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Successive extracts from the Whole plant of Eupatorium Odoratum. L (Asteraceae) were investigated for their anthelmintic activity against Pheretima posthuma and three concentrations (10, 50 and 100 mg/ml) of each extracts were studied in activity, which involved the determination of time of paralysis and time of death of the worm. Ethanolic extract exhibited significant anthelmintic activity at highest concentration of 100 mg/ml. Piperazine citrate in 10 mg/ml concentration as that of extract was included as standard reference and 1% Gum acacia in normal saline as control. The anthelmintic activity of ethanloic extract was significant followed by hydroalcoholic extract of Eupatorium Odoratum.Keywords: Anthelmintic activity, Eupatorium Odoratum, Pheretima posithuma, Piperazine citrate.

Studies showed that it is possible to chlorinate a substantial amount of zircon powder using chlorine in the presence of petroleum coke fines at a temperature of about 1 000 degrees C. Nearly 80% of ZrO can be chlorinated at 1 050 degrees C using 25 wt% petroleum coke during a reaction period of only 40 minutes. The extent of chlorination was not affected to any significant extent by the chlorine flow rate above 300 ml/min and the amount of petroleum coke in the charge above 25 wt%. The reaction rate in the temperature range studied was found to follow a topochemical reaction model with an activation energy of 18.3 kcal/mol., Studies showed that it is possible to chlorinate a substantial amount of zircon powder using chlorine in the presence of petroleum coke fines at a temperature of about 1 000 degrees C. Nearly 80% of ZrO can be chlorinated at 1 050 degrees C using 25 wt% petroleum coke during a reaction period of only 40 minutes. The extent of chlorination was not affected to any significant extent by the chlorine flow rate above 300 ml/min and the amount of petroleum coke in the charge above 25 wt%. The reaction rate in the temperature range studied was found to follow a topochemical reaction model with an activation energy of 18.3 kcal/mol.

Smaller deposits of high grade chalcopyrite or the concentrates of low grade and complex copper bearing ores, can be subjected to extraction processes other than smelting and converting. Low temperature chlorination and segregation processes are among those mentioned in the literature. Processes which have been carried out include: extraction of copper from a Brazilian chalcopyrite concentrate by slurry chlorination with chlorine gas at room temperature; dry chlorination by chlorine gas in the temperature range of 350 to 450 degrees C; and segregation roasting of the calcined concentrate in the temperature range of 650 to 800 degrees C. The effect of several experimental variables on the extraction of copper has been discussed for these three cases. Under optimum experimental conditions, 100% copper extraction was possible by either slurry or dry chlorination processes and around 86% Cu by segregation roasting. For all these cases, possible methods have been suggested to test the methods on a pilot scale in order to assess their technoeconomic viability., Smaller deposits of high grade chalcopyrite or the concentrates of low grade and complex copper bearing ores, can be subjected to extraction processes other than smelting and converting. Low temperature chlorination and segregation processes are among those mentioned in the literature. Processes which have been carried out include: extraction of copper from a Brazilian chalcopyrite concentrate by slurry chlorination with chlorine gas at room temperature; dry chlorination by chlorine gas in the temperature range of 350 to 450 degrees C; and segregation roasting of the calcined concentrate in the temperature range of 650 to 800 degrees C. The effect of several experimental variables on the extraction of copper has been discussed for these three cases. Under optimum experimental conditions, 100% copper extraction was possible by either slurry or dry chlorination processes and around 86% Cu by segregation roasting. For all these cases, possible methods have been suggested to test the methods on a pilot scale in order to assess their technoeconomic viability.

Attempts have been made to develop a process based on pressure leaching of copper converter slag with dilute sulphuric acid in which iron contamination can be minimised by oxidation and hydrolysis. Various parameters including leaching time, pulp density, particle size, concentration of acid and oxygen partial pressure were studied to optimise the solubility of metal values. Under optimum conditions about 90% Cu and more than 95% each of Ni and Co could be extracted with only 0.8% extraction of iron., Attempts have been made to develop a process based on pressure leaching of copper converter slag with dilute sulphuric acid in which iron contamination can be minimised by oxidation and hydrolysis. Various parameters including leaching time, pulp density, particle size, concentration of acid and oxygen partial pressure were studied to optimise the solubility of metal values. Under optimum conditions about 90% Cu and more than 95% each of Ni and Co could be extracted with only 0.8% extraction of iron.

The manganese ore was roasted with ammonium sulphate with addition of ammonium chloride and sodium sulphate. Varying proportions of the reagents and different roasting times and temperatures were investigated. About 75% manganese can be recovered using the optimal procedure with fine particle size, followed by water leaching. Roasting pellets rather than powder brought less iron into solution., The manganese ore was roasted with ammonium sulphate with addition of ammonium chloride and sodium sulphate. Varying proportions of the reagents and different roasting times and temperatures were investigated. About 75% manganese can be recovered using the optimal procedure with fine particle size, followed by water leaching. Roasting pellets rather than powder brought less iron into solution.

