just about the time of this discovery of this previously unsuspected volatilization loss, in the smelting of the suboxide, the Kansas City company made a contract with the Western Union Telegraph Co. for its supply of copper sulphate and the Hunt and Douglas

Jan 11, 1914

By I. H. Warren

Hydrometallurgy has a well-known and established role in the production of manganese metal (1) and manganese compounds, including battery grade manganese dioxide (2). Also, several leaching processes have been developed and applied at various times for the recovery of manganese from low grade ores (3). Recently, interest in the hydrometallurgy of manganese has been stimulated by the problems presented in the processing of sea nodules (4). Fumes and dusts from the ferro-manganese furnaces also offer a potential field for the application of manganese leaching techniques (5). In treating low grade manganese ores a major requirement of the leaching process is that it should dissolve manganese from a mixture of manganese and iron oxides. The most selective reagent for this purpose is ammonium carbamate applied after preliminary pyrometallurgical reduction of the ore (6). Since soluble carbamate complexes of the postulated type (7) are formed by both manganous and ferrous, but by neither manganic nor ferric, ions, selectivety of manganese leaching is achieved by control of the reduction operation to yield, essentially MnO and Fe304? A potential improvement to this two step process would be to simultaneously reduce and selectively leach the manganese in one hydrometallurgical operation. This possibility has been experimentally studied and the results are described below. During the preliminary study of the stability of solutions of ammonium, ferrous and manganous carbamates at elevated temperatures, it was observed, contrary to the observation of Welsh (8) at lower temperatures, that

Jan 1, 1971

By E. Yigit

Hydrometallurgical production of zinc from concentrates is generally carried out by H2SO4 leaching. For the smooth operation of this process, some harmful components must be removed from the leach solution before the electrolysis stage; MgO is one of these compounds. Zinc concentrate samples from the Mount Isa mines were used in leaching tests. Leaching tests with an aqueous solution of H2SO4 and SO2 gas were carried out at various temperatures, pH levels and leaching durations. While duration and the pH value of the solution were effective parameters in H2SO4 leaching, temperature and pH were more effective in SO2 leaching for the removal of MgCO3 and CaCO3. Removal of MgCO3 and CaCO3 was generally achieved at low pH values. The results show that SO2 gas can be a good alternative to sulfuric acid in the leaching of MgCO3 and CaCO3

Jan 1, 2010

By Lawson F

Narrowly sized samples of synthetic chalcocite (Cu2S) and covellite (CuS) were leached in oxygenated acidic sulphate-chloride solutions. Chalcocite leaches via a two-stage process, the first stage being much more rapid than the second. The apparent activation energies are 42.7 kJ/mol and 68.2 kJ/mol respectively, indicating that both stages are controlled by surface chemical reactions. The leaching of covellite is also chemically rate limited with an apparent activation energy of 77.0 kJ/mol. The shrinking core model is able to be used to describe the leaching processes of both copper sulphides. Over 97% of copper present in chalcocite dissolved in 0.5 M H2SO4i0.5 M NaCl solution under one atmosphere pressure of oxygen at 85¦C in three hours from particles 36 pm in diameter while only 85% of the contained copper in covellite particles 13 pm in diameter dissolved in the same solution under the same pressure at 90¦C in six hours. About 99% of the sulphide sulphur in chalcocite was oxidised to elemental sulphur while only 90% of sulphide sulphur was converted to elemental sulphur during covellite leaching. The mechanism proposed for the two-stage leaching of chalcocite particles is as follows: in the first stage, cuprous ions rapidly diffuse to the external reaction surface where they lose electrons and go into solution as a complex cupric chloride ion CuCl'. The second leaching stage begins when covellite is formed at the external surface. Copper again goes into solution as the CuCl* ion. The mechanism for leaching synthetic covellite is similar to the second leaching stage for chalcocite except that the leaching rate of synthetic covellite is substantially lower than that of the newly formed 'covellite'. Elemental sulphur formed on the particle surface retards the dissolution processes during both chalcocite and covellite leaching. The presence of chloride ions is necessary if essentially all of the copper from chalcocite is to go into solution. When no chloride ions were present, the leaching rate of both newly formed and synthetic covellite was almost zero. The main role of chloride ions in both chalcocite and covellite leaching appears to be to disrupt the passivation film of sulphur on the particle surface.

