WO2024231946A1

WO2024231946A1 - An improved hydrometallurgy method

Info

Publication number: WO2024231946A1
Application number: PCT/IN2024/050480
Authority: WO
Inventors: Vinod Chintamani Malshe; Divya Madhukar MHATRE
Original assignee: Vinod Chintamani Malshe
Priority date: 2023-05-06
Filing date: 2024-05-04
Publication date: 2024-11-14

Abstract

An improved hydrometallurgy method, particularly a modified acid leaching step; wherein the acid leaching step comprises subjecting zinc calcine obtained from roasting step to acid leaching by using dilute sulfuric acid characterized in that (a) filtering the reaction mixture to obtain an acid insoluble residue; and (b) subjecting the acid insoluble residue of step (a) to acid leaching by treating it with 30 to 60 % of sulfuric acid in presence of reducing agent at temperature in range of 10°C to 100°C followed by filtering the reaction mixture to obtain an acid insoluble residue enriched with silver to at least 2000 ppm. The method further comprises step (c) of alkali digestion and step (d) to enriched silver further. Jarosite may be obtained by precipitation from the filtrate of step (b) with sodium or potassium sulphate that is free from heavy metal and safe to dispose of.

Description

TITLE OF THE INVENTION:

An improved hydrometallurgy method

This application claims priority from Indian Patent Application No. 202321032230 filed on 6^th May 2023.

TECHNICAL FIELD OF THE INVENTION:

The present invention relates in general to an improved hydrometallurgical method, particularly modifying leaching step of hydrometallurgy method.

Particularly, the present invention relates to the effective and efficient leaching step of hydrometallurgy method, which efficiently recovers valuable metals and leads to formation of Jarosite without any residual heavy metal in it, thereby making Jarosite safe to dispose of even in agricultural land or to be used as a pigment in interior and exterior paints, as a pigment for industrial solvent based paints, as a filler in plastics, as a raw material for preparation of high purity iron oxide or as an additive to cement for replacement of gypsum.

Particularly, the present invention enriches Silver (Ag) concentration to at least 2000 ppm in the insoluble residue besides being simple.

BACKGROUND OF THE INVENTION:

The hydrometallurgy method for production of Zinc from complex zinc ore, particularly, froth flotation, Roasting-leaching -purification-electrowinning (RLE) zinc hydrometallurgy technology is the conventional technology. This method comprises five steps:

1. Roasting: Zinc, lead, cadmium, barium, and iron are the main content of zinc concentrate. Due to formation of malodorous, corrosive and poisonous hydrogen sulfide in acidic pH, it cannot be directly used as the feed of the leaching process. Therefore, as the first step of the RLE process, the task of roasting is to convert sulfide concentrate to calcine. Calcine is mainly composed of oxides of zinc, iron, lead, cadmium, barium and calcium. Other elements present as oxides are silica and alumina. Traces of silver and gold are also found, silver being 70 to 100 ppm. Calcine and roasting dust are collected and fed to the subsequent leaching process. Leaching: In the leaching process, dilute sulfuric acid or spent acid from the electro winning process is used as solvent to dissolve zinc in calcine, such that the zinc ions are liberated and those enter the solution. There are various types of leaching technologies depending on the reaction conditions and number of stages. Each stage is composed of several leaching reactors. Calcine is first leached in acidic solution (i.e. dilute sulfuric acid) in order to leach zinc out of zinc oxide. The remaining calcine is then leached in strong sulfuric acid to leach the rest of the zinc out of zinc oxide and zinc ferrite. A good quantity of iron which was present in ferrous form also dissolves forming ferrous sulphate. Iron which has been oxidized to ferric form in the calciner at high temperature does not readily dissolve. The result of this process is a solid and a liquid; the liquid contains zinc and ferrous iron and is often called leach product; the solid is called leach residue and contains precious metals (usually lead and silver). Some manufacturers remove the acid insoluble part and sell it as a lean source of silver. There is also iron in the leach product from the strong acid leach, which is removed in an intermediate step, in the form of Jarosite. Newer processes use air oxidation to remove ferrous iron from solution by air oxidation and precipitating it as Goethite and Haematite. There is still cadmium, copper, arsenic, antimony, cobalt, germanium, nickel, and thallium in the leach product. Therefore, it needs to be purified. Solution purification: As zinc concentrate is not pure, zinc calcine obtained by roasting contains not only ZnO but also compounds of other elements, e.g., Fe, Si, and As. In the leaching process, besides the extraction of zinc ions, these impurity ions are also liberated and enter the leaching solution. Part of these impurities are deposited in the leaching process, e.g., Fe and Si. The precious metal residue is collected for further refining. The remaining impurities in the leaching solution, which have to be processed using specific technology, are removed in the subsequent purification process. After leaching and purification, the zinc sulphate solution is of high purity. Electro winning: The zinc electrowinning process consists of several series of electrolytic cells which extract pure metal zinc from electrolyte (zinc sulphate solution). By running a direct current through the cell, electrons are transferred between the anode and the cathode. A hydrolysis reaction takes place on the anode; OH- loses electrons and releases oxygen, while zinc ions take up electrons and are gradually deposited on the surface of the cathode. The deposited zinc on the plate is stripped periodically, and 5. Casting: Zinc obtained in electrowinning is cast as ingot or other zinc products in the casting process.

First, zinc concentrate is transformed to zinc oxide in the roasting process. Then, in the leaching process, the solid oxidized concentrate ores are treated in a sulfuric acid solution to liberate the valuable metal ions from concentrated ores. Due to the impurity and heterogeneity of the concentrate ores, the other associated metallic ions in the concentrate ores are simultaneously dissolved into the acid solution. Therefore, the resulting leaching solution contains ions of impurity metals harmful to the electrowinning process in which the pure valuable metal is recovered by electrodeposition. The recovered zinc metal is then casted into final zinc products with different specifications, e.g., zinc ingot.

Jarosite is a basic hydrous sulphate of potassium and ferric iron (Fe-III) with a chemical formula of KFe3(SO4)2,(OH)6 or NaFe3(SO4)2(OH)6. It is often produced as a byproduct during the extraction of zinc from ore as well as purification and refining of zinc. This byproduct is a complex compound of iron and contains zinc along with traces of other metals. Several manufacturers do not filter the acid insoluble portion of the leaching process and allow this to become a part of Jarosite. Due to toxic ingredients like lead, zinc, nickel, manganese, cobalt, copper, cadmium, etc., Jarosite is universally considered a hazardous waste. It is not allowed to be used as a land fill. It needs to be stocked above ground. All rain water falling on the heaps needs to be collected and purified before discharge. Several manufacturers have stocked up millions of tons over last 50 years and are left with no space to store further. Recently the government has allowed addition of 2% Jarosite to cement. The cement industry is not very enthused to use it because Jarosite contains 30% moisture and drying is an expensive process. About fifty percent of zinc is present in the ore concentrate, which is roasted at 900°C temperature. After that, the leaching process is carried out where Jarosite gets produced as waste material. Conventionally, Jarosite is dumped in various landfills, lined impermeable ponds and the like. There have been several efforts to “dispose” of this waste. Typical analysis of Jarosite is given in the Table 1 below.

