AU6770387A - Process for recovering metal values from jarosite solids - Google Patents

Description

PROCESS FOR RECOVERING METAL VALUES FROM JAROSITE SOLIDS

Field of the Invention

This invention relates to a process for recovering metal values from the jarosite containing wastes from electrolytic zinc recovery plants.

Background of the Invention

Electrolytic zinc processes are used to treat complex zinc-containing ores that cannot readily be treated by pyrometallurgical recovery. The usual steps in such an electrolytic process include: (a) concentrating the zinc ores; (b) roasting the zinc concentrate to eliminate sulfur and produce zinc calcine; (c) leaching the zinc calcine to provide an impure zinc sulfate solution; (d) separating the iron present usually by forming a jarosite precipitate; (e) purifying the zinc sulfate solution; and (f) subjecting the zinc sulfate solution to. electrolysis_< to recover the zinc -metal. Such a process is described in U.S. Patent 4, 128,617 (1978) of DeGuire et al. which is incorporated herein by reference. A simplified process flow sheet for an electrolytic zinc plant is shown in Fig. 1.

Additional details of various process modifications can be found in "The Encyclopedia of Chemical Tech¬ nology", Kirk-Oth er, Vol. 24, pp. 812-824, 3rd Ed, incorporated herein by reference.

As indicated above, a common method of removing the iron present in the leachate is through the formation of a "jarosite" precipitate. Jarosite, MFe₃ (SO.) ₂(OH)_β where M is a monovalent ion, usually an alkali metal (general sodium or potassium) or ammonium, is commonly formed by adding a source of ammonium or sodium ions to the leach solution and maintaining the solution at an appropriate pH by the addition of base. This process is also shown in Fig. 1. Various modifications to the so-called "jarosite process" are discussed in the article entitled "The Jarosite Process - Past, Present and Future", V. Arregui et al. , Lead-Zinc-Tin, TMS-AIME World Symposium on Metallurgy and Environmental Control, 1980, JM Cigon, T.S. Mackey and T.J. O'Keefe, Ed., pp. 97-123, incorporated herein by reference. 5 There are a number of problems associated with the formation of jarosite waste material. The jarosite can contain valuable metals such as silver, zinc, copper, lead, indium, etc. which require numerous expensive process steps to recover. The jarosite can also contain 0 toxic species which can be leached into the environment by rain and groundwater. Therefore to avoid environ¬ mental contamination it is usually necessary to store the jarosite wastes in sealed lagoons which are expensive to build. ₅ A number of methods for treating such wastes have been disclosed. Steintveit et al. in Norwegian Patent 142,406 (1980) disclose a process for leaching iron- containing waste with chloride-containing, acidic solu¬ tion at a temperature between 50°C and the boiling point _Q of the solution. An alkali metal chloride or an alkaline earth metal chloride is used as the source of the chlor¬ ide with calcium chloride being disclosed as the pre¬ ferred material. The pH of the solution is adjusted during the leaching step so that the iron remains as a ₅ precipitate while the valuable metals are leached into the hot solution. The pH is adjusted to between about 2 and 4 preferably with calcium hydroxide. It is disclosed that these conditions allow extraction of up to 95 percent of the lead and silver content of the waste. _Q However, Applicants have found that unpredictably with some jarosite wastes this process provides recoveries of less than 20 percent of the silver present.

U.S. Patent No. 4,054,638 of Dreulle et al. (1977) is directed to a process for recovering metals from 5 sulfated residues from electrolytic zinc plants. The residue is digested preferably at a temperature between 95° and 115°C with hydrochloric acid in the presence of calcium chloride. This leaching process dissolves the metals present, including the iron, by forming the corresponding metal chlorides. Consequently, the process requires that the iron chloride be removed by extraction by an organic solvent. This process has a disadvantage of solubilizing the iron and requiring a separate separa¬ tion step. There is no suggestion or disclosure of using superatmospheric pressure for this leach.

U.S. Patent 4,070,437 of Van Ceulen (1978), dis¬ closes a process for recovery of metals from jarosite sludges. The process involves leaching the jarosite with an acidic calcium chloride solution, preferably formed by mixing hydrochloric acid and calcium hydroxide or calcium carbonate. The leaching is preferably carried out close to the boiling point of the leaching medium. Insoluble calcium sulfate is formed and is separated by filtration. This process has the disadvantage of solubilizing essen¬ tially all of the iron in the jarosite.

Another waste which contains metal values is zinc ferrite-containing materials. Modern electrolytic zinc processes commonly use a two-step leaching process as depicted in Fig. 1. The second leaching step involves a hot acid leach to dissolve zinc ferrite present. How¬ ever, prior to the development of the two-step leach, a single neutral leach was used which caused much of the zinc ferrite and associated metal values to be discarded as wastes. Therefore, there are existing waste lagoons which contain substantial quantities of zinc ferrite and other metal values. The term "ferrite" is used herein to refer to a combined metal oxide-ferric oxide material, e.g. zinc ferrite (ZnO'Fe-O-.) .

A number of processes have been developed for the purpose of recovering this zinc. One such process is disclosed by Rastas et al. in U.S. Patent No. 3,959,437 (1976) . Rastas et al. disclose a process in which the ferrite of a non-ferrous metal, as well as the oxide of the non-ferrous metal, is subjected to a neutral leach which dissolves most of the oxide but leaves the ferrite substantially unaffected. The non-ferrous values in the solution are recovered and the undissolved ferrite material is further treated in a "conversion" stage with sulfuric acid-bearing solution at atmospheric pressure and at a temperature of about 80°C to about 105°C in the presence of alkali or ammonium ions. Under these condi¬ tions, the non-ferrous metals dissolve as sulfates, while iron is simultaneously precipitated as an insoluble complex sulfate, i.e., jarosite. U.S. Patent No. 4,355,005 of Rastas et al. (1982), U.S. Patent No. 4,366,127 of Rastas et al. (1982), as well as U.S. Patent No. 4,383,979 of Rastas et al. (1983) each disclose modifications to the process disclosed in the '437 patent.

Steintveit in U.S. Patent 3,684,490 (1972) discloses a method for treating jarosite residue in which the residue is subjected to leaching with sulfuric acid at a temperature of 50° to 95°C and an acid concentration of 10 to 70 grams per liter (hereinafter g/1) . These leaching conditions are intended to decompose any zinc ferrites present and provide for a greater recovery of the zinc.

U.S. Patent No. 3,691,038 of Von Roepenack et al. (1972) discloses a method for recovering zinc from oxides containing zinc and iron. The oxide is leached with sulfuric acid at a temperature of 95° to 100°C with an excess of sulfuric acid to solubilize the zinc and iron. Alkali metal or ammonium ions are added to the liquid phase along with a zinc-containing oxidic material at a temperature of 95° to 100°C to precipitate jarosite.

U.S. Patent No. 4,192,852 of Pammenter et al. (1980) discloses a process for treating zinc plant residues containing zinc ferrite and precipitating the iron as a jarosite. The sulfate solution containing ferric iron, free acid and non-ferrous metals is cooled, partially neutralized and then heated to a temperature not exceeding the boiling point at atmospheric pressure in the presence of sodium, potassium or ammonium ions. U.S.

Patent No. 4,305,914 of Pammenter et al. (1981) discloses a process similar to that in the '852 patent.

U.S. Patent 4,128,617, of DeGuire et al. (1978), describes a three-step process for the treatment of zinc calcine containing zinc oxide, zinc sulfates, and zinc ferrites. The first step involves the neutral leaching of the zinc calcine with an effective amount of aqueous sulfuric acid containing solution. The leach residue is subjected to hot acid leaching with sulfuric acid fol¬ lowed by jarosite precipitation by alkali with the subsequent recycling of the jarosite-containing pulp. The preferred temperature range for the hot acid leaching is from about 80°C to the boiling point and preferably the temperature is greater than 90°C. There is no suggestion of the use of pressure.

The processes described hereinabove have one or more of the disadvantages of (1) having low rates of extrac¬ tion of metal values, (2) providing low levels of re¬ covery of certain metal values, (3) having poor filter- ability of the iron-containing residue, and/or (4) solubilizing large amounts of iron.

Several processes disclosed in the art have used elevated temperature and pressure leaching steps in treating zinc plant residues.

U.S. Patent No. 3,143,486 of Pickering et al. (1964) discloses a process for the extraction of zinc from zinc ferrite containing residue. The process involves sub¬ jecting the residue to a first-stage leaching treatment under non-oxidizing conditions in a^' closed vessel in the presence of excess sulfuric acid at a temperature between 140°C and 260°C. Zinc is dissolved as well as ferrous sulfate which is stable at the temperatures and acidities used. Ferric iron is precipitated as a basic sulfate. The leachate is then subjected to a second-stage leaching treatment at 140°C to 260°C under oxidizing conditions to oxidize the ferrous sulfate to ferric and precipitate the ferric material. Similarly, U.S. Patent No. 3,493,365 of Pickering et al. (1970) discloses a two-step high temperature method of treating zinc plant residue con- taining zinc ferrite. This process differs from that of the '486 patent in that in the second step a source of a cation selected from the group consisting of sodium, potassium and ammonium is added in order to precipitate

05 the iron from the liquor as a jarosite material.

A process for treating sulfide ores which involves a two-step leach is disclosed by U.S. Patent 4,266,972 of Redodno-Abad et al. , (1981) . The first leach uses

10 sulfuric acid under an oxygen atmosphere at 150° to 250°C. Zinc and copper are solubilized with lead, the noble metals, and iron oxide remaining as a residue. After a solid liquid separation, the filtrate is adjusted to a pH of 1.5 to 2. Sodium chloride, calcium chloride

15 and ferric chloride are added to precipitate calcium sulfate. The leach is conducted at a temperature between 60°C and 90°C with the silver, lead, and gold being solubilized as the chlorides. The iron oxide remains as a residue. After a solid/liquid separation, the silver,

20 lead, and gold are recovered by cementation with zinc, with the liquid being subjected to an extraction to recover the zinc.

