CA1160055A - Method for the recovery of valuable metals from finely-divided pyrite - Google Patents
Method for the recovery of valuable metals from finely-divided pyriteInfo
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- CA1160055A CA1160055A CA000373342A CA373342A CA1160055A CA 1160055 A CA1160055 A CA 1160055A CA 000373342 A CA000373342 A CA 000373342A CA 373342 A CA373342 A CA 373342A CA 1160055 A CA1160055 A CA 1160055A
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting processes
- C22B1/04—Blast roasting
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting processes
- C22B1/06—Sulfating roasting
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/04—Obtaining lead by wet processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/06—Refining
- C22B13/08—Separating metals from lead by precipitating, e.g. Parkes process
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/26—Refining solutions containing zinc values, e.g. obtained by leaching zinc ores
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/02—Working-up flue dust
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- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Organic Chemistry (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Manufacturing & Machinery (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Geochemistry & Mineralogy (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
ABSTRACT OF THE DISCLOSURE
A process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases is disclosed wherein a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; or the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
A process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases is disclosed wherein a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; or the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
Description
Outokumpu Oy, Ou-tokumpu A method for the recovery of valuable metals from finely-divided pyrite The present invention relates to a process for the recovery of valuable metals from impure, finely ground pyrite, either directly or by smelting the pyrite by means of, for example, non-oxidizing gases at an elevated temperature in order to produce iron matte with a valuable-metal content ~nd dust-bearing gases. This invention relates in particular to a multiple-stage process for the recovery of valuable metals from the roasting residue obtained from the oxidizing roasting of pyrite or from the iron matte, dusts and gases obtained from the smelting of pyrite.
From Finnish Patent 32 465 there is known the smelting of finely ground pyrite in a flash smelting furnace at an elevated temperature by means of non-oxidizing gases. The retention time of the suspension in the reaction shaft is approx. 1-2 s, during which time the solid materials heat up and melt in accordance with known chemical reactions. Aftex leaving the reaction shaft the gas contains sulfur dioxide in excess, and therefore the gas has to be rèduced in the rising shaft .. . ... . . ... . .. ..
of the flash smelting furnace in order to recover the elemental sulfur, as described in, for example, Finnish Patents 44 797, 45 948 and 45 037. The reaction products obtained are iron ma-tte, iron slag and gas, from which the sulfur is recovered, and also dust.
The iron matte contains var:ious compounds of iron, copper, zinc, cadmium, lead, cobalt, nickel, gold and silver. The dusts separated from the gases, for their part, contain various compounds of lead, zinc, arsenic, cadmium, copper, cobalt, nickel, mercury and selenium.
The pyrite used as raw material contains varying amounts of different valuable metals, and in a typical case the valuable-metal contents can be, for example:
Cu 1-1.5 % Cd 0.001-0.1 %
Zn ~-4 Ag 0.001-0.1 Pb 1-2 Au 0.0001-0~01 As 0.5-2 Hg 0.001-0.1 Co 0.001-0.1 Se 0.001-0.1 Ni 0.001-0.1 In a flash smelting furnace, the distribution of these valuable metals is as follows:
Zn 30-40 % in matte, the remainder in dusts Pb 5-15 - " -Cd 30-40 _ ~ _ Cu 85-95 - " -Co 85-95 - ' Ni 85-95 - " -Au and Ag approx. 100 % in the matte Hg and Se approx. 95 % in the gases As approx. 95 % in the gases, the remainder in the dusts After leaving the e]ectric filter, the gases are directed to the sulfur condensation towers, in which the elemental sulfur is recovered by known methods.
s Thereafter, the valuable metals present in the matte and the dust can be recovered by various methods. Thus, there are numerous previously known methods for the recovery of various valuable metals, and the problem has been how to select those very methods which are best suited for use in conjunction with each other so as to obtain as a final result a maximally economical and disturbance-free total process for the recovery of valuable metals from impure pyrite.
The object of the present invention is thus to produce a total process, better than previous ones, for the recovery of valuable metals from finely ground pyrite.
Thus, in one aspect the present invention provides a process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps: a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leac~ed and the obtained solution is separated in order to recover the valuable metals from thc solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases; c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; and d) the sulfur is condensed from the gases.
In another aspect, the invention provides a process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated tempera-ture in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps: a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production; b) the dusts which contain valuable metals are separated from the gases; c) the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
The invention is described below in greater detail with reference to the accompanying drawings, in which Figures 1-3 depict flow diagrams of three different alternative total processes, Figure 4 depicts in greater detail the recovery of sulfur from the gases of a sulf.ur smelting furnace, and Figure 5 depicts one alternative method .for the sulfating roasting of the roasting residue.
In the alternative process according to Figure 1, the sulide-oxide matte from the flash smelting furnace is granulated during stage 1 and ground during stage 2~ Part of the matte is dead roasted during stage 3, the temperature being 950-1050C. The following reactions occur during this stage:
1) FeS t 1.75 2 ->0 5 Fe20 ~ S0 - 3a -
From Finnish Patent 32 465 there is known the smelting of finely ground pyrite in a flash smelting furnace at an elevated temperature by means of non-oxidizing gases. The retention time of the suspension in the reaction shaft is approx. 1-2 s, during which time the solid materials heat up and melt in accordance with known chemical reactions. Aftex leaving the reaction shaft the gas contains sulfur dioxide in excess, and therefore the gas has to be rèduced in the rising shaft .. . ... . . ... . .. ..
of the flash smelting furnace in order to recover the elemental sulfur, as described in, for example, Finnish Patents 44 797, 45 948 and 45 037. The reaction products obtained are iron ma-tte, iron slag and gas, from which the sulfur is recovered, and also dust.
The iron matte contains var:ious compounds of iron, copper, zinc, cadmium, lead, cobalt, nickel, gold and silver. The dusts separated from the gases, for their part, contain various compounds of lead, zinc, arsenic, cadmium, copper, cobalt, nickel, mercury and selenium.
