FI85878C - Process for reducing melting of materials containing base metals - Google Patents

Description

1 85878

This invention relates to the field of extraction metallurgy of copper, nickel and cobalt. More specifically, it relates to a process for the reductive smelting of various raw material and intermediate materials containing these metals.

In general, raw material and intermediate materials containing base metals can be divided into two main categories.

10 The second category represents a large number of raw materials and intermediate products containing non-ferrous base metals and significant amounts of iron. The main purpose of the reduction smelting of such materials is to separate the non-ferrous parent metals into a highly concentrated liquid phase containing these metals in metal and / or sulfide form and removing most of the iron in the form of slag containing as low concentrations of non-ferrous metals as is economically appropriate.

Such metallurgical technology is currently used in the processing of nickel laterite ores, partially roasted sulphidin nickel concentrates, tanned roasted copper concentrates (calcines) and various and multi-metallic raw or intermediate materials, including dusts. With the exception of very few obsolete plants, these 25 technologies are now practiced mainly in electric furnaces * ·:: under reducing conditions.

‘Although reductive electric smelting proves to be very effective in separating non-ferrous metals from iron, this process involves high capital costs and: :: 30 operating costs. Besides, in most cases it requires very careful and expensive preparation of the feed. For example, nickel laterite ores must be good. : preheated and preheated; nickel sulphide concentrates must be partially or thoroughly purified from sulfur 35 and agglomerated; copper calcines must contain the lowest possible sulfur content, etc. In addition, a non-ferrous metal producer who chooses to use a reductive electric smelting process is confronted with the choice of whether to buy electricity or build his own power plant. In the former case, the producer becomes financially dependent on the energy producer, while in the latter case, the capital costs are even higher.

Another disadvantage of reducing electric melting technology stems from the fact that the best thermal efficiency 10 of electric furnaces is achieved when the feed is a dense, low-porosity material, for example in the form of briquettes or granules, while the materials fed to electric furnaces are mostly fine. When finely divided material is reduced-melted in an electric furnace, numerous undesirable phenomena occur, such as poor heat transfer to the feed, unnecessarily high slag overheating and heat flow to the refractory walls of the furnace, premature and useless burning of solid reducing agent added to the feed, etc. for example, flame smelting has a limited ability to utilize the energy obtained from the oxidation of sulfide sulfur in non-ferrous metal concentrates and is therefore not an energy efficient process. In addition, reductive electric smelting is an inflexible process that is very sensitive to changes in feed and slag composition, especially with respect to iron and silica.

* The second category of materials containing non-ferrous base metals represents a large number of intermediates that do not contain significant amounts of iron, if any.

::: 30 The main purpose of the reduction smelter in this case is to produce a smelter which is practically oxygen-free and which is suitable as either a final product or an intermediate product. : for processing, for example for cleaning or alloying with other metals, as the case may be.

il 3 85878 Of the various materials in this class, the following may be mentioned: nickel oxide, cobalt oxide, copper oxide, copper sulphide or nickel sulphide or a mixture of NiCu metal compounds with these sulphides (for example, concentrates from slowly cooled and then ground nickel-copper converter metals) hydrometallurgical concentrates such as partially oxidized copper cement, hydroxides and / or carbonates of nickel and cobalt and many others. These materials are in most cases 10 also quite finely divided. Although they can be processed in many different ways, in most cases they are also reduced and melted in electric ovens, unless the old technology is used, i.e., for example, furnace melting or melting in converters. A roasting operation to remove sulfur is required prior to reductive smelting.

Thus, it is apparent that the reduction smelting of finely divided materials has a wide range of applications in the metallurgy of non-ferrous base metals. It is carried out in most cases in electric furnaces and this process is characterized by a number of serious shortcomings which have already been mentioned above.

The main impetus for the development of new processes for the reduction smelting of materials containing non-ferrous base metals is to reduce capital costs and operating costs and to provide operational flexibility * ·:: to process a wide range of materials without having to take any major technical or operational action. The new process must be energy-efficient and non-polluting, and it must be able to take full advantage of the fact that most of the materials to be treated are already fine-grained or can be easily ground.

: In recent years, numerous new reduction smelting processes have been developed to produce iron from oxide ores and; \ 35 concentrates. The five most promising of them have been studied in Article 1 ·. h 4 85878 kelissa J. J. Moore, "An Examination of the New Direct Smelting Process for Iron and Steelmaking"; published in the Journal of Metals, June 1982, pages 39-48. Each is a continuous two-stage process with (1) pre-5 heating and partial pre-reduction followed by (2) final reduction and smelting and the resulting product is a high carbon molten iron .

Of these five processes, the Inred process is of particular interest because it has some features in common with the present invention and is claimed to be suitable for processing oxide materials that also contain non-ferrous metals. This process has been described in detail previously in H. J. Elvander, I. A. Edenwall, and S. C.-J. Hellestan "Boliden Inred Process 15 for Smelting Reduction of Fine-grained Iron Oxides and

Concentrates ", published in Ironmaking and Steelmaking 1979, Vol. 6, No. 5, pages 235-244 and later in HI Elvander," The Inred Process - A Revolutionary Method to Produce Hot Metal "; published in Iron and 20 Steel Engineer, April 1982, pages 57 to 80.

In this process, the flame melting furnace and the electric melting furnace are combined into a single reactor with the flame melting furnace mounted above the electric furnace. In the first step of the process, coal powder and fluxes are mixed with finely divided iron oxide and the mixture is injected into a flame melting furnace in a stream of gaseous oxygen.

* Coal partially burns and partially carbonizes to coke, while the oxide melts and pregelatinizes to wustite (FeO).