None of these processes which use high temperature and pressurized leaches discloses or suggests that __. jarosite-containing wastes can be advantageously treated in such a manner. In fact, most disclose the use of an oxidizing atmosphere to form ferric iron which will precipitate. These patents also disclose that potassium ions can be added to a zinc ferrite leach solution in

30 order to precipitate potassium jarosite. As discussed in detail hereinbelow, it has been found that the recovery of metal values can be unpredictably affected by the presence of potassium ions.

Accordingly, there is a need for a process to treat

35 jarosite and ferrite containing wastes from zinc recovery processes in order to recover metal values which are contained in the waste materials and render the residue suitable for disposal as a nonhazardous waste. There is also the need for a process which will not be subject to the unpredictable effect of potassium ions.

Summary of the Invention

It has now been found that the above described disadvantages of known processes can be minimized or eliminated by the instant invention. According to the present invention, a process is provided for recovering metal values from jarosite wastes from a zinc recovery plant said process comprising leaching the waste with an acidic solution of calcium chloride at or above the atmospheric boiling point of the solution and under at least the atmospheric pressure.

One of the embodiments of the instant invention comprises a process for recovering metal values from jarosite-σontaining wastes from a zinc plant using an acidic solution of calcium chloride at a temperature above the boiling point of the solution and under super- atmospheric- pressure.

Another embodiment of the instant invention comp¬ rises a process for recovering metal values from a jarosite-containing waste wherein the process comprises contacting the waste at a temperature of between about 110°C and 300°C and a pressure of at least the auto- geneous pressure with a solution containing between 1.0 and 5.0 molar calcium chloride. The contacting solution has a pH of between about 2.0 and 3.5. This pH is maintained by the addition of a calcium compound selected from the group consisting of calcium oxide, calcium hydroxide, calcium carbonate and mixtures thereof.

In a further embodiment, the instant invention comprises a process in which jarosite-containing wastes are leached with an acidic solution of calcium chloride at greater than atmospheric pressure and a temperature greater than the atmospheric boiling point of the solu¬ tion to form a liquid leachate and a solid residue. The liquid leachate is separated from the solid residue and contacted with a reducing metal to reduce silver cations contained in the leachate to metallic silver. The metallic silver is then separated from the liquid solu¬ tion. The liquid solution is sulfided by mixing the solution with a sulfide compound to precipitate the lead contained in the solution as lead sulfide. The solid

05 lead sulfide is separated from the liquid phase. Sub¬ stantially all of the zinc in the remaining liquid phase is recovered using a zinc recovery process to provide a liquid solution substantially free of zinc. The result¬ ing liquid solution is heated with calcium oxide to

10 provide a vapor containing ammonium hydroxide and bottoms which contain calcium chloride.

In another embodiment, the instant invention com¬ prises contacting a ferrite which contains metal values with a leach solution of sulfuric acid and ammonium ions.

15 The leach solution containing the ferrite is heated at a temperature above about 90°C for up to about 12 hours to form a solid containing ammonium jarosite and a liquid

.-phase. The solid, phase is mixed with an acidic leaching

20 solution containing calcium chloride. This mixture is heated at superatmospheric pressure and at a temperature above the atmospheric boiling point of the solution to solubilize a substantial portion of the metal values.

__. Brief Description of the Drawing

Fig. 1 shows a typical process flowsheet for an electrolytic zinc plant;

Fig. 2 shows the effect of potassium on metals extraction; _3Ω Fig. 3 shows the effect of temperature and potassium concentration on silver extraction;

Fig. 4 shows the flowsheet for a high temperature calcium chloride leach?

Fig. 5 shows a process flowsheet for the recovery of __. metal values from jarosite;

Fig. 6 shows a process flowsheet for a ferrite leach followed by a high temperature calcium chloride leach; Fig. 7 shows a process flowsheet for metals recovery from ferrite and jarosite in association with a zinc plant; and

Fig. 8 shows a process flowsheet of a preferred embodiment.

Description of the Preferred Embodiments

The process of this invention comprises a method for recovering metal values from jarosite containing wastes from electrolytic metal recovery plants. The jarosite waste is leached at or above the boiling point of the leach solution under pressure with an acidic calcium chloride solution. Optionally, non-jarosite wastes such as zinc ferrite can be subjected to a preleach step in which the iron is converted to jarosite material and the zinc is substantially solubilized.

The use of a high temperature, pressurized leaching process for leaching jarosite has been found to provide several unexpected advantages over the lower temperature, atmospheric pressure processes . described in the prior art. It has been found that the final leach slurry unexpectedly filters two to three times faster than comparable slurries obtained with a 95°C to 100°C leach such as that described by Steintveit in Norwegian Patent 142,406 (Supra). Additionally, the leach times are significantly less than those required when leach temper¬ atures below 100°C are used. This procedure has also been found to provide higher extractions of metal values from certain types of wastes than obtained with the lower temperature leaches.

It has also been found that certain materials un¬ expectedly interfere with the recovery of metal values from jarosite. In particular, it has been observed that with some jarosite wastes the presence of potassium ions can result in significantly lower recoveries of metal values such as silver and lead from the leach solution. Fig. 2 is a graphic presentation of experimental data showing the effect of potassium on the amount of metal extracted. This effect is particularly evident at lower leaching temperatures. Thus, using higher leaching temperatures than disclosed in the prior art appears to unexpectedly compensate for the presence of potassium in the leaching solution. Graphic evidence of this pre¬ viously unrecognized phenomenon is provided in Fig. 3

05 where the effect of increasing the extraction temperature from 130°C to 180°C is shown. The results also show that only 10 percent of the silver was extracted at 95°C when 5 grams per liter of potassium were present. With 10 grams per liter of potassium, 80 percent of the silver

10 was extracted at i80°C.

The detrimental effect of ionic potassium is not observed with every jarosite material. Differences in the effect of potassium concentration on the recovery of silver have even been observed on samples of jarosite

15 obtained from different locations in the same waste lagoon. Although extensive work has been done to deter¬ mine the basis of this phenomenon, the reason remains unknown. A method for predicting the magnitude of this

20 .effect has not been discovered as yet. Therefore, practicing the present invention assures that the re¬ covery of metal values can be maximized even with varia¬ tions in the magnitude of the potassium effect.

The leaching process for ainmonium jarosite can be ₂_. represented by the following reaction scheme.

(1) 2NH₄Fe₃(S0₄)₂(OH)_e + CaCl₂ + 3Ca(OH)₂ *-

6Fe(OH)₃ + 2NH₄C1 + 4CaS0₄

The overall leach process, including regeneration of CaCl₂ , may be represented by combining the above reaction ,₀ with the reaction occurring during the "lime boil" operation,

(2) 2NH₄C1 + Ca(OH)₂—s- CaCl_j + 2NH-. + 2H₂0 to get the following overall reaction,

(3) 2NH₄Fe₃ (S0₄)₂(0H)₆ + 4Ca(OH)₂ ■>•

₃₅ 6Fe(OH)₃ + 4CaS0₄ + 2NH-. + 2H₂0.

The calcium chloride leaching process of the instant invention is normally conducted at or in excess of the atmospheric boiling point of the leach solution. As used herein, the terms "atmospheric boiling point" and "boil¬ ing point" are used interchangeably to refer to the temperature at which the solution boils under the parti¬ cular atmospheric pressure to which it is being sub¬ jected. Thus the boiling point for a solution of the same composition can vary depending upon the geographical elevation at which it is heated. Ordinarily, the boiling point of the leach solution is in excess of 110°C. It has been found that leaching at a temperature of at least the boiling point of the solution and preferably above 110°C provides improved results in extracting metal values from the jarosite-containing waste materials. The boiling point of the leaching solution can be increased by increasing the concentration of calcium chloride in the solution. Therefore, it is contemplated that the instant invention encompasses the embodiment of heating a solution at or above 110°C even though the boiling point of the solution is in excess of 110°C. Leaching tempera¬ tures from the boiling point to about 300°C can be used with temperatures of between about 120°C and 225°C being preferred and temperatures of between about 130°C and 200°C being most preferred. The pressure on the leaching solution has not been found to be critical as long as the pressure is sufficient to prevent the solution from boiling. Ordinarily, autogeneous pressure is maintained in the system, i.e. greater than 1 atmosphere and prefer¬ ably at least about 1.1 atmospheres. As used herein, the term "autogeneous pressure" means that pressure which develops in a closed system when the system is heated.

The concentration of calcium chloride should be maintained above about 0.5 molar in the leaching solution to achieve the best extraction and solubilization of metal values. The concentration can range from about 0.5 molar up to the saturation point for calcium chloride for the particular temperature and solution. It is preferred that the concentration of calcium chloride be between about 1 molar and about 4 molar in the leaching solution with the most preferred concentration of calcium chloride between about 2 and 3.5 molar.

Calcium chloride can be added to the leaching solution directly or can be formed in situ by using a chloride source and a calcium source. For example, 5 materials such as hydrogen chloride, sodium chloride and ammonium chloride can be used as the chloride source. Useful sources of calcium include lime, hydrated lime, and calcium carbonate. The particular materials used depends upon the economics. The use of the calcium oxide 0 or calcium hydroxide as the calcium source has the additional advantage that the pH of the solution can be adjusted to the desired range by their addition. To avoid the buildup of certain cations, e.g. sodium, in a continuous system, it may be necessary to bleed a small 5 stream out of the recycle or to remove the cation by other means such as ion exchange.