Table 1: Typical Analysis of Jarosite

India manufactures about 7.9 lakh tons of zinc metal at one site (Chandaria in Rajasthan). It produces about 500 tons of residues from leaching which poses environmental hazard. The plant is operating for last 20 years and has stocked up about 8 million tons of the waste in a solidified form.

There have been several efforts to “dispose” of this waste. Several agencies employed for safe disposal have studied the composition in detail and have recommended methods of safe disposal of this material. Most recommendations from reputed institutions including IIT Chennai, NEERI Nagpur includes the use of Jarosite in cement manufacturing or direct addition to last stage of cement manufacturing along with mineral gypsum. Journal of Hazardous Materials, Volume 137, Issue 3, 11 October 2006, Pages 1589-1599 titled "Hazardous Jarosite use in developing non- hazardous product for engineering application" discloses recycling the hazardous Jarosite released from zinc industries in developing non-hazardous products which can ultimately be used in building applications. Other environmental studies have recommended to safely dump it underground and capping it. (Proceedings of Indian geotechnical conference 15-17-2011, Kochi (Paper No. L-306)).

In several locations of Hindustan Zinc, it is stocked over ground taking care of prevention of leachate from entering the ground water due to potential contamination with zinc, lead and other harmful elements . (https ://www.hzlindia.com/ sustainability/environment-management/waste management/)

Presently Jarosite is treated with lime and cement to minimize flying of Jarosite and spreading around, leaching of heavy metals and the treated waste is called “Jarofix”. Jarofix is stocked in high density polyethylene (HDPE) lined disposal yards. The utilization of Jarofix is explored as a sub-grade and embankment material for road making during widening of State Highway near Chittorgarh (Rajasthan). However, the annual production of Jarofix is about 5 lakh metric tons while the unutilized accumulated Jarofix was about 15 lakh metric tons at Hindustan Zinc Limited, Chittorgarh, Rajasthan [ Sinha, A. K., et al. "Characterization of Jarofix waste material for the construction of road." (2013) and Sinha, A. K., et al. "Design, Construction & Evaluation of Jarofix Embankment and Sub Grade Layers." (2012).]. It must have increased by at least a million tons in last ten years. The material is occupying costly agricultural lands and has become an environmental hazard. Contamination of heavy metals in the environment is of major concern because of their toxicity and threat to human life and environment (Er. Nitisha Rathore et al., IJCIET, Volume 5, Issue 11, November (2014), pp. 192-200).

As can be seen from Table 1, a significant quantity of zinc remains un-extracted and it has become an environmental threat and a huge disposal problem. The world production of zinc is about 15 million tons/ year. Each ton produces about 0.16 tons of Jarosite waste. This represents a significant loss of valuable metals like Zinc, Lead and Silver.

All studies reported so far and the practices followed by the manufacturers all over the world are following methods which suitably “dispose” the waste safely.

There have been no attempts made in the existing art till date to modify acid leaching step of the hydrometallurgy method to recover effectively and efficiently metals like silver, lead, cadmium, copper, nickel, manganese and cobalt which have commercial value and importance and lead to formation of Jarosite without any residual heavy metals.

Due to increase in environmental concerns because of improper management of both hazardous and non-hazardous wastes released from different industrial processes prioritized the necessity for the innovation research further.

Thus, an attempt has been made by the applicants to improve hydrometallurgy method by modifying the acid leaching step to effectively and efficiently recover valuable metals before formation of Jarosite without any residual heavy metal and enrichment of Silver (Ag) concentration to at least 2000 ppm besides being simple. The process uses some materials available in the plant and these can be returned to the process. There are other possibilities as well.

OBJECTS OF THE INVENTION:

A primary object of present invention is to provide an improved hydrometallurgy method by modifying the acid leaching step to recover effectively and efficiently metals like silver, lead, cadmium, copper, nickel, manganese and cobalt and lead to formation of Jarosite without any residual heavy metals. It is another object of the invention to provide the improved hydrometallurgy method by modifying the acid leaching step to recover silver by increasing the concentration of silver to at least 2000 ppm in an insoluble residue.

It is still another object of the invention to provide the improved hydrometallurgy method by modifying the acid leaching step to form Jarosite without leaving any residual heavy metal at the end thereby making Jarosite to be safe to dispose even in agricultural land.

It is still another object of the invention to provide the improved hydrometallurgy method by modifying the acid leaching step wherein the chemicals may be recovered and recycled in the leaching method.

SUMMARY OF THE INVENTION:

Conventionally the acid leaching step of hydrometallurgy method comprises subjecting the zinc calcine obtained from roasting step to acid leaching step by using dilute sulfuric acid followed by precipitation of Jarosite and discarding Jarosite waste containing precious metals like silver, lead, cadmium, copper, nickel, manganese and cobalt by filtration. However, it is found in the present invention that when the acid insoluble residue is recovered by filtration followed by washing with demineralized water in step (a) and subsequently the acid insoluble residue is subjected to acid leaching with 30 to 60 % of sulfuric acid in the presence of a reducing agent at temperature in the range of 10°C to 100°C under intense agitation followed by filtering the reaction mixture to obtain an acid insoluble residue enriched with silver to least 2000 ppm and a filtrate in step (b). However, if the acid insoluble residue of step (b) is enriched with silver to less than that of 2000 ppm then acid insoluble residue of step (b) is subjected to alkali digestion. It is found that the alkali insoluble residue is enriched with silver to above 2000 ppm when the acid insoluble residue of the step (b) is subjected to alkali digestion at temperature in the range of 160°C to 180°C. However, if the alkali insoluble residue is enriched with silver to less than that of 2000 ppm then the alkali insoluble residue is subjected to step (d) where the alkali insoluble residue is mixed with the acid insoluble residue of step (a) subjecting to the steps (b) and (c) till a concentration of silver in an insoluble residue obtained in step (d) increases to at least 2000 ppm and above. It is found that the concentration of silver in the insoluble residue so obtained according to the present invention is of at least 2000ppm or above. At this concentration level, silver can be easily recoverable and thus the residue has good commercial value. The method according to the invention may comprise precipitation of Jarosite from the filtrate obtained in step (b) with sodium or potassium sulphate wherein said Jarosite may be free from heavy metal and safe to dispose of.

