FR2478672A1 - Metal recovery from pyrite, opt. after smelting - by oxidising and sulphating, roasting and leaching (BR 22.9.81) - Google Patents

Abstract

The process is used for direct treatment of finely crushed impure pyrites or for treatment of the matte and dust-laden gas from smelting the pyrites with a non-oxidising gas at elevated temp. It comprises (a) (i) oxidising roasting of part of the pyrites or the matte, (ii) sulphatising roasting of the product with the rest of the pyrites or matte to produce sulphates of the metal values and iron oxide and (iii) leaching to produce a soln. contg. the metal values and an insoluble iron oxide residue used for iron prodn.; (b) separating the metal value-contg. dust from the gas and either (i) passing the dust to step (a) (ii) followed immediately by arsenic removal or (ii) subjecting the dust to a separate sulphatising roasting and leaching operation followed by disposal of the arsenic-contg. iron oxide residue; and (c) condensing sulphur from the gas. The process recovers Fe, Cu, Zn, Pb, Co, Ni, Cd, Ag, Au and S from impure pyrites (e.g. used for sulphuric acid prodn.) and iron matte (e.g. from arc melting of pyrites).

Description

 The present invention relates to a process for the recovery of valuable metals from impure, finely divided pyrite, either directly or by melting the pyrite using, for example, non-oxidizing gases at a high temperature for the purpose to produce an iron matte with a proportion of valuable metal, and dust-bearing gases. The invention relates in particular to a multi-phase process for the recovery of valuable metals from roasting residues originating from the oxidizing roasting of pyrite or iron matte, dust and gases obtained from the pyrite fusion.

 According to Finnish Patent No. 32,465, the melting of finely ground pyrite in an arc melting furnace at elevated temperature using non-oxidizing gases is known. The fixing time of the suspension in the reaction tank is approximately 1-2 seconds, during which time the solid materials are heated and melt according to known chemical reactions. After leaving the reaction tank, the gas contains excess sulfur dioxide, and for this reason the gas must be reduced in the outlet tank of the arc melting furnace, in order to recover the elemental sulfur. , as described, for example, in Finnish patents No 44,797, No 45,948 and No 45,037. The reaction products obtained are iron matte, pig iron, gas, from which sulfur is recovered, and also dust.

 Iron matte contains different compounds of iron, copper, zinc, cadmium, lead, cobalt, nickel, gold and silver. Dust separated from gases, on the other hand, contains different compounds of lead, zinc, arsenic, cadmium, copper, cobalt, nickel, mercury and selenium.

Pyrite used as raw material contains different amounts of different valuable metals, and in a typical case the contents of valuable metals can be, for example
Cu 1-1.5 só Cd 0.001-0.1,>
Zn 2-4 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 an arc melting furnace, the distribution of these valuable metals is as follows:
Zn 30-40% in matte, the rest in dust
Pb 5-15 - I '-
Cd 30-40 - I '-
Cu 85-95 - '-
Co 85-95 - "
Ni 85-95 - I '-
Au and Ag approximately 100% in the matte
Hg and Se approximately 95% in gases
As approximately 95% in gases, the rest in dust.

 After leaving the electric filter, the gases are directed to sulfur condensation columns, in which the elemental sulfur is recovered according to known methods.

 After this, the valuable metals present in the mat and in the dust can be recovered by various processes.

Thus, there are many known methods for the recovery of different valuable metals, and the problem has been how to choose the processes that are best suited for use over others, so as to obtain a final result process for recovering valuable metals from impure pyrite, economical to the maximum and generally free of troubles.

One of the aims of the present invention is to propose a global process, more efficient than those existing, for the recovery of valuable metals from finely ground pyrite. It is characterized by the fact that
a) a part of the pyrite or of the iron mat which contains valuable metals is first subjected to an oxidizing roasting and then, with the rest of the pyrite or of the iron mat, to a sulphating roasting in the purpose of transforming the valuable metals present in the pyrite and its roasting residues, or in the iron matte in the form of sulphate, and the iron in the form of oxide, after which the sulphates undergo leaching and the solution obtained is separated, in order to recover the valuable metals from the solution, and the insoluble iron oxide is used for the manufacture of steel
b) the dust containing the valuable metals is separated from the gases,
c) the separated dust is sent directly to the sulphation screen at point a, after all the arsenic present has been removed first, or the separated dust is sent to a separate sulphate screen for the purpose of transforming the metals of value in a sulphate form, and iron in an oxide form, after which the sulphates are leached in accordance with point a, and the insoluble iron oxide charged with arsenic is rejected, and
d) the sulfur is condensed from the gases.