For effective leaching with calcium chloride, it is important that the pH of the solution be less than about 5. It is preferred that the pH be less than about 4 and 0 most preferred that the pH be less than about 3. For best results, the extraction should be carried in a pH range of about 2.0 to 3.0 During the calcium chloride leach of the jarosite, the pH of the solution can de¬ crease to a pH of less than 1.0 unless adjusted by the 5 addition of a base. Although any base, such as sodium hydroxide or sodium carbonate can be used, it is advan¬ tageous as discussed hereinabove to use a calcium con¬ taining base such as calcium oxide, calcium hydroxide or calcium carbonate. _Q Other things being equal, the higher the leaching temperature, the shorter the leaching time required to achieve a particular level of extraction of metal values. The time required for the extraction depends upon the concentration of calcium chloride in the leach solution, the temperature at which the leaching is conducted, the pH of the solution and the particular waste feed being treated. Ordinarily, the jarosite leaching requires less than about 6 hours with leach times in the range of about 5 minutes up to about 2 hours being preferred. At the higher leaching temperatures, it is expected that in excess of about 75 percent and preferably in excess of 85 percent of the silver and lead present in the jarosite is solubilized during the preferred leaching time.

The jarosite-containing waste materials suitable for use as feed to the calcium chloride leaching process comprise materials which contain the metal values such as zinc, silver, lead, indium. The preferred feeds are jarosite residues from an electrolytic zinc process. The material called jarosite has the formula MFe-. (SO. ) ₂ (OH) _g. There are a variety of jarosite type compounds which contain different ions in place of the M, for example, potassium, lead, silver, sodium, rubidium or ammonium, Ammonium jarosite is a principal effluent from an elec¬ trolytic zinc process.

Other iron containing materials can be used as feeds to produce jarosite e.g. zinc ferrite (ZnFe₂0₄) . The zinc ferrite used can be any such material containing metal values. The source of zinc ferrite is ordinarily waste lagoons from zinc plants.

An alternative source of zinc ferrite material is the dust formed during steelmaking in an electric arc furnace (EAF) . The dust primarily comprises zinc fer¬ rite, zinc oxide, and various forms of ferric oxide. The dust can also contain lead, cadmium and chromium and is therefore usually classified as a hazardous waste. The dust can additionally contain metal values such as silver. Thus, the dust can be fed into the process of the instant invention to allow recovery of metal values and/or removal of the toxic metals.

In order to be used in the present process, the EAF ferrite is subjected to a leach operation in which the ferrite is contacted with sulfuric acid and an ammonium source to precipitate ammonium jarosite which contains some of the metal values. The ammonium jarosite can then be conveyed to the calcium chloride leaching process. In the practice of the instant leaching process, the jarosite 5 is added to the leaching vessel 6 along with calcium chloride as represented in Fig. 4. The calcium chloride can be makeup or be recycled from downstream process steps or a mixture of the two sources. During the high temperature leaching process, the pH of the solution tends to decrease due to generation of acid. The pH is adjusted to the preferred leaching range by adding a basic material, preferably calcium oxide, calcium carbonate, and/or calcium hydroxide. The solu¬ tion is heated to the leaching temperature, preferably by steam, with agitation of the mixture. After the approp¬ riate leaching time, the resulting slurry is subjected to a liquid-solid separation 7 to provide solid tails 8 and a liquid leachate 9. The solid tails contain iron oxides, gypsum, and/or unreacted gangue while the liquid leachate contains solubilized zinc and other metal values. The tails are discarded while the leachate is subjected to further downstream treatment as discussed in more detail hereinbelow. Optionally, a portion of the leachate is recycled 10 to the calcium chloride leach step to effect a build-up of metals in the pregnant solution.

The leachate from the calcium chloride leaching process can be treated by any suitable method to remove the metal values such as copper, silver, or gold which might be present. When silver and copper are present, it is preferred that a metal cementation process be used to remove the silver and/or copper. As represented in Fig. 5, this can be accomplished by treating the leachate with any material which is a suitable reductant for the metal values. Metals which have higher reduction potentials than silver, such as lead or zinc, can be used to form a silver cement. Presently for economic reasons, it is preferred to use zinc dust. Zinc dust can be added to precipitate a solid consisting of a mixture of silver, lead, copper (depending upon the presence of the parti¬ cular metal in the solution) and zinc. The solid product which is separated from the liquid phase can then be subjected to additional standard metallurgical processes^' to recover the pure metal values. The separation of the solid product from the liquid phase can be accomplished by standard solid-liquid separation techniques such as filtration or centrifugation.

The liquid from the metal cementation process contains dissolved reducing metal, calcium chloride, lead, zinc and other noncementable impurities The liquid can then be subjected to additional processing steps to separate the reducing metal and other metals such as lead and zinc (if not used as the reducing metal) from the liquid. Optionally, the liquid can be subjected to a sulfide precipitation procedure to recover the lead and any trace of heavy metals which precipitate as sulfides under these conditions. The sulfide precipitation can be accomplished by adding a sulfide-containing material such as hydrogen sulfide to the liquid to form solid lead sulfide. The liquid can then be separated from the solid lead sulfide cake by conventional solid-liquid separation techniques. Alternatively, the pH of the liquid can be adjusted to precipitate metals such as lead and zinc as their hydroxides. For example, calcium oxide can be added to precipitate lead hydroxide and zinc hydroxide. In another embodiment, the liquid from the metal cement¬ ation process is subjected to an extraction process to separate zinc as described herein below. The raffinate can be partially neutralized to a pH of about 8-9 to form a lead hydroxide precipitate which is separated from the liquid. The liquid can then be subjected to a lime boil as discussed below. The particular choice of procedures to be used to treat the liquid will depend upon the metals present, their concentrations and the economics of the various alternatives. The solids recovered can be disposed of in an appropriate manner or can be subjected to additional processing steps to recover any metal values present. The liquid from the lead removal process can be sub¬ jected to zinc recovery procedures. Any standard method of recovery suitable for removing zinc from such a solution can be used. A preferred method is solvent extraction in which the aqueous solution containing zinc is contacted with an extractant which transfers the zinc from the aqueous solution to the organic phase. The second phase can be a solvent which is not miscible with the aqueous solution or it can comprise the extractant. Alternatively, the liquor containing the zinc can be passed through an ion exchange resin to remove the zinc. A preferred method involves the use of extractants such as di-2-ethylhexyl phosphoric acid dissolved in a hydro¬ carbon solvent such as kerosene. In the use of such materials, the pH is adjusted to determine the phase in which the zinc ultimately resides. In a preferred method of operation, the zinc species is selectively extracted into the organic phase by maintaining the pH in an appropriate range depending on the extractant used. The pH can be adjusted with a base such as calcium oxide, calcium hydroxide, calcium carbonate, etc. Zinc is then stripped from the loaded organic with a strong acid solution such as spent electrolyte from a zinc plant. Strip liquor containing the zinc can be transferred to the zinc plant for electrowinning.

The raffinate from which the zinc has been removed can be subjected to additional processing to recover the calcium chloride and ammonia. A base can be added to increase the pH of the solution to a basic range, prefer¬ ably above about 9. Although bases such as NaOH or KOH can be used, it is preferred that a basic calcium compound such as calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof be used to avoid build-up of undesired materials such as potassium ions in the system if recycle is used. Preferably a "lime boil" operation can be used in which lime is added to the raffinate to increase the pH to a basic range, preferably above about 9, and the mixture heated to form ammonia which can be removed by volatilization. Preferably, a temperature of at least the boiling point of the solution is used. The ammonia formed is removed, usually in the vapor, and can be used to form jarosite in the ferrite treating process described hereinbelow- or can be trans¬

05 ferred to the zinc plant. The brine solution from the lime boil process contains calcium chloride as well as residual metal values. This brine solution can be recycled for use in the calcium chloride leaching pro¬ cess.

10 An alternative method for recovering the zinc is to treat the liquor from the sulfiding procedure with a base such as calcium oxide to precipitate zinc hydroxide. In this operation (not shown in Fig. 5) , sufficient base is added to provide a solution pH of between about 8 and 10. _₅- Bases which are useful in this process include calcium hydroxide, calcium oxide, sodium hydroxide, etc., pro¬ vided there is no detrimental build-up of cations in the system. The zinc hydroxide precipitate is separated from the liquor by standard liquid-solid separation techniques ₀ such as filtration or centrifugation. The zinc hydroxide solid is preferably transferred to a zinc plant for recovery of the zinc. The liquor from this precipitation process is transferred to a lime boil process as des¬ cribed hereinabove. 5 Depending upon the level of lead and/or zinc present in the liquor from the silver cementation process, the basic precipitation step can be optionally performed on the liquor from the silver cementation step. This will result in a mixture of lead hydroxide and zinc hydroxide 0 solids being formed. After a solid/liquid separation, these solids can be disposed of in any manner appro¬ priate. The liquor from this precipitation step is transferred to a lime boil process as described herein- above. 5 When zinc ferrite (ZnFe-,0.) is a significant com¬ ponent of the waste, a sulfuric acid (sulfuric) leach step is used. In this leach there is a simultaneous leach, of the ferrite and a precipitation of the iron in the form of jarosite. A method for this simultaneous leach and precipitation is described by Rastas et al. in U.S. Patent 3,959,437 (1976). Rastas, however, is only concerned with solubilizing the zinc in the ferrite and

05 does not treat the jarosite to recover metal values.

In the present invention, the ferrite material is combined with sulfuric acid and an ammonium source at a temperature above about 50°C and below the decomposition temperature of ammonium jarosite. Preferably, the

10 temperature is between about 80°C and about 170°C. The sulfuric acid used is preferably spent electrolyte from an electrolytic zinc plant and is in the ferrite leach solution to the extent of between about 10 and about 60 grams per liter, preferably between about 30 and about 50

•,5 g/1. The ammonium compound is preferably recovered from the lime boil process step described hereinabove. This leach is ordinarily conducted at a temperature of about 80°C to 100°C. However, as discussed hereinabove, the leach can be accomplished at a temperature above the

2 boiling point of the solution under pressure. It is contemplated that the leach is conducted under auto¬ geneous pressure at the elevated leaching temperatures.