A two stage chloride leach process is developed. In the first stage most of the lead and zinc are reacted with dilute FeCl solutions at a temperature of 70 degrees C where a major portion of the lead remains in the residue and zinc reports in solution. After taking out lead chloride by brine leaching, the residue is treated for copper dissolution. A concentrated FeCl leaching at boiling temperature as well as ambient aqueous chlorine leaching have been separately tried to dissolve out the copper from the residue., A two stage chloride leach process is developed. In the first stage most of the lead and zinc are reacted with dilute FeCl solutions at a temperature of 70 degrees C where a major portion of the lead remains in the residue and zinc reports in solution. After taking out lead chloride by brine leaching, the residue is treated for copper dissolution. A concentrated FeCl leaching at boiling temperature as well as ambient aqueous chlorine leaching have been separately tried to dissolve out the copper from the residue.

The effects of varying concentration of ferric chloride, duration of leaching, duration of reduction-roasting, temperature and nature of reducing agent on metal recovery were determined. Under constant experimental conditions a decrease in copper recovery, but an increase in nickel and cobalt recovery has been observed above a roasting temperature of 750 degrees C. Under optimum conditions 80% copper, 95% nickal and 80% cobalt could be recovered by reducing the slag at 850 degrees C with 10 wt% furnace oil, followed by leaching with ferric chloride., The effects of varying concentration of ferric chloride, duration of leaching, duration of reduction-roasting, temperature and nature of reducing agent on metal recovery were determined. Under constant experimental conditions a decrease in copper recovery, but an increase in nickel and cobalt recovery has been observed above a roasting temperature of 750 degrees C. Under optimum conditions 80% copper, 95% nickal and 80% cobalt could be recovered by reducing the slag at 850 degrees C with 10 wt% furnace oil, followed by leaching with ferric chloride.

The dissolution rate follows a parabolic equation which fits well with the model of Crank-Ginstling and Brounschtein. The activation energy of the dissolution process is 90 +12.5 kj mol-1. Inward diffusion of the ferric ion through a layer of elemental sulphur which forms on the zinc sulphide surface is suggested as the rate-controlling step., The dissolution rate follows a parabolic equation which fits well with the model of Crank-Ginstling and Brounschtein. The activation energy of the dissolution process is 90 +12.5 kj mol-1. Inward diffusion of the ferric ion through a layer of elemental sulphur which forms on the zinc sulphide surface is suggested as the rate-controlling step.

It was found that fine (25-32 micrometre) powder of pure TiO2, mixed in a 3:2 weight ratio with equally finely powdered pure graphite, can be 88% chlorinated in 75 min., even at the low temperature of 873 K. The rate of chlorination increased with the proportion of graphite, up to approximately 40 wt%. Chlorination, whether in the low (773-873 K) or high (973-1 273 K) temperature range, followed the topochemical reaction model. In the lower temperature range, the rate of chlorination was proportional to Cl2 pressure and in the higher range to its square root; the activation energy in the lower temperature range was 130 kJ/mole and in the higher range only 12 kJ/mole. The rate-determining step at 773-873 K may be the formation of the solid intermediate TiCl2, by the reaction of TiO2 with one molecule of Cl2 in the presence of carbon. At 973-1 273 K, the rate-controlling step may be the reaction of highly mobile TiCl2 with a half-molecule of chlorine to form gaseous TiCl3., It was found that fine (25-32 micrometre) powder of pure TiO2, mixed in a 3:2 weight ratio with equally finely powdered pure graphite, can be 88% chlorinated in 75 min., even at the low temperature of 873 K. The rate of chlorination increased with the proportion of graphite, up to approximately 40 wt%. Chlorination, whether in the low (773-873 K) or high (973-1 273 K) temperature range, followed the topochemical reaction model. In the lower temperature range, the rate of chlorination was proportional to Cl2 pressure and in the higher range to its square root; the activation energy in the lower temperature range was 130 kJ/mole and in the higher range only 12 kJ/mole. The rate-determining step at 773-873 K may be the formation of the solid intermediate TiCl2, by the reaction of TiO2 with one molecule of Cl2 in the presence of carbon. At 973-1 273 K, the rate-controlling step may be the reaction of highly mobile TiCl2 with a half-molecule of chlorine to form gaseous TiCl3.

