TECHNICAL FIELD
[0001] The present disclosure relates to the field of metallurgy. More specifically, the present disclosure relates to a method for hydrometallurgical recovery of metal values 5 from industrial waste and produce saleable compound crystal.
BACKGROUND
[0002] Nowadays, a significant amount of zinc is globally produced through hydrometallurgical process. The hydrometallurgical process is often referred to as Roast-Leach-Electrowinning (RLE) process. A copious amount of waste is generated through 10 the Roast-Leach-Electrowinning process. This waste is processed for obtaining zinc. One of the waste products which is produced during the course of the RLE process is jarosite. Moreover, jarosite is primarily a complex compound of iron which contains zinc with traces of other metals. Traditionally, jarosite is dumped in various landfills, lined impermeable ponds and the like. However, the disposal of the copious amount of jarosite 15 has presented problems involving the land acquisition and use of dumping ground. Nowadays, the cost of handling the copious amount of jarosite proves to be costly. However, jarosite, if treated can provide significant amount of zinc, iron, lead, silver and other saleable products.
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[0003] Typically, jarosite is treated through a chemical and pyro metallurgical process for extraction of zinc and other metals. The traditional method includes acid leaching of the jarosite in presence of sulfuric acid to dissolve the ferrites in the jarosite. Further, the traditional method includes treating the resulting leach slurry with zinc sulfide under reducing conditions to consume excess acid from leaching. The typical method further 25 includes recovering excess zinc sulfide and elemental sulfur produced in the reaction of zinc sulfide with acid and ferric iron by zinc flotation.
[0004] In one of the prior arts with patent number US5348713 A, the recovery method of zinc, lead, copper and precious metals from zinc plant residue is disclosed. The 30
3
process includes leaching the residue with return zinc spent electrolyte, neutralizing residual acid and reducing ferric iron in the solution by addition of zinc sulphide concentrate in the presence of a limited quantity of oxygen, flotation of the resulting slurry to separate unreacted zinc sulphide, treatment of flotation tailings with sulphur dioxide and elemental sulphur to further leach iron, zinc and impurity elements and 5 precipitate copper, flotation of the resulting slurry to separate a copper sulphide concentrate, thickening, filtering and washing of the flotation tailings followed by addition of lime and sodium sulphide to activate lead sulphate and flotation of a lead concentrate from the residue. Iron and impurity elements are precipitated from the copper flotation tailings thickener overflow solution by addition of zinc hydroxide sludge, lime 10 and oxygen to produce a high strength, iron free zinc sulphate solution.
[0005] In another prior art with patent number US3966569 A, a method of recovering metal from metalliferous waste is provided. Metal can be recovered from metal-containing waste by means of liquid extraction processes. The waste is leached with 15 sulphuric acid. The resulting metal sulphate solution is contacted with an organic solution of a reagent, so as to extract iron and zinc to the organic solution. The organic solution is washed with sulphuric acid in two steps, viz. with weaker acid in the first step, so as to transfer zinc to the washing solution, and with stronger acid in the second step, so as to transfer iron to the washing solution. Zinc and iron have now been separated, and 20 are recovered from the washing solutions in known ways, for example by crystallization. If the waste contains other metals, in addition to iron and zinc, additional selective liquid extraction processes are added, before or after the iron-zinc-extraction.
[0006] In yet another prior art with patent number US4366127 A, a hydrometallurgical 25 process for the recovery of lead, silver and gold, as well as zinc, from impure jarosite residues of an electrolytic zinc process is provided. The impure jarosite residue of an electrolytic zinc process is leached in a sulfuric-acid-bearing solution in order to produce a leach residue which contains lead, silver and gold and a ferrisulfate-bearing solution and to separate them from each other, where after the leach residue is sulfidized and froth-30 flotated in order to recover a combined concentrate which contains lead, silver and gold, and the ferrisulfate-bearing solution is fed to a ferritic treatment stage, in which
4
ferrisulfate and ferrites react in the presence of ions of alkali and ammonium at 80°C - 105° C. and form pure jarosite and zinc sulfate.
[0007] The above mentioned prior arts are found to be inefficient and bear several disadvantages. These prior arts are unable to utilize jarosite in an efficient manner in 5 generating metal values. In addition, these prior art references focus only on extraction of zinc, lead and silver by treating jarosite, while neglecting the production of saleable compound crystals like ferrous sulfate crystals that possesses high demand and market value. Moreover, these prior arts do not provide an efficient method of treating jarosite to generate high economic value for the industry. 10
[0008] In light of the above stated discussion, there is a need for a method and system that overcomes the above stated disadvantages.
OBJECT OF THE DISCLOSURE 15
[0009] A primary object of the present disclosure is to provide a method for recovery of metal values from industrial waste.
[0010] Another object of the present disclosure is to provide a method for treating waste produced in zinc hydrometallurgical smelters.
[0011] Yet another object of the present disclosure is to generate values through 20 efficient treatment of the industrial waste.
[0012] Yet another object of the present disclosure is to provide a saleable divalent iron compound.