The ferrite leach is conducted for a period suffi¬ cient to precipitate jarosite and solubilize zinc.

25 Preferably the leach time is about 12 to 36 hours. Commonly, at least about 70 percent of the zinc present in the ferrite is solublized with less than five percent of the iron and essentially no silver or lead solu¬ bilized. The solid jarosite containing metal values and _Q any remaining untreated ferrite is separated from the liquid phase by standard solid-liquid separation tech¬ niques such as thickening, filtration or centrifugation. The liquor from the ferrite leach step which contains zinc sulfate is preferably conveyed to a zinc plant for 5 recovery of the zinc. The solid jarosite is subjected to the calcium chloride leach process described hereinabove. Referring now to Fig. 4, a process involving a calcium chloride leach of jarosite in a system at a temperature above the boiling point of the solution is represented. Preferred embodiments of this process are hereinafter described. Sufficient jarosite 5 is added to

05 the leaching vessel 6 to provide a slurry containing about 10 to about 40 weight percent jarosite. A chloride source such as calcium chloride is added to the slurry. Alternatively, other chloride sources can be used de¬ pending upon the economics of the process. The pH is

10 maintained in the desired range of about 1.5 to about 3.5 by the addition of lime, i.e. calcium oxide, and/or calcium hydroxide, calcium carbonate. The slurry is heated to a temperature ordinarily above the boiling point of the solution, normally above 110°C and pre- - _. ferably above about 120°C. The leaching zone or leaching vessel must be capable of withstanding the equilibrium pressure at the temperature selected. While the vessel can be pressurized with an inert gas, in ordinary opera¬ tion it is maintained under autogeneous pressure, i.e. 0 the pressure established in the vessel by the vaporiza¬ tion of volatile components such as water at the leaching temperature. Ordinarily, the pressure in the leaching zone will be between about 5 to about 210 psig with the preferred range being from about 15 to about 130 psig. 5 The leaching operation can be carried out in either a batch or a continuous mode. The particular choice of operation will depend upon the leaching time necessary to extract the desired amount of the metal values. In a batch operation, the necessary quantity of calcium oxide, 0 calcium carbonate and/or calcium hydroxide can be intro¬ duced initially instead of being added incrementally during the leaching process. The appropriate amount of base can be readily calculated by one skilled in the art based upon the jarosite content of the feed material ₅ added to the leach. The initial quantitative addition of the calcium oxide is operationally advantageous since it eliminates the equipment required to add the material incrementally throughout the leaching process as well as the need to monitor the pH of the system. We have found that there is no difference in the final level of the metal values extracted when the calcium oxide is added initially as opposed to incrementally during the leaching operation as long as the final pH of the leaching solu¬ tion is within the desired range.

At the conclusion of the leaching process, the liquid-solid slurry is separated 7 into solid tails 8 and a liquid leachate 9. This can be accomplished by any known method for solid-liquid separations, although with this type of slurry, it is preferred that the separation be accomplished by filtration accompanied by appropriate washings. Surprisingly, it has been found that slurries from the high temperature leaches filter two to three times faster than the slurries from a leach conducted at 95°C to 100°C. This is an unexpected advantage of operating at the higher leach temperatures. The solid tails which contain gypsum and iron oxides are discarded. Optionally, a slip stream from the leachate is recycled 10 to the calcium chloride leaching process to allow further concentration of metal values. The remaining leachate liquid is subjected to further processing for recovery of metal values.

The recovery of the metal values can be accomplished by any known method useful for separating such materials. A particular separation scheme is set forth in Fig. 5. The leachate 9 from the calcium chloride leach 6 is contacted 13 with a metal 14 capable of reducing silver. Preferably zinc metal is used although lead is also suitable and the choice is generally a matter of econo¬ mics. The metal, ordinarily in the form of fine parti¬ culates, is contacted with the leachate preferably at a temperature between about 40°C and about 80°C. The silver cement 15 is separated 17 from the liquid by filtration although centrifugation could also be used. The silver is recovered from the silver cement by conven¬ tional means such as smelting. As shown in Fig. 7, lead can be separated from the liquor from the cementation process by sulfiding 51 with hydrogen sulfide or any appropriate sulfide source to precipitate lead sulfide 53. Ordinarily, greater than about 95 weight percent of the lead is precipitated. The lead sulfide is separated from the liquid by filtration although centrifugation can also be used.

The liquid from the sulfiding step is treated 19 to recover zinc. Solvent extraction and precipitation are preferred methods of separation although any method suitable for recovering zinc 21 from such a stream can be used. The zinc-solvent extraction 55 process, depicted in Fig. 7 and Fig. 8, involves contacting the aqueous solution containing the zinc with an extractant which will bind the zinc and allow its extraction into an organic phase. A preferred extractant is di-2-ethylhexyl phosphoric acid dissolved in an appropriate diluent. In operation, the extractant is mixed with the zinc feed solution. Since the extraction of zinc lowers the " aqueous pH, a base such as NaOH, CaO, or CaCO-. can be added to maintain the desired pH. At a pH of between about 2 and 3 the zinc is selectively extracted into an organic phase 56 in contact with the aqueous solution. This organic phase containing the zinc is separated from the aqueous phase and the zinc is then stripped 57 into a strong sulfuric acid solution. The solution 59 contain¬ ing the zinc can be transported to a zinc plant for recovery of the zinc.

In another process sequence (shown as a part of the scheme in Fig. 8) , the liquid from the cementation step is subjected to solvent extraction to remove the zinc as described hereinabove. The raffinate liquor from the zinc solvent extraction is partially neutralized, prefer¬ ably with calcium oxide, to a pH of between about 8 and 9 to precipitate lead hydroxide. The lead hydroxide solid is removed, for example by filtration 65 or centrifuga¬ tion and disposed of or the lead can be recovered by known procedures. The liquid 23 recovered from this separation is sub¬ jected to a lime boil 25 by adding lime to provide a pH above about 9 and heating to about the boiling point of the mixture. Ammonia is vaporized and passed through a condensor to provide an a_mmonium hydroxide stream which

05 can be recycled to the ferrite leach or to the zinc plant to produce ammonium jarosite_* The liquid condensate can be neutralized with sulfuric acid to produce ammonium sulfate suitable for fertilizer.

Alternatively (not shown) , the zinc can be precipi¬

10 tated as zinc hydroxide by adding calcium oxide to the liquor from the sulfiding process. Sufficient calcium oxide is added to provide a solution pH in the range of about 8 to 10. The precipitated zinc hydroxide is separated by liquid-solid separation means preferably _ιg filtration. The recovered zinc hydroxide can be conveyed to a zinc plant for recovery of the zinc.

As shown in Fig. 8, the liquor from the zinc separa¬ tion process is subjected to a lime boil treatment with calcium oxide. This lime boil step is- conducted at a

„_n temperature effective to form ammonia and calcium chlo¬ ride commonly above room temperature. Sufficient calcium oxide and/or calcium hydroxide is added to provide a pH of above about 9. This process step provides an ammonia overhead vapor stream and a brine stream containing

₂5 calcium chloride and/or a precipitate of lead hydroxide. The brine stream is preferably recycled to the calcium chloride leach step to allow reuse of the chloride values. The ammonia, which can be dissolved in water or used directly, is preferably recycled to a zinc plant or ₀ to a ferrite preleach process.

In an alternative procedure (not shown) , the liquor from the cementation step can be combined with sufficient calcium oxide to cause the combined precipitation of zinc hydroxide and lead hydroxide. Calcium oxide is added to ₅ provide a pH in the range of about 8 to 10. The metal hydroxides are separated by usual solid-liquid separation techniques preferably by filtration. These hydroxides can then be returned to a zinc plant for recovery of the zinc. The liquor from the precipitation process is then subjected to a lime boil as discussed hereinabove. The resulting brine solution is preferably recycled to the calcium chloride leach with the ammonium hydroxide formed preferably being recycled to a zinc plant or sold as a by-product.

When zinc ferrite is treated, a sulfuric acid ferrite leach operation is used as discussed hereinabove. As shown in Fig. 6, an aqueous slurry of the ferrite solids 31 is combined with sulfuric acid and ammonium hydroxide. Sufficient ferrite solids are added to provide a slurry containing about 30 weight percent ferrite. Sulfuric acid is added to the extent necessary to provide a final acid level of about 10 to about 60 grams per liter. Preferably, the sulfuric acid used is spent electrolyte from a zinc plant. Sufficient ammonia is added to accomplish precipitation of the ammonium jarosite. Suitable ammonium sources include without

_* limitation ammonia, ammonia water^', ammonium hydroxide and ammonium salts such as ammonium sulfate. Preferably, the ammonia is received from the lime boil operation des¬ cribed hereinabove. Preferably, ammonium ions are present at the end of the leach to the extent of about 10 g/1 to about 20 g/1 of the solution. About 15 grams of ammonia per liter of slurry is used in this operation. The ferrite leach 33 is accomplished at a temperature of about 95°C for a period of about 12 to 24 hours. This resulting slurry is subjected to a thickening liquid- solid separation procedure. The addition of a small amount of "seed" jarosite can be added to aid in the formation of solid jarosite. The liquid 35 from this ferrite leach is rich in zinc and can be conveyed to a zinc recovery process. The solid material 37 is predomi¬ nantly jarosite which contains zinc and other metal values. This material is conveyed to a calcium chloride leach 6 as described hereinabove. In Fig. 7 there is shown an embodiment of the instant invention in which the product 41 from a ferrite leach 33 is combined with additional jarosite 43 in a repulp tank 45 and the resulting pulp 47 is fed to a calcium chloride leach 6. The subsequent metal recovery steps have been described in detail hereinabove. Also shown in this embodiment is the movement of various streams to and from a zinc plant 49. The zinc-rich liquid 35 from the ferrite leach is transferred to the zinc plant for recovery of the zinc. Also, the strip liquor from the zinc extraction is transferred to the zinc plant. Ammonia from the lime boil is used in the zinc plant and/or the ferrite leach step to form jaro¬ site. The spent sulfuric acid electrolyte from the zinc plant is used in the ferrite leach and in zinc stripping. As discussed hereinabove, an alternative method of zinc separation which is not shown is the precipitation of zinc hydroxide. This zinc hydroxide can also be conveyed to the zinc plant for recovery of the zinc metal.