The pyrite used as raw material contains varying amounts of different valuable metals, and in a typical case the valuable-metal contents can be, for example:
Cu 1-1.5 % Cd 0.001-0.1 %
Zn ~-4 Ag 0.001-0.1 Pb 1-2 Au 0.0001-0~01 As 0.5-2 Hg 0.001-0.1 Co 0.001-0.1 Se 0.001-0.1 Ni 0.001-0.1 In a flash smelting furnace, the distribution of these valuable metals is as follows:
Zn 30-40 % in matte, the remainder in dusts Pb 5-15 - " -Cd 30-40 _ ~ _ Cu 85-95 - " -Co 85-95 - ' Ni 85-95 - " -Au and Ag approx. 100 % in the matte Hg and Se approx. 95 % in the gases As approx. 95 % in the gases, the remainder in the dusts After leaving the e]ectric filter, the gases are directed to the sulfur condensation towers, in which the elemental sulfur is recovered by known methods.
s Thereafter, the valuable metals present in the matte and the dust can be recovered by various methods. Thus, there are numerous previously known methods for the recovery of various valuable metals, and the problem has been how to select those very methods which are best suited for use in conjunction with each other so as to obtain as a final result a maximally economical and disturbance-free total process for the recovery of valuable metals from impure pyrite.
The object of the present invention is thus to produce a total process, better than previous ones, for the recovery of valuable metals from finely ground pyrite.
Thus, in one aspect the present invention provides a process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps: a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leac~ed and the obtained solution is separated in order to recover the valuable metals from thc solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases; c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; and d) the sulfur is condensed from the gases.
In another aspect, the invention provides a process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated tempera-ture in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps: a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production; b) the dusts which contain valuable metals are separated from the gases; c) the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
The invention is described below in greater detail with reference to the accompanying drawings, in which Figures 1-3 depict flow diagrams of three different alternative total processes, Figure 4 depicts in greater detail the recovery of sulfur from the gases of a sulf.ur smelting furnace, and Figure 5 depicts one alternative method .for the sulfating roasting of the roasting residue.
In the alternative process according to Figure 1, the sulide-oxide matte from the flash smelting furnace is granulated during stage 1 and ground during stage 2~ Part of the matte is dead roasted during stage 3, the temperature being 950-1050C. The following reactions occur during this stage:
1) FeS t 1.75 2 ->0 5 Fe20 ~ S0 - 3a -
2) Fe304 ~ 0-25 o2~ 1.5 Pe2 3
3) FeO + 0.25 O2~ 0.5 Fe203
4) PbS t~ 1.5 2~ PbO ~ S02 3b -
5) ZnS -~ 1.5 2 + Fe2O3 ~ ZnO ~e2O3 + SO2
6) 2 2 2 3 > Cu2O Fe23 + S2 The sulfides of Co, Ni and Cd also oxidize, forming respective oxides or ferrites. The calcine of this stage is directed to the subsequent stage 4, i.e. the sulfating roasting (FI Patent 41 874), to which the final matte is also introduced. The sulfating temperature is 65~-750 C. The iron compounds are oxidized mostly to hematite and partly to magnetite and water-soluble sulfates. The compounds of Cu, Zn and Pb are sulfated to sulfates, the copper and zinc sulfates being water soluble, and to some extent also to zinc and copper ferrites, which are not soluble in water.
Co, Ni and-Cd are also sulfated to their respective sulfates.
The calcine is leached during the subsequent stage 5 in the iron residue washing water, which is obtained from stage 6, whereby the water-soluble sulfates dissolve. The iron residue Fe2O3 is directed to the chloride leaching cycle. The acid sulfate solution is directed to the recovery of metals.
The dusts emerging from the flash smelting furnace are directed from the electric filters to the sulfating roasting 12, in which the temperature is 620 -700 C. The necessary Na is added in the form of, for example, Na2SO4. During this stage, most of the iron sulfide oxidizes to hematite and to some extent also to magnetite and water-soluble sulfates. The sulfides of Cu and Zn are in part sulfated to sulfates and ir part oxidized to water insoluble ferrites. Pb, Co, Ni and Cd are sulfated, lead sulfate being water insoluble. The calcine is leached during the subsequent stage 13 in the washing water of the precipitate obtained from the iron removal stage 17, the washing water coming from stage 18, and thereby the sulfates of the non-ferrous metals present in the dusts are caused to dissolve. The precipitate is directed to the strong acid leach stage (SAL) 1~, and to this stage the necessary amourlt of return acid is - introduced from the zinc electrolysis of stage 22 in order to dis~olve the ferrites of the dust calcine. The temperature of the stage is 90-100 C, and its H2SO4 concentration is 30-150 g/l. The leach residue is washed during s-tage 15 and directed to the chloride leach cycle. The acid sulfate solution is directed to the metal recovery.
The sulfate solution obtained from the matte leaching stage 5 is directed to the Cu cementation stage 7, in which the copper is cemented by means of iron and thereby cement copper is recovered. The sulfate solution is directed to the iron removal stage 8, during which that part of the electrolysis return acid which is not used for the leach during stage 14 is neutralized by means of, for example, limestone, and the iron in a ferrous form is removed by oxidizing it by means of air, the temperature being 85-95 C, and by precipitating it in the form of goethite by using, for example, lime as a neutralizing agent. The produced gypsum and iron precipi-tate is washed during the sub-sequent stage 9 and removed, the washing water of stage 9 is directed to stage 8 or 10, and the sulfate solution is directed to the metal hydroxide precipitation stage 10, during which Zn, Co, Ni and Cd are precipitated in the form of hydroxides by means of, for example, lime. The precipitate is used as a neutralizing agent in the iron removal stage 17 of the dust cycle. Part of the metal-free sodium sulfate solution is used for the precipitation of lead carbonate, when the lead is recovered in the form of lead carbonate, otherwise it is vaporized in a form suitable to be directed to the grinding of matte during stage 2, through which the sodium sulfate necessary for the sulfating roasting of matte is also ob-tained.
The dust of the sulfate matte is directed from the calcine leaching stage 13 to the copper cementing stage 16, during !
which the precipitation takes place by means of, Eor example, iron. The solution is directed to the iron removal stage 17, to which the solution of the SAL stage is introduced, and par of the washing water of the precipitate washing stage 15 is neutralized by means of, for example, lime and the metal hydroxide precipitate from stage 10. The iron and the gypsum residue are washed with water during stage 18, and the washin~
water from this stage is used in the water leach stage 13. The Zn sulfate solution is directed to the subsequent solution purification stage 19, during which Cu, Co and Ni are cemented by known methods ~FI Patent 52 595 and Canadian Patent No. 1,113,253) by means of zinc dust, using arsenic trioxide as the reagent, and cadmium is cemented by means of Zn dust during stage 20 by methods known per se (FI Patent 50 715).
The solution is concentrated by evaporation during stage 21. The concentrated sulfate solution is directed to the zinc electrolysis 22, which operates by the electrowinning method. The zinc is recovered rom stage 22, and the return acid is used for -the purposes described above.
The iron residue from stage 6 is directed to the chloride leaching stage 23. The iron residue is leached in a sodium-calcium chloride solution, whereby the water-insoluble lead sulfate is brought to a water-soluble form.