In the second step of the process, the pre-reduced hot ::: 30 material containing the formed coke is collected in: Y: an electric furnace. Molten FeO and coke bodies fall on the surface of the charge previously accumulated in the electric furnace and form a large amount of semi-molten, highly viscous pasty mass. Due to the endothermic pel- * of wustitis. From 35 shrinkage reactions to metallic iron, the mass cools rapidly to 1,150 ° C. The upper balance of the charge in the electric furnace consists of an iron sponge, two non-reducing molten material, and quicklime. The final reduction of FeO and melting of the iron sponge is performed at a lower level around the 5 electrodes. The hot metal collects under a slag bath containing a partially submerged layer of coke.

The waste gas from the partial combustion of coal in the flame smelting zone and the carbon monoxide rising from the electric furnace are burned with secondary and tertiary oxygen at the top of the flame smelting chamber. As a result, the temperature of the gases leaving the flame melting chamber is about 1,900 ° C. The heat from these gases is utilized to produce dry steam, which in turn is used to generate the electrical energy needed for the electric furnace and oxygen plant. The efficiency of converting heat from steam to electrical energy is about 35%.

However, when processing the total oxygen consumption, 65% of the Fe hematite concentrate is about 40 and 60% by weight of the 20 concentrates, respectively. At these values, the process can be energy self-sufficient.

The two main features of the inred process should be mentioned: (a) coal powder is the only substance used as both fuel and reducing agent and (b) the conditions in concentrate-25 coal-oxygen jets are such that solid •: carbon is always present in excess. Only part of the feed coal is burned. Thus, the aeration ratio in the showers is less than 50% and therefore the combustion gases of the shower must consist almost exclusively of carbon monoxide. Under these conditions the combustion is: 30 very inefficient and the temperature of the jet flame is much lower .1 ♦ 1 · than it could have been if the aeration ratio had been reasonably close to 100%.

. ·. j Two variants of the Inred process were proposed as methods for processing abrasive sulfide materials containing pyrite, pyrrotite, Kalbopyrite, 4 ···· • · • · · • · * · · · · · · • · «m 6 85878 leaded sand, pentlandite, etc., as described in U.S. Patents 4,087,274 and 4,388,110.

According to U.S. Patent 4,087,274, the process consists of three successive steps performed in three successively connected reaction zones arranged vertically on top of each other. In the first zone, which is the upper part of the reactor, the metal sulphide-containing material is incinerated (oxidized) with oxygen and the resulting hot metal oxide-containing material then falls downwards into the second zone.

In the second zone, the finely divided solid carbonaceous reducing agent is fed into a stream of oxygen-containing gas and combusted at an aeration rate low enough to cause only partial combustion of the reducing agent while converting part of it to coke. The metal oxide-containing material is partially reduced as it falls downwardly through this zone and enters, together with the coked reducing agent, into a third zone to form a substantially solid product (sinter) containing partially reduced material and solid carbon.

The third zone is essentially the same electric furnace as in the Inred process described above. In this zone, the final reduction and sinter smelting take place.

25 It is obvious that oxidizing conditions are achieved: in the upper part of the reactor, while strongly reducing V: conditions prevail in the two lower zones.

Any reducing gas formed in these two zones is combusted by supplying oxygen to the upper levels of the reactor, immediately below and inside the first zone. This modification does not provide for the selective recovery of non-ferrous metals and can only be used for: processing ferrous sulfide concentrates. In addition, there is a huge excess of heat in the first • * ·! ./ zone because of the sulfi- *. In addition to the combustion of 35 dimer minerals, the reduction gases entering in zones 2 and 3 are also combusted there.

• · · • · • · • · * * • 85878

Another modification of the Inred process according to U.S. Patent 4,388,110 represents a fundamental deviation from the sketch described in the aforementioned articles and U.S. Patent 4,087,274 in that carbonaceous fuel and / or reducing agent is no longer added to the Inred reactor, no metal oxide preaxiation, and the formation of a coke-containing sinter does not occur and only very limited amounts, if any, of molten non-ferrous metals are formed in the reactor. This modification simply consists of autogenous flame (flash) melting of the sulfide materials 10 in the presence of excess oxygen and silica flux to form a low sulfur silicate melt, followed by separation of any non-ferrous phases from the solicate phase, if any. ferrous metals by its selective reduction in one or more stages in at least one additional furnace. Thus, in addition to the Inred reactor, at least one separate furnace is required to perform selective reduction of the molten silicate phase. It is obvious that this modification is not simply a method for reduction smelting, but for the production of silicate slag containing non-ferrous metals, which slag must then be reduced in a manner known per se in order to selectively recover these metals. 25 The curiosity of metallurgists has long been maintained by the • · · attractiveness of developing a one-step process for the direct production of molten metal *. · In a reducing flame obtained from the partial γ (incomplete) combustion of fuel. Examples of such mental:: ': 30 interesting ideas can be found in U.S. Patents .γ; 774,930, 817,414, 1,847,527, 4,421,552, CA Patent 864,451 • 1 and SU Copyright Papers 86,983 and 199,397. U.S. Pat. No. 4,421,552 discloses a process for producing copper metal from a tanning-baked copper-ferrous metal-calcined (calcine) tanning-calcined calcine by a reducing flame (flash). ) by smelting using pulverized coke, coal or other reducing agents (fuel). Calcine is charged to the reducing flame smelting zone together with oxygen and reducing agent fed through the same burners with sufficient amounts of oxygen and reducing agent to effect reduction and melting of the copper content of the calcine and to produce molten iron silicate slag. The slag must be subjected to further processing before discharge, as it has a copper content of about 5% and a slow cooling-grinding-foaming technique to clean the slag has been considered to obtain the slag.

Unfortunately, none of these reducing single-stage flame (flash) smelting methods have paved the way for industrial application to produce either iron or non-ferrous metals. The main difficulties associated with these one-step processes include the following. A high degree of reduction of fine oxides of non-ferrous metals is extremely difficult to achieve, even with combustion gases with a very high reduction potential. On the other hand, the increase in the reduction potential of the flue gases in the flame 20 requires that the aeration ratio of the fuel be calculated so that only a very low level of fuel combustion is achieved. However, the more incomplete the combustion, the less efficient it is in terms of utilizing the thermal power of the fuel 25 and the lower the temperature of the theoretical flame.

