CA1203083A - Process for the recovery of non-ferrous metals from sulphide ores and concentrates - Google Patents

Abstract

-a-ABSTRACT OF THE DISCLOSURE
This invention relates to a process to recover non-ferrous metal such as zinc, lead, copper, and precious metals either directly from their ores or from bulk or dirty con-centrates produced from their ores. The development of a variable metallurgical process which would be capable of treating low grade concentrates, bulk concentrates or the ore directly is of interest for massive fine grained sulphide ore found throughout the world. Thus the present specification discloses a method for recovering zinc from sulphidic ores and concentrates comprising the following consecutive steps:
a) roasting said material in a subdivided form in a fluidized bed reactor at a temperature from 620 to 700°C, with 20 - 60% excess air or oxygen over the stoichiometric for an average retention time of over 1 hour and at a superficial special velocity of the gas within the bed of 25 - 150 cm per sec at reactor temperature resulting in an atmosphere comprising SO2, SO3, water vapour, and remaining O2 and N2 from the air, to obtain a calcine containing the zinc primarily as sulphates and oxysulphates and most of the iron as hematite;
b) leaching said calcine with water or dilute sulphuric acid solution at a temperature below 80°C in such a manner that the sulphates, oxysulphates, and oxides of zinc are leached out in part from the calcine into the water or dilute sulphuric acid solution;
c) subjecting the leach pulp resulting from step (b) to a liquid-solid separation step to yield a leach solution suitable for purification-metal recovery steps for zinc;

- b -d) leaching the solid residue resulting from step (c) with hot, strong sulphuric acid solutions, at a temperature above 80°C but below the boiling point of the solution and with sulphuric acid solution stronger than used in step (b), in such a manner that most of the zinc ferrite and unreacted sulphide of zinc are converted to the sulphates of iron and zinc which dissolve in the leach solution;
e) subjecting the pulp resulting from step (d) to a solid-liquid separation step to obtain a leach solution containing said sulphates of iron and zinc;
f) recycling said sulphates of zinc and iron obtained in step (e) in part into said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3;
g) washing the solid residue resulting from step (e) with water;
h) subjecting any pulp resulting from step (g) to a solid-liquid separation step to obtain a washed leach residue and a wash liquor containing zinc and iron sulphates and sulphuric acid; and i) recycling said sulphates of zinc and iron resulting from step (h) in part to step (b) and/or in part to the said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3.

Description

~2~ 33 ThiS invention xelates to a process to recover non-ferrous metal such as zinc, lead, copper, and precious metals either directly ~rom their ores or from bulk or dirty concentrates produced from their ores. The develop-ment of a viable metallurgical pxocess which would becapable of treating low grade concentrates, bulk concentrates or the ore directly is of interest for massive fine grained sulphide ores found throughout the world.
Metallurgical processes presently in practice require concentrates of high grade in the metal for which the process was designed to recover. Decreases in grade of a few weight percent can make the processing uneconomic. Concentrations of other metal in these concentrates can make their treat-ment technically unfeasible. For example, lead concentration in zinc concentrates must be below 3 weight percent to be treated in conventional dead roast, electrolytic zinc plants.
In addition, recoveries of these impurity metals are usually low in present metallurgical plants. Hence the value of the ore is significan~ly reduced when thé ore responds poorly to conventional selecti~e flotation practices for producing separate high grade concentrates. Such is the case for most of the worldls fine ~rained massive sulphide ores.
Poor recoveries to separate high grade concentrates and high cross-contAm;n~tion in these concentrates have resulted in the development of only a few of the highest grade deposits.
This invention affords a solution to the problems impeding development in that relatively low grade concentrates including bulk concentrates or ore without preco~c~ntration can be treated, with resultant high metal recoveries~ The only requirements for the process are that smelter feed, either ore or concen~rate be relatively sulphide rich and that magnesium and sodium be sufficiently low to avoid excessive buildups within the recycle stream in the process.
Most ores and concentrates from massive sulphide orebodies will meet these requirements without pretrea~mentO Simple bulk flotation, gravity separationt or acid wash pretreat-'~' 12~36~83 ~ 2 ment will be xequired for some ores and concentrates.
This invention comprises impro~ements and refine-ments to the technology described in our U.S. Patent 4,224,122 and correspondin~ Canadian Patent 1,121,605, issued April 13~ 1982, e~tending its application to the treatment of high lead bearing ores and concentrates and lower ~rade æinc ore and concentrates. The disclos ures of those patents are incoxporated herein by reference.
The earlier patents disclosed a method for recovering non-ferrous metals hy thermal treatment of solutions containing non-ferrous and iron sulphates. It was a necessary condition of that invention that the non-ferrous metals to be recovered be solu~le in sulphuric acid, thus re-stricting its application to such me~als as zinc and copper, and other metals for which the sulphate salts and oxides are sulphuric acid soluble. The treatment of ores and concentrates containing large (>4 wt percent) amounts of lead in addition to acid soluble metals was not disclosed in this patent. Also, the invention as described in the earlier patents has difficulty in treating, economically, low grade concentrates and ores in that metal recoveries decrease rapidly when the grade of the ore or concentrate decreases below 20 wt percent combined acid soluble non-ferrous metal content.
The primary object of this invention is to provide a process for recovering zinc from sulphidicsources such as directly from ores and low-grade zinc-beaxing concen-trates regardless of zinc or iron contents and with less difficulty than previous methods involving complex iron~
removal processing.
Another object of this invention is to provide such a zinc-recovering method as above, wherein the lead content of the ore or concentrate is not a process limiting para-meter in that it does not affect the zinc~recovery method.
Yet another object of this invention is to provide such a zinc-recovering method as above, wherein other valuable non-ferxous metals such as lead, copper, silver, ?~ ~J