Chlorine gas is added for the oxidation of the noble metals, and lime is used as a neutralizer and for the removal of sulfates. In order to prevent the concentration of certain impurities, e.g. iron, in the solution cycle, part of the solution is removed, and in order to maintain the Na level, NaCl is added.
The obtained Fe203 is washed during stage 24 and fed directly to the iron production. The chloride solution from stage 19, containing Pb, Ag and Au, is directed to the second leach 15, during which the iron residue of the strong acid leach of the sulfated dusts is leached. In this stage also, lime is used for sulfate removal and pH control. Air blasting is used for removing the excess chlorine gas. If iron has dissolved during the previous leach 23, it is precipitated together with the arsenic of stage 14, carried by moisture. The arsenic-bearing iron residue of the stage is taken to the disposal area after the washing stage 26. The chloride solution, which contains some copper and zinc in addition to the lead, silver and gold, is combined with the washing water from stage 26 and directed to stage 27, during which the noble metals s and the copper are separated from the solution by, for example, cementation by means of zinc dust. The lead and ~inc are precipitated during the subsequent stage - 6a -28 by means of, for example, hydrogen sulfide. The precipitate is washed during stage 29, and the washing water is combined with the chloride solution and directed to the evaporation stage 33, from which the solution, from which the excess NaCl has been removed,is recycled to stage 23.
The sulfides of lead and zinc from stage 29 are dissolved in water by oxidation with air iIl an autoclave during stage 31.
The lead sulfate precipitate is washed during stage 32, and the washing water is directed to the autoclave. The zinc sulfate solution is directed to the metal hydroxide precipitation stage 10. The lead sulfate is converted to lead carbonate during stage 33 by leaching the residue in a sodium sulfa-te solution (from stage 8) by means of sodium carbonate. The lead carbonate precipitate is washed during stage 34 and the washing water is recycled to the previous stage. The solution from stage 23 contains sodium sulfate, and it is used, alternatively with -the NaS04 solution of stage 10, for the grinding of matte.
The roasting cycle for the dusts can be combined with the matte roastin~ cycle, when the arsenic has first been removed and thus cannot pollute the iron ore. The double roasting replaces the strong acid leach of the dusts, and increases the yield of sulfate.
In the alternative method illustrated in Figure 2, the leaching cycles for the roasting residues of matte and dust have been combined, whereby the strong acid leach of the dust line is eliminated, and the arsenic is removed from the dusts during a separate sta~e, ~and tshe calcine leach residue from the stage 9 is directed to the sulfating roasting ~ of matte.
In this manner, only one iron residue is obtained, and this residue is directed to the chloride leaching cycle. The sulfate solutions obtained from the leaching of the products of roasting can now also be -treated in one cycle, in which case the solution is directed via the copper cementation stage 10 to the iron precipitation stage 11, to which is fed, as a neutralizing agent in addition to the lime addition, the precipitate of the zinc hydroxide precipitation stage 18 subsequent to the neutralization 16 of the return acid obtained from the zinc electrolysis. Th~
sulfate solution now passes via the Cu, Co, Ni and Cd removal stages 13 and 14 to the evaporation s-tage 15 and further to t:he zinc electrolysis. Since there is now only one Fe203 residue, the separate chloride leaching stage for the dusts is eliminated and the chloride solution cycle differs from -the process alternative of Figure 1, subsequent to the washing stage for the precipltate obtained from the lead su]fide and zinc sulfide precipitation stage,in that the sulfates are removed from th~
solution during stage 25 by means of a lime addition, and the obtained residual gypsun precipitate is washed and the solutions are combined, and the excess NaCl is removed from the solution prior to the evaporation stage. The lead carbonate is recovered as in the alternative illustrated in Figure 1.
If it is desired, in accordance with Figure 3, to combine the leaching cycles of the residues from the roasting of matte and dusts and not -to carry out a separate arsenic removal from the dusts, the roasting residue from the sulfating roasting of the dusts is leached during the strong acid leach (SAL) 9 in order to dissolve the ferrites. However, in this case the electrolysis return acid is not fed to the SAL as in the process alternative of Figure 1, but a pure H2SO4 solution is used in order to maintain a suitable sodium concentration in the sulfate solution, and thus the sodium cannot interfere with the Zn electrolysis. In other respects the sulfate solution is pw:ified as in the process of Figure 1, i.e. the solutions from the leaching stages 5 and 8 for the matte and the dusts are combined and the sulfate solution passes via the copper removal stage 11, the Fe removal stage 12, the Cu, Co, Ni cementation stage 14, and the cadium removal stage 15 to evaporation and further to ~he Zn electrolysis 17, the return acid of which is neutralized during a separate stage 18 by means of, for example, limestone, and during the subsequent stage 19 the Zn(OH)2, which is used as a neutralizing agent during the iron removal stage 12,is precipitated. The Na2SO4 solution is evaporated and used, alternatively with the Na2SO4 solution obtained from the lead carbonate precipitation stage, for the grinding of matte.
The chloride leaching cycle is the same as in the alternative of Figure 1, i.e. the Fe residue obtained from the dust cycle is directed to the waste disposal area owing to the arsenic present in it.
Figure 4 illustrates in greater detail how dusts are separated from the dusty gases from the flash smelting furnace and how the sul~ur is removed from the gases as elemental sulfur.
Figure 5 shows in greater detail one alternative method for the sulfating roasting of the roasting residues, using pyrite for fulfilling the heat requirement and for creating a suitable gas atmosphere in the roasting rurnace.
Roasting residue is produced in the sulfuric acid industry, which roasts pyrite in order to obtain sulfur dioxide, the raw material of sulfuxic acid production. By means of the present invention, such roasting residues and impure pyrite can be exploited effectively.
The invention is described below with the aid of an example.
Example The basis for the calculations is roasting and leaching stages 7000 h/a and solution purification and metal recovery stages 7700 h/a. Pyrite concentrate was fed into a flash smelting furnace at a rate of 82 t/h. Oil was added at 3.1 t/h, and carbon was fed into the gas reduction zone at ll.S t/h.
Matte was obtained from the furnace at 55 t/h and fly dust at 12 t/h. Elemental sulfur was obtained from the smelting gases at 14 t/h.