* · This dependence is shown quantitatively in the figure of the drawing using methane and oxygen as an example. In Figure 1, the fuel aeration ratio is the actual amount of oxygen Y is fed per given amount of methane divided by the theoretical amount of oxygen required to perform complete Y combustion of this methane multiplied by 100; the net heat available at 1,600 ° C and shown in Figure 1 ·. : shows curve B, is the amount of heat in J / g-mol CH., ♦ · 4 λ · * available as a result of combustion at this temperature;

35 flue gas reduction potential shown in curve A

• · · · «· • · · • · · i · · I · · I · (9 85878 in Figure 1, is the volume-to-concentration ratio between H2 + CO and 1 ^ 0 + CC ^; and the theoretical the flame temperature shown by curve C in Figure 1 is the temperature that results from the combustion of methane with oxygen.All data in Figure 1 are given for an absolute pressure of 101 kPa and an initial temperature of 20 ° C for methane and oxygen. fuels other than methane, such as oil or coal.

Figure 1 shows that any attempt to improve the reduction of fine metal oxides in a flame by enhancing the reduction potential of the combustion gases, i.e., reducing the fuel aeration ratio, inevitably involves a dramatic decrease in the efficiency of direct utilization of the thermal power of the fuel. For example, burning methane at a 50% aeration ratio results in only about 20% of the net heat that may be available at an aeration ratio close to 100%. In addition, the theoretical flame temperature at a 50% aeration ratio is about 2,275 ° C, which is about 500 ° C units lower than the theoretical flame temperature caused by an aeration ratio close to 20%. This lower theoretical flame temperature results in a reduction of the heat transferred by the radiation from the flame by a factor of 2,2.

Thus, huge, uneconomical amounts of fuel as well as oxygen would be required for the process evaluated in U.S. Patent 25,421,552 or any other process in which reduction of fine non-ferrous metal oxides V: would be attempted in a reducing flame obtained from incomplete fuel combustion. Naturally ·; · exhaust gases which still contain significant H2 and / or; : '; 30 CO concentrations, can be afterburned and the heat thus generated can be partially recovered, for example in the form of electrical power, as is done in the case of the Inred process. However. . even this improvement path in the total! ./ female utilization of the heat capacity of the fuel can be economically uncertain for the non-ferrous metal producer, as it requires considerable capital costs and operating costs. In addition to having to deal with much larger volumes of exhaust gases and their afterburner gases, the process produces much larger amounts of dust, which these large gas volumes 5 generate. It is very expensive to produce electrical energy by burning the fuel with oxygen at a very low aeration ratio and then recovering the chemical heat contained in the exhaust gases by afterburning, although this electrical energy can be used to produce oxygen. Such 10 energy recovery is not only thermodynamically inefficient but also highly capital intensive.

A non-ferrous metal producer would greatly benefit from a new reduction smelting process that would require a minimum amount of capital expenditure and use minimum amounts of fuel and oxygen; a process that would generate minimum amounts of gases dust; a process that would be flexible in terms of its suitability for processing a variety of fine-grained materials and the use of any fuel available at the lowest cost; a process to be carried out in a single reactor of a simple design and as inexpensive as possible; a process that would directly produce the desired slag without the need for any slag cleaning operation; and a process that is easy to adjust and use.

The present invention relates to a process for the reduction smelting of a finely divided material containing at least one parent metal from the group consisting of copper, nickel and cobalt, at least in part in the oxide form, characterized in that a) the material is injected together with fuel and oxygen and any fines. with the fuel in a confined space while burning the fuel to produce a non-reducing high temperature flame. 35 si; »*» * 11 85878 (b) superheating the finely divided particles containing the parent metal in a substantially non-reducing flame to a temperature above the highest melting point of the reduced product to be manufactured; 5 c) the superheated particles are applied substantially evenly to a thin layer of particulate coke floating on the surface of the molten bath of reduced products, the thin coke layer and the adjacent atmosphere forming a reduction zone; D) the oxides of the parent metals are reduced in the reduction zone, whereby the reduced products are obtained in a liquid state while passing all the required heat to the reduction zone only in the form of free heat of superheated particles and radiation from a non-reducing flame; (e) the reduced products are sieved through a layer of coke into a molten bath; (f) the reduced products are removed from the molten bath; and 20 g) feeding the solid, granular, carbonaceous material to a thin layer of granular coke to replenish the coke consumed in the reduction.

When a fuel, e.g., a hydrocarbon fuel other than the sulfur or iron content of the finely divided material, is used as the main heat source for the process, the aeration ratio of the fuel-treated flame should be close to 100%. When a finely divided sulfide material is also present in the flame, it can provide some of the required heat, but the flame is in any case unreduced with respect to the metal oxide to be produced.

'' Figure 1 is a graph illustrating the dependences of the fuel aeration ratio on: (A) the reduction potential of the combustion gases, * (B) the net heat available at 1600 ° C, and (C) the theoretical flame temperature.

12 85878

Figures 2A and 2B are schematic views of an ideal furnace suitable for carrying out the process of the present invention.