Jan 1, 1989

Zinc sulphide concentrate from Mount lsa, Australia, was leached in acidic sulphate-chloride solutions using ferric ions as oxidants under atmospheric pressure. The parameters of temperature, particle size and concentrations of ferric and chloride ions were evaluated.The shrinking core model can be used to describe the leaching which is limited by the electrochemical processes in the initial leaching stage, and by both surface reaction and pore diffusion processes subsequently. The apparent activation energy measured in the initial leaching stage is 49 kJ/mole over the temperature range 60 to 90°C.About 99 per cent of the contained zinc in the -45+38 µm particle sample dissolved at 90°C in a 0.5 M H2SO4/0.1 M Fe2(SO4)3/0.8 M NaCI solution within two hours. Some nine per cent of sulphide sulphur was oxidised through to sulphate; the remainder was convened to elemental sulphur.The mixed sulphate-chloride system has advantages over separate chloride and sulphate systems due to the low reagent cost, high leaching rate and possible selective leaching of zinc sulphide concentrate with high lead contents.

Jan 1, 1994

By Charles Eichrodt

THE ore that is being treated by the present plant lies between the leached zone, or capping, and the mixed sulfide and oxide zone. The principal copper minerals are chalcanthite (CuSO4.5H20), brochantite (CuS04 3Cu (OH)2) and atacarnite, (CuC12.3Cu (OH)2). There are, in addition, some other minerals such as cuprite and kröhnkite (CuSO4. Na2SO4.2H20). These minerals, with many others, occur in a greatly crushed granodiorite rock. Most of the copper-bearing minerals occur in the cracks and veinlets, but there are disseminated values and it is probably the relative proportion of these that accounts for some otherwise unexplainable variations in extraction. The entire mineralization of the ore is complex and varies with depth. This variation, the problems presented thereby, and the effect of various constituents of the ore will be taken up in detail. An analysis of an unweighted composite of the ore treated during 1927 follows:

Jan 1, 1930

By R. Menge, N. Basin, M. Sibony, J. Lecadet, S. Schmidt

In 1996, Metaleurop replaced its conventional sintering and blast furnace lead smelting activity in Nordenham, Germany by Sirosmelt/Ausmelt technology. Following some alterations to the process and the equipment, the bath smelting plant reached its design capacity for the treatment of lead concentrates, lead sulphate paste from spent batteries, secondaries and internal by-products. A reliable metallurgical process has been established, and thus higher process stability has been achieved, increasing simultaneously the refractory life. Nevertheless, improvements are needed to control the process with a better automation system and reliable expert-systems, because of the intensity of the smelting process. The introduction of the Sirosmelt process for the pyrometallurgical production of lead instead of the prior conventional process has also demonstrated the following advantages: reduced emissions of heavy metals and gaseous compounds, reduction of the specific energy consumption, minimization of residues and a reduction of staff and maintenance costs. The environmental performance of the plant meets the requirements of the German regulations.

Jan 1, 2000

By Wm. E. Milligan

LEAD stocks at the beginning of 1943 were comfortable when compared with those of other base metals such as copper, zinc and tin. This situation was early recognized by W.P.B. and other Governmental agencies and every effort was made to encourage the substitution of lead for these latter metals in all essential uses. Despite this apparently healthy condition, lead was not denied participation in the various measures inaugurated by the Government to relieve metal shortages. Thus in the Mexican trade agreement, effective Jan. 20, the tariff on two important classes of lead importss was halved. Similarly, in the setting up of B quota premiums, a B base was allotted to lead, with all pro¬duction above this figure receiving an additional payment of 2.75c. per lb., or a total payment of 11.75c. per lb. At about the same time it was announced that all quota transactions would be based upon the lead content as shown by the smelter settlement sheets. This basis was 86 percent of the lead content for all ores shipped to custom milling plants and 95 per cent for ores and concentrates shipped to custom smelting plants. The importance of this action can be appreciated when in the Coeur d'Alene it was estimated as equivalent to a 5 per cent increase compared to the previous base.