The analysis of the pyrite was as follows:
S 47 %
Fe 42 %
Cu 1~15 %
Zn 3.2 Pb 1.3 As 0.5 Co 0.02 Ni 0.002 Cd 0~006 Ag 0.004 Au 0.0008 Hg 0.01 Se 0.002 The matte was granulated in water and ground so that 80 %
was ~ 74 ~m and 100 % ~ 290 ~m. The amount of sodium in the grinding was 0.94 t/h. To the oxidizing roasting matte was fed at 34.5 t/h, and to the sulfating roasting matte was fed at 20.5 t/h and dead-roasted product at 33.4 t/h. The calcine from the sulfating roasting of matte, the analysis of calcine being as follows:
S 2.63 %
Fe 60.6 %
Cu 1.69 %
Zn 1.66 %
Pb 0.28 %
Na 0O54 %
was leached ùsing washing water at 35.1 m3/h, and residue was obtained at 49.9 t/ho The sulfate solution was directed to the cementation of copper, to ~hich scrap iron was added at 1 t/h. This stage yielded 90-percent cement copper at 0.91 t/h. The sulfate solution at 23.9 m3/h was fed to the iron removal stage, to which limestone was added at 1.82 t/h and lime at 2.22 t/h. The amount of return acid was 9.9 m3/h, water arrived from the S~L stage at 6.0 m3/h and washing water at 19.8 m3/h~
The yield of goethite-gypsum precipitate was 9.9 t/h, and the sulfate solution, 53.0 m3/h, was directed to the metal hydroxide precipitation stage. Calclum at 1.89 t/h was added -to this stage, and solution arrived -, at 4.4 m3/h from the lead sulfide-zinc sulfide leaching stage. The amount of hydroxide precipitate was 6.1 t/h and the amount of Na2SO4 solution was 53.3 m /h.
.
The dusts from the flash smelting furnace, 12 t/h, were fed to the sulfati.ng roasting, to which the Na2SO4 addition was 0.2 t/h. The calcine, 12 01 t/h, which contained S 8.9 %
Fe21.9 %
Cu0.61 %
Zn:'.5.8 %
P~8.4 %
Na0.53 %
was leached in the water arriving from the iron precipitate washing stage at 15.3 m3/h, and residue was obtained at 7.5 t/h, and it was directed to the strong acid leach, to which return acid arrived at 14.8 m3/h, and the amoun-t of water arriving at the residue washing stage was 13.3 m3/h. The yield of residue was 6.9 t/l~, and it was directed to the chloride leaching cycle.
The sulfate solution at 12.8 m3/h was directed to copper cementation, to which scrap iron was fed at 0.076 t/h. 90-percent cement copper was obtained from this stage at 0.052 t/h~ and sulfate so].ution was fed to the iron removal stage at 11.6 m /h, and to this stage metal hydroxide precipitate was fed from the leaching cycle for the roasted matte at 6.1 t/h, the lime addition was 0.08 t/h, and solution arrived from the strong acid leach at 14.8 m3/h, as well as washing water fr~,m the strong acid leach at 6 m3/h and iron residue washing.water at 13.9 m3/h. The yield of goethite-gypsum precipitate was 4.9 t/h, and the sulfate solution at 33.3 m3/h was directed to the Cu-Co-Ni cementation s-tage, to which zinc dust was fed at 78 kg/h, As2O3 at 2 kg/h, and the yield of Cu-Co-Ni precipitate was 0.~7 t/h. Solution was fed to the cadmium removal stage at 3~.3 m3/h and zinc dust at 3 k~/h. rrhe yield of cadmium cementate was 3 k~/h. Sulfate solution entered the evaporation stage at 33.3 m3/h, and the output was 24.5 m3/h and was fed to the zinc electrolysis, from which zinc cathodes were obtained at a rate of 2.5 t/h and return acid at 24.5 m3/h.
To the first stage of the chloride leaching cycle, the leaching of roasted matte, roasted matté leach residue entered at 45.4 t/h, chloride solution at 80 m3/h, washing water at 20 m3/h, C12 at 0.29 t/h, lime at 0.46 t/h, and NaCi at 0.47 t/h. Fe203, a suitable raw material for iron production,was obtained at 45.5 t/h, and it contained:
Fe 6/.5 Zn 0.12 %
Cu o.og Pb ~.04 ~
S ~.2 %
and chloride solution at 100 m3/h was directed to the leaching stage for dust calcine. Residue from the roasting of dust was obtained at 6.3 t/h, and washing water at 14.8 m3/h and lirne at 0.41 t/h. Reiectable iron residue was obtained at 5.6 t/h and chloride solution at 11.5 m3/h. ~ext, Cu, Ag and Au were cemented out from the chloride solution. Zinc dust was used at 0~04 t/h~
an~ silver-gold-copper cementate was obtained at 0.01~ t/ho Chloride solution entered the lead and zinc precipitation stage at 117.2 rn /h and hydrogen sulfide was added at 0.31 t/h, the yield of ~h'-, ZnS precipitate was ln 3 t/h and the arnount of washing w~ter was 2.9 m3/h. The amount of outlet solution was 4.5 m3/h, and chloride solution passed to the evaporat:ion stage at 112.7 m3/h, and its output was ~ m3/h. The PbSr ZnS
precipitate was leached in an autoclave, to which w~shing water was fed at 4s5 m3/h. L~ach residue was obtained at 1.6 t/h, and rhe zinc sul~ate solution, 4.4 m3/h, was directed to the metal hydroxide precipitation stage of the matte cycle. The leach residue ~-as (~irected to the lead carbonate precipitation stage, t:o which re~idual sodium s~llfate solution from the rnatte cycle arrived at 5 m3/h, and the Na2C03 addition was 0.6 t/h and the amount of washing water 4~5 m3/h. The sodium sulfate solution, 9 m3/h, was direc-ted -to the grinding of matte, and lead carbonate was obtained at 1.4 t/h, which was calcinated to lead oxide, the amount of which was 1.16 t/h, and the lead oxide was reduced to metallic lead, the yield being 1.0 t/h.
Co, Ni and-Cd are also sulfated to their respective sulfates.
The calcine is leached during the subsequent stage 5 in the iron residue washing water, which is obtained from stage 6, whereby the water-soluble sulfates dissolve. The iron residue Fe2O3 is directed to the chloride leaching cycle. The acid sulfate solution is directed to the recovery of metals.