More specifically, in the method of the invention: (a) at least one artificial material containing at least one non-ferrous parent metal, fuel and / or at least one finely divided sulfide material, which may also contain one or more non-ferrous parent metals, finely divided, is injected through the burner 10 for the slagging of iron, if present in significant amounts in said materials, and gaseous oxygen in a preheated combustion zone at a weight ratio of oxygen to fuel and / or finely divided sulfide material reasonably close to that required to perform said fuel 15 and / or sulphide material most efficiently; (b) burning the fuel and / or sulfide material to produce a substantially unreduced high temperature flame and superheat any condensed particles of said materials to a temperature substantially higher than the highest melting point of the reduced product produced in the process; (c) injecting (or perhaps sprinkling or spreading) the superheated condensed particles from the flame down to the reduction zone into a thin layer of granular coke floating on the surface of an already existing molten bath of said reduced product while separating the combustion gases from the superheated and directed upward; (d) reducing the oxides of non-ferrous base metals in said materials and / or formed in a flame to the corresponding metals in a thin layer of granular coke having a predominance of iron oxides, if present in said materials, and / or is formed in a flame, •: ··· is reduced only selectively to the degree of reduction, “·. 35 required for discarded molten, iron-containing • slag. to form; * · M »♦ · * S, • · r 13 85878 (e) sieving said reduced non-ferrous base metals and / or parent metal sulphides, including ferrous sulphide when required, and ferrous slag, if present; through a coke bed to said molten 5 bath; (f) continuously or intermittently removing a portion of the final reduced molten product, which is highly concentrated with respect to at least one non-ferrous parent metal, and iron-containing slag when said materials contain iron; and (g) feeding a solid granular carbonaceous material having a particle size at least 10 times larger than said finely divided materials to replenish the granular coke consumed in the reduction.

15 The following definitions apply to this specification and claims.

"Injection" means that solid, finely divided materials, including flux, are fed as a cohesive jet into a preheated combustion zone as a suspension of fine solid particles in a gas phase containing gaseous oxygen and may also contain gaseous fuel and any other gas that may be used in the process; «:" Burner "means any burner or device constructed for the same purpose · · · V: 25 capable of feeding the above suspension as well as any combination of the above *: 1 suspension with any non-gaseous fuel to a preheated combustion belt. • · · into the zone as a high-speed, cohesive jet so that any of the above-mentioned components of the jet se-. . with each other in any way. Mixing • · · • · · I .. 'can occur when the ingredients enter the burner, the burner itself. immediately after they have come out of the burner or in any combination thereof; "Finely divided material containing at least one non-ferrous parent metal" means a powdered material or mixture of materials containing one or more metals, such as copper, nickel, and cobalt 5, which may contains minor impurities of titanium, zinc, arsenic, antimony, selenium, tellurium, lead, bismuth, tin, etc., said metal or metals in one or a combination of the following chemical forms; oxides, oxides plus sulfides, oxides plus 10 metals, oxides plus metals and sulfides, metals plus sulfides, and sulfates, carbonates and hydroxides. This material may also contain ferrous precious metals, any combined metals, metalloids and non-metals and minor amounts, as well as stone. When iron is present, it may be in the form of ferrites of ferrous oxides and / or non-ferrous parent metals.

"Fuel" means any flowable fuel, such as natural or other gas, oil, or other liquid hydrocarbon, solid, finely divided, pulverized carbonaceous material, or any combination thereof. In some special cases, elemental sulfur may also form part of the fuel; : "Fine sulphide material" means any sulphide material, such as ore concentrates and l '· 25 different metal stones, which react with oxygen and may in some cases be used as a substitute for fuel; "Finely iron slag flux" means any siliceous and / or calcareous flux which forms either iron silicate or 30 iron calcareous slag; "Most efficient combustion" means that the fuel. * and / or the thermal power of the sulphide material when burned with “ϊ oxygen” is used up to a quantity reasonably close to the maximum. In the case of fuel, this is achieved when the aeration ratio of the fuel (actual amount of oxygen supplied 85878 divided by the theoretical amount of oxygen required to achieve complete combustion of all combustible substances multiplied by 100) is reasonably close to 100%, e.g. about 90-130 %. In the case of sulphide materials, this is achieved when the sulfur is oxidized to sulfur dioxide and / or the flue gases contain very little, if any, free oxygen, the thermal equilibrium requirements of the process being fully satisfied by the oxidation of sulphide sulfur; 10 "Fine" means that the substances so described have a state of comminution such that the average cross-sectional dimension of the individual particles is less than n.

500 micrometers, e.g., 100% to 10 mesh and 80% to 200 mesh (U.S. Standard Screen Kit); 15 "A thin layer of granular coke" means a layer that is about 1 to 5 cm thick, with a maximum coke particle size of about 15 to 25 mm, and the minimum particle size of the coke fed to the bed is at least about 5 mm; "Selective reduction of iron oxides to the 20 ° C reduction required to form molten, discarded, iron-containing slag" means that Ferri iron oxide, whether free or bound to ferrites, is reduced primarily to ferro-iron oxide, also known as slag requires no further processing

IM

A: 25 operations such as slag refining to recover non-ferrous metals and therefore can be directly discarded; : "Discardable slag containing iron". means iron-containing slag in which the concentrations of copper, nickel and / or cobalt are sufficiently low, so. . that, in economic terms, the slag is worth disposing of. The composition of the slag depends in part on the • • t | ** * on the composition of the materials and the product phase in contact with the slag. Slag may require treatment 35 for reasons other than copper, nickel or cobalt content »· ·! • · ·: · * i

···· * I

a * | 16 85878. For example, when treating copper-zinc-tin materials, slag may very well have to be treated to recover zinc and / or tin before it can be discarded.

5 "Solid, granular carbonaceous material" means any coking material having a particle size between about 5 and 25 mm.

The present invention is based on numerous observations made in the course of experimental work aimed at developing a simple, economical and flexible process for the reduction smelting of a large number of materials containing non-ferrous base metals in oxide and / or sulphide form as well as sometimes in partial metal form. What was the most valuable and unexpected invention was that when finely divided and superheated particles containing non-ferrous metals in oxide form were continuously sprayed or precipitated on a thin layer of hot granular coke, they were rapidly reduced to the corresponding metals, the metals melted and sieved. floating on top of a molten bath.

It has also been found that when iron oxides in the form of hematite, magnetite and / or non-ferrous parent metal ferrites were present among and / or within the finely divided, superheated particle residues, the rapid reduction of all oxide forms of non-ferrous parent metals progressed well: and completely as without the oxide form of iron and. It- · ;. except, this reduction was highly selective oxy-. . ·. with the iron being reduced mainly to the wustite form only.