According to the present invention, there is provided an improved hydrometallurgy method, particularly a modified acid leaching step; wherein the acid leaching step comprises subjecting the zinc calcine obtained from roasting step to acid leaching step by using dilute sulfuric acid characterized in that

(a) filtering the reaction mixture to obtain an acid insoluble residue and a filtrate followed by washing the acid insoluble residue with demineralized water and combining washings with the filtrate; and

(b) subjecting the acid insoluble residue obtained from step (a) to acid leaching by treating it with 30 to 60 % of sulfuric acid in the presence of reducing agent at temperature in the range of 10°C to 100°C under intense agitation followed by filtering the reaction mixture to obtain an acid insoluble residue enriched with silver and a filtrate, and washing the acid insoluble residue followed by combining washings with the filtrate; said acid insoluble residue of step (b) enriched in a concentration of silver to at least 2000 ppm

Typically, the reducing agents used in the step (b) is selected from oxalic acid, sodium bisulfite, sodium meta bisulfite, sodium hydro sulfite, iron scrap, hydrazine hydrate, gaseous hydrogen, gaseous sulfur di oxide, zinc metal powder, iron metal powder, aluminum metal powder, any other metal powder that will reduce ferric form to ferrous form, sodium borohydride, or in combination thereof.

Typically, the reducing agents used in step (b) is 1 to 20 % of acid insoluble residue.

Typically, the reducing agents used in step (b) is oxalic acid, sodium bisulfite, sodium meta bisulfite, hydrazine hydrate, gaseous hydrogen, gaseous sulfur di oxide, iron metal powder, and zinc metal powder.

According to the present invention, there is provided the modified acid leaching step which comprises a step (c) of subjecting the acid insoluble residue obtained in step (b) to alkali digestion at temperature in the range of 160°C to 180°C and autogenous pressure for at least 3 to 6 hours under intense agitation followed by filtering the reaction mixture to obtain a filtrate and an alkali insoluble residue; said alkali insoluble residue of step (c) enriched in a concentration of silver to at least 2000 ppm.

According to the present invention, there is provided the modified acid leaching step which comprises a step (d) of loading the alkali insoluble residue obtained in step (c) with the acid insoluble residue obtained from the step (a) followed by repeating the steps (b) and (c) till a concentration of silver in an insoluble residue obtained in step (d) increases to at least 2000 ppm and above.

Typically, the step (c) is performed on the acid insoluble residue obtained in step (b) with silver less than 2000 ppm to obtain the alkali insoluble residue obtained in step (c) enriched with silver to at least 2000 ppm.

Typically, the step (c) is performed on the acid insoluble residue obtained in step (b) with silver to at least 2000 ppm to obtain the alkali insoluble residue obtained in step (c) enriched with silver to above 2000 ppm.

Typically, the step (c) is performed on the acid insoluble residue obtained in step (b) with silver to at least 2000 ppm to obtain the alkali insoluble residue in step (c) enriched with silver to above 2000 ppm followed by performing the step (d) to obtain the alkali insoluble residue in step (d) enriched with silver more than step (c).

Typically, the step (d) is performed on the alkali insoluble residue obtained in step (c) followed by step (b) with silver to less than 2000 ppm to obtain the insoluble residue in step (d) enriched with silver to at least 2000 ppm.

Typically, the method may comprise precipitation of Jarosite from the filtrate obtained in step (b) with sodium or potassium sulphate; said Jarosite may be free from heavy metal and safe to dispose of. This Jarosite is referred to as clean Jarosite and may be suitable for use as pigments in architectural paints for interior or exterior and industrial solvent based paints due to their low lead contents, filler in plastics, as a raw material for preparation of high purity iron oxide, as an additive to cement for replacement of gypsum, etc.

DETAILED DESCRIPTION OF THE INVENTION:

The terms “a,” “an,” “the” and similar referents used in the context of describing the invention following claims are to be construed to cover both the singular and the plural, unless otherwise indicated herein or clearly contradicted by context. Recitation of ranges of values herein is merely intended to serve as a shorthand method of referring individually to each separate value falling within the range. Unless otherwise indicated herein, each individual value is incorporated into the specification as if it were individually recited herein. All methods described herein can be performed in any suitable order unless otherwise indicated herein or otherwise clearly contradicted by context. The use of any and all examples, or exemplary language (e.g., “such as”) provided herein is intended merely to better illustrate the invention and does not pose a limitation on the scope of the invention otherwise claimed. No language in the specification should be construed as indicating any non-claimed element essential to the practice of the invention.

Certain embodiments of this invention are described herein, including the best mode known to the inventors for carrying out the invention. Of course, variations on these described embodiments will become apparent to those of ordinary skill in the art upon reading the description. The inventor expects skilled artisans to employ such variations as appropriate, and the inventors intend for the invention to be practiced otherwise than specifically described herein. Accordingly, this invention includes all modifications and equivalents of the subject matter recited in the claims appended hereto as permitted by applicable law. Moreover, any combination of the below-described elements in all possible variations thereof is encompassed by the invention unless otherwise indicated herein or otherwise clearly contradicted by context.

Specific embodiments disclosed herein can be further limited in the claims using consisting of or and consisting essentially of language. When used in the claims, whether as filed or added per amendment, the transition term “consisting of excludes any element, step, or ingredient not specified in the claims. The transition term “consisting essentially of limits the scope of a claim to the specified materials or steps and those that do not materially affect the basic and novel characteristic(s). Embodiments of the invention so claimed are inherently or expressly described and enabled herein.

The hydrometallurgical method for extracting zinc generally comprises one of the steps after roasting the sulfides to convert into oxides followed by neutral leaching of zinc oxide (calcine) by dissolving in sulfuric acid and using spent electrolyte in an acid leaching of the zinc ferrites by using spent electrolyte and concentrated sulfuric acid to obtain a residue containing lead and silver among the other metals and also a solution rich in zinc sulphate and ferric sulphate.