The dusts emerging from the flash smelting furnace are directed from the electric filters to the sulfating roasting 12, in which the temperature is 620 -700 C. The necessary Na is added in the form of, for example, Na2SO4. During this stage, most of the iron sulfide oxidizes to hematite and to some extent also to magnetite and water-soluble sulfates. The sulfides of Cu and Zn are in part sulfated to sulfates and ir part oxidized to water insoluble ferrites. Pb, Co, Ni and Cd are sulfated, lead sulfate being water insoluble. The calcine is leached during the subsequent stage 13 in the washing water of the precipitate obtained from the iron removal stage 17, the washing water coming from stage 18, and thereby the sulfates of the non-ferrous metals present in the dusts are caused to dissolve. The precipitate is directed to the strong acid leach stage (SAL) 1~, and to this stage the necessary amourlt of return acid is - introduced from the zinc electrolysis of stage 22 in order to dis~olve the ferrites of the dust calcine. The temperature of the stage is 90-100 C, and its H2SO4 concentration is 30-150 g/l. The leach residue is washed during s-tage 15 and directed to the chloride leach cycle. The acid sulfate solution is directed to the metal recovery.
The sulfate solution obtained from the matte leaching stage 5 is directed to the Cu cementation stage 7, in which the copper is cemented by means of iron and thereby cement copper is recovered. The sulfate solution is directed to the iron removal stage 8, during which that part of the electrolysis return acid which is not used for the leach during stage 14 is neutralized by means of, for example, limestone, and the iron in a ferrous form is removed by oxidizing it by means of air, the temperature being 85-95 C, and by precipitating it in the form of goethite by using, for example, lime as a neutralizing agent. The produced gypsum and iron precipi-tate is washed during the sub-sequent stage 9 and removed, the washing water of stage 9 is directed to stage 8 or 10, and the sulfate solution is directed to the metal hydroxide precipitation stage 10, during which Zn, Co, Ni and Cd are precipitated in the form of hydroxides by means of, for example, lime. The precipitate is used as a neutralizing agent in the iron removal stage 17 of the dust cycle. Part of the metal-free sodium sulfate solution is used for the precipitation of lead carbonate, when the lead is recovered in the form of lead carbonate, otherwise it is vaporized in a form suitable to be directed to the grinding of matte during stage 2, through which the sodium sulfate necessary for the sulfating roasting of matte is also ob-tained.
The dust of the sulfate matte is directed from the calcine leaching stage 13 to the copper cementing stage 16, during !
which the precipitation takes place by means of, Eor example, iron. The solution is directed to the iron removal stage 17, to which the solution of the SAL stage is introduced, and par of the washing water of the precipitate washing stage 15 is neutralized by means of, for example, lime and the metal hydroxide precipitate from stage 10. The iron and the gypsum residue are washed with water during stage 18, and the washin~
water from this stage is used in the water leach stage 13. The Zn sulfate solution is directed to the subsequent solution purification stage 19, during which Cu, Co and Ni are cemented by known methods ~FI Patent 52 595 and Canadian Patent No. 1,113,253) by means of zinc dust, using arsenic trioxide as the reagent, and cadmium is cemented by means of Zn dust during stage 20 by methods known per se (FI Patent 50 715).
The solution is concentrated by evaporation during stage 21. The concentrated sulfate solution is directed to the zinc electrolysis 22, which operates by the electrowinning method. The zinc is recovered rom stage 22, and the return acid is used for -the purposes described above.
The iron residue from stage 6 is directed to the chloride leaching stage 23. The iron residue is leached in a sodium-calcium chloride solution, whereby the water-insoluble lead sulfate is brought to a water-soluble form.
Chlorine gas is added for the oxidation of the noble metals, and lime is used as a neutralizer and for the removal of sulfates. In order to prevent the concentration of certain impurities, e.g. iron, in the solution cycle, part of the solution is removed, and in order to maintain the Na level, NaCl is added.
The obtained Fe203 is washed during stage 24 and fed directly to the iron production. The chloride solution from stage 19, containing Pb, Ag and Au, is directed to the second leach 15, during which the iron residue of the strong acid leach of the sulfated dusts is leached. In this stage also, lime is used for sulfate removal and pH control. Air blasting is used for removing the excess chlorine gas. If iron has dissolved during the previous leach 23, it is precipitated together with the arsenic of stage 14, carried by moisture. The arsenic-bearing iron residue of the stage is taken to the disposal area after the washing stage 26. The chloride solution, which contains some copper and zinc in addition to the lead, silver and gold, is combined with the washing water from stage 26 and directed to stage 27, during which the noble metals s and the copper are separated from the solution by, for example, cementation by means of zinc dust. The lead and ~inc are precipitated during the subsequent stage - 6a -28 by means of, for example, hydrogen sulfide. The precipitate is washed during stage 29, and the washing water is combined with the chloride solution and directed to the evaporation stage 33, from which the solution, from which the excess NaCl has been removed,is recycled to stage 23.
The sulfides of lead and zinc from stage 29 are dissolved in water by oxidation with air iIl an autoclave during stage 31.
The lead sulfate precipitate is washed during stage 32, and the washing water is directed to the autoclave. The zinc sulfate solution is directed to the metal hydroxide precipitation stage 10. The lead sulfate is converted to lead carbonate during stage 33 by leaching the residue in a sodium sulfa-te solution (from stage 8) by means of sodium carbonate. The lead carbonate precipitate is washed during stage 34 and the washing water is recycled to the previous stage. The solution from stage 23 contains sodium sulfate, and it is used, alternatively with -the NaS04 solution of stage 10, for the grinding of matte.
The roasting cycle for the dusts can be combined with the matte roastin~ cycle, when the arsenic has first been removed and thus cannot pollute the iron ore. The double roasting replaces the strong acid leach of the dusts, and increases the yield of sulfate.
In the alternative method illustrated in Figure 2, the leaching cycles for the roasting residues of matte and dust have been combined, whereby the strong acid leach of the dust line is eliminated, and the arsenic is removed from the dusts during a separate sta~e, ~and tshe calcine leach residue from the stage 9 is directed to the sulfating roasting ~ of matte.
In this manner, only one iron residue is obtained, and this residue is directed to the chloride leaching cycle. The sulfate solutions obtained from the leaching of the products of roasting can now also be -treated in one cycle, in which case the solution is directed via the copper cementation stage 10 to the iron precipitation stage 11, to which is fed, as a neutralizing agent in addition to the lime addition, the precipitate of the zinc hydroxide precipitation stage 18 subsequent to the neutralization 16 of the return acid obtained from the zinc electrolysis. Th~
sulfate solution now passes via the Cu, Co, Ni and Cd removal stages 13 and 14 to the evaporation s-tage 15 and further to t:he zinc electrolysis. Since there is now only one Fe203 residue, the separate chloride leaching stage for the dusts is eliminated and the chloride solution cycle differs from -the process alternative of Figure 1, subsequent to the washing stage for the precipltate obtained from the lead su]fide and zinc sulfide precipitation stage,in that the sulfates are removed from th~
solution during stage 25 by means of a lime addition, and the obtained residual gypsun precipitate is washed and the solutions are combined, and the excess NaCl is removed from the solution prior to the evaporation stage. The lead carbonate is recovered as in the alternative illustrated in Figure 1.