It was further found that when finely divided and superheated runs intended for iron loading * ‘30. . tea particles were continuously sprayed in appropriate amounts • · · ** 1: on top of the coke bed together with superheated particles «1 ♦ * ♦ 1 1 containing non-ferrous metals and iron oxides, ···» · they react very rapidly with wustite to form. ** '35 melts iron silicate or iron calcareous slag.

• · · * · · • 1 1 • · • · · i7 85878

The formation of slag in situ, together with the reduction in the thin layer of coke, was found to give a very low content of non-ferrous metals to the slag and therefore to make the slag directly discarded. The slag sieved under the 5 coke layers very quickly and the separation of the non-ferrous metal phase and the slag was excellent because practically no piece of non-ferrous metal phase was found to be mechanically migrated from the slag.

Another very surprising and valuable finding was that after wustite was obtained by selective reduction of iron oxide and then loaded with silicon-containing flux and sieved under a coke bed in the form of molten slag, it was no longer suitable for further reduction by coke floating on the slag surface. This was the case even when the spraying was stopped. The exact mechanism of this unexpected phenomenon remains unknown at present, although some factors that are likely to contribute to it can be mentioned. These are the facts that in practice (a) the molten bath 20 under the coke bed remains largely immobile (b) only a small amount of coke sinks into the slag because the coke layer is thin and (c) the temperature at the slag-coke interface is relatively low (as opposed to e.g. .

!. ·: These and possibly other factors prevent more reduction of wustite from the slag, which results in excellent iron separation from non-ferrous metals.

The above observations as well as other novel features of the process are described below in the Examples of various embodiments of the present invention. Because the present invention relates to a wide variety of base metal-containing materials. . processing, its description is given in the context of specific examples · · · representing at least some of these materials. However, the essential principles of the process are described first, using the reduction of fine oxides of non-ferrous metals simply as a suitable description.

• »# ie 85878

In its most general and simplest form, the process of the present invention consists of three indivisible main steps that take place continuously and simultaneously in the same furnace: metals on the surface of the molten bath; and (c) reduction of oxides to the corresponding metals in a thin layer of 10 granular coke.

Step (a) was performed as follows with reference to Figures 2A and 2B of the drawing. The feed of oxides 11, fuel 13 and oxygen 15 is injected through burner 17 into a preheated furnace 19. The aeration ratio of fuel in the feed jet 15 is kept reasonably close to 100%. This, according to Figure 1, provides the most efficient fuel combustion. The reduction potential of the flue gases in the flame can be either zero when the aeration ratio is 100% or higher, or it can be a small value, such as about 0.05 to 0.10, which corresponds to an aeration ratio of almost 95%. In general, however, the reduction potential of the combustion gases is not greater or less than what it is for the equilibrium weight of the simultaneous presence of the metal and its oxide and gas phase, e.g. the Ni-NiO gas phase. Thus, the flame is substantially non-reducing with respect to the metal oxide and may rather be oxidizing to the corresponding metal.

The amount of fuel in the feed jet is determined by the overall heat balance of the reduction smelting process, as the heat generated in the combustion of fuel materials during the treatment of oxide materials is by far the most important energy source to be processed. si requires. In general, the amount of fuel must be such that • · 1 .. * the oxide particles in the flame can be heated »· * · *. to a temperature significantly higher than the melting point of the reduced molten metal. For example, in the case of nickel oxide 35 din, the amount of fuel must be sufficient to heat the nickel oxide particles in the flame to a temperature well above 1,450 ° C and approx. 1,600 to 1,800 Temperatures of ° C meet this condition. Such a temperature is easily reached, since the theoretical flame temperature obtained, for example, from the combustion of methane (natural gas) with a aeration ratio close to 100% (curve C in Fig. 1) is about 1000 ° C higher.

According to step (b), the superheated oxide particles are sprayed or sprinkled from the flame 21 onto a thin layer of granular coke 10 23 floating on the surface 1a of the molten bath. Injection is an essential feature of this invention. It is desirable that the injection cover the surface 27 of the thin layer of granular coke 23 as much as practicable and that the superheated particles be distributed to the pin-15 as evenly as possible. On the other hand, it is recommended that the walls 29 of the furnace 19 be clean of the syringe. In this way, the risk of either forming a layer of refractory material such as nickel oxide on the walls or dissolving the refractory portions of the furnace wall with a low melting point and corrosive material such as copper oxide is avoided.

The simplified diagram of Fig. 2B shows an apparatus comprising two burners 17 arranged • • approximately horizontally in such a way that the flames 21 of the burners: 25 do not interfere with each other. The superheated particles fall out of the flames 21 onto the coke layer 23 due to gravity, while the combustion gases are separated from the particles and directed upwards.

: Y: There are numerous ways to achieve uniform spraying and distribution of the superheated particles on the contact layer 23. For example, the burners 17 can be rocked * »♦ vertically and / or horizontally or they can perform any other type of movement with predetermined work and which are useful for the process. Furthermore, they 35 may be mounted on the furnace ceiling 31 instead of the end walls 29 or side walls 28878 28 and may be fixed or movable moving along the walls 29 or the roof 31. The furnace 19 itself may have any configuration, for example a circular burners 17 arranged to produce a vortex of the same type as that used in the Inred process or oven 19 can rotate in the same manner as a top-blowing rotary converter. Other configurations are possible.

All of these and other possible furnace and burner configurations are considered to be within the scope of this invention, provided that the process of this invention is performed therein.

The reduction of the oxides 11, step (c) takes place in a thin layer of slag coke 23 floating on the surface of the molten bath 25. The success of this step depends on achieving an equilibrium relationship between the rate of reduction, the rate of heat transfer to the coke bed, and the rate of sieving of the reduction metals through the coke bed.