According to the present invention, an improved hydrometallurgy method, particularly a modified acid leaching step; wherein the acid leaching step comprises subjecting the zinc calcine obtained from roasting step to acid leaching step by using dilute sulfuric acid characterized in that

(a) filtering the reaction mixture to obtain an acid insoluble residue and filtrate and washing the acid insoluble residue followed by combining washings with the filtrate; and

(b) subjecting the acid insoluble residue obtained in the step (a) to acid leaching by treating it with dilute sulfuric acid in the presence of reducing agent at temperature in the range of 10°C to 100°C under intense agitation followed by filtering the reaction mixture to obtain an acid insoluble residue having a concentration of silver in the insoluble residue increased to at least 2000 ppm and filtrate, and washing the acid insoluble residue followed by combining washings with the filtrate.

In one of the embodiments, the reducing agent may be added to the acid leaching step of the existing plant in the last stages of extraction to avoid additional equipment. The concentrate so produced may be filtered, washed, optionally dried and taken to the step of alkali digestion if further enrichment in silver up to 2000 ppm is required.

According to the preferred embodiment of the present invention, the acid insoluble residue obtained in step (b) is enriched with silver to at least 2000 ppm then the step (c) and step (d) are optional. If performed, the step (c) and step (d) enable to enrich the residue with silver further, particularly to desired concentration, more particularly to above 2000 ppm. In one of the embodiment of the present invention, the improved hydrometallurgy method, particularly the modified acid leaching step comprises a step (c) of subjecting the acid insoluble residue obtained in the step (b) to alkali digestion at temperature in the range of 160°C to 180°C and autogenous pressure for at least 3 to 6 hours under intense agitation followed by filtering the reaction mixture to obtain a filtrate and an alkali insoluble residue having a concentration of silver in the insoluble residue increased to at least 2000 ppm.

According to one of the embodiments of the present invention, the acid insoluble residue obtained in step (b) is enriched with silver to less than 2000 ppm then the step (c) is performed to obtain the alkali insoluble residue enriched with silver further, particularly to at least 2000 ppm and step (d) is optional. If performed, the step (d) enables to enrich the residue with silver further, particularly to desired concentration, more particularly to above 2000 ppm.

In another embodiment of the present invention, the improved hydrometallurgy method, particularly the modified acid leaching step further comprises a step (d) of loading the alkali insoluble residue obtained in the step (c) with the acid insoluble residue obtained in the leaching step (a) followed by repeating the steps (b) and (c) till a concentration of silver in an insoluble residue increases to at least 2000 ppm.

According to another embodiment of the present invention, the alkali insoluble residue obtained in step (c) followed by step (b) is enriched with silver to less than 2000 ppm then the step (d) is performed to obtain the insoluble residue obtained in step (d) enriched with silver to at least 2000 ppm.

Typically, 30% to 60% dilute sulfuric acid is used in step (b).

The dilute sulfuric acid is typically obtained from enrichment of the 10% acid obtained from the cell house.

The reducing agent used in step (b) is in the range of 1 to 20 %. The amount of reducing agent used in the step (b) is dependent on the ferric iron in the insoluble residue and zinc ferrite.

The reducing agents used in step (b) is selected from oxalic acid, sodium bisulfite, sodium meta bisulfite, sodium hydro sulfite, iron scrap, hydrazine hydrate, gaseous hydrogen, gaseous sulfur di oxide, zinc metal powder, iron metal powder, aluminum metal powder, any other metal powder that will reduce ferric form to ferrous form, sodium borohydride, or in combination thereof.

The sodium bisulfite reducing agent has the advantage of reduced usage of sodium sulfate required for precipitation of Jarosite.

Zinc and iron are substantially removed from the acid insoluble residue by the process of step (b). It gets enriched in silica, lead, cadmium and silver.

The alkali used in the step (c) is sodium hydroxide or potassium hydroxide.

The residue of step (c) becomes lean in silica and alumina and it is enriched in lead, cadmium and silver.

Sodium or potassium silicate obtained in the digestion step (c) is used for precipitation of silica. Silica is a useful filler for rubber, polymers, tooth paste, paints and catalysts. The co-product sodium or potassium sulphate is useful for precipitation of Jarosite.

Step (d) can be repeated further till the residue having silver (Ag) is enriched to desired concentration.

The residue obtained at the end of step (d) having silver (Ag) enriched to at least 2000 ppm or above.

The alkali insoluble residue of step (c) gets enriched in lead, cadmium and silver. A part of the alkali insoluble residue can be withdrawn as a product and the balance can be recycled to the acid leaching step (b).

The hydrometallurgy method can be improved by modifying the acid leaching step as per the step (b) of the present invention. The Jarosite produced using filtrate obtained in step (b) by precipitation may be free from heavy metal and safe to dispose of. It may contain residual zinc as zinc sulphate. It is a micro nutrient to plants. This Jarosite may not need any special efforts to dispose and can be spread in the fields. It can also be used as pigments in architectural paints for interior or exterior and industrial solvent based paints due to their low lead contents, filler in plastics, as a raw material for preparation of high purity iron oxide, as an additive to cement for replacement of gypsum, etc. The present invention used acid insoluble residue containing lead, cadmium and silver among the other metals obtained in the acid leaching step of the hydrometallurgy method.

In step (a), the acid insoluble residue is washed with demineralized water till the filtrate becomes sulphate free followed by optionally drying the residue at temperature in the range of 110°C to 130°C.

The acid insoluble residue of step (a) is tested by XRF analysis to find out metal contents thereof. The results of the same are illustrated in Table 2.

Typically, the acid insoluble residue obtained in the step (a) is subjected to step (b) of the present invention.

In the step (b), the residue of the step (a) is treated with 30 % sulfuric acid followed by heating to 100°C, adding a 4 to 6 % sodium bisulfite aqueous solution dropwise for three hours maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate.

Typically, the acid insoluble residue and the filtrate obtained in the step (b) is weighed and solid content of the same is determined. The elemental analysis of the acid insoluble residue and the filtrate obtained in step (b) is carried out by XRF analysis and the same is illustrated in Table 3.

According to Table 3, the silver concentration in the acid insoluble residue obtained in step (b) using sodium bisulfite as reducing agent is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) is subjected to the step (c) of alkali digestion.

In step (c) of the present invention, the acid insoluble residue of step (b) is subjected to alkali digestion in the presence of 20 % solution sodium hydroxide at temperature in the range of 160 °C to 180°C at autogenous pressure for at least 3 hours at 1000 to 1200 rpm followed by filtering the reaction mixture to obtain an alkali insoluble residue and a filtrate. Typically, the alkali insoluble residue and the filtrate obtained in step (c) are weighed and solid content of the same is determined. The elemental analysis of the alkali insoluble residue and the filtrate obtained in step (c) is carried out by XRF analysis and the same is illustrated in Table 4.