Metasediments on the periphery of the Gangpur group comprise mainly a psammitic-pelitic assemblage of quartz-mica schists, garnet-staurolite schists with bands pf quartzite, silicified dolomites, calc-silicate rocks and carbonaceous phyllites; these mixed-facies units, cut by basic and granitic intrusions, show greenschist- to amphibolite-facies metamorphism. In the Sargipalli area, base-metal mineralisation is dominated by Pb sulphides with subordinate Cu, Zn and Ag. The mixed-facies lithological sequence at Sargipalli provides favourable host rocks for sulphide minerals and a WNW-ESE-trending shear plane, parallel to the axial plane of the regional synformal fold, could have provided a passage for mobilisation of mineralising fluids, a process possibly facilitated by the late-stage granitic intrusions. Although the Sargipalli deposit is apparently small and isolated, with total resources of 6 540 000 tonnes averaging 5.75% Pb, 0.2-0.4% Cu, 0.3-0.8% Zn and 50 g/t Ag disseminated as minute grains, minor fracture fillings, an intricate pattern of thin stringers at the crests and troughs of flexures, and massive bands alternating with host rocks, it is concluded that similar metamorphic sequences on the periphery of the Gangpur group that have been subjected to E-W shearing with superimposed N-S faults and late-phase granitic invasion should be explored for base metals in greater detail., Metasediments on the periphery of the Gangpur group comprise mainly a psammitic-pelitic assemblage of quartz-mica schists, garnet-staurolite schists with bands pf quartzite, silicified dolomites, calc-silicate rocks and carbonaceous phyllites; these mixed-facies units, cut by basic and granitic intrusions, show greenschist- to amphibolite-facies metamorphism. In the Sargipalli area, base-metal mineralisation is dominated by Pb sulphides with subordinate Cu, Zn and Ag. The mixed-facies lithological sequence at Sargipalli provides favourable host rocks for sulphide minerals and a WNW-ESE-trending shear plane, parallel to the axial plane of the regional synformal fold, could have provided a passage for mobilisation of mineralising fluids, a process possibly facilitated by the late-stage granitic intrusions. Although the Sargipalli deposit is apparently small and isolated, with total resources of 6 540 000 tonnes averaging 5.75% Pb, 0.2-0.4% Cu, 0.3-0.8% Zn and 50 g/t Ag disseminated as minute grains, minor fracture fillings, an intricate pattern of thin stringers at the crests and troughs of flexures, and massive bands alternating with host rocks, it is concluded that similar metamorphic sequences on the periphery of the Gangpur group that have been subjected to E-W shearing with superimposed N-S faults and late-phase granitic invasion should be explored for base metals in greater detail.

The recovery of nickel and chromium from the alloy nichrome, which is extensively used in the metallurgical industry in heating elements, was investigated. Ni and Cr in solutions of dissolved nichrome were analysed using atomic adsorption spectrophotometry. Concentrations of HCl ranged from 0.5 to 6.0 M, mixtures of 1.0 M HCl and 1.0-3.5 M NaCl were also used and the effect of temperature (40-100 degrees C) on dissolution in 6.0 M HCl for different periods was studied. The results were used to elucidate the process mechanism by applying Langmuir's adsorption isotherm. It was confirmed that both H+ and Cl- ions are responsible for the dissolution. In simple HCl, H+ is rate-controlling while in mixtures of HCl and excess NaCl Cl- is responsible for determining the dissolution rate. In low HCl (0.04-0.22 M) and high NaCl concentrations (1.0-4.0) the observed constancy in the rate of reaction was attributed to saturation of the active sites on the nichrome surface by chloride ion., The recovery of nickel and chromium from the alloy nichrome, which is extensively used in the metallurgical industry in heating elements, was investigated. Ni and Cr in solutions of dissolved nichrome were analysed using atomic adsorption spectrophotometry. Concentrations of HCl ranged from 0.5 to 6.0 M, mixtures of 1.0 M HCl and 1.0-3.5 M NaCl were also used and the effect of temperature (40-100 degrees C) on dissolution in 6.0 M HCl for different periods was studied. The results were used to elucidate the process mechanism by applying Langmuir's adsorption isotherm. It was confirmed that both H+ and Cl- ions are responsible for the dissolution. In simple HCl, H+ is rate-controlling while in mixtures of HCl and excess NaCl Cl- is responsible for determining the dissolution rate. In low HCl (0.04-0.22 M) and high NaCl concentrations (1.0-4.0) the observed constancy in the rate of reaction was attributed to saturation of the active sites on the nichrome surface by chloride ion.