SUMMARY
[0013] In an aspect, the present disclosure provides a method of recovery of metal values and 25 saleable compound crystals from industrial waste. The method includes a first step of leaching the industrial waste under a pre-defined acid leaching condition. The industrial waste is leached at a first pre-defined temperature range of about 90 °C to 100 °C and for a first pre-defined time interval of about 1 hours to 6 hours. The method includes a second step of reducing an acid leached filtrate separated from the acid leached product. 30 The acid leached filtrate is reduced at a second pre-defined temperature range of about 70
5
°C to 95 °C and for a second pre-defined time interval of about 2 hours to 8 hours. Further, the method includes a third step of treating a ferric reduced filtrate with zinc concentrate. The ferric reduced product is treated at a third pre-defined temperature range of about 70 °C to 95 °C and for a third pre-defined time interval of about 2 hours to 8 hours. The method includes a fourth step of crystallizing an acidic reduced filtrate 5 separated from the acidic reduced product. The crystallization of the acidic reduced filtrate is performed to obtain a plurality of divalent iron compound crystals present in the acidic reduced filtrate. Furthermore, the pre-defined acid leaching condition corresponds to a pre-defined acidic level of about 100 g/l to 400 g/l. The industrial waste is leached to obtain an acid leached product. The reduction of the acid leached filtrate is performed by 10 utilizing metal sulfides sparingly soluble or insoluble in sulfates to obtain a reduced product. The ferric reduced filtrate is separated from the reduced product. The ferric reduced filtrate is treated with the zinc concentrate to obtain an acidic reduced product having an acidity level in a range of about 10 g/l to 50 g/l. Moreover, the plurality of divalent iron compound crystals obtained during crystallization has a pre-defined 15 percentage purity range of about 90 % to 95 %.
[0014] In an embodiment of the present disclosure, the industrial waste is a jarosite waste utilized at a pre-defined pulp density of about 100 g/l to 200 g/l in acid leaching stage.
20
[0015] In an embodiment of the present disclosure, the method further includes solid-liquid separation of the acid leached product to separate the acid leached filtrate and an acid leached residue. The acid leached residue has a first pre-defined quantity of valuable metals. The first pre-defined quantity of valuable metals includes about 6% to 10% by weight of lead, about 4% to 6% by weight of zinc, about 2% to 5% by weight of iron and 25 about 300 ppm to 1500 ppm of silver.
[0016] In an embodiment of the present disclosure, the method further includes leaching an acid leached residue by utilizing a caustic solution at a fourth pre-defined temperature range of about 25 °C to 50 °C. The caustic solution utilized for leaching has an alkalinity 30 of about 40 g/l to 200 g/l. The leaching of the acid leached residue is performed for a fourth pre-defined time interval of about 1 hours to 4 hours. Moreover, the leaching of
6
the acid leached residue is performed to obtain an alkaline leached product. The alkaline leached product includes an alkaline leached residue having a second pre-defined quantity of valuable metals. The second pre-defined quantity of valuable metals includes about 8% to 20% by weight of lead, about 5% to 15% by weight of zinc, about 3% to 15% by weight of iron and about 500 ppm to 4000 ppm of silver. The second pre-defined 5 quantity of valuable metals is recovered by feeding the alkaline leached residue to a pyro-metallurgical furnace
[0017] In an embodiment of the present disclosure, the metal sulphides used for ferric ion reduction of acid leached filtrate are lead concentrate or lead sulfide. The lead 10 concentrate or lead sulfide are in the excess of 0 - 50% of stoichiometric requirements.
[0018] In an embodiment of the present disclosure, the method further includes solid-liquid separation of the reduced product to separate the ferric reduced filtrate and a ferric reduced residue. The ferric reduced residue is collected for the recovery of metal values. 15
[0019] In an embodiment of the present disclosure, the zinc concentrate used for acidity reduction of ferric reduced filtrate is in the ratio of 1.5 - 5.0 g/g of acid.
[0020] In an embodiment of the present disclosure, the method further includes solid-liquid 20 separation of the acidic reduced product to separate the acidic reduced filtrate and an acidic reduced residue. The acidic reduced residue is collected for the recovery of metal values.
[0021] In an embodiment of the present disclosure, mother liquor obtained during 25 crystallization of the acidic reduced filtrate is recycled for further leaching or in reducing unit for reduction and crystallization.
STATEMENT OF THE DISCLOSURE
[0022] The present disclosure relates to a method of recovery of saleable compound 30 crystals from industrial waste. The method includes a first step of leaching the industrial
7
waste under a pre-defined acid leaching condition. The industrial waste is leached at a first pre-defined temperature range of about 90 °C to 100 °C and for a first pre-defined time interval of about 1 hours to 6 hours. The method includes a second step of reducing an acid leached filtrate separated from the acid leached product. The acid leached filtrate is reduced at a second pre-defined temperature range of about 70 °C to 95 °C and for a 5 second pre-defined time interval of about 2 hours to 8 hours. Further, the method includes a third step of treating a ferric reduced filtrate with zinc concentrate. The ferric reduced product is treated at a third pre-defined temperature range of about 70 °C to 95 °C and for a third pre-defined time interval of about 2 hours to 8 hours. The method includes a fourth step of crystallizing an acidic reduced filtrate separated from the acidic 10 reduced product. The crystallization of the acidic reduced filtrate is performed to obtain a plurality of divalent iron compound crystals present in the acidic reduced filtrate. Furthermore, the pre-defined acid leaching condition corresponds to a pre-defined acidic level of about 100 g/l to 400 g/l. The industrial waste is leached to obtain an acid leached product. The reduction of the acid leached filtrate is performed by utilizing lead 15 concentrate to obtain a reduced product. The ferric reduced filtrate is separated from the reduced product. The ferric reduced filtrate is treated with the zinc concentrate to obtain an acidic reduced product having an acidity level in a range of about 10 g/l to 50 g/l. Moreover, the plurality of divalent iron compound crystals obtained during crystallization has a pre-defined percentage purity range of about 90 % to 95 %. Furthermore, the acid 20 leached product may be treated in alkaline conditions to further improve the percentage of metal concentrations in residues.
BRIEF DESCRIPTION OF FIGURES
[0023] Having thus described the disclosure in general terms, reference will now be made to the accompanying figures, wherein: 25
[0024] FIG. 1 illustrates a process flow diagram for a recovery of saleable compound crystals from industrial waste, in accordance with an embodiment of the present disclosure;
[0025] FIG. 2A illustrates a system for recovery of saleable compound crystals from industrial waste, in accordance with an embodiment of the present disclosure; 30
8
[0026] FIG. 2B illustrates the system for the recovery of saleable compound crystals from industrial waste, in accordance with another embodiment of the present disclosure; and
[0027] FIG. 3 illustrates a flowchart for the recovery of saleable compound crystals from industrial waste, in accordance with various embodiment of the present disclosure. 5
[0028] It should be noted that the accompanying figures are intended to present illustrations of exemplary embodiments of the present disclosure. These figures are not intended to limit the scope of the present disclosure. It should also be noted that accompanying figures are not necessarily drawn to scale.