- In Fig. 8 is shown a preferred process scheme for the treatment of ferrite and/or jarosite wastes. A ferrite containing feed 31 is subjected to a leach 33 by contacting it with sulfuric acid and an ammonium source as described hereinabove. The source of the sulfuric acid is spent electrolyte 71 from a zinc plant 49. The ammonium source is ammonia or ammonium hydroxide 73 recycled from the lime boil 25. A flocculant 75 is added to the slurry and the solids are separated from the liquid. The liquid 77, containing zinc, is conveyed to a zinc plant for zinc recovery. The solids, containing ammonium jarosite and metal values, are transferred to a repulp tank 45 where these solids can be mixed with additional jarosite feed 43 and slurried with a calcium chloride source. The leach mixture 6 is heated to above 110°C under pressure and stirred. The pH of the leach is adjusted to the desired level by the addition of calcium oxide, calcium carbonate and/or calcium hydroxide. After leaching, the slurry is filtered 79 with washing to separate the solid tails 8 and the leachate 9. The leachate is contacted 13 with metallic zinc to form a silver/copper cement 15. The cement is removed, preferably by filtration 81, and the silver and copper are purified by standard metallurigical methods. The liquid is mixed 55 with an extractant with the pH of the

05 solution adjusted by adding calcium carbonate. The extractant binds the zinc and allows its extraction into an organic phase. The phases are separated and the zinc-containing organic phase 56 is contacted 57 with spent electrolyte 71 from a zinc plant to strip the zinc

10 from the organic phase into an aqueous phase. The aqueous phase 59 containing the zinc is transferred to a zinc plant for recovery of the zinc. The organic phase 83 containing the barren extractant is recycled to contact fresh zinc-containing solution. - -. The aqueous phase from which substantially all of the zinc has been removed is contacted with lime 85 to increase the pH to between about 8 and 9 to precipitate lead hydroxide. The solid lead hydroxide is separated from the liquid by filtration 65. The liquid phase is 0 mixed with lime and heated to boiling. The. liquid from the lime boil 25 is a calcium chloride-brine solution and is recycled 87 to the calcium chloride leaching step. The vapor containing ammonia is condensed 89 with cold water and the resulting ammonium hydroxide solution is 5 recycled 73 to the zinc plant and/or to the ferrite leach. Any uncondensed vapor can be passed into an acid scrubber using, for example, a sulfuric acid wash to form a solution of ammonium salt such as ammonium sulfate.

The following examples are given for illustrative 0 purposes only and are not to be a limitation on the subject invention.

Example 1 Three jarosite feed materials were used in the Examples. Sample 1 was taken from a first zinc plant 5 waste lagoon. Samples 2 and 3 were taken from different locations in a second zinc plant waste lagoon. The assay results are given in Table 1A. Table 1A

Element, % Sample 1 Stample 2 Sample 3

Zn 4.95 7.53 7.46

Ag oz/T 7.53 12.3 11.2

Fe 27.1 22.6 19.9 5 Pb 2.13 6.98 7.10

NH, 2.04 1.01 1.20 K ⁴ 0.054 0.364 0.337

Na 0.122 0.162 0.156

Cu 0.32 0.3 0.347

Mn 0.362 5.17 4.47 _n In 0.010 — 0.008 0

A series of runs were made in which the jarosite feed material was contacted with a leaching solution containing 330 grams per liter calcium chloride, for 5 hours unless otherwise noted. The pH was maintained in 5 the range of 1.8 - 3.5 by the addition of calcium hydrox¬ ide. The amount of solids in the leach was about 16.8 weight percent.

The leaching process was conducted at different temperatures as indicated in Table IB. When the tempera¬ 0 ture of the leach solution was at or below the boiling point of the solution, the leach was conducted in conven¬ tional glassware. When the leaching temperature was above the boiling point of the solution, the leach was conducted in a 2-liter Parr agitated autoclave. 5 The leachate was separated from solids by filtration through a Buchner funnel using low vacuum.

Table IB

'

; Leach

Run Temperature, Extraction, %

NO. °C Ag Pb Zn

Sample 1 Jarosite (5-10 g/1 K)

1 95 76 81 49

2 103 (boiling) 78 77 47

3 110 93 89 44

4 120 96 85 42

5 130 99 94 48

Sample 3 Jarosite (25 g/1 K)

6 95 72 84 28

7 110 72 (62) 28

8 120 73 (62) 29

9 130 73 88 29

10 140 72 89 31

11 150 74 80 30

12 180 (1 hr) 76 86 37

13 200 (2 hr) 75 84 34

¹⁴(a) 225 (1 hr) 75 87 35

15^(a) 325 74 80 31

(a) Run in the nickel bomb reactor for approximately 5 minutes. Bomb leach extractions are generally lower than comparable autoclave values. Description of nickel bomb reactor in Example 3.

Example 2 The effect of potassium on the recovery of metal values from certain feeds is shown in Table 2. The leaching conditions and apparatus of Example 1 were used. The conventional glassware was used for the 95°C to 103°C leaches with the Parr autoclave used for the leaches conducted at temperatures greater than 103°C- With Sample 2 feed, the recovery of metal values^* was signifi¬ cantly reduced when potassium was present. Increasing the leach temperature in Run 18 however, allowed recovery of the metal values even in the presence of potassium. Sample 3 feed showed no potassium effect. Table 2

Conditions: 95-103°C tests: 330 g/1 CaCl-,; 4-5 hours; pH 1.8-3.5.

>103°C tests: autoclave; 330 g/1 Cad-,; 1 hour; pH 1.8-3.5.

Leach K in Leach

Sample Run Temp Solution Extraction , % No. No. °C 2/1 Ag_ Pb Zn

1 95 5-10 76 81 49

2 103 (boiling) 5-10 78 77 47

3 110 5-10 93 89 44

4 120 5-10 96 85 42

5 130 5-10 99 94 48

16 95-100 0 72 84 41

17 95-100 5 13 35 27

18 180 10 79 88 (60)

19 95-100 0 73 86 34

6 95-100 25 72 84 28

12 180 25 76 86 37

Example 3

A series of runs was made with Sample 3 jarosite to determine the effect of leaching time in atmospheric leaches and in pressurized leaches at temperatures above the boiling point of the leach solution (at 130°C and 180°C) . The atmospheric leaches were conducted at 95-100°C using the procedure of Example 1. The atmos¬ pheric results are given in Table 3A. The pressurized leaches were conducted using Samples 1 and 3 in a 500- milliliter reactor fabricated from nickel pipe mounted in a horizontal configuration. Lifters were installed on the internal surface parallel to the axis. The reactor was rotated about its horizontal axis to provide agita¬ tion to the reaction mixture. A thermocouple set in a thermowell was ^' inserted into the reaction mixture to monitor the temperature. The contents of the reactor were heated to temperature within about 3 minutes by using external Meker gas burners. The leach mixture was maintained at temperature for the specified time. The reactor was then quenched in water to lower the tempera¬ ture rapidly. The pressurized results are given in Table 3B.

Table 3A

ple 3

Run (a) Sam Leaching Time, Extraction , % No. hours Ag Pb Zn

6 (b) 0 54 44 21

6 1. 58 56 22

6 3, 61

(c) 79 29

19 3, 57 76 30

6 5. 72 84 28

19 5. 73 86 34

19 8 73 82 32

19 12 75 84 36

19 24 74 85 33

(a) 330 g/1 CaCl -₉, PH 1.8-3.5

(b) Run 6 had 2255 g/1 K

(c) Run 19 had 0 g/1 K

Table 3B

Run Leaching Time, Extraction, % No. minutes Ag Pb Zn

Sample 1 Jarosite (130°C)

20 5 25 9.4 32 21 10 53 24 33 22 (a) 10 68 69 42 23 20 77 76 46

5 30 (autoclave) (84) (80) (45)

(b)

Sample 3 Jarosite (180°C)

24 (c) 0 72 57 28 25 2, 69 69 31 26 5 72 62 31 27 10 69 78 32 28 30 71 77 33

(a) 20% solids leach. Other leaches: 16.85 solids.

(b) Nickel bomb reactor, 330 g/1 CaCl2' P^H 1.8-3.5, 25 g/1 K.

(c) Heat to temperature, cool immediately. Approximately 3 min heat up. Example 4

A series of runs was made using Sample 1 jarosite to determine the effect of temperature on the amount of solids or pulp density which can be effectively leached. The leach was conducted using 330g CaCl₂/l, at 95 to 100°C, for 6 hours, and 1.8 to 3.5 pH. The results are given in Table 4A.

Sample 1 and Sample 3 jarosite materials were leached in the nickel bomb reactor of Example 3 with 33Og CaCl₂/l, pH 1.8-3.5 at temperature for 10 minutes. The leach of Sample 1 also contained 25 g/1 . These re¬ sidues were washed with hot CaCl₂ brine solution (330g CaCl^/D . The effect of a single wash is shown in Table 4B. The effect of cumulative washings on two residues is given in Table 4C.