The reduction rate depends on the nature of the non-ferrous metal oxides 11 20 and the size of the particles injected onto their coke layer 23. For example, nickel oxide is more difficult to reduce than copper oxide because nickel oxide and nickel metal have very high melting points (about 1,960 * * * Λ ··· *: and about 1,450 ° C in the same order), molten nickel • · · • · ·. * 25 has a relatively low porosity for oxygen (and carbon) and

The reduction potential of the Ni-NiO gas phase equilibrium is considerably higher than that of copper. In addition, the sieving rate of molten nickel * .j |: through the coke bed is lower than: *: * · for copper metal, because the surface tension of nickel is much higher than that of copper (approx. 19 mN / cm at 1 550 ° C and. ·, · 13 mN / cm at 1 100 ° C in the same order.

• ··

Ideally, the above equilibrium ratio • · · *. means that at the given reduction rate the heat transfer *: ** · and the sieving rate must be high enough so that ···:> # ^ 35 that the reduced metal is obtained in the molten state and it; ·· · f '• · · t- • · • · m »^

B

i il 21 85878 rapidly sieves below the coke layer 23. If this ratio is not in equilibrium, the granular carbon will be surrounded by a mixture of solid oxide and solid metal, resulting in the formation of a sinter that is characteristic of the Inred process and not of the process of this invention. As sinter is an obstacle to heat transfer, its formation overturns this process: the sinter layer gradually becomes thicker, the molten bath solidifies and as a result the whole process stops completely.

In this process, heat is transferred to the coke layer 23 mainly by means of superheated, condensed particles which are sprayed on it and by radiation from the flame and the walls of the furnace. If, in the nickel oxide example above, the coke layer 23 is maintained at about 1,500 ° C and the nickel oxide particles are heated in a flame to 1,600-1,800 ° C, the proportion of heat transferred to the reduction zone with the superheated solid nickel oxide particles is 60. - 70%. The remainder of the heat required for the endothermic process of the reduction of NiO and the smelting of the reduced nickel metal, as well as for the heating of the gaseous products of the reduction, enters the coke bed as radiation. According to the present invention, it is not desirable for the convection of the flue gases to play an important role in the heat transfer to the reduction zone, since this leads to excessive gasification of the coke bed 23 and worsens the conditions of the reduction process. For this reason .:. the gases are separated from the zone. By completing the process. with an aeration ratio close to 100% in the flame, the best fuel efficiency is obtained at the highest theoretical flame temperature, which in turn promotes the best possible heat transfer by radiation.

• · · * · * · The reduction rate naturally increases as the process is performed at a higher temperature. On the other hand, the reduction · · ** **: the speed can also be improved when the reduction smelters. ** ’. 35 materials consist of particles with a reasonably small size * * · Λ» 22 85378. The reduction rate is highest when the superheated condensed particles injected into the coke layer 23 are molten, as is the case, for example, with copper oxide (as opposed to nickel oxide).

5 Likewise, the proportion of heat transferred to the reduction zone with the superheated particles when they are molten increases up to about 80-95% depending on the flame temperature.

When the heat transfer is sufficient, the reduction and melting rate may become higher than the sieving rate.

10 In this case, sieving can become a limiting factor in the whole process. It was found that the sieving rate is accelerated in the presence of surfactants, including sulfur and oxygen. Very low concentrations of these substances dramatically reduce the surface tension of non-ferrous metals and 15 alloys and therefore contribute to an improvement in the sieving rate.

As a result of the reduction of the non-ferrous metal oxides (step (c) above), some coke is consumed and the gas phase consisting of CCl 4 and CO is discharged from the coke bed. To replenish the granular coke consumed by the reduction reactions, solid granular carbonaceous material is fed to the reduction zone. This can be done in a number of ways and two possible alternatives are illustrated in Figure 2.

": 25 One of the alternatives is to feed carbonaceous material 33 through the furnace roof 31 to the feed jets (lie kit 21). Another way is to feed this material with the feed mixture through the burners 17. In both cases .1 the carbonaceous material 33 is fairly evenly distributed. Since this material has a particle size between about 5 and 25 mm according to the present invention, only a small portion of it burns as it passes through the flame 21 and falls into the coke layer 23. This amount of combustion can be taken into account by adjusting the oxygen 15 and fuel 13. dictates that the aeration ratio in the flame remains at the level required by the process.

i, 23 85878

The carbon monoxide discharged from the coke bed 23 can be post-combusted in the furnace 19. This can be done either by supplying secondary oxygen 3 and / or air as shown in Figure 2 or by adjusting the fuel aeration ratio to more than 100% as required to complete carbon monoxide post-combustion. Afterburning contributes to the overall heat balance, leading to a further reduction in fuel consumption.

Example 1 A nickel-copper alloy was prepared using a mixture of industrially produced oxides of nickel and copper. The mixture contained 35% by weight: 35.1 Ni, 39.3 Cu, 3.9 Fe and 0.5 Co. The particle size distribution of the mixture was as follows: size, mm +212 -212 + 150 -150 + 75 -75 + 38 -38 15 wt% '0 26 T6 38 20 This mixture was continuously injected with natural gas and oxygen through a water-cooled burner into a preheated furnace to the desired temperature. During the experiments, the oven operated autogenously. To achieve this, it was heated externally, but only to the extent necessary to prevent (compensate) for heat loss from the furnace through its walls, bottom and roof. Thus, the furnace was substantially adiabatic and therefore the heat required for the reduction-smelting process performed therein was completely quenched inside the furnace by burning the mixture of fuel and oxygen with which the oxide mixture was injected.

The feed was sprayed onto the top of the furnace chamber, which served as a combustion zone. The lower part of the furnace below the combustion zone contained a receiving crucible. The hot particles from the oxide feed fell downwards from the feed jet (flame) to collect in the receiving crucible, while the hot combustion gases (waste) were removed from the furnace space through a vertical duct in the furnace roof.