Jan 1, 1944

By R. A. Perry

DURING 1943, supplies of lead, like those of most base metals, moved from a position of scarcity to one of ample supply for all possible war requirements. The principal worry in the market, as 1944 began, was that surplus stocks might increase too rapidly to maintain price stability. The problem of postwar stockpiling had definitely entered the picture. The end of the year finds the lead industry having met all use demands, but not without depleting stocks to a critical low in spite of heavy importations. The principal reason for this dangerously reduced stock position is lack of adequate man power, which, without question, has become the lead industry's number one problem. This condition became obvious about the middle of the year, after Selective Service placed in effect orders that no deferments would be given employees under 26 years of age, and to a complacent attitude toward the industry assumed by the War Manpower Commission. It has become progressively worse since then. The WPB continued its policy, started during 1943, of easing restrictions on materials to be used for maintenance, repairs, and operating supplies. The policy of premium payments established by the Metals Reserve Co. was continued and, in spite of expressions to the contrary, has proved quite satisfactory. In short, the Government continued its program of stimulating the lead industry in practically all

Jan 1, 1945

By L. D. Anderson

IN STUDYING Utah as a lead producing state one is immediately confronted by the fact that few, if any, of the ores of the state are valued for their lead contents alone. More correctly the ores from which this metal is recovered are characterized as silver lead. Even then the whole story is not told for many of these ores carry also gold, copper, zinc, bismuth, antimony, arsenic, tellurium, and other metals. It becomes apparent, therefore, that the production of lead i11 Utah is influenced by many factors. For example, i11 the early part of 1923 there was a marked increase in the production of the lead by reason of the efforts made by silver mines to produce as much silver as possible before the termination of purchases under the Pittinan Act in June, 1923. As nearly all silver ores carried more or less lead the production of that metal was increased with the increase in silver production. Very few, if any, ores of the State are mined for their lead contents alone. Important as Utah's contribution to the lead output of the country is nevertheless it is bound up to some extent with the production of gold, silver and copper.

Jan 1, 1925

By R. S. Dean

THE first attempt to establish an equilibrium diagram of the lead-antimony series was made by Roland-Gosselin1 in 1896. This investigation classified the system as a purely eutectiferous one, with the eutectic at 12.5 per cent. antimony and 228° C. In the next year, Stead2 placed the eutectic temperature at 247° C. The error in temperature made by Roland-Gosselin, however, has been perpetuated in the literature with a remarkable degree of persistency; this value was accepted by Campbell (1902), and Ewen (1910),3 and will be found in the last edition of Desch's3 Metallography (1922). Gonterman's investigation, made in 1907, was the most complete one available at the time the present investigation was started. He concluded' that the system was a simple eutectiferous one and placed the eutectic at 13 per cent. antimony and 247°C. Other investigations by Tammann (1907),5 Garbow (1909),6 Rudolfi (1910),7 Loebe (1911),8 and Leroux (1913)9 added little, if any-thing, to Gonterman's diagram. A number of investigations have been made of physical properties of the alloys. The electrical conductivity of the series was studied by Matthiessen (1860)10 and by A. W. Smith (1921);" the thermoforce by Rudolfi (1910),12 and Pelabon (1923);13 the specific heat by Durrer

Jan 2, 1926

By J. D. Little

"THE SALMO PROPERTIES of Canadian Exploration, Limited comprise forty-one Crown-granted mineral claims situated on Iron mountain between Sheep creek (on the north) and Lost creek, some eight miles southward from Salmo, in the Nelson mining division (see Figure 1). On these claims there are at present four operating mines. These are, from north to south, the Dodger, Feeney, and Emerald, all tungsten mines, and the Jersey, a lead-zinc mine, about 7,000 feet south of the Dodger. About midway between the Dodger and the Jersey is a fifth mine, the Emerald lead-zinc mine, which was the first producer in the area but has been inactive since 1926. The locations of the mines are shown in Figure 2.Records of prospecting and claim staking on Iron mountain -which owes its name to the presence of large iron-stained outcrops -go back to 1895. The first recorded owner of the claims that form the old Emerald lead-zinc mine was John Waldbeser who, in 1908, did considerable work and produced 426 tons of lead ore valued, at the then price of lead, at $7,000. This ore was chiefly lead carbonate and was taken from the surface outcrops of two bands of ore that form the Emerald orebodies. In 1909, the Emerald was the largest lead producer in the Nelson mining district."

Jan 1, 1953