Although some work has been reported in the literature on the extraction of copper from different copper-bearing sources by the segregation roasting route, studies relating to the kinetics and the mechanism have been rather limited. There is no universally accepted rate equation for the segregation reaction. An attempt has been made to understand the mechanism of the copper segregation process by means of an isothermal kinetic study based on the evolved gas analysis technique. In addition, empirical relationships were established to express the percentage recovery of copper as a function of process parameters determined through the application of statistically designed experimental procedure., Although some work has been reported in the literature on the extraction of copper from different copper-bearing sources by the segregation roasting route, studies relating to the kinetics and the mechanism have been rather limited. There is no universally accepted rate equation for the segregation reaction. An attempt has been made to understand the mechanism of the copper segregation process by means of an isothermal kinetic study based on the evolved gas analysis technique. In addition, empirical relationships were established to express the percentage recovery of copper as a function of process parameters determined through the application of statistically designed experimental procedure.

Low-grade complex sulphide ores from the Jin Chuan mine, containing pentlandite, chalcopyrite, pyrrhotite, serpentine and related gangue minerals, was upgraded by grinding and flotation to an intimate mixture of the three main minerals containing Cu, Ni, Fe and an appreciable amount of magnesium oxide. After calcining at 850 and 900 degrees C to ensure complete conversion of sulphides to oxides, the ground concentrate was roasted with coke for 1 hr at 750 degrees C. The coarse, metal-coated carbon particles were then screened out and the fines subjected to further segregation roasting at 900-1 100 degrees C. The metals from the coarse fraction were dissolved in HCl and analysed. First-stage roasting gave maximum Cu extraction. Ni and Fe yields increased with temperature, though at 1 100 degrees the product tended to sinter, making separation difficult. Overall metal recovery of 91% Cu and 76% Ni was achieved, though about 62% of the contained iron impurity was also extracted. The optimum charge concentrations for 5 g calcined concentrate were 1 g coke particles, 1 g sodium chloride, 0.7 g silica and 0.25 cm3 water., Low-grade complex sulphide ores from the Jin Chuan mine, containing pentlandite, chalcopyrite, pyrrhotite, serpentine and related gangue minerals, was upgraded by grinding and flotation to an intimate mixture of the three main minerals containing Cu, Ni, Fe and an appreciable amount of magnesium oxide. After calcining at 850 and 900 degrees C to ensure complete conversion of sulphides to oxides, the ground concentrate was roasted with coke for 1 hr at 750 degrees C. The coarse, metal-coated carbon particles were then screened out and the fines subjected to further segregation roasting at 900-1 100 degrees C. The metals from the coarse fraction were dissolved in HCl and analysed. First-stage roasting gave maximum Cu extraction. Ni and Fe yields increased with temperature, though at 1 100 degrees the product tended to sinter, making separation difficult. Overall metal recovery of 91% Cu and 76% Ni was achieved, though about 62% of the contained iron impurity was also extracted. The optimum charge concentrations for 5 g calcined concentrate were 1 g coke particles, 1 g sodium chloride, 0.7 g silica and 0.25 cm3 water.

Complete extraction of copper from Salobo chalcopyrite concentrate can be achieved in less than 10 minutes by chlorination of the dry concentrate with Cl2 at 450 degrees C and in 5 minutes by chlorination of an aqueous slurry containing 5 wt% concentrate at 30 degrees C. In general, slurry chlorination appears to be faster than dry chlorination. The kinetics of dry chlorination of copper sulphide are controlled by an adsorption mechanism with an activation energy of 6.1 kcal/mole. The rate is independent of chlorine partial pressure when pCl2 exceeds 0.4 atm. In slurry chlorination the dissolution of copper sulphide follows a topochemical reaction model and the rate of chlorination decreases with increase in pulp density and also temperature. Dissolved chlorine appears to take part in the reaction in atomic or ionic form., Complete extraction of copper from Salobo chalcopyrite concentrate can be achieved in less than 10 minutes by chlorination of the dry concentrate with Cl2 at 450 degrees C and in 5 minutes by chlorination of an aqueous slurry containing 5 wt% concentrate at 30 degrees C. In general, slurry chlorination appears to be faster than dry chlorination. The kinetics of dry chlorination of copper sulphide are controlled by an adsorption mechanism with an activation energy of 6.1 kcal/mole. The rate is independent of chlorine partial pressure when pCl2 exceeds 0.4 atm. In slurry chlorination the dissolution of copper sulphide follows a topochemical reaction model and the rate of chlorination decreases with increase in pulp density and also temperature. Dissolved chlorine appears to take part in the reaction in atomic or ionic form.