Table 4A

Solubilized, %

Pulp Density Zinc Silver Iron Lead

30% solids 49 25 1 52 30% solids 43 38 1 53 20% solids 49 76 1 90 15% solids 46 68 1 95 10% solids 46 72 1 90

Table 4B

Initial Extraction, %

Run Leach * Pb a) -ΛSL No. Solids Original Rewash Original Rewash Zn

Sample 1 jarosite (130°C)

22 20 68 83 69 79 42

29 30 68 84 60 78- 37

30 35 66 82 52 72 38

31 40 69 82 57 77 40

32 45 Slurry too thick to agitate

33 50 Slurry too thick to agitate

Sample 3 jarosite (180°C)

27 16.8 69 - 78 32 34 20 71 - 84 33 35 30 70 74 29 70 30 36 35 66 72 18 83 31 37 40 70 76 19 83 31 38 45 Slurry too thick to agitate 39 50 Slurry too thick to agitate

(a) % solids is the weight % of jarosite feed in the initial leach slurry.

Ca(OH) added to maintain the leach pH is not included.

Table 4C

Wash Cumulative Washing,

Cumulative Efficiency, %

Filter Cake Solution Displacements Aq Zn

Sample 1 residue 0 0 0

0.6 42.8 47.1

1.2 78.5 83.6

1.9 89.3 91.2

Sample 3 residue 0 0 0

1.6 66.0 60.0

3.25 96.9 94.1

4.9 (100) 98.6

6.6 (100) 99.4 Example 5

Closed-cycle process simulation runs were made. Sample 3 feed was used in Runs 40 and 41 and Sample 1 f ed was used in Run 42. The same CaCl₂ leach pro¬ cedure was followed as in Example 1 except the leach time for each was 30 minutes. The temperature for Run 40 was 180°C and was 130°C for Runs 41 and 42. Each leachate from the CaCl₂ leaches was separated from solid residue by filtration. In Run 40, a stainless steel autoclave was used which might have resulted in some cementation of silver during the leach. The "Cycle No.", e.g. C-l, refers to the number of times spent brine was recycyled to the CaCl- leach step. Results are given in Table 5A.

The filtrates were cemented with zinc to recover silver and copper. The cementation was accomplished by bringing the filtrate to temperature in a beaker. Zinc dust was added and the slurry was agitated for one hour. The resulting slurry was filtered and washed with three 50-ml portions of deionized water. The effect of reac- tion time and zinc requirement on the recovery of metals by cementation with zinc powder is shown in Table 5B.

Table 5A

Liquor Stoichiometric Barren

From Zn Addition for Solution,

Run Run No. Ag and Cu (II) Cementation r % g/1

Description (Cycle) Cementation Ag Cu Pb Ag Cu

A 40 5.9 84 42 0.004 0.042

B 41, (C-l) 2.8 60^a 98 18 0.004 0.001 c 41, CC-2) 14.2 90 66 3 0.001 0.002

D 41, (C-3) 0.7 95 24 0.1 0.001 0.002

Ξ 42, (C-l) 1.2 94 99.4 0.8 0.002 0.002

F 42, (C-2) 1.3 98 99 ²¹ h 0.001 0.004

G 42, (C-3) 7.3 98 97 99.6^b 0.001 0.002

Other conditions: Temperature 60°C

Time 1 hour pH 2.6-5.0 a) Low value due to low initial concentration in feed to cementation. b) Anomalous value. Tabl.e 5B

Reaction Barren

Stoichiomelbrie Zn Time , Cementation, % Solution, g/1 for Ag & Cu minutes Ag Cu Ag cu

15 62 73 0.008 0.042

1.0 30 71 70 0.006 0.046

60 86 66 0.003 0.051

15 67 68 0.007 0.048

1.5 30 76 68 0.005 0.049

60 90 74 0.002 0.040

15 71 70 0.006 0.045

2.0 30 81 77 0.004 0.035

60 90 85 0.002 0.023

Conditions: Temperature: 60°C •

Feed Solution: 0.021 g/1 Aιg, 0.152 < ?/l Cu,

3.15 g/1 Pb, 6.3 g/1 Zn, 0.012 g/1 Fe

(from autoclave leach of Sample 3,

330 g CaCl /l, 180°C for 1 hour,

25 g/1 , 25.1 wt. % solids, pH 1.8-

3.5).

Zinc Requirement: 1.0 stoich.= 0.163 g Zn/1

(0.0064 g/1 for Ag, 0.157 g/1 for Cu). Final pH: 5.1-5.2.

Example 6 Filtrates from the zinc cementations of Example 5 were each brought to temperature and contacted with hydrogen sulfide gas sufficient to provide a solution emf below 0.0 millivolts versus a standard calomel electrode (SCE) . This quantity of H₂≤ resulted in final solution pH's of 0.0 to 0.2. After the indicated reaction time, the slurry was filtered and the solid was washed with three 50-ml portions of deionized water.

Table 6

Designation From Run Solution Pb

Run NO. Time, emf, v Final Precipitation (Cycle) minutes Initial Final PH %

A 40 15 +327 -74 1.1 38

B 41, (C-l) 15 +297 -26 0.65 48

05 C 41, (C-2) 30 +290 -25 0.20 72

D 41, (C-3) 30 +348 -30 - 99.6

E 42, (C-l) 30 +246 +0 - 42

F 42, (C-2) 15 +304 -3 0.28 46

G 42, (C-3) Data not applicable.

Other conditions: Temperature: 60°C

10 Feed solution pH: 3.5-5, .0

Example 7 A calcium chloride leach of Sample 3 feed was conducted with a leach solution of 330 g/1 CaCl-, at an initial solids of 25.1 weight percent at a temperature of

15 about 180°C (±3°C) for one hour. The target pH was 1.8-3.5 with 200 lb/ton of feed of Ca(OH)₂ added initi¬ ally to maintain the pH within the target range. The leach mixture was filtered and the leachate solution, maintained at 22°C, was mixed with hydrated lime,

20 Ca(OH)₂, to adjust the pH to the indicated values. The amounts of zinc and lead hydroxides precipitated at the particular pH was determined. Results are given in Table 7A.

Table 7A

25

Analysis, g/i Preσipi .tation, %

PH Zinc Lead Zinc Lead

Feed 2.58 8.68 3.31 0.0 0.0

3.0 9.33 2.83 0.0 14.5

4.0 9.64 2.98 0 10.0

30 5.0 9.58 2.98 0 10.0

6.0 10.2 3.28 0 .9

7.0 9.7 3.13 0 5.4

7.5 5.4 1.4 37.8 57.7

8.0 2.06 .125 76.3 96.2

9.0 1.89 .055 78.2 98.3

9.5 1.47 .441 83.1 86.7

10.0 1.47 2.15 83.1 35.0

35 10.7 2.06 2.11 76.3 36.3 The filtrates from Example 6 were each mixed with hydrated lime to adjust the pH to a target final pH of

9.5. The temperature was maintained at about 60°C for 15 minutes. The solid zinc hydroxide was separated by filtration using a Buchner funnel. The solids were

»_. washed with three 50-ml portions of deionized water and Ub dried overnight at 100°C prior to assay. The results are given in Table 7B.

Table 7B

10 Ca(OH)₂ Zinc

Test From Required Feed Zinc Assay, g/1 Precipitation

Designation Test No. g/i PH Feed Final %

A 40 19.8 0.55 5.39 (1.15)^a (80)

B 41, (C-l) 7.2 1.0 4.77 1.47 69 c 41, (C-2) 18.0 0.55 5.16 1.00 80

41, (C-3) — 0.40 4.98 0.88 82 s E 42, (C-l) 30.0 0.30 6.24 1.18 .81

F 42, (C-2) — 0.30 7.47 2.21 70

G 42, (C-3) 19.7 1.37 3.35 1.89 62

a) Value calculated from solids assays. 0 Example 8 Filtrate from Example 7 was mixed with calcium hydroxide and heated at boil. The ammonia evolved was recovered in the distillate or by condensing or scrubbing the off-gas. The pH was maintained between about 8.8 and 5 10.5 with Ca(OH)₂- Results showing essentially complete NH₃ recovery is possible by boiling the solution at the indicated pH are provided in Table 8. Some of the zinc or lead present in the feed solution precipitated during the lime boil step and those values are given in Table 8 0 as residuals precipitated. Table 8

From Volume

Run No. Target Ca(OH) Reduction

Des:icrnation (Cycle) PH Required, ²g/l %

A 40 10. .5 _ 33

B 41, (C-l) 10. .5 3.2 * 27

C 41, (C-2) 8. .75 2.4 6.8

05 D 41, (C-3) 8. .75 - -

E 42, (C-l) 8. ,75 Q.O 8.8

F 42, (C-2) 10. ,5 1.9 10

G 42, (C-3) 8. .60 2.0 4.3

10 NH_Δ Residilals

NH. Assay, g/1 Voilatiϊized Precipitited, %

Designat: Lon Feed - Final Distillate % Zn Pb

A 1.76 0.01 3.02 99.8 48 5

B 1.47 - 3.94 100 83 10-12

C - - 11.4 (100) 4 1

D not .assayed - - -

15 E 2.94 1.06 21.1 72 48 1

F not iassayed - 81 7

G not iassayed — 74 4

Example 9

Ammonia recovery into a high concentration solution

20 from a CaCl₂ solution by boiling'with Ca(OΞ)₂ was deter¬ mined by preparing a synthetic feed solution. This solution contained 450 g/1 CaCl₂ and 50 g/1 NH3 as NH₄C1. A volume of solution was placed in a sealed distillation apparatus with provision for incremental sampling of the

25 distillate. Sufficient Ca(OH)₂ was added to liberate 70% of the contained H₃ and the distillation was performed with periodic collection and H₃ analyses of the distil¬ lates. The test data are summarized in Table 9. The distillation time of 43 minutes produced a 27% volume

30 reduction and a 53% recovery of H₃ in the distillate (75% of the theoretical recovery based on the Ca(OH)₂ addition) . Initial distillate contained over 300 g/1 NH₃, demonstrating that the lime boil step can recover a

35 high concentration NH_ product. The usable NH- concen¬ tration should be in the 200 g/1 range as shown by the Distillate No. 1 + 2 + 3.