 The invention will be better understood if reference is made to the description below, as well as to the accompanying drawings which form an integral part thereof.

 Figures 1 to 3 show diagrams of three different possible global processes.

 FIG. 4 shows in detail the recovery of sulfur from the gases of the sulfur melting furnace.

 FIG. 5 represents a possible method for the roasting of sulfation of the roasting residues.

In the process according to FIG. 1, the sulphide-oxide matte from the arc melting furnace is grained during phase 1 and ground during phase 2. Part of the matte is roasted to death during phase 3, the temperature being from 950 to 1,050 OC. The following reactions occur during this phase: 1) FeS + 1.75 02 - 0.5 Fe203 + S02 2) Fe304 + 0.25 02 - 1.5 Fe2O3 3) FeO + 0.25 02 - 0.5 Fe203 4) PbS + 1.5 02 - PbO + S02 5) ZnS + 1.5 02 + Fe203 - ZnO. Fe203 + 502 6) Cu2S + 1.5 02 + Fe203 - Cu2O. Fe203 + S02
The sulfides of Co, Ni, and Cd also oxidize, respectively forming oxides or ferrites. The roasting of this phase is directed to the next phase 4, i.e. the sulfating roasting (Finnish patent No 41.874), to which also the final matte is directed. The sulphating temperature is 650 to 750 "C. The iron compounds are mainly oxidized to hematite, and partially to magnetite and water-soluble sulfates.
Cu, Zn and Pb are sulphates to sulphates, copper and zinc sulphates being soluble in water, and for a small quantity also in copper and zinc ferrite which are not soluble in water.

Co, Ni and Cd are also sulfated to their respective sulfates. The roasting is leached during the next phase in the iron residue wash water, which is obtained from phase 6, where the water-soluble sulfates are dissolved. Iron residue
Fe203 are directed to the chloride wash cycle. The acid-sulfate solution is directed to the recovery of metals.

 The dust from the arc melting furnace is directed from the electric filters to the sulphating mesh 12, where the temperature is 620 to 700 OC. The necessary Na is added in the form of, for example, Na2SO4. During this phase, most of the iron sulfide oxidizes to hematite and, to a small extent, also to magnetite and water-soluble sulfates.

Cu and Zn sulfides are partly sulphates to sulphates, and partly oxides to ferrite insoluble in water. Pb, Co, Ni and Cd are sulphated, lead sulphate being insoluble in water. The roasting is leached during the next phase 13 in the washing water of the precipitate obtained from the iron recovery phase 17, the washing water coming from phase 18, and therefore the metal sulfates not ferrous present in dust dissolve. The precipitate is directed to the strong acid leaching phase (SAL), and for this phase, the necessary amount of return acid is introduced from the electrolysis of zinc in phase 22, for the purpose dissolve the ferrites of the dust mesh. The temperature of this phase is 90 to 100 OC, and its concentration of H2SO4 is 30 to 150 g / l. The leach residues are washed during phase 15 and directed to the chloride leach cycle. The acid sulfate solution is directed to the recovery of metal.