~L2~ 3 and gold present in the ores or concentrates are also rendered recoverable.
A further ob~ect of this invention is to provide a zinc-recovering method as above, wherein effective use of,the potential energy contained within -the sulphidic ores and concentrates is made.
A still further object of this invention is to provide a zinc recovering method as ahove, wherein sulphur is recovered in a valuable form from ~he sulphidic sources.
The last object of this invention i5 ~0 provide such a zinc-recovering method as above, wherein said method can be combined with a conventional dead roasting method to the overall bene~it of the zinc recovery process.
In one aspect of this invention there is provided a method for recovering zinc from sulphidic ores and concentrates comprising a) roasting said material in a subdi~ided form in a fluid-ized bed reactor at a temperature from 620 ~o 700C, with 20 - 60% excess air over the stoichiometric for an average retention time of over 1 hour and at a superficial spacial velocity ol the gas ~ithin the bed o 25 - lS0 cm per sec at reactor temperature resulting in an atmosphere comprising SO2, SO3, water vapour, rPm~;n;ng 2~ and N2 from the air, to obtain a calcine containing the zinc primarily as sulphates and oxysulphates and most of the iron as hematite;
b) leaching said calcine with water or dilute sulphuric acid solution at a temperature below 80C in such a manner that the sulphates, oxysulphates, and oxides of zinc are leached out in part from the calcine into the water or dilute sulphuric acid solution;
c) subjecting the leach pulp resulting ~rom step (b~ to a'li~uid-solid separation step to yield a leach solution suitable for purification--metal recovery steps for zinc;
d~ leaching the solid residue resulting from step (c~
with hotr strong sulphuric acid solutions, at a tempera ture above 80C but below the boiling point of the solution and with sulphuric acid solution stronger than , ~2~3~83 ~ 4 _ used in step (b), in such a manner that most of the zinc ~errite and unreacted sulphide of zinc are converted to the sulphates o~ iron and zinc which dissolve in the leach solution;
e) subjecting the pulp resulting from step (d) to a solid-liquid separation s~ep to obtain a leach solution cont~;n;ng said sulphates of iron and zinc;
f) recyclin~ said sulphates of zinc and iron obtained in step (e) in part into said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3;
g) washing the solid residue resulting from step (e) with fresh water or recycled water or solutions;
h) subjecting any pulp resulting from step (g) to a solid-liquid separation step to obtain a washed leach residue anda wash liquor containing zinc and iron sulphates and sulphuric acid; and i) recycling said sulphates o~ zinc and iron resulting from step (h) in part to step (b) and/or in part to the said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3.
The reference throughout this specification and claims to the term "excess air" indicates the use of air, or oxygen-enriched air, or the like, as will be clear to an~ person skilled in the art. Excess air (oxygen) will thus provide the required 20 to 60~ excess of oxygen over the stoichiometric requiremen~s of the fluidized bed reactor system, for oxidation of the various components mentioned therein.
In a further aspect o this invention there is provided a method set forth in the immediately preceding paragraph, in which calcine from a dead roaster and calcine from the sulphation fluidized bed reactor are leached with dilute sulphuric acid solution(s) and the hot, strong sulphuric acid leach solution is rec~cled to the sulphation roaster.
~he present invention will be more fully apprecia~

2~ 33 ted by the follo~ing detailed descxiptions of embodiments o~ the invention~ referring to the accompanying diagrams, ~n which:
Figure 1 is a schematic flow sheet illustrating an embodiment of the sulphation roasting process applied either directly to an ore or to a low grade zinc or bulk zinc-lead-copper-precious metal concentrate according to this invention;
Figure 2 is a schematic flow sheet illustrating an embodiment of the integrated sulphation roasting and dead roasting processes applied to an ore or a low grade ~inc-bearing concentrate or bulk concentrate and a high grade zinc concentrate, respectively, according to this invention.
In U.S. Patent 4,224,122 the difficulties and com-plexities encountered in conventional electrolytic zincplant practice of treating and separating the iron in solution after the ferrite residues are leached were shown to be substantially eliminated by thermal decomposition of the hot, strong sulphuric acid leach solution at 600 -750C in a fluidized bed reactor. use was made of theexcess heat produced from the oxidation and sulphation reactions of the ore or concentrate being roasted in the bed of this fluidized bed reactor. The zinc and other metals such as copper remain as solid sulphates and oxy-sulphates in the roaster calcines and can be xecoveredtotally or in par~ with a water or dilute sulphuric acid leach. All iron sulphates in the strong acid leach sol-ution are decomposed in the roaster to hematite and SO2 and SO3, and the sulphurlc acid to H2O, SO2, and SO3.
The hematite remains inert in the subsequent water or dilute sulphuric acid leach and the hot strong acid leach.
The process as described in U.S. Patent 4,224,122 relies on the excess heat generated in the roaster to thermally decompose all the hot strong acid leach solution and hot strong acid leach residue wash solutions while maint~;n;ng the temperature in the bed of the roas~er at 600 - 750C. Tf there is insufficient excess heat , ...