Table 9

Time Volume NH CaCl NH

Product minutes ml g/ g/i² Amount, g

Feed to lime boil 0 500 (50) (450) 25.0

Distillate No. 1 7 13.2 305 - 4.02

Distillate No. 2 14 17.2 173 - 2.97

Distillate No. 3 20 17.5 146 - 2.55

Distillate No. 4 29 27.0 70 - 1.88

Distillate No. 5 30 29.0 37 - 1.08

Distillate No. 6 43 30.5 20 - 0.62

Distillate total 134.4 13.12^a

Depleted feed 363 - (737)^b

Overall total 497

Distillate Volume, NH Evolved,

% of Feed % of ^r. Theoretical

Product Increment Cumulative Increment Cumulative

Feed to lime boil _ _ 1 3 0

Distillate No. 1 2.6 2.6 23 .0 23.0^*

Distillate No. 2 3.4 6.1 17 .0 40.0

Distillate No. 3 3.5 9.6 14, .6 54.5

Distillate No. 4 5.4 15.0 10, .7 65.2

Distillate No. 5 5.8 20.8 6. .2 71.4

Distillate No. 6 6.1 26.9 3. .5 75.0

Distillate total

Depleted feed

Overall total a ) Sufficient Ca( OH) was added to volatilize 17.5 g NH ( 100% efficiency) ; therefore, net ef iciency after 43 minutes was 75.0% ( 13.12/75 x 100 ) . b) Calculated value includes CaCl formed from Ca (OH) in lime boil.

Example 10 A zinc ferrite waste was subjected to simultaneous leach- jarosite precipitation. The ferrite was assayed and found to contain the following components in weight percent: Zn, 13.0 ; Fe , 33 .4 ; Pb , 0 .325 ; H₄ , 1. 12 ; K , 0.095 ; and Na, 0 . 104 . Silver was present in the amount of 5 .06 ounces per ton of waste. The following procedure was used to leach the solid ferrite waste. The ferrite was mixed with H₂S0₄ (150 g/1 H₂S0₄) and NH₄)₂S0₄ at 20 percent weight/weight solids and heated at 90-95°C. The H₂S0₄ was maintained at the target level indicated in Table 10A by the periodic addition of H₂S0₄ solution (150 g/1 H-,S0₄) . The mixture was agitated for 24 hours with samples taken periodically as indicated in Table 10B. The slurry was filtered to remove solids and the leachate was analyzed with the results given in Table 10B.

Table 10A

Run No. 43 44 45 46 47

Variable conditions

Leach solution

Initial H₂S0i' g/i 100 100 100 150 100 lb/ton 750 750 750 1125 750

Initial (NH₄ >₂so₄, g/i 50 150 84 84 84 lb/ton 375 1125 625 625 625

Target H-SO , g/i 30 30 22 58 30

Table 10B

Run No. 43 ^' 44 45 • 46 47

Elemerit Time, hr % Extracted

Zn 1 43 28 58 38 33

4 57 45 73 58 51

8.5 60 53 81 76 59

11.5 66 53 82 83 64

24 80 65 89 89 75

Ag 1 2.9 6.2 5.2 5.2 11

4 4.4 9 4.4 4.8 4

8.5 5.1 6.6 4.7 5.8 4

11.5 6.2 5.1 3.3 6 4

24 1.8 2 1.8 2.3 1

Fe 1 17 17 34 18 18

4 6.8 8.8 20 12 8

8.5 2.9 3.8 12 16 6

11.5 3.5 3 9.2 16 5

24 2.7 2.6 5 15 2

Pb 1 0 4.4 0 0 4

4 1.5 4.7 1.9 0 2

8.5 0 0 0.5 0 2

11.5 0 3.3 0 0 2

24 0 0 0 0 2 Example 11

Residues from Runs 45 and 46 of Example 10 were leached with CaCl₂ at 95°-100°C (45a and 46a) and 105°C

(45b and 46b) . The procedure of Example 1 was used to leach the residue from the ferrite leach. The results are presented in Tables 11A and lie

Tab. .e 11A

Run No j 45a 45b

Element Time, hr Extractions, %

Zn 1 11 52

3 7.4 36

5 8.3 38 g 1 22 99

3 19 98

5 24 98

Fe 1 0.1 0.5

3 0 0

5 2.2 0.4

Pb 1 18 54

3 4.2 50

5 0 35

The recovery for the overall extraction, i.e. the ferrite extraction (given in Table 10B) and the CaCl₂ extraction (given in Table 11A at two temperatures) for Run 45 are given in Table 11B.

Table 11B

5 hour Extractions Leach(45a) Leach(45b)

Element Zn Ag Fe Pb Zn Ag Fe Pb

Extractions %

Ferrite Leach 89 1.8 5 0 89 1.8 5 0 ^CaCl₂ ^Leach 8.3 24 2.2 0. 38. 98 0.4 35.

Overall extraction, % 90 25 7 0 93 98 5 35 Table llC

Test No , 46a 46b Element Time, hr Extractions , %

Zn 1 9.5 3.9 3 23 37 5 23 43

Ag 1 8.4 14 3 41 92 5 71 92

Fe 1 1.4 0.8 3 1.2 0 5 0.2 0.2

Pb 1 18 6.2 3 4.2 54 5 0 39

The recovery for the overall extraction , i . e. the ferrite extraction (given in Table 10B) and the CaCl₂ extraction (given in Table llC at two temperatures) for Run 46 are given in Table 11D .

Table 11D

5 hour Extractions Leachl ^'46a) Leach(46b)

Element Zn Ag Fe Pb Zn Ag Fe Pb Extractions % Ferrite Leach 89 2.3 15 0 89 2.3 15 0

CaCl Leach 23_ 71 0.2 0. 43. 92 0.2 2_L

Overall extraction, % 92 72 15 94 92 15 39

Example 12 Filtration rate determinations were performed on final slurries from 15 of the jarosite leach tests. The filtration data are summarized in Table 12. Rates were

2 determined using an Eimco 0.1-ft vacuum filter leaf apparatus, fitted with a medium-weave polypropylene filter cloth. The apparatus was top-loaded with leach slurry and 18 to 22 inches Hg of vacuum was applied. The filter cakes were washed first with hot CaCl- brine followed by water, and the washing rates determined. The anomalously high wash rates of Runs 48, 49, and 50 probably are due to cake cracking (channeling) and the values are not included in the group averages.

The filtration rate averages in Table 12 show that the elevated temperature leach slurries, with solids

2 rates of 80 to 230 lbs/ft /hour, filter two to three times faster than the slurries from 95-100°C leaching. Two of the leach slurries were flocculated with Percol 351 prior to filtration. Floσculant doses of approxi¬ mately 200 ppm (solids basis) were required to coagulate the slurry solids. The flocculated filtration Runs 51 and 52, produced inconsistent results as shown in Table 12.

Table 12

Filtration Rates

. Cake B _,ri•ne^b H O^b Thickness

Run No. Solid³ Filtrate Wash Wash (Inches)

95-: L00°C Leaches

53 9.6 4, .0 - - 1/4

54 50 24 7.0 2.1 1/8

51 99 , 30 10.6 9.0 3/16

55 31 9. .3 3.2 6.0 1/8

52 31 9. ,5 3.5 3.6 1/4

1 44 18 4.6 4.5 5/16

50 81 42. (32) (30) 3/16

Average 56 22 5.8 5.0 3/16

(excluding Run 53)

130° C Autoclave > Leaches

56 228 95 16 16 5/16

48 83 43 31 20 1/8

57 91 35 6.8 5.9 1/4

58 120 42 13 15 3/16

59 130 54 5.8 5.8 1/4

5 203 83 7.9 6.7 3/16

Average 142 59 9.9 9.9 3/16-1/4

180 and 225 °C Leaches

49 111 44 21 27 1/4

60 75 30 7.3 11 1/4 3 a) Solids filtration rate units: lbs/ft /hour. b) Liquid filtration rate units: gal/ft /hour. Example 13 Calcium carbonate was used for pH control in a leach of Sample 3 feed. The leach was conducted for 10 minutes at 180°C in the nickel bomb reactor described in Example 3. The run was made under the following conditions and the results are given in Table 13.

Feed Sample 3

Leach Solution 330 g/1 CaCl-,, Sp.Gr. 1.24,(6.0 N Cl) K in Solution 25 g/1 initial Initial % Solids 20.0% (wt) Target Leach pH 1.8-3.5 CaC0₃ required 283 lb/ton, added initially (Equiv. to 210 lb/ton Ca(0H)₂)

Temperature 180°C (performed in nickel bomb reactor)

Heat-up Time 3.5 minutes

Time 10 minutes

Final Slurry pH 5.0

Table 13

Extraction*, % Zn Ag Fe Pb 25.6 53.2 0.2 43.6

* metal extractions were calculated from feed and residue assays.

Example 14 Several leaches were conducted at different chloride levels. Leach solutions of 1.8 normal, 6 normal and 10 normal chloride were used. The conditions were similar to those of Example 1. The results are given in Table 14. Table 14

Leach Solution, g/1 CaCl₂ 100(1.8NC1~) 330 (6NC1~) 550(10NC1~)

Feed Sample 1 Sample 2 Sample 2

Extraction, % Ag 15 70 74

Pb 8 84 89

Zn 38 41 42 Example 15

A direct recycle of leach filtrate to the next stage of leaching was conducted to determine the effect of recycle and impurities build-up on the recovery of values. Five leach cycles were conducted using Sample 3 as feed. The first three cycles used the strongly agitated 2-liter autoclave, while the final two stages were performed in the less well agitated nickel bomb reactor. The chloride concentration of the solution was determined before and after each cycle and was maintained at 6 normal Cl (330g/l CaCl₂) by addition of CaCl-,, if required. The 16.8% solids leaches were performed for

^' one hour at 180°C and a pH of 1.8 to 4.0 by initial addition of Ca(0H)₂. The leach filtrate from one cycle was advanced to the next cycle with no intervening solution treatments other than reestablishing 6 N Cl concentration, if required.