The reagent, manufactured by Hoechst, has an empirical formula of C16H22O2, a molecular weight of 246 and includes a beta-diketone group. It was diluted to 10 vol% with kerosene. Aqueous solutions of 0.005-0.05 M zinc sulphate and 0.5 M ammonium sulphate were used. Zinc extraction efficiency was found to increase with pH to a maximum at 7.5. The decrease at higher pH was attributed to the formation of a non-extractable zinc ammine complex. Extraction was exothermic, decreasing as temperature rose, while stripping was endothermic. Efficiency of stripping also increased with sulphuric acid concentration up to 0.08 M, beyond which there was no further improvement. Hostarex DK16 was seen to have better loading capacity than DEHPA and LIX 34, but its extraction efficiency was low compared with LIX 34., The reagent, manufactured by Hoechst, has an empirical formula of C16H22O2, a molecular weight of 246 and includes a beta-diketone group. It was diluted to 10 vol% with kerosene. Aqueous solutions of 0.005-0.05 M zinc sulphate and 0.5 M ammonium sulphate were used. Zinc extraction efficiency was found to increase with pH to a maximum at 7.5. The decrease at higher pH was attributed to the formation of a non-extractable zinc ammine complex. Extraction was exothermic, decreasing as temperature rose, while stripping was endothermic. Efficiency of stripping also increased with sulphuric acid concentration up to 0.08 M, beyond which there was no further improvement. Hostarex DK16 was seen to have better loading capacity than DEHPA and LIX 34, but its extraction efficiency was low compared with LIX 34.

Carbon tetrachloride was found to be a very efficient chlorinating agent for anatase concentrate, even at the low temperatures that were investigated (250-400 degrees C). For example, at 400 degrees C about 90% of the anatase concentrate was found to have been chlorinated within 60 minutes., Carbon tetrachloride was found to be a very efficient chlorinating agent for anatase concentrate, even at the low temperatures that were investigated (250-400 degrees C). For example, at 400 degrees C about 90% of the anatase concentrate was found to have been chlorinated within 60 minutes.

Cu-Ni-Co matte obtained by pyritic smelting of Cu converter slag was leached at elevated pressures. The leaching studies were carried out using a factorial design for optimisation. Regression equations for the extraction of Cu, Ni and Co were determined, taking temperature, oxygen partial pressure and acid concentration as the variables. The regression equations (except for cobalt extraction at two hours) satisfactorily predicted the extractions for another factorial matrix chosen with narrow intervals of the three leaching variables., Cu-Ni-Co matte obtained by pyritic smelting of Cu converter slag was leached at elevated pressures. The leaching studies were carried out using a factorial design for optimisation. Regression equations for the extraction of Cu, Ni and Co were determined, taking temperature, oxygen partial pressure and acid concentration as the variables. The regression equations (except for cobalt extraction at two hours) satisfactorily predicted the extractions for another factorial matrix chosen with narrow intervals of the three leaching variables.

It was observed that Zn can be extracted using Hostarex DK-16 as solvent. The extraction is influenced by the pH and ammonium sulphate concentration of the aqueous phase. The decrease in Zn extraction with increasing concentration of ammonium salt is attributed to the formation of zinc ammine complexes, thereby decreasing the free metal concentration in the aqueous phase. Identical nature of the infra-red absorption spectra of the extracts obtained from sodium sulphate and ammonium sulphate media substantiated that only free metal ion is extractable by Hostarex DK-16. The composition of the extracted species has been found by slope analysis to be ZnR2., It was observed that Zn can be extracted using Hostarex DK-16 as solvent. The extraction is influenced by the pH and ammonium sulphate concentration of the aqueous phase. The decrease in Zn extraction with increasing concentration of ammonium salt is attributed to the formation of zinc ammine complexes, thereby decreasing the free metal concentration in the aqueous phase. Identical nature of the infra-red absorption spectra of the extracts obtained from sodium sulphate and ammonium sulphate media substantiated that only free metal ion is extractable by Hostarex DK-16. The composition of the extracted species has been found by slope analysis to be ZnR2.

By leaching the nodules with 8% sulphuric acid for 12 h at 100 degrees C, copper and nickel could be completely extracted along with 50% cobalt and 6.2% manganese. 94.2% of the iron was also dissolved. To increase manganese and cobalt extraction, charcoal was used as a reductant. The role of charcoal was to reduce the manganese dioxide and render it soluble as manganese sulphate as well as to facilitate solution of the cobalt minerals associated with the manganese dioxide. It was found that using 5 wt./wt.% charcoal (0.05 g of charcoal per g of nodule) almost complete extraction of copper, nickel, cobalt and manganese could be achieved. However, under these conditions all the iron present in the nodules was also leached., By leaching the nodules with 8% sulphuric acid for 12 h at 100 degrees C, copper and nickel could be completely extracted along with 50% cobalt and 6.2% manganese. 94.2% of the iron was also dissolved. To increase manganese and cobalt extraction, charcoal was used as a reductant. The role of charcoal was to reduce the manganese dioxide and render it soluble as manganese sulphate as well as to facilitate solution of the cobalt minerals associated with the manganese dioxide. It was found that using 5 wt./wt.% charcoal (0.05 g of charcoal per g of nodule) almost complete extraction of copper, nickel, cobalt and manganese could be achieved. However, under these conditions all the iron present in the nodules was also leached.