10
DETAILED DESCRIPTION
[0029] In the following description, for purposes of explanation, numerous specific details are set forth in order to provide a thorough understanding of the present technology. It will be apparent, however, to one skilled in the art that the present technology can be practiced without these specific details. In other instances, structures 15 and devices are shown in block diagram form only in order to avoid obscuring the present technology.
[0030] Reference in this specification to “one embodiment” or “an embodiment” means that a particular feature, structure, or characteristic described in connection with the 20 embodiment is included in at least one embodiment of the present technology. The appearance of the phrase “in one embodiment” in various places in the specification are not necessarily all referring to the same embodiment, nor are separate or alternative embodiments mutually exclusive of other embodiments. Moreover, various features are described which may be exhibited by some embodiments and not by others. Similarly, 25 various requirements are described which may be requirements for some embodiments but not other embodiments.
[0031] Although the following description contains many specifics for the purposes of illustration, anyone skilled in the art will appreciate that many variations and/or 30 alterations to said details are within the scope of the present technology. Similarly,
9
although many of the features of the present technology are described in terms of each other, or in conjunction with each other, one skilled in the art will appreciate that many of these features can be provided independently of other features. Accordingly, this description of the present technology is set forth without any loss of generality to, and without imposing limitations upon, the present technology. 5
[0032] FIG. 1 illustrates a process flow diagram 100 for a recovery of saleable compound crystals from industrial waste 102, in accordance with various embodiments of the present disclosure. Also, FIG. 1 illustrates the process flow diagram 100 for the recovery of metal values from the industrial waste 102. In an embodiment of the present 10 disclosure, the saleable compound crystal is ferrous sulfate crystals. Metal values include but may not be limited to a plurality of valuable metals and compounds. The method is configured for the recovery of the plurality of valuable metals by utilizing the industrial waste 102. The plurality of valuable metals includes but may not be limited to lead, zinc and silver. Moreover, the method is configured to produce a plurality of divalent iron 15 compound crystals by utilizing the industrial waste 102.
[0033] The industrial waste 102 is in the form of jarosite waste. In general, jarosite is a complex basic iron sulfate with a chemical formula of [NaFe3 (OH)6 (SO4)2]. In general, jarosite is obtained as the industrial waste during a hydrometallurgical refining of zinc 20 from zinc ore concentrate. In an embodiment of the present disclosure, jarosite is obtained in the form of precipitate. In another embodiment of the present disclosure, jarosite is obtained in the form of dust. In yet another embodiment of the present disclosure, jarosite is obtained in the form of sludge. In yet another embodiment of the present disclosure, jarosite is obtained in the form of residues. The industrial waste 102 25 utilized for the recovery of the plurality of valuable metals has a pre-defined chemical composition. The pre-defined chemical composition includes about 2.0 % to 4.0 % by weight of lead, about 1.0 % to 4.0 % by weight of zinc, about 100 to 300 parts-per million (hereinafter as “ppm”) of silver and about 22 % to 30 % by weight of iron.
30
10
[0034] Referring to FIG. 2A and FIG. 2B, a system 200 is configured to carry the process for recovery of the saleable compound crystals and metal values from the industrial waste 102. The system 200 includes an input unit 202, a leaching unit 204, a recovery unit 206, a reduction unit 208, a crystallization unit 210 and an output unit 212. The input unit 202 stores one or more input materials required for the recovery of the 5 plurality of valuable metals. In addition, the input unit 202 stores the one or more input materials required for the production of the plurality of divalent iron compound crystals. The one or more input materials include but may not be limited to jarosite, zinc concentrate, lead concentrate, zinc oxide, an acidic solution, and a basic or alkaline solution. Each of the one or more input materials are added in a pre-determined quantity 10 based on requirement for production of a pre-determined amount of the plurality of divalent iron compound crystals. The input unit 202 may be manually, mechanically or electrically operated. In an embodiment of the present disclosure, each of the one or more input materials is stored separately inside the input unit 202. The input unit 202 provides the one or more input materials required during the method for the recovery of 15 metal values.
[0035] Continuing with FIG. 1, the method includes a first step 104 of acid leaching the industrial waste 102. In general, the leaching of the industrial waste 102 is performed to extract the plurality of valuable metals. In an embodiment of the present disclosure, the 20 leaching unit 204 receives the industrial waste 102 from the input unit 202 (as shown in FIG. 2A and FIG. 2B). In addition, the leaching unit 204 receives the one or more input materials from the input unit 202. The industrial waste 102 received by the leaching unit 204 is in the form of jarosite waste having a pre-defined pulp density of about 100 g/l to 200 g/l, preferably about 130 g/l to 150 g/l. In general, pulp density associated with the 25 jarosite waste represents a pre-defined quantity of solid jarosite content per liter of liquid content. The leaching unit 204 includes an acid leaching unit 204a and a basic leaching unit 204b (as shown in FIG. 2A and FIG. 2B). In an embodiment of the present disclosure, the acid leaching unit 204a is utilized to perform the first step 104 of acid leaching of the industrial waste 102. 30
11
[0036] The leaching of the industrial waste 102 is performed under a pre-defined acid leaching condition at a first pre-defined temperature range for a first pre-defined time interval. In an embodiment of the present disclosure, the pre-defined acid leach condition is achieved by utilizing sulfuric acid solution. Moreover, the pre-defined acid leach condition corresponds to a pre-defined acidic level. In an embodiment of the present 5 disclosure, the pre-defined acidic level is 100 g/l to 400 g/l acidic, preferably about 180 g/l to 240 g/l. The first pre-defined temperature range for leaching the industrial waste 102 lies in a range of about 90 °C to 100 °C, preferably about 95 ºC to 100 ºC. In addition, the first pre-defined time interval for leaching the industrial waste 102 is about 1 hours to 6 hours, preferably about 2 hours to 4 hours. Moreover, the first step 104 of acid 10 leaching the industrial waste 102 is performed according to the following reaction:
[0037] 2 NaFe3 (SO4)2 (OH)6 + 6 H2SO4 → 3 Fe2 (SO4)3 + 12 H2O + Na2 SO4
[0038] ZnO.Fe2O3+ 4 H2SO4 → Fe2 (SO4)3 + ZnSO4 + 4 H2O
[0039] The first step 104 of acid leaching is performed to obtain an acid leached 15 product. In an embodiment of the present disclosure, the acid leached product is obtained in the form of a heterogeneous mixture. The acid leached product obtained during the first step 104 of acid leaching has a first pre-defined terminal acidity. In an embodiment of the present disclosure, the first pre-defined terminal acidity corresponding to the acid leached product is in a range of 70 g/l to 300 g/l. Further, an intermediate step of solid-20 liquid separation of the acid leached product is performed to separate an acid leached filtrate and an acid leached residue from the acid leached product.