 The sulphate solution obtained from stage 5 of leaching of the matte is directed to phase 5 of cementation of copper, during which the copper is cemented by means of iron, and therefore, of cementation copper is recovered. The sulphate solution is directed to phase 8 of iron extraction, during which the part of the acid from the electrolysis return which is not used for leaching during phase 14 is neutralized by means, for example, of limestone, and the iron in ferrous form is extracted by oxidizing it by means of air, the temperature entering from 85 to 95 OC, and by precipitating it in the form of goethite using, for example, lime as neutralizing agent. The gypsum produced and the iron precipitate are washed during the next phase 9 and extracted, the washing water from phase 9 is directed to phases 8 or 10, and the sulfate solution is directed to phase 10. of metal hydroxide precipitation, during which Zn, Co, Ni and Cd are precipitated in the form of hydroxides, for example by means of lime. The precipitate is used as a neutralizing agent for the iron extraction phase 17 of the dust cycle. Part of the metal-free sodium sulphate solution is used for the precipitation of lead carbonate, if the lead is recovered in the form of lead carbonate, otherwise it is vaporized in a form suitable for being directed to grinding of the matte during phase 2, and therefore, the sodium sulphate necessary for the sulphation roasting is obtained. The dust of the sulphate matte is directed from phase 13 of leaching of the roasting product to phase 16 of cementation of copper, during which precipitation occurs by means of, for example, iron.

The solution is directed to phase 17 of iron extraction where the solution of phase SAL is introduced, and part of the washing water of the phase 15 of washing the precipitate is neutralized, by means, for example, of lime and metal hydroxide precipitate from phase 10. The iron and gypsum residues are washed with water during phase 18, and the washing water from this phase is used in phase 13 in water leaching. The zinc sulphate solution is directed to the next phase 19 of purification of the solution, during which Cu, Co and Ni are cemented according to known methods (Finnish patent No 52.595 and patent application No 773589 ) by means of zinc dust, using arsenic trioxide as reagent, and the cadmium is cemented by means of Zn dust during phase 20 according to methods known per se (Finnish patent No 50.715).

 The solution is concentrated by evaporation during phase 21. The concentrated sulphate solution is directed to the electrolysis of zinc 22, which operates according to a known method of electrowinning.

Zinc is recovered from phase 22, and the acid in return is used for the uses described above.

The iron residues from phase 6 are directed to phase 23 chloride leaching. The iron residues are leached into a sodium-calcium chloride solution, whereby the water-insoluble lead sulfate is put into a water-soluble form. Chlorine is added for the oxidation of noble metals, and lime is used as a neutralizer and for the extraction of sulfates. In order to avoid the concentration of certain impurities, i.e. iron in the solution cycle, part of the solution is extracted, and in order to maintain the level of Na, NaCl is added . The Fe203 obtained is washed during phase 24 and directed directly for the production of iron. The chlorinated solution of phase 19, containing Pb, Ag and Au, is directed to the second filtering 15, during which the iron leach residues
with strong acid sulfated dust undergoes leaching. In this phase, too, lime is used to extract the sulphate and control the Ph. Blown air is used to extract the excess chlorine. If the iron has dissolved during the preliminary filtering 23, it is precipitated at the same time as the arsenic of phase 14, produced by humidity. The iron residues carrying arsenic from this phase are sent to the disposal zone after a washing phase 26. The chlorinated solution, which contains a little copper and zinc in addition to lead, silver and the gold, is mixed with the washing water coming from phase 26, and directed to phase 27, during which the noble metals and copper are separated from the solution, for example, by cementation using dust of zinc. Lead and zinc are precipitated during the next phase 28 using, for example, hydrogen sulfide. The precipitate is washed during phase 29 and the washing water is mixed with the chlorinated solution and directed to the evaporation phase 30, from which the solution from which the excess NaCl has been extracted is recycled to phase 23.

 The lead and zinc sulphides from phase 29 are dissolved in water by oxidation with air in an autoclave during phase 31. The lead sulphate precipitate is washed during phase 32, and the water is directed to the autoclave. The zinc sulphate solution is directed to the phase 10 of metal hydroxide precipitation. The lead sulphate is transformed into lead carbonate in phase 33 by excess residue delivery in a solution of sodium sulphate (coming from phase 8) using sodium carbonate.

The lead carbonate precipitate is washed during phase 34 and the washing water is recycled to the previous phase. The solution from phase 23 contains sodium sulphate, and it is used, optionally with the NaSO 4 solution from phase 10 for roasting the mat.

 The dust roasting cycle can be combined with the mat roasting cycle, if the arsenic was first extracted, and thus cannot pollute the iron ore. The double mesh replaces the strong acid leaching of the dust, and increases the yield of sulphates.