In the presence of any reductant, the maximum extractions of Cu, Ni and Co are obtained in between pH 9.0 and 10.0. Metal dissolution is sensitive to the leaching temperature. Apparent activation energies indicate the leaching reactions to be chemically controlled. The reactivity of the reductants is FeSO4 greater than MnSO4 greater than (NH4)2SO3 greater than FeS. Dissolution of Cu and Ni is faster in chloride media. Co extraction is better in carbonate medium, except when FeSO4 is used as the reductant. About 50% cobalt could be leached with low manganese extraction. Complete dissolution of Co was possible with about 35% Mn extraction., In the presence of any reductant, the maximum extractions of Cu, Ni and Co are obtained in between pH 9.0 and 10.0. Metal dissolution is sensitive to the leaching temperature. Apparent activation energies indicate the leaching reactions to be chemically controlled. The reactivity of the reductants is FeSO4 greater than MnSO4 greater than (NH4)2SO3 greater than FeS. Dissolution of Cu and Ni is faster in chloride media. Co extraction is better in carbonate medium, except when FeSO4 is used as the reductant. About 50% cobalt could be leached with low manganese extraction. Complete dissolution of Co was possible with about 35% Mn extraction.

A compact lead deposit with about 90% current efficiency can be achieved when a bath containing lead 30 g/l, glue 2g/l and beta-naphthol 0.1 g/l is used and electrolysis is carried out at pH 0.75-0.95, temperature 70 degrees C and a current density of 50 A/m2. Agitation of the electrolyte increased the current efficiency by about 5%, with improvement in the nature of the deposit., A compact lead deposit with about 90% current efficiency can be achieved when a bath containing lead 30 g/l, glue 2g/l and beta-naphthol 0.1 g/l is used and electrolysis is carried out at pH 0.75-0.95, temperature 70 degrees C and a current density of 50 A/m2. Agitation of the electrolyte increased the current efficiency by about 5%, with improvement in the nature of the deposit.

From the observations made by varying initial pH and total molarity of the leachant it can be concluded that both pH and concentration of the leachant greatly affect the dissolution rates of copper, nickel, cobalt and manganese., From the observations made by varying initial pH and total molarity of the leachant it can be concluded that both pH and concentration of the leachant greatly affect the dissolution rates of copper, nickel, cobalt and manganese.

Leaching of manganese nodules at elevated temperatures and pressure can be adopted for three-metal (Cu, Ni, Co) as well as for four-metal extraction (Co, Ni, Co and Mn). With the addition of charcoal to the feed solution and by using higher acid levels, the conditions employed for three-metals can be extended to four-metal extraction in the same autoclave. By leaching the nodules for 4h at 423K using 0.46g of H2SO4 per g of nodule and by keeping the oxygen partial pressure at 0.55 MPa, almost all the copper, nickel and about 88% of the cobalt can be extracted. To achieve four-metal extraction, the acid concentration was increased to 0.66 g/g of nodule and 0.05g of charcoal/g of nodule was added as a reductant. The extraction of copper, nickel, cobalt, manganese and iron were 77, 99.8, 88, 99.8 and 4.5 respectively., Leaching of manganese nodules at elevated temperatures and pressure can be adopted for three-metal (Cu, Ni, Co) as well as for four-metal extraction (Co, Ni, Co and Mn). With the addition of charcoal to the feed solution and by using higher acid levels, the conditions employed for three-metals can be extended to four-metal extraction in the same autoclave. By leaching the nodules for 4h at 423K using 0.46g of H2SO4 per g of nodule and by keeping the oxygen partial pressure at 0.55 MPa, almost all the copper, nickel and about 88% of the cobalt can be extracted. To achieve four-metal extraction, the acid concentration was increased to 0.66 g/g of nodule and 0.05g of charcoal/g of nodule was added as a reductant. The extraction of copper, nickel, cobalt, manganese and iron were 77, 99.8, 88, 99.8 and 4.5 respectively.