According to Table 4, the silver concentration in the alkali insoluble residue obtained in step (c) is less than 2000 ppm. Therefore, the alkali insoluble residue obtained in the step (c) is subjected to the step (d) of the present invention.

In step (d) of the present invention, the alkali insoluble residue obtained in step (c) is mixed with the residue obtained in step (a) followed by subjecting it to the treatment with 30 % dilute sulfuric acid followed by heating to 100°C, adding a 4 to 6 % sodium bisulfite aqueous solution drop- wise for 3 hours maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from the filtrate, washing the acid insoluble residue with water till the filtrate becomes sulphate free, the acid insoluble residue subjected to alkali digestion in the presence of 20 % solution of sodium hydroxide at temperature in the range of 160 to 180°C and autogenous pressure for at least 3 hours with intense agitation followed by filtering the reaction mixture to obtain an alkali insoluble residue and a filtrate.

The alkali insoluble residue and the filtrate obtained in the step (d) are weighed and solid content of the same is determined. The elemental analysis of the insoluble residue and the filtrate obtained in the step (d) is carried out by XRF analysis and the same is illustrated in Table 5.

According to Table 5, the silver concentration in the insoluble residue obtained in step (d) is more than 2000 ppm.

The insoluble residue having 0.0314 % silver when subjected to the step (b) using sodium bisulfite as reducing agent and step (c) of the present invention results into the alkali insoluble residue having 0.170 % silver. Thus, the residue obtained after treatment of step (b) and step (c) is 5.6 times more concentrated in silver than that of the residue of step (a). At the end of step (d), the residue obtained has 0.276 % silver. Thus, the residue obtained after treatment of step (d) is 9 times more concentrated in silver than that of the residue of step (a).

According to another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 30 % sulfuric acid followed by heating to 100°C, adding a 10 to 50 ml of 10 % oxalic acid aqueous solution drop wise for 30 minutes maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate.

Typically, the acid insoluble residue and the filtrate obtained in the step (b) is weighed and solid content of the same is determined.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 1 %, 2%, 3% and 5% oxalic acid as reducing agent is carried out by XRF analysis and the same is illustrated in Tables 6 to 9 respectively.

According to the present invention, the acid insoluble residues obtained in step (b) using 1 %, 2%, 3% and 5 % oxalic acid is enriched with silver to 0.151 %, 0.195 %, 0.195 % and 0.192 % respectively.

According to Tables 6 to 9, the silver concentration in the acid insoluble residue obtained in step (b) using oxalic acid as reducing agent is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) is subjected to the step (c) of alkali digestion.

In step (c) of the present invention, the acid insoluble residue of step (b) is subjected to alkali digestion in the presence of 20 % solution sodium hydroxide at temperature in the range of 160 to 180°C at autogenous pressure for at least 3 hours at 1000 to 1200 rpm followed by filtering the reaction mixture to obtain an alkali insoluble residue and a filtrate.

Typically, the alkali insoluble residue and the filtrate obtained in step (c) are weighed and solid content of the same is determined. The elemental analysis of the alkali insoluble residue and the filtrate obtained in step (c) is carried out by XRF analysis and the same is illustrated in Table 10.

According to the present invention, the acid insoluble residue obtained in step (b) is subjected to alkali digestion step (c) to obtain the insoluble residue enriched with silver to 0.463 % respectively. According to Table 10, the silver concentration in the insoluble residue obtained in step (c) is more than 2000 ppm i.e. 4630 ppm. Therefore, step (d) is not performed.

The insoluble residue having 0.0314 % silver when subjected to the step (b) of the present invention using oxalic acid as reducing agent results into the acid insoluble residue having 0.151 to 0.195 % silver. Thus, the residue obtained after treatment of step (b) is 4.8 to 6.2 times more concentrated in silver than that of the residue of step (a). At the end of step (c), the residue obtained has 0.463 % silver. Thus, the residue obtained after treatment of step (c) is 14.75 times more concentrated in silver than that of the residue of step (a). Thus, the step (d) is not performed.

According to yet another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 30 ml of 10 % Hydrazine hydrate aqueous solution maintaining temperature to 100 °C followed by heating at 100 °C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 3 % Hydrazine hydrate as reducing agent is carried out by XRF analysis and the same is illustrated in Table 11.

According to the present invention, the acid insoluble residues obtained in step (b) using 3 % Hydrazine hydrate as reducing agent is enriched with silver to 0.189 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

According to yet another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 50 ml of 5 % Sodium Meta Bisulphite aqueous solution maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate. Typically, the acid insoluble residue and the filtrate obtained in the step (b) is weighed and solid content of the same is determined.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 2.5 % Sodium Meta Bisulphite as reducing agent is carried out by XRF analysis and the same is illustrated in Table 12.

According to the present invention, the acid insoluble residues obtained in step (b) using 2.5 % Sodium Meta Bisulphite as reducing agent is enriched with silver to 0.185 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to the step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

According to still another embodiment of the invention, in the step (b), the residue of step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 5 % Iron metal powder maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 5 % Iron metal powder is carried out by XRF analysis and the same is illustrated in Table 13.

According to the present invention, the acid insoluble residues obtained in step (b) using 5 % Iron metal powder is enriched with silver to 0.168 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to the step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

According to still another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 20 % gaseous sulfur dioxide maintaining temperature to 100°C for at least one hour, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes free of dissolved iron and washings mixed with the filtrate.

The use of sulfur di oxide as a reducing agent in the present invention has advantage as it is available in a zinc plant in huge quantity. The contact of sulfur di oxide with the reaction mixture is an important process parameter. To ensure proper contact and dissolution of sulfur di oxide in acidic solution, the reaction mixture is filled in a tubular glass column of 3 meter height and 35 mm diameter.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 20 % gaseous sulfur di oxide is carried out by XRF analysis and the same is illustrated in Table 14.

According to the present invention, the acid insoluble residues obtained in step (b) using 20 % gaseous sulfur di oxide is enriched with silver to 0.155 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

According to still another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 20 % gaseous hydrogen maintaining temperature to 100°C for at least one hour, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes free of dissolved iron and washings mixed with the filtrate.

The contact of hydrogen gas with the reaction mixture is an important process parameter. To ensure proper contact and dissolution of hydrogen in acidic solution, the reaction mixture is filled in a tubular glass column of 3 meter height and 35 mm diameter. The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 20 % gaseous hydrogen is carried out by XRF analysis and the same is illustrated in Table 15.