~ 6 --available in the roasteX some ~f the hot strong acid leach solution or wash liquor must ~e neutralized and discarded, resulting in losses of valua~le metals such as z-inc and copper. Efficient washing is required to reduce ~he soluble zinc and ~ron sulphate and sulphuric acid content of the hot strong acid leach residu~ to enable economic recovery of precious metals and lead. For ores and concentrates containing less than 20 weight percent combined acid soluble non-ferrous metal content or less than 30 weight percent total non-ferrous metal content the roaster gener-ally produces insufficient excess heat to decompose all the hot strong acid leach solution and wash liquors even when excess heat production is m~;m;zed by predrying the ore or concentrate to very low, less than 1 weight percent, moisture contents. Simila~ly, some concentrates exhibit physical and chemical properties which result in the production of calcines, during fluidized bed roasting under sulphating conditions, cont~;n;ng greater than 15 percent of the zinc as zinc ferrite. Even when these concentrates contain greater than 30 weight percent zinc and excess heat is ~x;m; zed ~y predrying there is insu~ficient excess heat in the bed of the roaster to decompose all the hot strong acid leach solutionO
The present invention effects a significant increase in the utilization efficiency o~ the excess heat to the extent that ores directly (as mined); and very low grade concentrates; and ores and concentrates generating calcines containing up to 45 weight percent o~ -the zinc as zinc ferrite; can all be processed economically by thermally decomposing all the hot, strong acid leach solution. Wash liquors may be thermally decomposed with excess heat generated in the roas~er and/or recycled to the water or dilute sulphuric acid leach. ~hus much more water may be emplo~ed in the washing of the hot, strong leach residue, xesulting in lower zinc and iron sulphate and sulphuric acid contents o~ the residues and hence lowex reagent costs in any subsequent precious metal or lead recovery steps.
i ~3iQ183 ~ 7 -This enables the economic recovery- o~ lead and precious metals which were previousl~ too lo~ to warrant recovery.
Referring now to Figure 1, klle zinc bearing ore or concentrate is fed to a fluidized ~ed reactor. This ~eeding may ~e accomplished hy spraying as a slurry into the bed as shown in Figure 1 or feeding directly as a filter cake or pneumatically in a dry sub-divided form.
Slurry feeding offers several advantages over dry or filter cake feeding. Firstly, when the process is situated nearby the mine or mill from which the ores or concentrates are received, slurry feeding would eliminate most of the expensive filtering and drying required for ilter cake or dry feeding. Secondly, slurry feeding has been shown ~o result in the formation of calcine pellets in the bed of the roaster which result in a broader size distri~ution, better fluidization, and a lower calcine elutriation rate from the roaster thus resulting in higher metal reco~eries. Although this pellet formation occurs with water and/or hot strong acid leach solution as slurry medium, the pellets exhi~it better physical characteristics such as a slower attrition rate and contain lower ferrite contents when hot, strong acia leach solution is employed.
These better fluidization characteristics aid in the treatment o~ high lead contA; n; ng ores and concentrates.
In some cases, however, dr~ or filter cake feeding is preferred due to the generally lower ferrite content of calcines produced with these feed modes. Accordingly, wh~n the potential energy content of the ore or concentrate is low thereby resulting in only small amounts o excess heat produced during roasting, and when the ore or concentrate roasting generates significant ferrite it may be necessary to partially dewater the ore or concentrate prior to roasting and feed as a ~ilter cake or pneumatically dry, thus increasing the excess heat available for hot strong acid leach solution decvmposition. In these cases, the hot strong acid leach solution would be introduced separ-ately to the roaster in a suh~divided form.

.~

~Z~3~83 ~ 8 -The ro~sting is pex~ormed at a ~ed temperature between 620 and 700CI with 20 to 60~ excess air (oxy~en) over the stoichlometric requirement, for a reten~ion time of the calcine in the bed of greater than 1 hour, and at a super~icial spacial velocity of gas within the bed of 25 ~o 150 cm per sec. The spacial velocity m;n;mllm is required to ensure adequate fluidization and prevent sintering of calcines containing high lead values~ Spacial velocities greater than 150 cm per sec are not recommended due to high valuable metal losses due to elutriation of calcines from the reactor.
The temperature minimum is required to prevent the formation of excessive amounts of fexrous and/or ferric sulphates, maintain the kinetics of ~he sulphide roasting sufficiently fast so as to keep the retention time reasonably low, and ensure efficient conversion of the zinc su~phide minerals, namel~ spalerite or marmatite, to zinc sulphate. The temperature maximum is required to prevent the formation of significant quantities of zinc ferrite incorporating zinc from the recycledhot, strong acid leach solution. Also, temperatures above 700C may result in bed sintering when high lead cont~;n;ng ore or concentrate is roasted.
The excess air requirement is kept abo~e 20 percent to ensure sufficien~ air to maintain fluidization and ehsure eficient sulphation of sulphides at reasonable xetention times. Air in excess of 60 percen~ of s~oichiometric will result in excessive dilution of the off-gas stream resulting in higher costs for gas cleaning, gas cooling, and sulphuric acid manufacture~ Also the zinc ferrite content of the calcine will increase with increasing excess air. Accordingly, the object of the roasting operation is to produce a calcine with maximum zinc and copper sulphate composition and minimum sulphide and iron sulphate composition, with m~; mllm heat generation and SO2 strength in the off-gas.
All calcines produced in the roasting operation are discharged directly from the bed by mechani~ms such as bed ,. . .
i