The production of the nickel-copper alloy was carried out with a '35 95% fuel aeration ratio, which corresponds to a reduction potential of 24 85878 flue gases of about 0.07 (see page 1 of the drawing). This reduction potential is lower than that required for nickel oxide reduction at the temperatures used in this experiment (see below) and therefore nickel oxide reduction was not thermodynamically possible in the flame. On the other hand, copper-copper oxide, which was approx.

4 9% by weight of the oxide mixture, is thermodynamic unstable at high temperature. It dissociates in the air atmosphere at about 1 020 ° C as follows: 10 2 CuO CU2O + 0.5 C> 2

The pressure of the dissociating oxygen at flame temperatures well above 1,400 ° C is several orders of magnitude higher than 100 kPa. Oxygen from one of the dissociation reactions shown, which takes place at a high temperature in virtually zero time, is considered to be directly involved in the combustion process and therefore this oxygen was added to the gaseous oxygen when calculating and setting the above 95% aeration.

Thus, in this operation, the feeds of natural gas and gaseous oxygen were set to about 7.0 and 20% by weight of the solid feed, respectively, and the solids feed rate was 9.7 kg / h.

Before starting the run, a 4.6 kg nickel-copper-25 alloy base alloy was pre-melted from the receiving crucible using an approximately 3 cm thick layer of crushed coal (reducing agent) on its surface. During the experiment, the thickness of the coke layer was kept at about 2 to 4 feet by adding coke through the furnace roof.

The solid feed contained the addition of sulfur-containing material 30 in an amount of 7.5% by weight of the oxide mixture to ensure a good rate of sieving of the metal product through the coke. The composition of the sulfur-containing material was by weight:

Cu Ri Co Fe S SO 2 • 35 15.6 56.2 0.98 2.6 16.0 1.76 25 85878

The coal crush used in the run had the following composition in weight percent:

C C

total bound S Volatile Ash H 2 O 89.3 89.2 0.47 2.42 8.11 0.28 5 The particle size distribution was as follows: size, mm -13.5 + 4.75 -4.75 + 2.36 -2.36 +1.7 -1.7 wt% 28.9 48.6 12.3 10.2

Coke ash contained, by weight:

Fe SiO2 Al2®3 MgO CaO

10 7.4 54.5 25.6 5.1 1.9 This ash is a highly refractory material that must be made flowable to avoid its accumulation and the associated slowing of the sieving of the metal through the coke. This was achieved by forming an easily digestible, low viscosity slag. Thus, a flux containing 75% by weight of CaO and 25% of CaF2 by weight was added to the solid feed in an amount corresponding to 65% by weight of ash SiO 2 + A 2 ° 3 + + CaO 2 P% of the coke used during the reduction. The resulting slurry was estimated to have the following composition in weight percent: 34% CaO +

MgO, 3.8% SiO 2, 18% and 10% CaF 2. This slag has a liquefaction temperature below 1,250 ° C and a viscosity below - · - 0.2 Pa. s at 1,400 ° C. The liquefaction temperature and viscosity of this slag decrease substantially as the iron content of the slag in the form of FeO increases.

The solid feed described above was fed into the oven for 1.5 hours. The above-mentioned feed rate 2 of 9.7 kg / h corresponded to about 20 t / m per day, calculated from the cross-sectional area of the coke layer inside the receiving crucible.

During the experiment, the temperature of the waste gas was 1,470 to 1,480 ° C, the temperature above the receiving crucible was 1,480 to 1,500 ° C, and the temperature inside the bottom of the crucible was n.

1440 ° C.

26 85878

At the end of the run, pin samples were taken from the liquid reduced metal. The metal contained by weight:

Cu Ni Co Fe S O C

5 46.9 49.0 0.078 2.67 0.67 0.03 0.02 Thus, it was shown that substantially complete reduction was achieved using the method of this invention. In this experiment, the consumption of coke was about 9% by weight of the oxide feed. This consumption represents an excess of approximately 25% 10 compared to the amount required stoichiometrically for the reduction of the oxides. The waste gases contained low concentrations of H2 and CO with a reduction potential between only 0.10 and 0.15.

Example 2 This example demonstrates that the process of this invention can be successfully performed with a fuel aeration ratio of just over 100%.

In this operation, the solid feed had exactly the same composition as in Example 1, but the aeration ratio of the fuel 20 was 118%, i.e. neither nickel nor copper metal could be formed in the flame itself because it contained free oxygen. The feeds of natural gas and gaseous oxygen were set to about 7 and 26% by weight of the solid feed, respectively, and the feed rate of the solid was 25 10 kg / h or 21 t / m per day. The same coal crush was used as a reducing agent and the thickness of the coke bed was kept at n.

3-5 in the same manner as in Example 1. During the experiment, the temperature of the waste gas was about 1,500 ° C, the temperature above the receiving crucible was about 1,510 ° C and the temperature inside the bottom of the crucible was about 1,425 ° C.

. . At the end of the run, a pin sample was taken from the liquid reduced metal. The metal contained by weight:

Cu Ni Fe S O C

44.3 50.2 3.95 0.73 0.04 0.07 i: '35 27 8 5 8 78 In this experiment, the consumption of coke was about 11% by weight of the oxide feed and the reduction potential of the waste gas was between 0.01 and 0, 05. It is evident that in this experiment there was a significant post-combustion of CO rising from the coke bed, which resulted in a much lower waste gas reduction potential as well as a low waste gas temperature (combustion zone) and a temperature immediately above the receiving crucible (reduction zone). 1.