According to the present invention, the acid insoluble residues obtained in step (b) using 20 % gaseous hydrogen is enriched with silver to 0.160 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to the step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

According to yet still another embodiment of the invention, in the step (b), the residue of the step (a) is treated with 60 % sulfuric acid followed by heating to 100°C, adding 5 % zinc metal powder maintaining temperature to 100°C followed by heating at 100°C for at least 4 hours, filtering the reaction mixture to separate an acid insoluble residue from a filtrate, washing the residue with water till the filtrate becomes sulphate free and washings mixed with the filtrate.

The elemental analysis of the acid insoluble residues and the filtrates obtained in step (b) using 5 % zinc metal powder is carried out by XRF analysis and the same is illustrated in Table 16.

According to the present invention, the acid insoluble residues obtained in step (b) using 5 % zinc metal powder is enriched with silver to 0.180 %. Thus, the silver concentration in the acid insoluble residue obtained in step (b) is less than 2000 ppm. Therefore, the acid insoluble residue obtained in the step (b) may be subjected to step (c) of alkali digestion and / or step (d) to enrich insoluble residue with silver further.

The Jarosite produced using filtrate obtained in step (b) by precipitation with sodium or potassium silicate may be free from heavy metal and safe to dispose of. The elemental analysis of the Jarosite (i.e. commercially available) and Jarosite prepared from the filtrate obtained in step (b) is carried out by XRF analysis and the same is illustrated in Table 17 and 18 respectively. According to Table 18, Jarosite of the present invention may contain residual zinc as zinc sulphate. It is a micro nutrient to plants. This Jarosite may not need any special efforts to dispose of and can be spread in the fields. The clean Jarosite contains 0.1% PbO, corresponding to about 920 ppm lead. Recently, lead in architectural finishes has been restricted to 90 ppm. Thus, up to 10% clean Jarosite can be added to architectural finishes for interior use. There may be no restriction on its concentration in industrial finishes.

According to the present invention, the acid leaching step of the conventional hydrometallurgy method can be carried out under reducing conditions and filtering it before removing iron as Jarosite.

The present invention has following advantages:

1. Silver is available in a much more concentrated form. Easy and very less expensive to further concentrate it.

2. Very little change in the existing plant as a reducing agent will have to be added. Two are available in the existing plant i.e. SO2 and Zn metal powder. The quantity of reducing agent used in the step (a) or acid leaching step is very small i.e. Only to the extent of ferric iron present in the insoluble component.

3. Fresh investment required only for filtration and alkali digestion. The high price of silver and a good concentration of lead in the mineral will compensate for it. These steps would any way be off line.

4. Jarosite would be free of heavy metals and can be safely disposed in the fields as a source of zinc.

5. Clean Jarosite can also be used for manufacture of architectural finishes and industrial coatings, as filler in plastics, as additive to cement for replacement of gypsum or as raw material for preparation of high purity iron oxide.

The above invention can be illustrated with the below mentioned examples but not by way of limitations. In other word, exemplary illustrations of the operation of the present invention, the practice of its formulation and the rendering of the disclosed process are described in the following examples. In addition to the preferred modes of operation, a practitioner of sufficient skill in the art will appreciate that the meets and bounds of the present invention are not limited by the specific instances described herein, rather are defined by the equivalents provided by the claims of the present invention.

EXAMPLE 1

Step (a) - The acid insoluble residue as raw material

1 Kg of acid insoluble residue obtained in the acid leaching step of the hydrometallurgy method was collected from the plant of Hindustan Zinc Ltd Udaipur. The acid insoluble residue was washed with demineralized water till the filtrate becomes sulphate free (i.e. Sulphate test: A small portion of filtrate was taken in a test tube and to this, barium chloride solution was added to check if no white colored precipitate was observed. White colour precipitates was indication of presence of sulphate). The washed acid insoluble residue was dried at 120°C. The acid insoluble residue was tested by XRF analysis to find out metal contents thereof. The results of the same are illustrated in Table 2.

Table 2: XRF analysis of acid insoluble residue before treatment

The residue obtained at step (a) had silver concentration of 0.0314 %.

Step (b) - Acid leaching in the presence of reducing agent, 2 % sodium bisulfite (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 33.33 ml of 6 % sodium bisulfite aqueous solution (i.e. 2 gm of sodium bisulfite) was added dropwise for three hours maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

The acid insoluble residue and the filtrate were weighed and solid content of the same was determined.

Weight of Residue: 50.0 g

Weight of filtrate: 912 g

Solid content of filtrate: 15%

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 3.

Table 3: XRF analysis of the acid insoluble residue and filtrate obtained in step (b)

The acid insoluble residue obtained at step (b) had silver concentration of 0.0731 %.

Step (c): Alkali digestion

138 gm Sodium hydroxide (97%) was dissolved in 555 gm of water in an autoclave to make 20% solution. 50 gm of acid insoluble residue obtained from step (b) was added to 20 % solution sodium hydroxide and alkali digestion step was carried out by heating the reaction mixture at temperature of 170°C at an autogenous pressure (approx. 5 bar) for 3 hours at 350 rpm. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to obtain an alkali insoluble residue and a filtrate.

The alkali insoluble residue and the filtrate were weighed and solid content of the same was determined.

Weight of Residue: 18 g

Weight of filtrate: 1200 ml

Solid content of filtrate: 11.76%

The elemental analysis of the alkali insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 4.

Table 4: XRF analysis of the alkali insoluble residue and the filtrate obtained in step (c)

The alkali insoluble residue obtained in step (c) had silver concentration of 0.170 %.

Step (d): Repetition of Step (a) and Step (b)

The alkali insoluble residue (18 gm) obtained in the step (c) was mixed with 100 gm of the fresh residue obtained from step (a). This combined residue was subjected to step (b) and step (c) as mentioned in the Example 1 to obtain an alkali insoluble residue and a filtrate.

The alkali insoluble residue and the filtrate obtained in the step (d) were weighed and solid content of the same was determined.

Weight of Residue: 20 g

Weight of filtrate: 1192 g

The elemental analysis of the insoluble residue and the filtrate obtained in the step (d) was carried out by XRF analysis and the same is illustrated in Table 5.

Table 5: XRF analysis of the insoluble residue and the filtrate obtained in step (d)

The alkali insoluble residue obtained in step (d) had silver concentration of 0.276 %.

The residue of step (a) with 0.0314 % silver was subjected to the step (b) and step (c) and resulted into the alkali insoluble residue enriched with 0.170 % silver. Thus, the residue obtained after treatment of step (b) and step (c) was 5.6 times more concentrated in silver than that of the residue of step (a). At the end of step (d), the residue obtained had 0.276 % silver. Thus, the residue obtained after treatment of step (d) was 9 times more concentrated in silver than that of the residue of step (a).