If it is desired, in accordance with Figure 3, to combine the leaching cycles of the residues from the roasting of matte and dusts and not -to carry out a separate arsenic removal from the dusts, the roasting residue from the sulfating roasting of the dusts is leached during the strong acid leach (SAL) 9 in order to dissolve the ferrites. However, in this case the electrolysis return acid is not fed to the SAL as in the process alternative of Figure 1, but a pure H2SO4 solution is used in order to maintain a suitable sodium concentration in the sulfate solution, and thus the sodium cannot interfere with the Zn electrolysis. In other respects the sulfate solution is pw:ified as in the process of Figure 1, i.e. the solutions from the leaching stages 5 and 8 for the matte and the dusts are combined and the sulfate solution passes via the copper removal stage 11, the Fe removal stage 12, the Cu, Co, Ni cementation stage 14, and the cadium removal stage 15 to evaporation and further to ~he Zn electrolysis 17, the return acid of which is neutralized during a separate stage 18 by means of, for example, limestone, and during the subsequent stage 19 the Zn(OH)2, which is used as a neutralizing agent during the iron removal stage 12,is precipitated. The Na2SO4 solution is evaporated and used, alternatively with the Na2SO4 solution obtained from the lead carbonate precipitation stage, for the grinding of matte.
The chloride leaching cycle is the same as in the alternative of Figure 1, i.e. the Fe residue obtained from the dust cycle is directed to the waste disposal area owing to the arsenic present in it.
Figure 4 illustrates in greater detail how dusts are separated from the dusty gases from the flash smelting furnace and how the sul~ur is removed from the gases as elemental sulfur.
Figure 5 shows in greater detail one alternative method for the sulfating roasting of the roasting residues, using pyrite for fulfilling the heat requirement and for creating a suitable gas atmosphere in the roasting rurnace.
Roasting residue is produced in the sulfuric acid industry, which roasts pyrite in order to obtain sulfur dioxide, the raw material of sulfuxic acid production. By means of the present invention, such roasting residues and impure pyrite can be exploited effectively.
The invention is described below with the aid of an example.
Example The basis for the calculations is roasting and leaching stages 7000 h/a and solution purification and metal recovery stages 7700 h/a. Pyrite concentrate was fed into a flash smelting furnace at a rate of 82 t/h. Oil was added at 3.1 t/h, and carbon was fed into the gas reduction zone at ll.S t/h.
Matte was obtained from the furnace at 55 t/h and fly dust at 12 t/h. Elemental sulfur was obtained from the smelting gases at 14 t/h.
The analysis of the pyrite was as follows:
S 47 %
Fe 42 %
Cu 1~15 %
Zn 3.2 Pb 1.3 As 0.5 Co 0.02 Ni 0.002 Cd 0~006 Ag 0.004 Au 0.0008 Hg 0.01 Se 0.002 The matte was granulated in water and ground so that 80 %
was ~ 74 ~m and 100 % ~ 290 ~m. The amount of sodium in the grinding was 0.94 t/h. To the oxidizing roasting matte was fed at 34.5 t/h, and to the sulfating roasting matte was fed at 20.5 t/h and dead-roasted product at 33.4 t/h. The calcine from the sulfating roasting of matte, the analysis of calcine being as follows:
S 2.63 %
Fe 60.6 %
Cu 1.69 %
Zn 1.66 %
Pb 0.28 %
Na 0O54 %
was leached ùsing washing water at 35.1 m3/h, and residue was obtained at 49.9 t/ho The sulfate solution was directed to the cementation of copper, to ~hich scrap iron was added at 1 t/h. This stage yielded 90-percent cement copper at 0.91 t/h. The sulfate solution at 23.9 m3/h was fed to the iron removal stage, to which limestone was added at 1.82 t/h and lime at 2.22 t/h. The amount of return acid was 9.9 m3/h, water arrived from the S~L stage at 6.0 m3/h and washing water at 19.8 m3/h~
The yield of goethite-gypsum precipitate was 9.9 t/h, and the sulfate solution, 53.0 m3/h, was directed to the metal hydroxide precipitation stage. Calclum at 1.89 t/h was added -to this stage, and solution arrived -, at 4.4 m3/h from the lead sulfide-zinc sulfide leaching stage. The amount of hydroxide precipitate was 6.1 t/h and the amount of Na2SO4 solution was 53.3 m /h.
.
The dusts from the flash smelting furnace, 12 t/h, were fed to the sulfati.ng roasting, to which the Na2SO4 addition was 0.2 t/h. The calcine, 12 01 t/h, which contained S 8.9 %
Fe21.9 %
Cu0.61 %
Zn:'.5.8 %
P~8.4 %
Na0.53 %
was leached in the water arriving from the iron precipitate washing stage at 15.3 m3/h, and residue was obtained at 7.5 t/h, and it was directed to the strong acid leach, to which return acid arrived at 14.8 m3/h, and the amoun-t of water arriving at the residue washing stage was 13.3 m3/h. The yield of residue was 6.9 t/l~, and it was directed to the chloride leaching cycle.
The sulfate solution at 12.8 m3/h was directed to copper cementation, to which scrap iron was fed at 0.076 t/h. 90-percent cement copper was obtained from this stage at 0.052 t/h~ and sulfate so].ution was fed to the iron removal stage at 11.6 m /h, and to this stage metal hydroxide precipitate was fed from the leaching cycle for the roasted matte at 6.1 t/h, the lime addition was 0.08 t/h, and solution arrived from the strong acid leach at 14.8 m3/h, as well as washing water fr~,m the strong acid leach at 6 m3/h and iron residue washing.water at 13.9 m3/h. The yield of goethite-gypsum precipitate was 4.9 t/h, and the sulfate solution at 33.3 m3/h was directed to the Cu-Co-Ni cementation s-tage, to which zinc dust was fed at 78 kg/h, As2O3 at 2 kg/h, and the yield of Cu-Co-Ni precipitate was 0.~7 t/h. Solution was fed to the cadmium removal stage at 3~.3 m3/h and zinc dust at 3 k~/h. rrhe yield of cadmium cementate was 3 k~/h. Sulfate solution entered the evaporation stage at 33.3 m3/h, and the output was 24.5 m3/h and was fed to the zinc electrolysis, from which zinc cathodes were obtained at a rate of 2.5 t/h and return acid at 24.5 m3/h.