 In the process illustrated in FIG. 2, the cycles of filtering the roasting residues of the matte and the dust have been combined, and therefore the filtering of the dust with strong acid is eliminated, and the arsenic is extracted from the dust during a separate phase 7, and the filtering residues of the roasting product from phase 9 are directed to the sulphating roasting 4 of the mat.

 In this way, only one kind of iron residue is obtained, and these residues are directed to the chloride leaching cycle. The sulfate solutions obtained by leaching the roasting products can also be treated in a single cycle. In this case, the solution is directed via the phase 10 of cementation of the copper to the phase 11 of precipitation of iron, to which is added, as neutralizing agent, in addition to the addition of lime, the precipitate of the zinc hydroxide precipitation phase 18 following neutralization 16 of the return acid obtained from the electrolysis of zinc. The sulphate solution now passes via phases 13 and 14 of extraction of Cu, Co, Ni and Cd to phase 15 of evaporation and then to the electrolysis of zinc. Since there is only one kind of Fe 2 O 3 residue, the separate chloride leaching phase for the dust is eliminated and the chlorinated solution cycle differs from the procedure in FIG. 1, after the washing phase for the precipitate obtained from the precipitation phase of lead sulphide and zinc sulphide, in that the sulphates are extracted from the solution during phase 25 by the addition of lime, and the residual precipitate of gypsum obtained is washed, and the solutions are mixed, and the excess NaCl is extracted from the solution before the evaporation phase. The lead carbonate is recovered in substantially the same manner as that illustrated in FIG. 1.

 If it is desired, in accordance with FIG. 3, to combine the filtering cycles of the residues from the roasting of the matte and of the dusts, and not to carry out a separate extraction of arsenic from the dust, the residuals from the sulfating roasting dust undergoes leaching during strong acid leaching (SAL) 9, in order to dissolve the ferrites. However, in this case, the acid back from electrolysis is not introduced into the SAL, as is the case of the process illustrated in Figure 1, but a solution of pure H2SO4 is used in order to maintain a concentration as appropriate sodium in the sulfate solution, and thus, the sodium cannot interfere with the electrolysis of zinc. In other ways, the sulfate solution is purified as in the process of Figure 1, it that is to say, the solutions coming from phases 5 and 8 of filtering for the matte and the dust are mixed and the sulphate solution passes via phase 11 of copper extraction, phase 12 of iron extraction, phase 14 for cementing Cu, Co, Ni, and phase 15 of cadmium towards evaporation, and then electrolysis 17 of zinc, whose acid in return is neutralized during phase 18 separated by means, for example, of limestone, and during the next phase i9, Zn (OH) 2, which is used as a neutralizing agent for ant phase 12 of iron extraction, is precipitated. The Na2SO4 solution is evaporated, and optionally used with the Na2SO4 solution obtained from the precipitation phase of lead carbonate for roasting the matte.

 The chloride filtration cycle is the same as in the case of FIG. 1, that is to say, the iron residues obtained from the dust cycle are directed towards the waste discharge zone, because of arsenic present.

 Figure 4 illustrates in detail how the dust is separated from the gases from the arc melting furnace, and how the sulfur is recovered from the gases, as elemental sulfur.

 Figure 5 shows in detail a process for the roasting of sulphated roasting residues, using pyrite to fulfill the heating condition, and to create an appropriate gas atmosphere in the roasting oven.

 The roasting residues are produced in the sulfuric acid industry, which roasts pyrite in order to obtain sulfur dioxide, the basic material for the production of sulfuric acid. By means of the present invention, such scorch and impure pyrite residues can be effectively exploited.

 The invention will now be described with the aid of an example.

EXAMPLE
The basis for the calculations consists of the roasting and filtering phases of 7,000 h / a and the solution purification and metal recovery phases of 7,700 h / a. The pyrite concentrate was introduced into an arc melting furnace at a speed of 82 t / h.

Oil was added in an amount of 3.1 t / h, and carbon was introduced into the gas reduction zone in an amount of 11.5 t / h.