EXAMPLE 2

Step (a) - The acid insoluble residue as raw material

1 Kg of acid insoluble residue obtained in the acid leaching step of the hydrometallurgy method was processed according to the Example 1 step (a). The acid insoluble residue was tested by XRF analysis and the results of the same are found to be same as Table 2.

Step (b) - Acid leaching in the presence of reducing agent, 1 % oxalic acid (wt./wt. with respect to acid insoluble residue) 100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 10 ml of 10 % oxalic acid aqueous solution (i.e. 1 gm of oxalic acid) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 21.2 g

Weight of filtrate: 5.14 L

Solid Content of filtrate: 6.84 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 6.

Table 6: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with 1 % oxalic acid

The acid insoluble residue obtained in step (b) had silver concentration of 0.151 %.

EXAMPLE 3

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 2 % oxalic acid (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 20 ml of 10 % oxalic acid aqueous solution (i.e. 2 gm of oxalic acid) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 20.8 g

Weight of filtrate: 2.41 L

Solid Content of filtrate: 12.77 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 7. Table 7: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with 2 % oxalic acid

The acid insoluble residue obtained in step (b) had silver concentration of 0.195 %.

EXAMPLE 4

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 3 % oxalic acid (wt./wt. with respect to acid insoluble residue) 100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 30 ml of 10 % oxalic acid aqueous solution (i.e. 3 gm) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 20.0 g

Weight of filtrate: 3.6 L

Solid Content of filtrate: 10.37%

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 8.

Table 8: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with 3 % oxalic acid

EXAMPLE 5

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 5 % oxalic acid (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 50 ml of 10 % oxalic acid aqueous solution (i.e. 5 gm of oxalic acid) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 19.5 g

Weight of filtrate: 3.21 L

Solid Content of filtrate: 7.27 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 9. Table 9: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with 5 % oxalic acid

The acid insoluble residue obtained in step (b) had silver concentration of 0.192 %.

Step (c): Alkali digestion

138 gm Sodium hydroxide (97%) was dissolved in 555 gm of water in an autoclave to make 20% solution. 50 gm of acid insoluble residue obtained from step (b) was added to 20 % solution sodium hydroxide and alkali digestion step was carried out by heating the reaction mixture at temperature of 170°C at an autogenous pressure (approx. 5 bar) for 3 hours at 1000 rpm. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to obtain an alkali insoluble residue and a filtrate. The alkali insoluble residue and the filtrate were weighed and solid content of the same was determined.

Weight of residue after treament: 5.2 g

Weight of filtrate: 1.13 L

Solid Content of filtrate: 11.05 %

The elemental analysis of the alkali insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 10.

Table 10: XRF analysis of the alkali insoluble residue and filtrate obtained in step (c)

The alkali insoluble residue obtained in step (c) had silver concentration of 0.463 %. Thus, the silver concentration in the insoluble residue obtained in step (c) was more than 2000 ppm i.e. 4630 ppm. Therefore, step (d) was not performed. The residue of step (a) having 0.0314 % silver when subjected to the step (b) of the Examples 2 to 5 resulted into the acid insoluble residue having 0.151 to 0.195 % silver. Thus, the residue obtained after treatment of step (b) is 4.8 to 6.2 times more concentrated in silver than that of the residue of step (a). At the end of step (c) of the Example 5, the residue obtained has 0.463 % silver. Thus, the residue obtained after treatment of step (c) is 14.75 times more concentrated in silver than that of the residue of step (a). Thus, the step (d) was not performed.

EXAMPLE 6

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 3 % Hydrazine Hydrate (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 30 ml of 10 % Hydrazine hydrate aqueous solution (i.e. 3 gm of Hydrazine hydrate) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 23.5 g

Weight of filtrate: 5.50 L

Solid Content of filtrate: 5.35 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 11. Table 11: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with Hydrazine hydrate

The acid insoluble residue obtained in step (b) had silver concentration of 0.189 %.

EXAMPLE 7

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 2.5 % Sodium Metabisulphite (wt./wt. with respect to acid insoluble residue) 100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 50 ml of 5 % Sodium Meta Bisulphite aqueous solution (i.e. 2.5 gm of Sodium Meta Bisulphite) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 19.5 g

Weight of filtrate: 2.89 L

Solid Content of filtrate: 13.56 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 12.

Table 12: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with Sodium Meta Bisulphite

The acid insoluble residue obtained in step (b) had silver concentration of 0.185 %.

EXAMPLE 8

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 5 % Iron metal powder (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, Iron metal powder (5 gm) was added maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 28.6 g

Weight of filtrate: 2.9 L

Solid Content of filtrate: 11.12%

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 13. Table 13: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with Iron metal powder.

The acid insoluble residue obtained in step (b) had silver concentration of 0.168 %.

EXAMPLE 9

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent, 20 % Gaseous sulfur di oxide (SO2) (vol./wt. with respect to acid insoluble residue) 100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. To ensure proper contact of SO2 gas with the reaction mixture and dissolution of SO2 gas in acidic solution, the reaction mixture was filled in a tubular glass column of 3 meter height and 35 mm diameter with a sparger at the bottom. The reaction mixture was heated to 100° C. SO2 gas was passed till the bed expanded by 20%. The unreacted SO2 was scrubbed in an alkaline scrubber. About 20 litres SO2 was passed in the reaction mixture in about 1 hour by maintaining temperature of 100° C. The contents were discharged from the column and filtered. The residue was washed with demineralized water till the washings were free of dissolved iron.

Weight of residue after treament: 30 g

Weight of filtrateincluding washings: 2.9 L

Solid Content of filtrate: 11.12%

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 14.

Table 14: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with sulfur di oxide

The acid insoluble residue obtained in step (b) had silver concentration of 0.155 %.

EXAMPLE 10

Step (a) - The acid insoluble residue as raw material

Step (b) - Acid leaching in the presence of reducing agent 20 % hydrogen gas (vol./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. To ensure proper contact of hydrogen gas with the reaction mixture and dissolution of hydrogen gas in acidic solution, the reaction mixture was filled in a tubular glass column of 3 meter height and 35 mm diameter with a sparger at the bottom. The reaction mixture was heated to 100° C. Hydrogen gas was passed till the bed expanded by 20%. The unreacted hydrogen was scrubbed in an alkaline scrubber and was diluted with air before dispersal. About 20 litres hydrogen gas was passed in the reaction mixture in about 1 hour by maintaining temperature of 100° C. The contents were discharged from the column and filtered. The residue was washed with demineralized water till the washings were free of dissolved iron.