3~83 ~ g o~e~low~ This ensures contxol o~ Xetention time of calcines at the bed temperature xesultin~ ln hi~h con~
version o~ non-ferrous met~l sulphides to sulphates; and iron sulphides and sulphates to hematite. This calcine is cooled reasonably quickly in an atmo~phe~e from which roaster gas levels ha~e been reduced by mechanisms such as rotary star ~alves, a~r lvcks, counter-current air ~lows etc. The calcine cooling can be accomplished by man~ devices, some examples of which are rotary coolers, fluidized hed coolers, and direc~ quenchiny in water or dilute sulphuric acid soluti~n. These procedures are required tu mln;m;ze the amount of hematite resulphation to iron sulphates.
The cooled calcines are leached in water or dilute sulphuric acid but generally dilute sulphuric acid. This operation referred to as a "Neutral Leach" is carried out under the ~ollowing condit;ons: ambient temperature but not exceeding 80C, pH 7.0 or lower throughout the leach, and for a retention time of not less than one-half hour.
The pH at the termination of the leach must not be lower than 1.5 and preferably hi~her than 3.5 to ensure the pxecipitation of most of the arsenic and most of the iron, which incomes as sulphates to the leac'n, as hydroxides and basic sulphates. This iron precipitation is required to purify the leach solution of metal ions injurious to the conventional solution purification and zinc recovery operations which follow. Accordingly, the object of the "Neutral Leach" operation is to m~;m; ze leaching and dissolution of zinc and copper sulphates and oxysulphates and minimize iron and arsenic leaching and dissolutionO
The ~ater or dilute sulphuric acid used in the leach can be from any o~ the following sources; recycled wash solution from the hot strong acid leach residue wash operation~ recycled spent electrol~te fxom zinc electro-winning cells, recycled hot acid leach solution, or ~reshor recycled water ~rom within the plant. Preferably, all the water xe~uirement ~or the leach is satisfied by wash i ~2~ 3~83 1~ -solutions from the hot acid leach ~esidue w~sh circuit, thus permitting maximu~ wash ~ater Usaye and thus hi~h ~ashing ef~iciencyO If this solution contains too much acid such that the final pH of the "Neutral Leach" is below the pH required for effecti~e iron precipitation, then a neutralizing agent such as limestone, lime, or a calcine containing zinc oxide from a dead roasting operation must be added to raise the p~ to the required value at termination. This situation will arise when 1~ sulphation roast calcines which contain low zinc oxy-sulphate and high iron sulphate valves are leached and/or when the final pH re~uirement is above 3.5.
The pulp at termination of the "Neutral Leach" is subjected to a solid-liquid separation operatlon. This lS operation may b~ carried out in thickeners, Glarifiers and/or filters but preferably by thickeners and~or clarifiers only. The solution is then treated by con-ventional processes for zlnc, copper, and cadmium recover~
The preferred pH at the termination of the "Neutral Leach" depends on the zinc to copper ratio in the ore or concentrate. For ratios above 30:1 the preferred pH
will be greater than 3.5 thus ensuring iron precipitation to levels less than 20 mgpl in the leach solution. For ratios below 10:1 the preferred pH will be greater ~han l.S since a copper recovery operation such as cementation on scrap iron, solvent extraction and electrowinning, precipitation as copper sulphide, or direct electrowinning will be performed to recover copper before conventional zinc electrolyte purification with zinc dust. Final pH
adjustment to greater than 3.5 which causes the precipi-tation of the rP~a;n;ng iron to less than 20 mgpl in the solution will occur in this case after copper recovery but prior to conventional zinc electrolyte purification.
The residue from the ~Neutral Leach" in the form o~ a thickened pulp or filter cake is su~jected to a hot, strong acid leach. This operation is carried out under the following range o~ conditions: temperature between .~

~3~8~3 -; 11 ~
85C and the boiling point, ~ee sulphuxic acid concentra-tion of 5 - 1$0 gra~s per litex through~ut the leach; ana xetention time of greater then 0~5 hour. Preferred con-ditions are: temperature 95aCt ~ree sulphuric acid concentration of 25 - 40 grams per liter (.pH 0.1 - G.3) for 1.0 - 3.0 hour ~y addition of sulphuric acid followed by cessation of acid addition allowing the free acid concentration to decrease to less than 15 gpl at which time the leach is terminated. The object of this leach is to maximize zinc erritet copper ferrite, sphalerite, and chalcopyrite leaching and m;n;m; ze hematite leaching and sulphuric acid consumption. Zinc sulphate, iron sulphates, and copper sulphate, which dissolve, and elemental sulphur which precipita~es into the residue are the major products o the leach reactions.
The sulphuric acid reguired ~or the leaching operation is provided by spent electrolyte from the zinc electrowinning operation. Fresh sulphuric acid is added only if additional acid is required above what is available in the spent elec~rolyte. The leach pulp is subjected to solid-liquid separation in conventional apparatus such as thickeners, filters, centrifuges etc. The residue, either a filter cake or thickened pulp, is thoroughly washed o soluble sulphates with water, ei~her fresh or recycled, by conventional methods such as counter~
current decantation, displacement washing, or repulp washing.
Due to the procedures and conditions employed in the previous operations~ most of the silver, gold, and lead contained in the ore or concentrate i5 now in a form, and contained in a residue which has favourable physical and chemical characteristics,. for recovery by con~entional methods. These methods include such techniques as cyanide leaching, brine leaching, and ammoniacal~ammonium sulphate leachin~.
The hot, strong acid leach solution resulting from I

~ ~D3~3 - 12 ~
the solid li~uid separation ste~ contains siyni~icant concentrations of zinc, iron~ and co~per sul~hates and $ulphuric acid. This solution, in whole or in part, is treated according to the method descri~ed in U.S. Patent