The leaching of copper, nickel and cobalt from copper converter slags (from the Ghatsila plant of Hindustan Copper) by dilute sulphuric acid at elevated pressure and temperature was investigated. While copper was readily leached, nickel and cobalt in the magnetite phase were not leachable. Slow cooling of the slag (e.g. maintaining it at around the melting point of fayalite for about 6 hours) resulting in the transfer of nickel and cobalt from the magnetite phase to the rest of the slag matrix and about 90% of Ni and Co could be leached from such slags., The leaching of copper, nickel and cobalt from copper converter slags (from the Ghatsila plant of Hindustan Copper) by dilute sulphuric acid at elevated pressure and temperature was investigated. While copper was readily leached, nickel and cobalt in the magnetite phase were not leachable. Slow cooling of the slag (e.g. maintaining it at around the melting point of fayalite for about 6 hours) resulting in the transfer of nickel and cobalt from the magnetite phase to the rest of the slag matrix and about 90% of Ni and Co could be leached from such slags.

The results of an investigation carried out on leaching of manganese nodules in dilute HCl in the presence of various reductants, such as pyrite, sodium sulphite and carbon, so that Mn and Co can be extracted along with Cu, Ni and Zn without any appreciable dissolution of Fe, are described., The results of an investigation carried out on leaching of manganese nodules in dilute HCl in the presence of various reductants, such as pyrite, sodium sulphite and carbon, so that Mn and Co can be extracted along with Cu, Ni and Zn without any appreciable dissolution of Fe, are described.

The leaching behaviour of complex sulphide concentrates of lead, zinc and copper in iron III chloride media is predicted. Conditions for the preferential dissolution of one or two minerals and complete dissolution of all three minerals from three complex sulphide concentrates of different composition and origin are predicted and subjected to experimental verification., The leaching behaviour of complex sulphide concentrates of lead, zinc and copper in iron III chloride media is predicted. Conditions for the preferential dissolution of one or two minerals and complete dissolution of all three minerals from three complex sulphide concentrates of different composition and origin are predicted and subjected to experimental verification.

High degrees of sulphation were achieved at a reaction temperature of 400 degrees C. The effects of various roasting parameters, such as temperature, reaction time, particle size, moisture content and ratio of sulphur dioxide to air, were investigated. The sulphated nodules were leached with water to take soluble sulphates into solution. With the use of a 1:1 mixture of sulphur dioxide and air at a temperature of 400 degrees C and a reaction time of 60 min. it was possible to extract 82% Cu, 95% Ni, 99% Co and 82% Mn with little contamination by iron (1.2%)., High degrees of sulphation were achieved at a reaction temperature of 400 degrees C. The effects of various roasting parameters, such as temperature, reaction time, particle size, moisture content and ratio of sulphur dioxide to air, were investigated. The sulphated nodules were leached with water to take soluble sulphates into solution. With the use of a 1:1 mixture of sulphur dioxide and air at a temperature of 400 degrees C and a reaction time of 60 min. it was possible to extract 82% Cu, 95% Ni, 99% Co and 82% Mn with little contamination by iron (1.2%).

The reaction of galena, sphalerite and chalcopyrite with iron III chloride solution each follow the simple shrinking core model. In each system the rate is directly proportional to the square root of iron III chloride concentration and inversely proportional to the mineral particle diameter. The dissolution reactions are electrochemical in nature., The reaction of galena, sphalerite and chalcopyrite with iron III chloride solution each follow the simple shrinking core model. In each system the rate is directly proportional to the square root of iron III chloride concentration and inversely proportional to the mineral particle diameter. The dissolution reactions are electrochemical in nature.

It is observed that while more than 80% of Ni, Cu and Zn can be easily leached along with 30-35% of Fe in 1.0-1.5M HCl, extractions of Mn and Co are very low, less than 20%. By increasing both temperature of leaching and acid concentration the extractions of Mn and Co can be improved, but only at the cost of considerable dissolution of iron. Kinetic studies reveal that Mn and Co follow a parabolic rate law of dissolution from constant surface area, while Fe and Ni follow a rate law for diffusion through a non-porous product layer, and for Cu and Zn radom nucleation and growth of product are the rate controlling steps., It is observed that while more than 80% of Ni, Cu and Zn can be easily leached along with 30-35% of Fe in 1.0-1.5M HCl, extractions of Mn and Co are very low, less than 20%. By increasing both temperature of leaching and acid concentration the extractions of Mn and Co can be improved, but only at the cost of considerable dissolution of iron. Kinetic studies reveal that Mn and Co follow a parabolic rate law of dissolution from constant surface area, while Fe and Ni follow a rate law for diffusion through a non-porous product layer, and for Cu and Zn radom nucleation and growth of product are the rate controlling steps.