To the first stage of the chloride leaching cycle, the leaching of roasted matte, roasted matté leach residue entered at 45.4 t/h, chloride solution at 80 m3/h, washing water at 20 m3/h, C12 at 0.29 t/h, lime at 0.46 t/h, and NaCi at 0.47 t/h. Fe203, a suitable raw material for iron production,was obtained at 45.5 t/h, and it contained:
Fe 6/.5 Zn 0.12 %
Cu o.og Pb ~.04 ~
S ~.2 %
and chloride solution at 100 m3/h was directed to the leaching stage for dust calcine. Residue from the roasting of dust was obtained at 6.3 t/h, and washing water at 14.8 m3/h and lirne at 0.41 t/h. Reiectable iron residue was obtained at 5.6 t/h and chloride solution at 11.5 m3/h. ~ext, Cu, Ag and Au were cemented out from the chloride solution. Zinc dust was used at 0~04 t/h~
an~ silver-gold-copper cementate was obtained at 0.01~ t/ho Chloride solution entered the lead and zinc precipitation stage at 117.2 rn /h and hydrogen sulfide was added at 0.31 t/h, the yield of ~h'-, ZnS precipitate was ln 3 t/h and the arnount of washing w~ter was 2.9 m3/h. The amount of outlet solution was 4.5 m3/h, and chloride solution passed to the evaporat:ion stage at 112.7 m3/h, and its output was ~ m3/h. The PbSr ZnS
precipitate was leached in an autoclave, to which w~shing water was fed at 4s5 m3/h. L~ach residue was obtained at 1.6 t/h, and rhe zinc sul~ate solution, 4.4 m3/h, was directed to the metal hydroxide precipitation stage of the matte cycle. The leach residue ~-as (~irected to the lead carbonate precipitation stage, t:o which re~idual sodium s~llfate solution from the rnatte cycle arrived at 5 m3/h, and the Na2C03 addition was 0.6 t/h and the amount of washing water 4~5 m3/h. The sodium sulfate solution, 9 m3/h, was direc-ted -to the grinding of matte, and lead carbonate was obtained at 1.4 t/h, which was calcinated to lead oxide, the amount of which was 1.16 t/h, and the lead oxide was reduced to metallic lead, the yield being 1.0 t/h.
Claims (6)
1. A process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps:
a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; and d) the sulfur is condensed from the gases.
a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are fed directly to the sulfatizing roasting of step a) after any arsenic present therein has first been removed; and d) the sulfur is condensed from the gases.
2. A process according to Claim 1, in which the sulfates are leached in two stages:
a) first using water or dilute acid in order to dissolve the readily soluble sulfates and to separate them from the solid material; and b) the obtained solid is subjected to a chloride leach in order to dissolve the lead, silver and/or gold and to separate them from the insoluble iron oxide.
a) first using water or dilute acid in order to dissolve the readily soluble sulfates and to separate them from the solid material; and b) the obtained solid is subjected to a chloride leach in order to dissolve the lead, silver and/or gold and to separate them from the insoluble iron oxide.
3. A process according to Claim 2, in which zinc dust is added to the solution obtained from the chloride leach in order to cement gold and silver and to separate them from the solution, whereafter hydrogen sulfide is introduced into the solution in order to precipitate lead and zinc as sulfides, the sulfide precipitate is leached in an autoclave by air oxidation in order to dissolve the zinc, and the separated insoluble lead sulfate is converted, in a sodium sulfate solution by means of sodium carbonate, to lead carbonate, which is separated from the solution and decomposed thermally to lead oxide, which is ultimately reduced to lead.
4. A process for the recovery of metal values from impure finely-divided pyrite either directly or by first smelting the pyrite with the aid of non-oxidizing gases at an elevated temperature in order to produce an iron matte with a valuable-metal content and dusty gases, comprising in combination the following steps:
a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals form the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
a) part of the pyrite or iron matte which contains valuable metals is first subjected to an oxidizing roasting and thereafter, together with the remainder of the pyrite or iron matte, to a sulfatizing roasting in order to convert the valuable metals present in the pyrite and its roasting residue or in the iron matte to a sulfate form and the iron to an oxide form, whereafter the sulfates are leached and the obtained solution is separated in order to recover the valuable metals form the solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases;
c) the separated dusts are directed to a separate sulfatizing roasting in order to convert the valuable metals to sulfate form and the iron to oxide form, whereafter the sulfates are leached in accordance with step a) and the insoluble, arsenic-bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
5. A process according to Claim 4, in which the sulfates are leached in two stages:
a) first using water or dilute acid in order to dissolve the readily soluble sulfates and to separate them from the solid material; and b) the obtained solid is subjected to a chloride leach in order to dissolve the lead, silver and/or gold and to separate them from the insoluble iron oxide.
a) first using water or dilute acid in order to dissolve the readily soluble sulfates and to separate them from the solid material; and b) the obtained solid is subjected to a chloride leach in order to dissolve the lead, silver and/or gold and to separate them from the insoluble iron oxide.