4,~24,122.
Accordingly, a major portton of this liquor is recycled to the sulphation roaster in which the zinc ore or concentrate is being roas~ed. ~s discussed previously, this liquor can either be sprayed into the roaster in a subdivided form separately from the ore or concentrate feed or slurried with this feed be~ore introduction to the roaster. ~he zinc and copper in the solution reports as zinc sulphate and copper sulphate in the calcine and the iron sulphates as hematite in the calcine.
Any hot, strong acid leach solution surplus to the quantity for which excess heat is available in the bed for decomposition is sent to a spray dr~er situated in the hot roaster off~gas train. The gas incoming to the spray dryer may or may not have been partially cleaned of elutriated calcine in cyclones prior to entering the spray dryer. Any calcines collected in cyclones is re-circulated to the sulphation roaster. The gas, however, has been kept as hot as possible and is at a temperature above 500C and preferably above 600C at the entrance to the spray dryer. The gas temperature exiting the dryer is maintained abo~e 300~C. ~inc, copper and most of the iron in the solution sprayed in a subdivided form into the spray dryer are reco~ered as dry sulphates in the dryer settling chamber or in cyclones following the dryer.
Water evaporates into the o~f~as, and sulphuric acid decomposes to SO2, SO3. and gaseous H2O. Some of the iron sulphates may thermally deco~pose to hematite, S2 and SO3. The settling chamber and spray dryer cyclones also collect a si~nificant percentage of elutriated roaster calcines, thus effecting h;gher metal recoveries. All the material collected in the spray dryer and cyclones .
i ~..2~ 3 ~- 13 ~
is recirculated ei~her directl~ to the $ulph~tion roaster or to slurry feed tanks to be blended with the ore or concentrate and hot, strony acid recycle solution. The iron sulphates thermall~ decompose to hematite in the bed of the sulphation roaster. ~n this mannex, effective use has been made of both the excess heat a~ailabl~ in the bed of the sulphation roaster and all recoverable heat contained in the roaster off-gas.
The temperature of off-gas exiting the spray dryer must be maintained above 300C to prev~nt formation of metal sulphate hydrates, condensation of sulphuric acid, and condensation of ~olatile impurities such as mercury, chlorine, and fluorine.
Wash solution from the hot, strong acid leach residue wash operation is recirculated to the neutral leach as discussed previously and/or to the spray dryer for thermal decomposition and hence reco~ery of contained zinc and copper through reactions similar to those discussed above for the hot acid leach liquor.
In some cases, where sufficient hea~ is generated within the bed of the sulphation roaster, all the hot, strong acid leach liquor and wash solution may be thermally treated directly in the roaster. In these cases, the spray dryer may be substituted by conventional waste heat recovery and gas cleaning equipment such as waste heat boilers, cyclones, electrostatic precipitators, and hot gas filters.
Alternatively for these cases, most or all of the hot acid leach li~uor and wash liquor may be recycled to the spray dryPr to provide ~he re~uired cooling requirement and the remainder, if any, recycled directly to the roaster.
The excess heat generated in the roaster bed may be recover-ed by bed coils, thus producin~ steam ~or use in the leach-ing and purification operations ~eferring to Figure 2, the simplicity of integration o~ the present in~ention with a conventional dead roasting operation treating a relatively hi~h grade zinc concentrate ~2~3C183 s 14 ~
is shown. The zinc oxide produced in the de~d roaster serves as the neutralization ayent in the 'INeutral Leach", thus decreasin~ the ~eagent costs and resu~ting in a hot, strong acid leach residue of higher grade in silver, gold, and lead than i lime or limestone were employed as neutralizing a~ent ~oxming gypsum which precipi~ates into the residue. All ~inc and copper ferrites and sulphides present in the dead roast calcine are treated in the hot, strong acid leach, the leach liquor from which is all sent to the sulphation roaster and/or spray dryer for thermal decompositionO In this manner, all the iron entering the plantl in both ~eeds to the sulphation roaster and dead roaster, exits the process as hematite in the hot, strong acid leach residue. Also, lead and precious metals, contained in the zinc concentrate treated in the dead roaster, are recoverable from the hot, strong acid leach residue.
Example No. 1: Application to a Sulphation Roast-Leach-Electxowinning Pxocess for Zinc Containing Ores.
~n ore assaying 6.3~ Zn, 3.0% Pb, 0.4% Cu, 38.6~ Fe, 2.0 oz~mt Ag, and 0.04 oz/mt Au with principal mineral components being sphalerite, galena, chalcopyrite, and pyrite, was roasted in a continuous semi-pilot fluid-bed reactor at a rate of 150 - 380 kg/day of dry feed under conditions where zinc, lead and copper are selectively sulphated and iron forms the oxide hematite. Sulphation roasting was carried out at temperatures between 620 and 700C with air between 20 and 60% excess over stoichio-metric for retention times between 1.0 and 6.0 hours.
Superficial velocities of gas within the bed were cal-culated to be 25 - 100 cm/sec, controlled by the retention time, eed rate, and excess air settings fox the ~est.
The roastex was equipped with cyclones and a spray dryer cham~ex ~or collectin~ elutriated calcines from the o~f gases and recycling these to the roaster. The cleaned gases were cooled and neutralized in a wet scrubber.

l ' 3~1 33 ~ 15 _ Calcines from the overflow were leached continuously with dilute sulphuric acid solution at ambient temperature, 1.5 hours retention ti~e at ambient pH which was less than 4Ø The dilute sulphuric acid solution used in the leach was produced as ~ash solution from repulp washing of hot, strong acid leach residue. After 1.5 hours leachiny time, limestone and potassium permanganate were added to the leach pulp to increase the pH to 3.6 -4~0 and oxidize ferrous ion to ~erric and precipitate ferric ion as iron hydroxide.
After pH adjustment the leach pulp was thickened, the clear supernate decanted, and the thickened pulp leached in hot strong sulphuric acid. This acid was contained in a solution of 150 gpl sulphuric acid, approximating the acid content of spent electrolytP from zinc electrowinning cells. The leach was conducted for 2 hours at 95C at a pH maintained at less than 0.3 by continuous addition of the acid solution. ~fter 2 hours~
acid addition was terminated and the pH was allowed to climb to 0~4 - 0.6 which took about 1 hour. At this point the pulp was filtered hot, producing a wet filter cake and a hot, strong acid leach solution. This solution was rec~cled in part to the roaster, as a fine spray.
The hot acid leach xesidue was repulped,washed in fresh water and refiltered to produce a wash solution and a washed leach residue. Sufficient wash water was employed so as to provide wash solution for the dilute sulphuric acid leach to satisfy the water re~uirement such that the dilute acid leach supernate contained 150 gpl zinc and to provide excess wash solution ~hich when mixed with hot acid leach solution, excess to roaster cooling requirements and sprayed into a spray dryer, was suf~icient to cool the roaster off gas to 3~5C All calcines exiting the spray dryer ~e~e reçycled to the sulphation roasterO
The ~ashed hot acid leach residue was repulped in water, pH adjusted with lime, and the silver and gold recovered by leaching with sodium cyanide~ Techni~ues ~{~ 7 ~0;~13 ~ 16 ~
employed ~ere entirely conyentional.
Residue fxom the cyanide leach was then l~ached in a saturated NaCl-CaC12 brine extracting ~he remaining lead and sil~er from the hematite residue.
The results are summari~ed as shown below on Tables I and II.
I

33~

." . .~ _ r O 'C~ ~r a) ~o ~a-, r~ U
u o rl ,~ a~ O~o C n -n~r~ ~ N
~~ O r ~rl O O
W ~ ~ ~