[0040] The acid leached filtrate includes a trivalent iron compound, a divalent zinc compound. The trivalent iron compound and the divalent zinc compound are derivatives 25 of sulfate ion. In an embodiment of the present disclosure, the trivalent iron compound is ferric sulfate. In an embodiment of the present disclosure, the divalent zinc compound is zinc sulfate. The acid leached residue includes the plurality of valuable metals agglomerated after the solid-liquid separation. The acid leached residue has a first pre-defined quantity of valuable metals. In an embodiment of the present disclosure, the first 30 pre-defined quantity of valuable metals is about 6 % to 10 % by weight of lead, about 4 %
12
to 6 % by weight of zinc, about 2 % to 5 % by weight of iron and about 300 ppm to 1500 ppm of silver.
[0041] Further, the plurality of valuable metals agglomerated after the solid-liquid separation is filtered and collected. In an embodiment of the present disclosure, the acid 5 leached residue is filtered and collected from the acid leaching unit 204a and allowed to transfer directly to the recovery unit 206 (as shown in FIG. 2A). In another embodiment of the present disclosure, the acid leached residue is transferred from the acid leaching unit 204a to the basic leaching unit 204b for performing basic leaching (as shown in the FIG. 2B). The acid leached residue is further transferred to the basic leaching unit 204b 10 to improve the quantity of the plurality of valuable metals.
[0042] Going further, the method includes a second step 106 of basic leaching the acid leached residue. The second step 106 of basic leaching is performed under an alkaline condition. In an embodiment of the present disclosure, the basic leaching unit 204b is 15 utilized for performing the second step 106 of basic leaching. The basic leaching unit 204b includes a caustic solution having an alkalinity of about 40 g/l to 200 g/l, preferably 80 g/l to 100 g/l. The second step 106 of basic leaching is performed by utilizing a pre-determined quantity of the caustic solution at a fourth pre-defined temperature range. In an embodiment of the present disclosure, the fourth pre-defined temperature range is 20 about 25 °C to 50 °C, preferably 30 °C to 40 °C. In addition, the second step 106 of basic leaching the acid leached residue is performed for a fourth pre-defined time interval. In an embodiment of the present disclosure, the fourth pre-defined time interval is about 1 hours to 4 hours, preferably 2 hours to 3 hours.
25
[0043] The second step 106 of basic leaching of the acid leached residue is performed to obtain alkaline leached product. In an embodiment of the present disclosure, the alkaline leached product is obtained in the form of a heterogeneous mixture. Further, an intermediate step of solid-liquid separation is performed to separate out an alkaline leached filtrate and an alkaline leached residue from the alkaline leached product. The 30 alkaline leached filtrate includes the traces of trivalent iron compound and the divalent
13
zinc compound. The second step 106 of basic leaching is performed repetitively to obtain the alkaline leached product. Soluble silica level increases during repetitively recycling of liquid in second step 106; resisting the basic leaching and solid-liquid separation. In an embodiment of the present disclosure, soluble silica level is decreased by recycling back the alkaline leached filtrate to basic leaching unit 204b. The recycling back of the 5 alkaline leached filtrate to basic leaching unit 204b is done after a conventional silica precipitation process known to those skilled in prior art and adding a required quantity of caustic soda for maintaining the basicity. In another embodiment of the present disclosure, soluble silica level is decreased by performing part bleeding the alkaline leached filtrate to effluent treatment plants followed by the addition of make up caustic in 10 remaining filtrate recycled for second step 106.
[0044] The alkaline leached residue includes the plurality of valuable metals. The plurality of valuable metals associated with the alkaline leached residue is agglomerated after the solid-liquid separation. In an embodiment of the present disclosure, the alkaline 15 leached residue is filtered and collected from the basic leaching unit 204b. The alkaline leached residue has a second pre-defined quantity of valuable metals. The second pre-defined quantity of valuable metals consist of about 8 % to 20 % by weight of lead, about 5 % to 15 % by weight of zinc, about 3 % to 10 % by weight of iron and about 500 ppm to 4000 ppm of silver. 20
[0045] Further, the alkaline leached residue containing the plurality of valuable metals is subjected to a third step 108 of recovery. In an embodiment of the present disclosure, the third step 108 of recovery is performed by the recovery unit 206 (as shown in FIG. 2A and FIG. 2B). The recovery unit 206 receives the alkaline leached residue containing the 25 plurality of valuable metals from the leaching unit 204. In an embodiment of the present disclosure, the recovery unit 206 receives the acid leached residue containing the plurality of valuable metals direct from the acid leaching unit 204a. In another embodiment of the present disclosure, the recovery unit 206 receives the alkaline leached residue containing the plurality of valuable metals direct from the basic leaching unit 204b. Furthermore, 30 the recovery unit 206 utilizes a pyro-metallurgical furnace for the recovery of the metal values.