 Matte was obtained from the oven in an amount of 55 t / h, and flying dust for 12 t / h. Elemental sulfide is obtained from the fusion gases for 14 t / h.

The analysis of 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
At 0.0008
Hg 0.01
Se 0.002
The mat was granulated in water and milled so that 80% had dimensions less than 74 microns and 100 e / - dimensions less than 290 microns. The amount of sodium in the ground grains was 0.94 t / h. An oxidizing mesh is supplied with
matte in an amount of 34.5 t / h and a sulphation grid is fed with matte in an amount of 20.5-t / h and product toasted to death in an amount of 33.4 t / h gratings from the sulphating mesh of the matte, the analysis of these products being as follows:
S 2.63%
Fe 60.6%
Cu 1.69%
Zn 1.66%
Pb 0.28%
Na 0.54 X was elixivified using washing water in an amount of 35.1 m3 / h, and the residues were obtained in an amount of 49.9 t / h.

 The sulfate solution was sent to the copper cementation, to which iron debris was added for 1 t / h.

This phase produced 90% copper-carburizing in an amount of 0.91 t / h. The sulphate solution, in an amount of 23.9 m / ha was sent to the iron extraction phase, to which limestone was added in an amount of 1.82 t / h, and lime for 2 , 22 t / h. The amount of acid in return was 9.9 m3 / h, that of the water arriving from the SAL phase of 6.0 m3 / h, and that of the washing water of 19.8 m3 / h.

 The product of the goethite-gypsum precipitate was 9.9 t / h, and the sulfate solution, 53.0 m3 / h, was sent to the metal hydroxide precipitation phase. Calcium was added for 1.80 t / h at this phase, and the solution arrived, for 4.4 m3 / h, from the lead sulfide-zinc sulfide leaching phase. The amount of hydroxide precipitate was 6.1 t / h, and the amount of Na2SO4 solution was 53.3 m3 / h.

The dust from the arc melting furnace, 12 t / h, was sent to the sulphating grid, for which the addition of Na2SO4 was 0.2 t / h. Roasting products, 12.1 t / h, which contain
S 8.9%
Fe 21.9%
Cu 0.61%
Zn 15.8%
Pb 8.4%
Nua 0.53% was filtered in water from the washing phase of the iron precipitate, in an amount of 15.3 m3 / h, and the residues were obtained for 7.5 t / h, and were directed to strong acid leaching, where the return acid reached 14.8 m3 / h, and the amount of water arriving at the residue washing phase was 13.3 m3 / h .The residue yield was 6.9 t / h, and it was directed to the chloride leach cycle
The sulphate solution - for 12.8 m3Xh, was directed to the cementation of copper, where iron waste was introduced for 0.076 t / h. CjO% carburizing copper was obtained from this phase at 0.052 t / h, and the sulphate solution was introduced into the iron recovery phase, in an amount of 7.6 m3 / h, and in this phase , metal hydroxide p precipitate was added, from the leaching cycle for the toasted matte, in an amount of 6.1 t / h, the addition of lime was 0.08 t / h, and the solution arrived from the strong acid leaching at 14.8 m3 / h, as well as the washing water from the strong acid leaching at 6 m3 / h, and the washing water from the residues of iron at 13.9 m3 / h. The yield of the geothite-gypsum precipitate was 4.9 t / h, and the sulphate solution for 33.3 m3 / ha was directed to the Cu-Co-Ni cementation phase where zinc dust was added to 78 kg / h, As 2 O 3 for 2 kg / h, and the yield of the Cu-Co-Ni precipitate was 0.37 t / h. The solution was sent to the cadmium recovery phase for 33.3 m3 / h, and the zinc dust for 3 kg / h. The cementation cadmium yield was 3 kg / h. The sulfate solution was directed to the evaporation phase, at 33.3 m3 / h and the output was 24.5 m3 / h and was sent towards the electrolysis of zinc where zinc cathodes were obtained at a speed of 2.5 t / h, and of the return acid at 24.5 m3 / h.