The results in Table 15 show a steady decrease in the apparent Pb extractions with leach cycling. From cycle 1 to 4, the concentration of Pb in solution in¬ creased from 15 to 45 g/1 (calculated) , which approaches Pb saturation in 6 N CaCl₂ brine at 50 to 70°C. Leach slurries were cooled to this temperature range prior to removal from the reactor and filtration. It was found that the reduced Pb extractions in cycles two through five were due to saturation of the cooled solutions, causing bCl₂ to crystallize in the residue solids. The standard residue washing procedure, two to three cake displacements with 80°C CaCl₂ brine followed by an equal amount of water , was not effective in totally removing the PbCl₂ from the residues , thus producing low extrac¬ tion values.

The true leach cycle extractions for Pb ( "Pb-corr" column in Table 15 ) were determined by repulping the residues in 80 to 90 °C CaCl₂ to dissolve the residual PbCl₂ prior to reassay of the solids . The corrected Pb extractions show that leach cycling does not affect the efficiency of the CaCl- leach significantly , provided that the leach slurry is filtered and washed under conditions (temperature and wash volume) which assure complete dissolution of marginally soluble species such as PbCl- .

Table 15

Run Cycle Reactor Extraction, % ,_τ. Filtrate Assay, g/i

No. No. Type Zn Ag Pb Pb-corr Zn Ag Pb

12 1 A 37 76 86 86 7.0 0.071 (15)

12a 2 . A 31 76 80 87 11.6 0.135 (26)

3 A 35 75 76 84 18.9 0.191 (36)

^12b(2) 12c ' 4 B 32 72 72 84 (25.4) (0.264) (45)

12d 5 B 31 72 45 81 (21.1) (0.253) (35)

Values in parentheses are calculated.

(1) Extractions calculated from rewashed residues . The Pb-corr ( corrected) column presents the actual extraction achieved in the leach.

(2) Filtrate was diluted to 67% strength prior to advancing to 12d leach. Dilution occurred when washing slurry from bomb reactor .

Example 16 A process simulation using closed-cycle steps was performed using Sample 1 as feed. The following process steps were included: CaCl₂ leaching, Ag/Cu cementation with zinc , Pb precipitation as sulfide , Zn precipitation as hydroxide , H3 evolution by lime boil, and recycle of the processed CaCl₂ solution to the next stage of leach¬ ing. The values (Ag, Pb , Zn) were removed from the leach solutions prior to recycle to the next leach. The results are given in Table 16 . Table 16

Cycle Element

Process Step No. Zn Ag Pb Cu Fe NH,

4

Extractioni, % :

CaCl Leach 1 46 94 94 47 13 (?)

2 48 97 92 51 0.01

3 38 96 61^a 18 0.01

Average 44 96 82 39 4

Highest 48 97 94 51 0.01 b Precipitation,%:

Ag/Cu 1 94 0.8 99

Cementation 2 98 21 99

3 98 99.6 97

Average 97 41 98

Highest 98 99.6 99

Precipitation,%:

Lead 1 22 42 84

Precipitation 2 5 46 53

3 0 0 0

Average 13 44 68

(excluding Cycle 3) Highest 22 46 84

Precipitation,%:'

Zinc 1 81 - 55 -

^'Precipitation^' 2 70 - 0.4 -

3 62 - 9 -

Average 71 22

Highest 81 55

Precipitation,%: Evolution,%

Lime Boil 1 48 - 0.6 72

2 81 - 7 -

3 74 - 4 -

Average 68 4 -

Highest 81 7 72 a) Corrected value after rewash of residue. Extraction calculated from original residue (incompletely washed) was 12.0%. b) % of specie in inlet liquor to particular process step.

It will be understood that the above description of the present invention is susceptible to various modifica¬ tions, changes and adaptations and the same are intended to be comprehended within the meaning and range of equivalents of the appended claims.

Claims (20)

What Is Claimed Is:

1. In a process for recovering metal values from a jarosite-containing waste by leaching said waste with an acidic solution of a metal chloride, the improvement comprises leaching said waste with a solution comprising calcium chloride at a temperature above the atmospheric boiling point of the solution and under a pressure of at least the autogeneous pressure.

2. The process of Claim 1 wherein said temperature is in excess of about 110°C.

3. The process of Claim 1 wherein said solution has a pH of about 1.5 to about 3.5.

4. The process of Claim 1 wherein said temperature is between about 120°C and about 300°C.

5. The process of Claim 1 wherein said calcium chloride concentration is between about 1.0 molar and the saturation point of the solution.

6. The process of Claim 1 wherein said temperature is between about 150°C and about 220°C.

7. The process of Claim 1 wherein potassium is present at a concentration greater than about 0.5 grams per liter of said solution.

8. The process of Claim 3 wherein said pH is maintained by adding a compound selected from the group consisting of calcium oxide, calcium hydroxide, calcium carbonate and mixtures thereof.

9. The process of Claim 1 wherein chloride is provided by adding calcium chloride to said solution.

10. A process for recovering metal values from a jarosite-containing waste wherein said process comprises contacting said waste at a temperature of between about 120°C and about 200°C and a pressure of at least the autogeneous pressure with a solution containing between about 2.0 and about 4.0 molar calcium chloride wherein said solution has a pH of between about 1.5 and about 3.5 which is maintained by the addition of a calcium compound selected from the group consisting of calcium oxide, calcium hydroxide, calcium carbonate and mixtures there¬ of.

11. The process of Claim 10 wherein said jarosite- containing waste comprises ammonium jarosite formed by subjecting a material containing zinc ferrite to a leach said process consisting essentially of:

(a) mixing an aqueous slurry of said material with sulfuric acid and a source of ammonium ions to form a sulfuric acid leach mixture;

(b) heating said sulfuric acid leach mixture to form a solid containing ammonium jarosite and a liquid containing zinc sulfate; and

(c) conveying said solid into contact with said solution of calcium chloride.

12. The process of Claim 10 wherein said leaching provides a liquid leachate and a solid residue and wherein said process further comprises:

(a) separating said liquid leachate from said solid residue;

(b) contacting said liquid leachate with a reducing metal to reduce silver cations contained in said leachate to metallic silver and then recovering said metallic silver from the liquid phase;

(c) recovering zinc from the liquid phase of step (b) using a zinc recovery process to provide a liquid solution substantially free of zinc; and (d) adjusting the pH of the liquid solution from step (c) to above about 9 by adding a basic material and then heating the solution to provide a vapor containing ammonia.

13. The process of Claim 12 wherein said zinc is recovered in step (c) by a process comprising extracting said zinc by contacting the liquid phase containing said zinc with an extractant to remove said zinc from the liquid phase to a second phase.

14. The process of Claim 12 wherein the basic material from step (d) is selected from the group con¬ sisting of calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof.

15. The process of Claim 14 wherein said basic material consists essentially of calcium oxide.

16. The process of Claim 12 wherein said reducing metal is selected from the group consisting of lead and zinc.

17. The process of Claim 16 wherein said reducing metal is lead and whersin the process further comprises the steps (i) mixing the liquid phase remaining in step (b) after removing said metallic silver with a sulfide compound to precipitate the lead as lead sulfide, and (ii) removing substantially all of said lead sulfide from the solution and then conveying said substantially lead-free solution to the zinc recovery process.

18. The process of Claim 16 wherein said reducing metal is lead and wherein the process further comprises the steps of:

(a) contacting the substantially zinc-free liquid phase from step (c) with sufficient calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof to provide a solution pH of between about 8 and about 9 to precipitate lead hydroxide;

(b) removing the precipitated lead hydroxide from the liquid phase; (c) adding sufficient calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof to the liquid phase from step (ii) to increase the pH to above about 10; and

(d) increasing the temperature of the resulting solution to about the boiling point of said solution.

19. The process of Claim 12 wherein said zinc is recovered by contacting said liquid phase from step (b) with calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof to precipitate said zinc as zinc hydroxide and separating said zinc hydroxide precipitate from the liquid phase.

20. A process for recovering metal values from a jarosite-containing waste, said process comprising:

(a) mixing an aqueous slurry containing a material comprising zinc ferrite, sulfuric acid and a source of ammonium ions to form a sulfuric acid leach mixture;

(b) heating said leach mixture to form a solid phase containing ammonium jarosite and a liquid phase containing zinc sulfate;

(c) separating said solid phase and said liquid phase;

(d) contacting said separated solid phase with a solution comprising between 1.0 and 5.0 molar calcium chloride at a temperature between about 110°C and about 300°C and a pressure of at least the autogeneous pressure of the solution wherein said solution has a pH of between about 1.5 and about 3.5 to form a liquid leachate and a solid residue;

(e) separating said liquid leachate from said solid residue;

(f) contacting said liquid leachate with a reducing metal selected from the group consisting of lead and zinc to reduce silver cations contained in said leachate to metallic silver;

(g) separating said metallic silver from the liquid phase as a cement;

(h) recovering zinc contained in said liquid phase by contacting said liquid phase with an extractant which selectively removes said zinc from said liquid phase to a second phase;

(i) contacting the substantially zinc-free liquid phase from step (h) with sufficient calcium oxide, calcium hydroxide, calcium carbonate or mixtures thereof to provide a solution pH of between about 8 and 9 to precipitate lead hydroxide;

(j) removing the precipitated lead hydroxide from the liquid phase; (k) adding sufficient calcium oxide, calcium hydroxide, calcium carbonate' or mixtures thereof to the liquid phase from (j) to increase the pH to above about 10;

(1) increasing the temperature of the resulting solution to form a vapor containing ammonia and a liquid residue containing calcium chloride brine;

(m) condensing the vapor from step (1) to provide a solution containing ammonium hydroxide; and

(n) recycling the liquid residue from step (1) to the calcium chloride leaching of step (d) .