14
[0046] Going further, the acid leached filtrate is collected after the intermediate step of solid-liquid separation. In an embodiment of the present disclosure, the filtrate is collected as the acid leached filtrate from the acid leaching unit 204a. The acid leached filtrate collected includes the trivalent iron compound and the divalent zinc compound. In 5 an embodiment of the present disclosure, the trivalent iron compound is ferric sulfate and the divalent zinc compound is zinc sulfate. Moreover, the trivalent iron compound and the divalent zinc compound are homogenously dissolved in the acid leached filtrate. In an embodiment of the present disclosure, the acid leached filtrate obtained from the acid leaching unit 204a is transferred to the reduction unit 208. 10
[0047] Further, reduction of the acid leached filtrate separated from the acid leached product is performed. The method includes a fourth step 110 of ferric reduction of the acid leached filtrate separated from the acid leached product. In an embodiment of the present disclosure, the reduction unit 208 is utilized for reducing the acid leached filtrate. 15 The reduction unit 208 includes a ferric reduction unit 208a and an acidity reduction unit 208b (as shown in FIG. 2A and FIG. 2B). In an embodiment of the present disclosure, lead concentrate (sulfidic in nature) is utilized as a reducing agent for the ferric reduction of the acid leached filtrate. Lead concentrate is utilized as the reducing agent in a pre-defined ratio higher than stoichiometric calculations. In an embodiment of the present 20 disclosure, the pre-defined ratio of lead concentrate utilization is 0% to 50% higher than the stoichiometric calculations. In an embodiment of the present disclosure, the acid leached filtrate subjected to reduction has the first pre-defined terminal acidity of about 70 g/l to 300 g/l.
25
[0048] The fourth step 110 of ferric reduction of the acid leached filtrate is performed at a second pre-defined temperature range for a second pre-defined time interval. The second pre-defined temperature range for reducing the acid leached filtrate is about 70 °C to 95 °C, preferably 80 °C to 90 °C. Moreover, the second pre-defined time interval for reducing the acid leached filtrate is about 2 hours to 8 hours, preferably about 4 hours to 6 30 hours. The lead concentrate as the reducing agent reduces the trivalent iron compound
15
present in the acid leached filtrate. Moreover, the lead concentrate reduces the trivalent iron compound to obtain a divalent iron compound.
[0049] The ferric reduction unit 208a performs reduction of the acid leached filtrate to obtain a reduced product. The reduced product obtain in the ferric reduction unit 208a 5 includes the divalent iron compound, the divalent zinc compound and metal values. In an embodiment of the present disclosure, the divalent iron compound is ferrous sulfate and the divalent zinc compound is zinc sulfate. Moreover, the reduced product has the first pre-defined terminal acidity of 70 g/l to 300 g/l. The fourth step 110 of ferric reduction of the acid leached filtrate is performed according to the following reaction: 10
[0050] PbS + Fe2 (SO4)3 → PbSO4 + 2 FeSO4 + S
[0051] Further, the reduced product obtained during the ferric reduction step is a heterogeneous mixture containing a ferric reduced filtrate and a ferric reduced residue. The method includes an intermediate step of solid-liquid separation of the reduced 15 product to separate the ferric reduced filtrate and the ferric reduced residue. The solid-liquid separation of the reduced product is performed through a suitable separation media. The suitable separation media is utilized for separating the ferric reduced filtrate and the ferric reduced residue. The ferric reduced filtrate includes the divalent iron compound and the divalent zinc compound. In addition, the ferric reduced residue includes metal 20 values. The ferric reduced residue is further transferred for the recovery of metal values. In an embodiment of the present disclosure, the ferric reduced residue is collected from the suitable separation media associated with the ferric reduction unit 208a and transferred to the recovery unit 206. Furthermore, the ferric reduced filtrate obtained after the solid-liquid separation is transferred for further reduction. 25
[0052] The ratio is calculated considering sulphides in reducing agent. In an embodiment of the present disclosure, the reducing agent is lead concentrate which include galena (PbS). In another embodiment of the present disclosure, the sulfidic material is zinc concentrate. In yet another embodiment of the present disclosure, the 30 sulfidic material is ZnS. In yet another embodiment of the present disclosure, the sulfidic
16
material is Fe
xSy. Moreover, the cationic sulphides which are sparingly soluble or insoluble in sulphate media are selected to avoid impurity addition.