In the first phase of the chloride leaching cycle, the leaching of the roasted matte was sent at 45.4 t / h, chlorinated solution at 86 m3 / h, washing water at 20 m3 / h , Cl2 at 0.29 t / h, lime at 0.46 t / h, and NaCl at 0.47 t / h. Fe203, a suitable starting material for the production of iron, was obtained at 45.5 t / h, and contained
Fe 67.5 X
Zn 0.12%
Cu n. .09%
Pb 0.04%
S 0.2% and the chloride solution at 100 m3 / ha was directed to the leaching phase for dust roasting products. Residues from roasting of dust were obtained at 6.3 t / h and washing water at 14.8 m3 / h and lime at 0.41 t / h. Scrap iron residues were obtained at 5.6 t / h, and the chlorinated solution at 11.5 m3 / h. Then,
Cu, Ag and Au were cemented with the chloride solution. Zinc dust was used for 0.04 t / h and cementitious silver-gold-copper was obtained in an amount of 0.019 t / h. The chlorine solution was sent to the lead and zinc precipitation phase in an amount of 117.2 m3 / h, and hydrogen sulfide was added for 0.31 t / h, the yield of the precipitate of PbS, ZnS was 1.3 t / h, and the quantity of washing water was 2.9 m3 / h. The amount of the final solution was 4.5 m3 / h, and the chloride solution has passed through the evaporation phase, in an amount of 112.7 m3 / h, and at its outlet of 80 m3 / h. The PbS, ZnS precipitate was leached in the autoclave, to which washing water was added for 4.5 m3 / h. The leach residues were obtained in an amount of 1.6 t / h, and the zinc sulphate solution, 4.4 m3 / h, was directed to the metal hydroxide precipitation phase of the matte cycle . The leach residue was directed to the lead carbonate precipitation phase, where the residual sodium sulfate solution from the matte cycle reached 5 m3 / h, and the addition of Na2CO3 was 0.6 t / h and the quantity of washing water of 4.5 m3 / h. The sodium sulphate solution, 9 m3 / ha was directed to the roasting of the mat and the lead carbonate was obtained in an amount of 1.4 t / h, which was calcined into lead oxide, the amount of which 1.16 t / h, and the lead oxide was reduced to metallic lead, the yield being 1.0 t / h.

Claims (3)

 1. Process for the recovery of valuable metals from finely ground impure pyrite, either directly or by first melting the pyrite with the aid of non-oxidizing gases, at a high temperature, in order to produce iron mat with a proportion of valuable metals and dust gases, characterized in that:  a) a part of the pyrite or of the iron mat which contains valuable metals is first subjected to an oxidizing roasting and then, with the rest of the pyrite or of the iron mat, to a sulphating roasting in the purpose of transforming the valuable metals present in the pyrite and in these roasting residues or in the iron matte in the form of sulphate, and the iron in the form of oxides after which the sulphates undergo leaching and the solution obtained is separated for the purpose of recovering valuable metals from the solution, and the insoluble iron oxide is used for the production of iron,  b) dust containing valuable metals is separated from gases,  c) the separated dust is introduced directly into the sulphating screen of point a, after all the arsenic present has first been extracted, or  the separated dust is sent to a separate sulphating screen in order to transform the valuable metals into a sulphate form and the iron into an oxide form, after which the sulphates are leached in accordance with points a and l insoluble iron oxide carrying arsenic is rejected, and  d) the sulfide is condensed from the gases.  2. Method according to claim 1, characterized in that the sulfates undergo leaching in two phases  a) first by using water or diluted acid in order to dissolve the easily soluble sulfates, and to separate them from the solid matter and  b) the solid obtained is subjected to chloride leaching in order to dissolve the lead, silver and / or gold and to separate them from the insoluble iron oxide.  3. Method according to claim 2, characterized in that zinc dust is added to the solution obtained from the chloride leaching in order to cement ltor and silver and to separate them from the solution, after which hydrogen sulfide is directed into the solution in order to precipitate lead and zinc as sulfation, the sulfate precipitates undergo leaching in an autoclave by air oxidation in order to dissolve the zinc, and the separated insoluble lead sulphate is transformed, in a solution of sodium sulphate, by means of sodium carbonate, into carbonate of lead, which is separated from the solution and thermally decomposed into lead oxide which is finally reduced to lead.