Weight of residue after treament: 27 gm

Weight of filtrate: 2.9 L

Solid Content of filtrate: 9 % The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 15.

Table 15: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with hydrogen gas.

The acid insoluble residue obtained in step (b) had silver concentration of 0.16 %.

EXAMPLE 11

Step (a) - The acid insoluble residue as raw material

1 Kg of acid insoluble residue obtained in the acid leaching step of the hydrometallurgy method was processed according to the Example 1 step (a). The acid insoluble residue was tested by XRF analysis and the results of the same are found to be same as Table 2. Step (b) - Acid leaching in the presence of reducing agent, 5 % Zinc Metal Powder (wt./wt. with respect to acid insoluble residue)

100 gm of acid insoluble residue obtained from step (a) was treated with 60 % sulfuric acid till pH of the reaction mixture changed to 1. The reaction mixture was heated to 100°C. To this boiling mixture, 5 gm of zinc metal powder was added slowly over 3 hours maintaining temperature to 100°C. The reaction mixture was further heated at 100°C for 4 hours. The reaction mixture was cooled down to 40°C. The reaction mixture was filtered to separate an acid insoluble residue and a filtrate. The filtered residue was washed with water till the filtrate becomes iron and sulphate free. The washing was combined with the filtrate.

Weight of residue after treament: 19.5 g

Weight of filtrate: 2.8 L

Solid Content of filtrate: 13 %

The elemental analysis of the acid insoluble residue and the filtrate was carried out by XRF analysis and the same is illustrated in Table 16.

Table 16: XRF analysis of the acid insoluble residue and filtrate obtained in step (b) with hydrogen gas

The acid insoluble residue obtained in step (b) had silver concentration of 0.180 %.

The Jarosite was produced using filtrate obtained in step (b) of Example 1 by precipitation with sodium or potassium silicate may be free from heavy metal and safe to dispose of. The elemental analysis of the Jarosite (i.e. commercially available) and Jarosite prepared from the filtrate obtained in step (b) was carried out by XRF analysis and the same is illustrated in Table 17 and 18 respectively.

Table 17: XRF analysis of the Jarosite (i.e. commercially available)

Table 18: XRF analysis of the Jarosite prepared from the filtrate obtained in step (b) of the present invention

According to Table 18, Jarosite of the present invention may contain residual zinc as zinc sulphate. It is a micro nutrient to plants. This Jarosite may not need any special efforts to dispose of and can be spread in the fields. The clean Jarosite contains 0.1% PbO, corresponding to about 920 ppm lead. Recently, lead in architectural finishes has been restricted to 90 ppm. Thus, clean Jarosite may be suitable for use as pigments in architectural paints for interior or exterior and industrial solvent based paints due to their low lead contents, filler in plastics, as a raw material for preparation of high purity iron oxide, as an additive to cement for replacement of gypsum, etc.

Thus, the present invention recovers Silver to the fullest and may simultaneously generate Jarosite free from heavy metal and safe to dispose of. Jarosite so obtained may contain residual zinc as zinc sulphate. It may be used as a micro nutrient to plants. This Jarosite may not need any special efforts to dispose of and can be spread in the fields.

Claims

I claim:

1. An improved hydrometallurgy method, particularly a modified acid leaching step; wherein the acid leaching step comprises subjecting the zinc calcine obtained from roasting step to acid leaching step by using dilute sulfuric acid characterized in that

(b) subjecting the acid insoluble residue obtained from step (a) to acid leaching by treating it with 30 to 60 % of sulfuric acid in the presence of reducing agent at temperature in the range of 10 °C to 100 °C under intense agitation followed by filtering the reaction mixture to obtain an acid insoluble residue enriched with silver and a filtrate, and washing the acid insoluble residue followed by combining washings with the filtrate; said acid insoluble residue of step (b) enriched in a concentration of silver to at least 2000

PPm

2. The method as claimed in claim 1, wherein the reducing agents used in the step (b) is selected from oxalic acid, sodium bisulfite, sodium meta bisulfite, sodium hydro sulfite, iron scrap, hydrazine hydrate, gaseous hydrogen, gaseous sulfur di oxide, zinc metal powder, iron metal powder, aluminum metal powder, any other metal powder that will reduce ferric form to ferrous form, sodium borohydride, or in combination thereof.

3. The method as claimed in claim 1, wherein the modified acid leaching step comprises a step (c) of subjecting the acid insoluble residue obtained in the step (b) to alkali digestion at temperature in the range of 160°C to 180°C and autogenous pressure for at least 3 to 6 hours under intense agitation followed by filtering the reaction mixture to obtain a filtrate and an alkali insoluble residue. said alkali insoluble residue of step (c) enriched in a concentration of silver to at least 2000 ppm.

4. The method as claimed in claims 1 and 3, wherein the modified acid leaching step comprises a step (d) of loading the alkali insoluble residue obtained in the step (c) with the acid insoluble residue obtained in the step (a) followed by repeating the steps (b) and (c) till a concentration of silver in an insoluble residue increases to at least 2000 ppm and above.

5. The method as claimed in claims 1 to 3, wherein performing the step (c) on the acid insoluble residue obtained in step (b) with silver less than 2000 ppm to obtain the alkali insoluble residue obtained in step (c) enriched with silver to at least 2000 ppm.

6. The method as claimed in claims 1 to 3, wherein performing the step (c) on the acid insoluble residue obtained in step (b) with silver to at least 2000 ppm to obtain the alkali insoluble residue obtained in step (c) enriched with silver to above 2000 ppm.

7. The method as claimed in claims 1 to 6, wherein performing the step (c) on the acid insoluble residue obtained in step (b) with silver to at least 2000 ppm to obtain the alkali insoluble residue in step (c) enriched with silver to above 2000 ppm followed by performing the step (d) to obtain the alkali insoluble residue in step (d) enriched with silver more than step (c).

8. The method as claimed in claims 1 to 5, wherein performing the step (d) on the alkali insoluble residue obtained in step (c) followed by step (b) with silver to less than 2000 ppm to obtain the insoluble residue in step (d) enriched with silver to at least 2000 ppm.

9. The method as claimed in any of the claims 1 to 8, wherein the method may comprise precipitation of a Jarosite from the filtrate obtained in step (b) with sodium or potassium sulphate; said Jarosite may be free from heavy metal and safe to dispose of or may be used as pigments in architectural paints for interior or exterior and industrial solvent based paints, filler in plastics, as a raw material for preparation of high purity iron oxide or as an additive to cement for replacement of gypsum.