[0053] The method includes a fifth step 112 of acidic reduction of the ferric reduced filtrate. The fifth step 112 of acidic reduction is performed by treating the ferric reduced 5 filtrate with zinc concentrate. The fifth step 112 of acidic reduction is performed for reducing the acidity of the ferric reduced filtrate. In an embodiment of the present disclosure, the acidity reduction unit 208b is utilized for performing the fifth step 112 of acidic reduction. The ferric reduced filtrate obtained after the solid-liquid separation is transferred to the acidity reduction unit 208b. In an embodiment of the present 10 disclosure, zinc concentrate (ZnS) is utilized as the acid neutralizing agent in the acidity reduction unit 208b due to cheaper and easily available source. In another embodiment of the present disclosure, zinc oxide is utilized as the acid neutralizing agent in the acidity reduction unit 208b. In yet another embodiment of the present disclosure, suitable metal oxides are utilized as the acid neutralizing agent in the acidity reduction unit 208b. 15
[0054] The ferric reduced filtrate is treated with the zinc concentrate in the acidity reduction unit 208b to obtain the acidic reduced product. In an embodiment of the present disclosure, zinc concentrate is utilized in a pre-defined quantity in a range of about 1.5 to 5.0 grams per gram of acid, preferably about 2.0 to 4.0 grams per gram of 20 acid. The acidity reduction unit 208b treats the reduced product at a third pre-defined temperature range for a third pre-defined time interval. The third pre-defined temperature range for treating the ferric reduced filtrate is about 70 °C to 95 °C, preferably 80 °C to 90 °C. In addition, the third pre-defined time interval for treating the ferric reduced filtrate is about 2 hours to 8 hours, preferably 4 hours to 6 hours. Furthermore, the acidic 25 reduced product has a second pre-defined terminal acidity. In an embodiment of the present disclosure, the second pre-defined terminal acidity is 10 g/l to 50 g/l, preferably 20 g/l to 30 g/l. The fifth step 112 of acidic reduction is performed according to the following reaction:
[0055] ZnO + H2SO4 → ZnSO4 + H2O 30
17
[0056] The acidic reduced product obtained during the acidic reduction step is a heterogeneous mixture containing an acidic reduced filtrate and an acidic reduced residue. The method includes an intermediate step of solid-liquid separation of the acidic reduced product to obtain the acidic reduced filtrate and an acidic reduced residue. The solid-liquid separation of the acidic reduced product is performed through a suitable separation 5 media. The suitable separation media is utilized for separating the acidic reduced filtrate and the acidic reduced residue. The acidic reduced residue is further transferred for the recovery of the metal values. In an embodiment of the present disclosure, the acidic reduced residue is collected from the suitable separation media associated with the acidity reduction unit 208b and transferred to the recovery unit 206. Furthermore, the ferric 10 reduced filtrate obtained after the solid-liquid separation is transferred for crystallization.
[0057] Further, the method includes a sixth step 114 of crystallizing the acidic reduced filtrate separated from the acidic reduced product. In an embodiment of the present disclosure, the acidic reduced filtrate is transferred from the reduction unit 208 to the 15 crystallization unit 210 (as shown in FIG. 2A and FIG. 2B). The crystallization unit 210 receives the acidic reduced filtrate. The acidic reduced filtrate includes the divalent iron compound and traces of the divalent zinc compound. The crystallization unit 210 performs crystallization of the acidic reduced filtrate for forming plurality of divalent iron compound crystals. Furthermore, the crystallization unit 210 includes an evaporator and 20 a crystallizer for performing the crystallization of the divalent iron compound. Moreover, mother liquor obtained during the crystallization of the acidic reduced filtrate is recycled and sent back to the leaching unit 204a as leaching solution when the iron levels are low. Moreover, the flowsheet has an option that the same mother liquor can also be recycled to reduction unit 208 for necessary reductions and followed by crystallization. 25
[0058] Referring to FIG. 2A and FIG. 2B, the output unit 212 is connected to the crystallization unit 210. In an embodiment of the present disclosure, the output unit 212 is connected to the crystallization unit 210 through any connecting medium. In another embodiment of the present disclosure, the output unit 212 is mechanically connected to 30 the crystallization unit 210. In addition, the output unit 212 is configured to collect a final output from the crystallization unit 210 in the form of crystals. In an embodiment of the
18
present disclosure, the output unit 212 collects the plurality of divalent iron compound crystals. In addition, the output unit212 is configured to store the plurality of divalent iron compound crystals in a storage space of the output unit 212. The output unit 212 may be any type of unit or container having a pre-defined amount of space for the storage of the plurality of divalent iron compound crystals. Also, the output unit 212 may be 5 operated manually, mechanically or electrically for enabling the storage of the plurality of divalent iron compound crystals.
[0059] It may be noted that in FIG. 2A and FIG. 2B, the input unit 202 is connected to the leaching unit 204; however, those skilled in the art would appreciate that there are 10 more number of input units connected to the leaching unit 204 for providing the one or more input materials. Also, It may be noted that in FIG. 2A and FIG. 2B, the output unit 212 is connected to the crystallization unit 210 for collecting the plurality of divalent iron compound crystals; however, those skilled in the art would appreciate that there are more number of output units connected to the crystallization unit 210 for collecting the 15 plurality of divalent iron compound crystals.
[0060] FIG. 3 illustrates a flowchart 300 for the recovery of the metal values from the industrial waste, in accordance with various embodiments of the present disclosure. It may be noted that to explain the process steps of flowchart 300, references will be made 20 to the system elements of the FIG. 1, the FIG. 2A, and the FIG. 2B. Also, it may be noted that the flowchart 300 may have lesser or more number of steps.
[0061] The flowchart 300 initiates at step 302. Following step 302, at step 304, the leaching unit 304 is allowed to perform the leaching of the industrial waste under the pre-25 defined acid leaching condition at the first pre-defined temperature range for the first pre-defined time interval. The industrial waste is subjected to leaching to obtain the acid leached product. At step 306, the ferric reduction unit 308a is allowed to reduce the acid leached filtrate separated from the acid leached product. The acid leached filtrate is reduced to obtain the reduced product. Further, at step 308, the acidic reduction unit 30 308b is allowed to treat the ferric reduced filtrate with the zinc concentrate. The ferric
19
reduced filtrate is treated with the zinc concentrate to control acidity to achieve suitable ferrous ions concentration. Further, at step 310, the crystallization unit 310 is allowed to crystallize the acidic reduced filtrate separated from the acidic reduced product. The acidic reduced filtrate is crystallize to obtain the plurality of divalent iron compound crystals. The flowchart 300 terminates at step 312. 5
[0062] The present invention is described below in greater details with the aid of examples. The following examples of the technical aspect of the present invention will be further explained, but the content of the following is not intended to limit the scope of the present invention. 10
[0063] Example 1 illustrates about the acidic leaching of the industrial waste. Example 2 illustrates about the basic leaching of the acid leached residue. Example 3 illustrates about the reduction of the acid leached filtrate separated from the acid leached product. Also, example 3 illustrates about the treatment of the ferric reduced filtrate to control 15 acidity. Further, example 4 illustrates about the crystallization of the divalent iron compound present in the acidic reduced filtrate.