O
) a >
I r~ J
-rl rl lr~rl r rl ~!~
H ~ ~rl S C~
l_ ~rl U ~
a, Lt E- r u ~
r ~r ~
C '~ O ~ ~ O CO
1~ . o ~ ~ ~
o~O ~rl O
o ~;1 E-~

~ C) u~ ~, tao ~ Ln u~
1) $~ ~
o -r ~
~1 J l o o U-1 Lr) O
O

TABLE II
Results of Tests at 685C Roasting Temperature Zinc Zinc Re- SO2 Content Roast Losses to covery into of Off-Gas Retention Zinc Losses Hot,Strong Dilute Acid Gases AirVelocity Time to Scrubber Acid Leach Leach Super- (Vol% dry-Excess (%) (cm/sec) _(hours~ (~) Residue (%) nate (%) basis) 3 4.7 3.1 91~9 11.8 ~0 105 1 8.2 3.0 88.8 10.1 3 5.3 2.5 92.0 10.4 3 5.3 ~.8 gl.6 9.1 Silver and gold extractions in the cyanide leach ranged between 53 - 58~ and 60 - 69% respect~'vely and appeared to be independent of the roasting conditions studied.
Lead extractions into the brine solution were 87 - g3%.
Silver extractions were 40 - 70~ of the silver r~m~;ning after cyanide leaching.
Copper recoveries into the dilute acid leach supernate were similar to zinc recoveries in the range of 87 - 91%.

3~ 3 Example. No. 2; ~pplic~ti~n to a S~lph.ation Ro~st-Leach-Electrowinning Proces~s for Zinc Containlng Ores~ Slurry Fed ~o the Roaster, An ore exactly similar to example 1 was slurried to 70 wt% in water and treated in a manner similar to example 1. Sulphation roasting conditions were fixed a~
685C, 40~ excess air o~er stoichiometric, 3.2 hours retention time, and superficial velocity of 85 cm/sec.
Leaching procedures and conditions were identical to example 1. The ma~or departure from example 1 was in the recycle of the hot, strong acid leach solution and the calcine produced and collec~ed in the spray dryer. That part of the hot, strong acid leach solution which was recycled di.rectly to the roaster as a spray in example 1 was mixed with the slurry feed in this test. Also, the spray dryer calcine was recycled to the slurry feed tank and slurried with the feed prior to introduction into the roaster. The results are summarized as follows:
Zinc loss to the scrubber 1.8 Zinc loss to the hot~ strong acid leach residue 3.2%
Zinc recovery to the dilute acid leach supernate 95.0%
Lead content of the hot, strong acid leach residue 3.9 wt%
S2 content of ~he roaster off-gas (dry basis) 10,4 vol%
Copper recovery to the dilute acid leach supernate 91~0~
Silver recovery in the cyanide leach ~1.8%
Gold recovery in the cyanide leach 68.9%
Lead recovery in the brine leach 93.0%
Silver reco~ery from the cyanide leach residue in the brine lPach 60.0%

Example No. 3: Application ta a Sulphation ~oast~Leach Electrowinnin~ Procass for ~ Bulk Zinc I,ead-Coppe~Sil~er Concentrate.
A bulk concentrate assaying 30.0% Zn, 9.0% Pb, 8.Q% Cu, 18.2% Fe, 7.9 ozfmt Ag, 0.02 oz/mt Au with principal mineral components being sphalerite, galena, chalcopyrite, and pyrite, was processed in a manner exactly similar to example No. 2 excepting gold values did not warrant inclusion of a cyanide leach stage. Eor purposes of comparison, tests were also performed with dry feeding as in example 1 and also at a low superficial velocity of 25 cm/sec.
The results are summarized as follows:
Test No. 1: Slurry Feeding - 85 cm/5ec superficial ~locity Zinc loss to the scrubber 1.3 Zinc loss to the hot, strong acid leach residue 3.0 Zinc recovery to the dilute acid leach supernate 95.6 Lead content of the hot, strong acid leach residue 15.7 wt%
S2 conten~ of the roaster off-gas (dry basis~ 4.7 vol%
Copper recovery to the dilute acid leach supernate 93.2%
Lead recovery in the brine leach 97.7%
Silver recovexy in the brine leach82.3~
Test No. 2: Dry Feeding ~ 85 cm/sec superficial velocity Zinc loss to the scrubber 6.8%
Zinc loss to the hot, strong acid leach residue 2 6%
Zinc recovery to the dilute acid leach supernate 90.2%
Lead content of the hot, stron~ acid leach residue 20.3 wt%
SO2 content o~ the roaster off~as (dry basis~ 4.5 vol%

~3 ~2~31E~3 - 21 ~
Copper ~ecovery to the dilute acid leach supernate 89~1%
Lead recovery in the brine leach 91.2 5ilvery recovery in the br~ne leach 79~1~
S Test No. 3: Dry Feeding -- 25 cm~sec superficial velocity Zinc loss to the scrubber 4.3%
Zinc loss to the hot, strong acid leach residue 3 9%
Zinc reco~ery to the dilute acid leach supernate 91.8%
Lead content of the hot, ~t~ong acid leach residue 18.3 wt%
S2 content of the roaster of~gRs (dry basis) 4.8 vol%
Copper xecovery to the dilute acid leach supernate 90.1%
Lead recovery in the brine leach 92.2%
Silver recovery in the brine leach 77.0%
In test No. 3, problems were experienced wit~ bed fluidization in the sulphation roaster due to calcine sintering after about lO hours running indicating insuf-ficient super~icial velocity for this bulk concen~rate when ~ed dry.
Example No. 4: Application to a Sulphation Roaster-Leach-Electrowinning Process for Zinc~
Lead Bulk Concentrates.
A concentrate assaying 29.6~ Zn, 25.1% Pb, 0.4% Cu, 9.4~ Fe, 28.1 oz~mt Ag, and 0.3 oz~mt Au with principal mineral components being sphalerite, galena,chalcopyrite, and pyrite was processed in a manner exactly similar to example 2, excepting no cyanide leach or brine leach was performed on the hot, strong acid leach residue since the lead grade ~as sufficiently high to qualify the residue as a high grade lead concentrate suitable for conventional lead smelters.