6. A process according to Claim 4, in which zinc dust is added to the solution obtained from the chloride leach in order to cement gold and silver and to separate them from the solution, whereafter hydrogen sulfide is introduced into the solution in order to precipitate lead and zinc as sulfides, the sulfide precipitate is leached in an autoclave by air oxidation in order to dissolve the zinc, and the separated insoluble lead sulfate is converted, in a sodium sulfate solution by means of sodium carbonate, to lead carbonate, which is separated from the solution and decomposed thermally to lead oxide, which is ultimately reduced to lead.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| FI800851A FI63781B (en) | 1980-03-19 | 1980-03-19 | FOERFARANDE FOER AOTERVINNING AV VAERDEMETALLER UR FINMALEN PYIT |
| FI800851 | 1980-03-19 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1160055A true CA1160055A (en) | 1984-01-10 |
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Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000373342A Expired CA1160055A (en) | 1980-03-19 | 1981-03-18 | Method for the recovery of valuable metals from finely-divided pyrite |
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| Country | Link |
|---|---|
| BR (1) | BR8101567A (en) |
| CA (1) | CA1160055A (en) |
| ES (1) | ES8202368A1 (en) |
| FI (1) | FI63781B (en) |
| FR (1) | FR2478672B1 (en) |
| GR (1) | GR71642B (en) |
| IT (1) | IT1170818B (en) |
| MX (1) | MX155472A (en) |
| PT (1) | PT72519B (en) |
Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1999047714A1 (en) * | 1998-03-13 | 1999-09-23 | Lewis-Gray, Elizabeth, Beatrice, Gail | Apparatus and methods for recovering valuable metals |
| WO2015178752A1 (en) * | 2014-05-21 | 2015-11-26 | Minera Pecobre, S.A. De C.V. | Hydrometallurgical process for the recovery of copper, lead and/or zinc |
| WO2023119017A1 (en) * | 2021-12-23 | 2023-06-29 | Lifezone Ltd | Hydrogen-based valorisation of metal-containing feed materials to extract metals |
Families Citing this family (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| IT1136589B (en) * | 1981-04-29 | 1986-09-03 | Tonolli Costruzioni Spa | PROCEDURE FOR THE TREATMENT OF LEAD SULPHATE COMPOUNDS TO RECOVER METALLIC LEAD AND A HIGH-PURITY SULPHATE, AND PLANT FOR ITS REALIZATION |
| DE3328708A1 (en) * | 1983-08-09 | 1985-02-28 | Bayer Ag, 5090 Leverkusen | METHOD FOR PRODUCING SULFUR DIOXIDE |
| ES2117028T3 (en) * | 1991-04-12 | 1998-08-01 | Metallgesellschaft Ag | PROCEDURE FOR THE TREATMENT OF MINERAL WITH VALUE OF RECOVERABLE METALS INCLUDING ARSENIC-CONTAINING COMPONENTS. |
| FI922722A (en) * | 1991-06-14 | 1993-12-12 | Riotinto Minera Sa | PROCESS FOER AOTERVINNING AV ICKE-JAERNHALTIGA METALLER FRAON PYRITSLAGG |
| FR2705102B1 (en) * | 1993-05-12 | 1995-08-11 | Rhone Poulenc Chimie | PROCESS FOR TREATING COMPOSITIONS CONTAINING PRECIOUS METALS AND OTHER VALUABLE ELEMENTS FOR THEIR RECOVERY. |
| US5364444A (en) * | 1993-07-08 | 1994-11-15 | North American Pallidium Ltd. | Metal leaching and recovery process |
| AUPQ078399A0 (en) * | 1999-06-04 | 1999-06-24 | Tox Free Systems Limited | Recovery of gold from gold sulphides |
| RU2599316C2 (en) * | 2014-09-04 | 2016-10-10 | Водопьянов Сергей Юрьевич | Process line for production of iron trichloride from sulphur wastes |
Family Cites Families (7)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US2206117A (en) * | 1938-03-03 | 1940-07-02 | Phelps Dodge Corp | Recovery of metal values from fume deposits |
| FI44797C (en) * | 1967-02-22 | 1972-01-10 | Outokumpu Oy | Methods of producing elemental sulfur from pyrite or pyrite concentrates |
| DE2353154B2 (en) * | 1973-10-19 | 1979-03-01 | Wilhelm 6800 Mannheim Westermann | Regenerative electric winding capacitor |
| SE396968B (en) * | 1975-07-01 | 1977-10-10 | Boliden Ab | PROCEDURE FOR EXTRACTING NON-IRON METALS FROM SULPHIDY MATERIALS BY ROASTING AND LACHING |
| FI58354C (en) * | 1977-11-28 | 1981-01-12 | Outokumpu Oy | FOERFARANDE FOER BEHANDLING AV RAOMATERIAL INNEHAOLLANDE ARSENIK OCH METALL |
| US4177067A (en) * | 1978-04-17 | 1979-12-04 | Texasgulf Inc. | Recovery of silver, copper and zinc from partially roasted pyrite concentrate by ferric sulphate leaching |
| US4168969A (en) * | 1978-04-17 | 1979-09-25 | Texasgulf Inc. | Recovery of silver, copper, zinc and lead from partially roasted pyrite concentrate by acid chloride leaching |
-
1980
- 1980-03-19 FI FI800851A patent/FI63781B/en not_active Application Discontinuation
-
1981
- 1981-02-17 PT PT7251981A patent/PT72519B/en unknown
- 1981-03-10 ES ES500248A patent/ES8202368A1/en not_active Expired
- 1981-03-17 BR BR8101567A patent/BR8101567A/en not_active IP Right Cessation
- 1981-03-18 IT IT4804681A patent/IT1170818B/en active
- 1981-03-18 CA CA000373342A patent/CA1160055A/en not_active Expired
- 1981-03-18 MX MX18643581A patent/MX155472A/en unknown
- 1981-03-19 FR FR8105912A patent/FR2478672B1/en not_active Expired
- 1981-11-27 GR GR63981A patent/GR71642B/el unknown
Cited By (7)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1999047714A1 (en) * | 1998-03-13 | 1999-09-23 | Lewis-Gray, Elizabeth, Beatrice, Gail | Apparatus and methods for recovering valuable metals |
| US6613271B1 (en) | 1998-03-13 | 2003-09-02 | Alexander Hamilton Lewis-Gray | Apparatus and methods for recovering valuable metals |
| AP1940A (en) * | 1998-03-13 | 2009-01-19 | Lewis Gray Elizabeth Beatrice Gail | Apparatus and methods for recovering valuable metals. |
| WO2015178752A1 (en) * | 2014-05-21 | 2015-11-26 | Minera Pecobre, S.A. De C.V. | Hydrometallurgical process for the recovery of copper, lead and/or zinc |
| US10633721B2 (en) | 2014-05-21 | 2020-04-28 | Penoles Tecnologia, S.A. De C.V. | Hydrometallurgical process for the recovery of copper, lead or zinc |
| WO2023119017A1 (en) * | 2021-12-23 | 2023-06-29 | Lifezone Ltd | Hydrogen-based valorisation of metal-containing feed materials to extract metals |
| GB2614283A (en) * | 2021-12-23 | 2023-07-05 | Lifezone Ltd | Hydrogen-based valorisation of metal-containing feed materials to extract metals |
Also Published As
| Publication number | Publication date |
|---|---|
| ES500248A0 (en) | 1982-01-16 |
| FR2478672B1 (en) | 1988-07-08 |
| MX155472A (en) | 1988-03-17 |
| FI800851A (en) | 1981-09-20 |
| GR71642B (en) | 1983-06-20 |
| PT72519A (en) | 1981-10-12 |
| PT72519B (en) | 1982-02-10 |
| FR2478672A1 (en) | 1981-09-25 |
| IT1170818B (en) | 1987-06-03 |
| ES8202368A1 (en) | 1982-01-16 |
| BR8101567A (en) | 1981-09-22 |
| FI63781B (en) | 1983-04-29 |
| IT8148046A0 (en) | 1981-03-18 |
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