[0064] Example 1
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[0065] In the acid leaching unit 204a, 150 grams (hereinafter as “g”) of industrial waste sample was leached in a 2 litre beaker with a working volume of 1 litre at initial acidity of 100 g/l to 400 g/l with sulfuric acid. The industrial waste sample utilized for acidic leaching includes about 2.98% zinc by weight, about 3.48% lead by weight, about 210 parts-per million of silver and about 25% iron by weight. In addition, the industrial waste 25 sample has a pulp density of about 100 g/l to 200 g/l. The acid leaching of the industrial waste product is performed for about 2 to 6 hours. Moreover, the acidic leaching of the industrial waste sample was carried out at atmospheric pressure. In addition, near boiling reaction temperature is more than 90°C during the acidic leaching of the industrial waste sample. The leaching efficiency of zinc lies in a range of about 12.7% to 69.5% by 30 weight. In addition, the leaching efficiency of iron lies in a range of about 59.6% to 95.7% by weight. Moreover, all the lead and silver content of the industrial waste sample
20
agglomerated as the residue. The residue assay includes about 3.40% to 4.48% zinc by weight, about 4.80% to 12.40 % lead by weight, about 410 to 880 ppm of silver and about 4.50% to 20.75 % iron by weight. The residue obtained during the acidic leaching of the industrial waste sample can be used in basic leaching for further enrichment. The leached filtrate contains 21.25 g/l to 37.69 g/l total iron content in the form of trivalent iron 5 compound. The trivalent iron compound is ferric sulphate.
[0066] Example 2
[0067] In the basic leaching unit 204b, 150 g of the residue of the acid leached product 10 (similar to residue of Example 1) was leached in 2 liter beaker with 1 liter working volume at a temperature range of 30 ºC to 50 ºC and at atmospheric pressure. Leachant having 40 g/l to 200 g/l alkalinity was prepared with Caustic soda and leaching was carried out for about 1 to 4 hours. The alkaline leached product containing residue is obtained during the basic leaching. The residue assay includes about 5.50% to 9.80% 15 zinc by weight, about 13.10% to 23.58% lead by weight, about 1050 to 2200 ppm of silver and about 8.50 % to 11.10 % iron by weight. Furthermore, the residue assay was allowed to pass through the recovery unit 206 in which a set of pyro-metallurgical operations were performed to extract valuable metals from the residue. In addition, filtrate is obtained during basic leaching. The filtrate includes silica as a major element. 20 Moreover, the filtrate was recycled for basic leaching by precipitating or bleeding silica.
[0068] Example 3
[0069] The filtrate of the acid leached product (same as described in Example 1) 25 contains trivalent iron compound. The trivalent iron compound is ferric sulfate. In the ferric reduction unit 208a, about 1 liter acid leached product was utilized for ferric reduction. The filtrate of the acid leached product includes about 37.69 g/l of ferric ion concentration (hereinafter as “Fe+++”). In addition, the filtrate of the acid leached product includes about 1.80 g/l of ferrous ion concentration (hereinafter as “Fe++”). 30 Moreover, the filtrate of terminal acidity of about 175 g/l was utilized. Lead concentrate is utilized as the reducing agent in a pre-defined ratio higher than the stoichiometric concentration. In an embodiment of the present disclosure, the pre-defined ratio of lead
21
concentrate utilization is 0% to 35% higher than the stoichiometric concentration. The 75g to 113g of lead concentrate, which includes up to 5% zinc by weight, 56 - 65% lead by weight and about 500 ppm to 3000 ppm of silver was added as reducing agent. The reaction was carried out to convert trivalent iron compound into divalent iron compound with the conversion rate of about 62 % to 93%. In addition, the reaction was carried out 5 at the atmospheric pressure and at temperature range of 70 °C to 90 °C. The reduced product was obtained in the form of filtrate and residue. The filtrate associated with the reduced product consists of about 23 g/l to 35 g/l of ferrous with output acidity of about 112 to 133 g/l. In addition, minor reduction in solution acidity was observed.
10
[0070] Further, the filtrate associated with the reduced product was treated with zinc concentrate. The reaction was carried out by adding about 1350 g to 2400 g zinc concentrate by weight to the reduced product at the atmospheric pressure and in the temperature range of about 80 °C to 85 °C. In addition, the terminal acidity was maintained between 20 g/l to 40 g/l throughout the reaction. Moreover, the oxidic content 15 in zinc concentrate was leached out and the remaining zinc concentrate was sent back to zinc smelter for metal recovery. The filtrate obtained contains about 30 g/l to 50 g/l of ferrous concentration (hereinafter as “Fe++”) suitable for ferrous sulfate crystallization.
[0071] Example 4 20
[0072] To generate ferrous sulphate crystals, the 1 liter filtrate containing ferrous ions (similar to Example 3) was evaporated in the crystallization unit 210. The filtrate was evaporated to increase iron content up to 90 g/l to 125g/l. The super saturated solution obtained during the evaporation was then cooled at room temperature in crystallizer and 25 ferrous sulphate crystals were obtained and were separated by centrifugation. About 89-113 g ferrous sulphate hepta-hydrate crystals of more than 90% purity was generated from 1liter liquor. The first pass iron recovery was more than 80% during the crystallization. Moreover, mother liquor consisting of about 20 g/l to 22 g/l ferrous was recycled. The nature of ferrous sulfate crystals obtained is FeSO4.7H2O that was further 30 converted in to other forms of FeSO4 crystals up to anhydrous crystals.