i ,1 , .
I

Th.e results are summarized ~.s ~ollo~s:
Zinc loss to the sc~ubber 1.5 Zinc loss to the hot~. strong acid leach residue 1.2 Zinc recovery to the d~lute acid leach supernate 97.3%
Lead content of the hot, strong acid leach residue 49.1 w~%
S2 content of the roaster o~f-gas (dry basis) 1.8 vol%
Copper recovery to the dilute acid leach supernate 97.5%
Silver content o~ the hot, strong acid leach residue 54.9 oz/mt Gold content of the hot, skrong acid leach residue 0.6 oz/mt It was determined during this test that if superficial velocities were decreased below 35 cm/sec accretions would develop on the walls of the sulphation roaster. This resulted in increased losses to the scrubber and unfavourable effects on bed fluidization resulting in eventual collapse.

:' I

Claims (32)

The embodiments of the invention in which an exclusive property or privilege is claimed are defined as follows:

1. A method for recovering zinc from sulphidic ores and concentrates comprising the following consecutive steps:
a) roasting said material in a subdivided form in a fluidized bed reactor at a temperature from 620 to 700°C, with 20 - 60% excess air over the stoichiometric for an average retention time of over 1 hour and at a superficial special velocity of the gas within the bed of 25 - 150 cm per sec at reactor temperature resulting in an atmosphere comprising SO2, SO3, water vapour, and remaining O2 and N2 from the air, to obtain a calcine containing the zinc primarily as sulphates and oxysulphates and most of the iron as hematite;
b) leaching said calcine with water or dilute sulphuric acid solution at a temperature below 80°C
in such a manner that the sulphates, oxysulphates, and oxides of zinc are leached out in part from the calcine into the water or dilute sulphuric acid solution;
c) subjecting the leach pulp resulting from step (b) to a liquid-solid separation step to yield a leach solution suitable for purification-metal recovery steps for zinc;
d) leaching the solid residue resulting from step (c) with hot, strong sulphuric acid solutions at a temperature above 80°C but below the boiling point of the solution and with sulphuric acid solution stronger than used in step (b), in such a manner that most of the zinc ferrite and unreacted sulphide of zinc are converted to the sulphates of iron and zinc which dissolve in the leach solution;

e) subjecting the pulp resulting from step (d) to a solid-liquid separation step to obtain a leach solution containing said sulphates of iron and zinc;
f) recycling said sulphates of zinc and iron obtained in step (e) in part into said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite SO2 and SO3;
g) washing the solid residue resulting from step (e) with water;
h) subjecting any pulp resulting from step (g) to a solid-liquid separation step to obtain a washed leach residue and a wash liquor contain-ing zinc and iron sulphates and sulphuric acid;and i) recycling said sulphates of zinc and iron resulting from step (h) in part to step (b) and/or in part to the said fluidized bed reactor in a sub-divided form to convert the sulphates of iron to hematite, SO2 and SO3.

2. The method as defined in claim 1 wherein the ore or concentrate is introduced into the fluidized bed reactor as a slurry.

3. The method as defined in claim 1 wherein the ore or concentrate is introduced into the fluidized bed reactor as a filter cake.

4. The method as defined in claim 1 wherein dry ore or concentrate is introduced pneumatically into the fluidized bed reactor.

5. The method as defined in claim 1 wherein the wash liquor resulting from step (h) is recycled to step (b).

6. The method as defined in claim 1 or 4 wherein the sulphates of iron and zinc obtained in step (e) are mixed in whole or in part, with the ore or concentrate and introduced into the fluidized bed reactor as a slurry.

7. The method as defined in claim 1, 2 or 3 wherein of gas from the fluidized bed reactor is fed to a H2SO4 plant.

8. The method as defined in claim 1, 2 or 3 wherein off gas from the fluidized bed reactor is fed to a spray dryer.

9. The method as defined in claim 1, 2 or 3 wherein off gas from the fluidized bed reactor is fed to a spray dryer, and wherein off gas from the spray dryer is fed to a H2SO4 plant.

10. The method as defined in claim 1 wherein part of the sulphates of iron and zinc obtained in step (e) are dried in a spray dryer, fuelled by off gas from the fluidized bed reactor, prior to recycle to the fluidized bed reactor.

11. The method as defined in claim 1 wherein part of the wash liquor resulting from step (h) is dried in a spray dryer, fuelled by off gas from the fluidized bed reactor, prior to recycle of the sulphates of zinc and iron to the fluidized bed reactor.

12. The method as defined in claim 10 or 11 wherein the sulphates of zinc and iron are slurried with the ore or concentrate before introduction into the fluidized bed reactor.

13. The method as defined in claim 10 or 11 wherein the sulphates of zinc and iron are introduced into the fluidized bed reactor in a dry form.

14. The method as defined in claim 10 or 11, wherein said temperature of the gas phase in the spray dryer is maintained above 300°C.

15. The method as defined in claim 1, 2 or 3, wherein said sulphation roasting is carried out at 685°C, with 40% excess air (oxygen), for an average retention time of 3 hours, and at a superficial spacial velocity of gas within the roaster of 50-100 cm/sec.

16. The method as defined in claim 1 wherein the step (b) is carried out such that the pH of the pulp at the termination of the leach is between 3.6 and 5.2.

17. The method as defined in claim 1 wherein the step (d) is carried out at a pH of less than 0.3, at a temperature of 95°C, for greater than 2.0 hours.

18. The method as defined in claim 1 to 3 wherein calcine from a dead roaster is leached in step (b) along with the calcine from the fluidized bed sulphation reactor.