The various parameters chosen for leaching studies were: amount of glucose, time, pH, temperature, concentrations of ammonia and ammonium salt, and particle size. The percentage extraction of copper, nickel and cobalt decreases when the glucose concentration is increased to more than 0.4 gramme per gramme of nodule at 338 K and 0.2 g/g at 358 K. The dissolution of copper has been found to be dependent more on the pH of the solution than on the total molarity of ammonia and the ammonium ion. Leaching temperatures above 373 K are disadvantageous for the extraction of copper and cobalt, whereas nickel extraction is about 95% even at 433 K. The rate of leaching for all three metals is independent of the particle size in the range studied. All the copper, about 90% of the nickel and 60% of the cobalt can be extracted under the following leaching conditions: temperature 358 K, time 4 h, initial ammonia concentration 2.5 M, ammonium chloride concentration 0.37 M and glucose 0.2 gramme per gramme of nodule., The various parameters chosen for leaching studies were: amount of glucose, time, pH, temperature, concentrations of ammonia and ammonium salt, and particle size. The percentage extraction of copper, nickel and cobalt decreases when the glucose concentration is increased to more than 0.4 gramme per gramme of nodule at 338 K and 0.2 g/g at 358 K. The dissolution of copper has been found to be dependent more on the pH of the solution than on the total molarity of ammonia and the ammonium ion. Leaching temperatures above 373 K are disadvantageous for the extraction of copper and cobalt, whereas nickel extraction is about 95% even at 433 K. The rate of leaching for all three metals is independent of the particle size in the range studied. All the copper, about 90% of the nickel and 60% of the cobalt can be extracted under the following leaching conditions: temperature 358 K, time 4 h, initial ammonia concentration 2.5 M, ammonium chloride concentration 0.37 M and glucose 0.2 gramme per gramme of nodule.

Roasting of copper converter slag containing 4.03% copper, 1.98% nickel and 0.48% cobalt with ammonium sulphate open to the atmosphere has been carried out in order to achieve sulphation of copper, nickel and cobalt followed by leaching of the metal values as soluble sulphates with water. The effect of temperature, time and amount of ammonium sulphate has been studied. Under atmospheric conditions, using 2.5 times the stoichiometric requirement of ammonium sulphate, the recoveries of copper, nickel and cobalt were found to be 85%, 81% and 85% respectively. Similar studies were carried out with sulphuric acid. At 150 degrees C and a roasting time of 60 minutes with the stoichiometric amount of sulphuric acid, recoveries of copper, nickel and cobalt were 95%, 90% and 99% respectively, along with a contamination of 60-80% iron. Removal of most of the iron from leach liquors has been effected with ammonia liquor and lime as precipitants., Roasting of copper converter slag containing 4.03% copper, 1.98% nickel and 0.48% cobalt with ammonium sulphate open to the atmosphere has been carried out in order to achieve sulphation of copper, nickel and cobalt followed by leaching of the metal values as soluble sulphates with water. The effect of temperature, time and amount of ammonium sulphate has been studied. Under atmospheric conditions, using 2.5 times the stoichiometric requirement of ammonium sulphate, the recoveries of copper, nickel and cobalt were found to be 85%, 81% and 85% respectively. Similar studies were carried out with sulphuric acid. At 150 degrees C and a roasting time of 60 minutes with the stoichiometric amount of sulphuric acid, recoveries of copper, nickel and cobalt were 95%, 90% and 99% respectively, along with a contamination of 60-80% iron. Removal of most of the iron from leach liquors has been effected with ammonia liquor and lime as precipitants.

Copper converter slag that contained 2-4% total copper was used in an investigation on recovery directly in the form of metal. The slag was preroasted and then segregation roasted in the presence of a carbonaceous material and an alkali or alkaline earth chloride. Finally the result was screened to recover the segregated material. The effect of roasting time, temperature, percentage and type of carbonaceous material and type of chloride was studied. It was found possible to obtain more than 90% of the copper metal by treating the preroasted slag (-200 mesh BSS) with 3 wt% carbon plus 1 wt% sodium chloride at 1023 K for 60 min. The role of moisture and hydrogen in the reduction of copper chloride to metallic copper was verified thermodynamically. More than 90% of the copper can be extracted by sieving the segregated mass at + 44 mesh, resulting in an overall recovery of 80% of the total copper., Copper converter slag that contained 2-4% total copper was used in an investigation on recovery directly in the form of metal. The slag was preroasted and then segregation roasted in the presence of a carbonaceous material and an alkali or alkaline earth chloride. Finally the result was screened to recover the segregated material. The effect of roasting time, temperature, percentage and type of carbonaceous material and type of chloride was studied. It was found possible to obtain more than 90% of the copper metal by treating the preroasted slag (-200 mesh BSS) with 3 wt% carbon plus 1 wt% sodium chloride at 1023 K for 60 min. The role of moisture and hydrogen in the reduction of copper chloride to metallic copper was verified thermodynamically. More than 90% of the copper can be extracted by sieving the segregated mass at + 44 mesh, resulting in an overall recovery of 80% of the total copper.