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[0073] The present disclosure provides several advantages over a prior art. The present disclosure provides a method for recovery of metal values from industrial waste. In addition, the present disclosure ergonomically treats the industrial waste to generate metal values. The present disclosure reduces the zinc industry waste by utilizing the zinc industry waste to produce valuable products. Furthermore, the present disclosure 5 provides an efficient process for manufacturing saleable iron compound possessing high market value thus increasing the revenue of the zinc industry.
[0074] The foregoing descriptions of specific embodiments of the present technology have been presented for purposes of illustration and description. They are not intended to 10 be exhaustive or to limit the present technology to the precise forms disclosed, and obviously many modifications and variations are possible in light of the above teaching. The embodiments were chosen and described in order to best explain the principles of the present technology and its practical application, to thereby enable others skilled in the art to best utilize the present technology and various embodiments with various 15 modifications as are suited to the particular use contemplated. It is understood that various omissions and substitutions of equivalents are contemplated as circumstance may suggest or render expedient, but such are intended to cover the application or implementation without departing from the spirit or scope of the claims of the present technology. 20
[0075] While several possible embodiments of the invention have been described above and illustrated in some cases, it should be interpreted and understood as to have been presented only by way of illustration and example, but not by limitation. Thus, the breadth and scope of a preferred embodiment should not be limited by any of the above-25 described exemplary embodiments.

CLAIMS
What is claimed is:
1. A method of recovery of metal values and saleable compound crystals from industrial waste, the method comprising:
leaching the industrial waste under a pre-defined acid leaching condition at a 5 first pre-defined temperature range for a first pre-defined time interval, wherein the pre-defined acid leaching condition is achieved by utilizing sulfuric acid solution, wherein the pre-defined acid leaching condition corresponds to a pre-defined acidic level of about 100 g/l to 400 g/l, wherein the first pre-defined temperature range is about 90 °C to 100 °C, wherein the first pre-defined time interval is about 1 hours to 6 10 hours and wherein the industrial waste is leached to obtain an acid leached product;
reducing an acid leached filtrate separated from the acid leached product, wherein the reduction of the acid leached filtrate is performed at a second pre-defined temperature range for a second pre-defined time interval, wherein the second pre-defined temperature range is about 70 °C to 95 °C, wherein the second pre-defined 15 time interval is about 2 hours to 8 hours and wherein the reduction of the acid leached filtrate is performed by utilizing metal sulfides sparingly soluble or insoluble in sulfates to obtain a reduced product;
treating a ferric reduced filtrate with zinc concentrate, wherein the ferric reduced filtrate is separated from the reduced product, wherein the treatment of the 20 ferric reduced filtrate is performed at a third pre-defined temperature range for a third pre-defined time interval, wherein the third pre-defined temperature range is about 70 °C to 95 °C, wherein the third pre-defined time interval is about 2 hours to 8 hours and wherein the ferric reduced filtrate is treated with the zinc concentrate to obtain an acidic reduced product having an acidity level in a range of about 10 g/l to 50 g/l; and 25
crystallizing an acidic reduced filtrate separated from the acidic reduced product, wherein the crystallization is performed to obtain a plurality of divalent iron compound crystals present in the acidic reduced filtrate and wherein the plurality of divalent iron compound crystals has a pre-defined percentage purity range of about 90 % to 95 %. 30
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2. The method as recited in claim 1, wherein the industrial waste is a jarosite waste utilized at a pre-defined pulp density of about 100 g/l to 200 g/l in acid leaching stage.
3. The method as recited in claim 1, further comprising solid-liquid separation of the 5 acid leached product to separate the acid leached filtrate and an acid leached residue, wherein the acid leached filtrate mainly containing a trivalent iron compound and a divalent zinc compound, wherein the acid leached residue has a first pre-defined quantity of valuable metals and wherein the first pre-defined quantity of valuable metals comprises about 6% to 10% by weight of lead, about 4% to 6% by weight of 10 zinc, about 2% to 5% by weight of iron and about 300 ppm to 1500 ppm of silver.
4. The method as recited in claim 1, further comprising leaching an acid leached residue by utilizing a caustic solution at a fourth pre-defined temperature range of about 25 °C to 50 °C, wherein the caustic solution utilized for leaching has an alkalinity of 15 about 40 g/l to 200 g/l, wherein the leaching of the acid leached residue is performed for a fourth pre-defined time interval of about 1 hours to 4 hours with initial pulp density of 100 – 200 g/l, wherein the leaching of the acid leached residue is performed to obtain an alkaline leached product, wherein the alkaline leached product comprises an alkaline leached residue having a second pre-defined quantity of valuable metals, 20 wherein the second pre-defined quantity of valuable metals comprises about 8% to 20% by weight of lead, about 5% to 15% by weight of zinc, about 3% to 15% by weight of iron and about 500 ppm to 4000 ppm of silver and wherein the second pre-defined quantity of valuable metals is recovered by feeding the alkaline leached residue to a conventional pyro-metallurgical furnace. 25
5. The method as recited in claim 1, wherein the metal sulfide used for ferric ion reduction of acid leach filtrate is lead concentrate or lead sulfide, wherein the lead concentrate or lead sulfide are in the excess of 0 – 50% of stoichiometric requirements. 30
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6. The method as recited in claim 1, further comprising solid-liquid separation of the reduced product to separate the ferric reduced filtrate mainly containing a divalent iron compound and a ferric reduced residue, wherein the ferric reduced residue is collected for the recovery of metal values.
5
7. The method as recited in claim 1, wherein the zinc concentrate used for acidity reduction of ferric reduced filtrate is in the ratio of 1.5 – 5.0 g/g of acid.
8. The method as recited in claim 1, further comprising solid-liquid separation of the acidic reduced product to separate the acidic reduced filtrate and an acidic reduced 10 residue, wherein the acidic reduced residue is collected for the recovery of metal values.
9. The method as recited in claim 1, wherein mother liquor obtained during crystallization of the acidic reduced filtrate is recycled for further leaching or in 15 reducing unit for reduction & crystallization.