19. The method as defined in claim 1 to 3 wherein zinc and copper are recovered.

20. The method as defined in claim 1 to 3 wherein zinc is recovered.

21. The method as defined in claim 1 to 3 wherein lead and precious metals, are recovered.

22. The method as defined in claim 1 to 3 wherein lead is recovered.

23. The method as defined in claim 1 to 3 wherein precious metals are recovered.

24. The method as defined in claim 1, 2 or 3 wherein the wash water for step (g) is recycled water or solution.

25. The method as defined in claim 1, 2 or 3 wherein said sulphation roasting is carried out at 660-690°C, with 30-45% excess air (oxygen), for an average retention time of 2-4 hours, and at a superficial spacial velocity of gas within the roaster of 40-100 cm/sec.

26. In a method for recovering zinc from material selected from the group consisting of sulphidic ores and concentrates and for rendering copper, lead, silver, and gold in said material more readily recoverable, the improvement comprising:
(a) roasting said material in a subdivided form in a fluidized bed reactor at a temperature from 620 to 700°C, with 20 - 60% excess air over the stoichiometric for an average retention time of over 1 hour at a superficial spacial velocity of the gas within the bed of 25 - 150 cm per sec at reactor temperature resulting in an atmosphere comprising SO2, SO3, water vapour, and remaining O2 and N2 from the air, to obtain a calcine containing zinc and copper as sulphates and oxysulphates, ferrites, oxides, and unreacted sulphides, iron as hematite, and lead, silver, and gold in forms amenable to recovery;
(b) leaching said calcine with water or dilute sul-phuric acid solution at a temperature below 80°C in such a manner that sulphates, oxysulphates, and oxides of zinc and copper are leached out from the calcine into the water or dilute sulphuric acid solution;
(c) subjecting the leach pulp resulting from step (b) to a liquid-solid separation step to yield a leach solution from which zinc and copper can be recovered;
(d) leaching the solid residue resulting from step (c) with hot, strong sulphuric acid solutions, at a tempera-ture above 80°C but below the boiling point of the solution and with sulphuric acid solution stronger than used in step (b), in such a manner that zinc and copper ferrites and unreacted sulphides of zinc and copper are converted to sulphates of iron, copper, and zinc which dissolve in the leach solution;

(e) subjecting the pulp resulting from step (d) to a solid-liquid separation step to obtain a leach solution containing said sulphates of iron, copper, and zinc;
(f) recycling said sulphates of zinc, copper, and iron obtained in step (e) in part into said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3 and render the sul-phates of zinc and copper to the calcine for recovery by means of steps (b) and (c);
(g) washing the solid residue resulting from step (e) with water;
(h) subjecting any pulp resulting from step (g) to a solid-liquid separation step to obtain a washed leach residue containing lead, silver and gold in forms amenable to recovery and a wash liquor containing zinc, copper, and iron sulphates and sulphuric acid; and (i) recycling said wash liquor resulting from step (h) in part to step (b) and/or in part to said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3 and enable recovery of the sulphates of zinc and copper by means of steps (b) and (c).

27. The method as defined in claim 26 wherein zinc and copper are recovered from the leach solution resulting from step (c).

28. The method as defined in claim 26 wherein zinc is recovered from the leach solution resulting from step (c).

29. The method as defined in claim 26 wherein lead, silver, and gold are recovered from the washed leach residue resulting from step (h).

30. The method as defined in claim 26 wherein lead is recovered from the washed leach residue resulting from step (h).

31. The method as defined in claim 26 wherein silver and gold are recovered from the washed leach residue resulting from step (h).

32. A method for recovering zinc from material selected from the group consisting of sulphidic ores and concentrates and for rendering any copper, lead, silver, and gold in said material more readily recoverable comprising:
(a) roasting said material in a subdivided form in a fluidized bed reactor at a temperature from 620 to 700°C, with 20 - 60% excess air over the stoichiometric for an average retention time of over 1 hour and at a superficial spacial velocity of the gas within the bed of 25 - 150 cm per sec at reactor temperature resulting in an atmosphere comprising SO2, SO3, water vapour, and remaining O2 and N2 from the air, to obtain a calcine containing most of the zinc and copper as sulphates and oxysulphates, minor amounts of zinc and copper as ferrites, oxides, and unreacted sulphides, most of the iron as hematite, and most of the lead silver, and gold in forms amenable to recovery;
(b) leaching said calcine with water or dilute sul-phuric acid solution at a temperature below 80°C in such a manner that most of the sulphates, oxysulphates, and oxides of zinc and copper are leached out from the calcine into the water or dilute sulphuric acid solu-tion;
(c) subjecting the leach pulp resulting from step (b) to a liquid-solid separation step to yield a leach solution from which zinc and copper are recovered by conventional methods;
(d) leaching the solid residue resulting from step (c) with hot, strong sulphuric acid solutions, at a temperature above 80°C but below the boiling point of the solution and with sulphuric acid solution stronger than used in step (b), in such a manner that most of the zinc and copper ferrites and unreacted sulphides of zinc and copper are converted to sulphates of iron, copper, and zinc which dissolve in the leach solution;
(e) subjecting the pulp resulting from step (d) to a solid-liquid separation step to obtain a leach solu-tion containing said sulphates of iron, copper, and zinc;
(f) recycling said sulphates of zinc, copper, and iron obtained in step (e) in part into said fluidized bed reactor in a subdivided form to convert the sulphates or iron to hematite, SO2 and SO3 and render the sul-phates of zinc and copper to the calcine for recovery by means of steps (b) and (c);
(g) washing the solid residue resulting from step (e) with water;
(h) subjecting any pulp resulting from step (g) to a solid-liquid separation step to obtain a washed leach residue containing most of the lead, silver and gold in forms amenable to recovery by conventional methods and a wash liquor containing zinc, copper, and iron sulphates and sulphuric acid; and (i) recycling said wash liquor resulting from step (h) in part to step (b) and/or in part to said fluidized bed reactor in a subdivided form to convert the sulphates of iron to hematite, SO2 and SO3 and enable recovery of the sulphates of zinc and copper by means of steps (b) and (c).