10 Comparative experiments

In contrast to the good results obtained in Examples 1 and 2, very poor results were obtained when flame reduction was attempted using aeration ratios of about 65% and 55% fuel. Aeration of 65% fuel gave 15 non-flowable products, although the surface temperature of the product was 1,440-1,450 ° C. Dust and immersion products were recovered by the following analyzes as weight percentages:

Ni Cu Co Fe O

Crucible product 45.9 42.0 0.59 2.36 9.51 20 Dust 9.07 76.8 0.13 8.70 4.32 XRD analysis showed that the crucible product was mainly: a mixture of copper-nickel alloy and nickel oxide; while the dust was a mixture of copper-nickel alloy, copper-copper oxide, and ferrite. Total reduction of the mixed oxide feed. The 25 degree was only 66%.

A second flame reduction experiment was then performed n.

With a fuel aeration ratio of 55%. The required fuel consumption with this aeration was calculated to be 22% methane by weight of the solid feed. The solids feed rate 30 was 8.8 kg / h while all other essential parameters remained essentially the same as in the previous flame reduction experiment.

At the end of this experiment, suction pin samples could not be obtained because the contents of the crucible were again not flowable.

35 Crucible product sampled by drilling and dust 28 85878 contained 5.0 and 4.5% oxygen, respectively. The overall reduction rate was 75%. XRD analysis showed that in addition to the copper-nickel alloy, the crucible product contained nickel oxide and the dust contained copper-copper oxide and fer-5 rite, i.e., the phase composition of the final materials obtained in the second experiment was the same as in the first.

Thus, it was experimentally shown that even with an aeration ratio of only 55%, the reduction of nickel and copper oxide could not be completed in the flame. In addition, at a 65% and 55% aeration ratios, the fuel consumption required to maintain a sufficient process temperature is 2.3 and 3.7 times higher, respectively, than the fuel consumption at a 100% aeration ratio. It could perhaps be said that with a much finer oxide particle size of 15 and an aeration ratio of about 45%, a much better degree of flame reduction could be achieved, as the reduction potential with this aeration would be almost double that with 55% aeration (Figure 1). However, the fuel consumption was then about 9 times the consumption of fuel 20 at a 100% aeration ratio, i.e. about 56% by weight of the oxide feed used in the example above. In addition, a huge volume of waste gases per unit weight of solid feed would be produced, resulting in a very large amount of dust. Finally, the exhaust gases would retain a predominant part of the chemical energy of the spent fuel 25 and this would make it necessary to use some system for afterburning and waste heat recovery. In this case, the flame reduction smelting process would become even less economically attractive for the production of non-ferrous metals.

Another embodiment of the present invention consists in applying a reduction smelting process to a material containing non-ferrous parent metals and iron - all primarily in the oxide form. The main goal of such a process is to reduce and recover non-ferrous metals to a molten phase that is highly enriched for these metals, while iron oxides are selectively reduced to the degree of reduction required to form a discardable molten iron-containing slag. One example of such a material is the product of the oxidation roasting of copper calcine - sulfine-5 dicopper concentrate. In this case, impure copper and iron silicate slag (or iron calcareous slag) are the end products of the process.

Those skilled in the art of extraction metallurgy are well aware of the dramatic difference between the thermodynamic conditions required for the reduction of free copper oxide and the conditions required for the reduction of copper oxides from iron silicate slags to ensure that the slags can be discarded. The free copper oxide is reduced to the copper metal with CO-CO 2 or H 2 O alloys containing very low concentrations of CO or H 2. For example, the gas phase in equilibrium with the Cu 2 O-Cu mixture at 1,250 ° C contains only about 0.005% v / v CO. On the other hand, in order to obtain an iron silicate slag containing 25% SiO 2 and containing 2.5% Cu, it is required that the slag be equilibrated with a mixture of CO 2 -CO 2 containing about 9.5% by volume of CO, which corresponds to CO 2 / CO 2. ~ ratio of about 0.1. The same slag at the same temperature contains 1% Cu with a C0 / CO2 ratio of about 0.75 and an additional increase in the · C0 / CO2 ratio of up to about 2.0 reduces the copper content of the slag to only about 0.6%.

. The dependence shown above between the copper content of the slag and the reduction potential (C0 / CO2 ~ ratio) of the gas phase is key to the reduction smelting process which results in the direct production of copper metal simultaneously with the production of discardable slag in the same furnace. The dependence shown in Fig. 1 30 and the above description of this dependence applied to the prior art as well as the experimental results presented in the previous examples show that the simultaneous production of crude copper and slag is not economically and technically feasible in a process designed to be carried out simultaneously. heating and reduction in flame. In particular, it makes no sense at all to produce crude copper and discardable slag containing, for example, 0.6% Cu, in a process similar to that discussed in U.S. Patent 4,421,552.

In contrast, in the process of the present invention, in which heating and reduction are separated in both time and place, the goal of producing crude copper and discardable slag in the same furnace is achieved in a simple and economical manner because both heating is performed under conditions best suited to each of these conditions. two steps of the proposed process of the invention. More specifically, the heating is performed in a substantially non-reducing high temperature flame, while the reduction is performed in a thin layer of granular coke. This is illustrated in the following example.

Example 3

The mixture of copper calcine and silicon-containing flux was continuously injected with natural gas and oxygen in the same manner and in a similar furnace as in the previous examples. The compositions by weight of these materials were as follows:

Cu Ni Fe S SiO2 Al2-3 £ a0

Calcine 33.0 1.11 34.3 0.35 2.50 0.94 0.94 0.50 V'25 Fluxes - - 0.07 - 99.2 0.43 0.01 0.01

The amount of flux added was 15% by weight of calcine.

This run was performed with a 100% fuel aeration ratio. Therefore, the natural gas and oxygen feed 30 was set to 6.2 and 23.2 wt% of the mixed solid feed. The solids feed rate was 9.9 kg / h or 2 21 t / m per day. The same coke breeze was used and fed to the furnace in the same manner as described in Examples 1 and 2. No special flux was used to slag the coke ... .35 ash.

tl 31 85878

During the run, the temperature of the waste gas was 1,430 to 1440 ° C, the temperature above the receiving crucible was 1,440 to 1,450 ° C and the temperature inside the bottom of the crucible was about 1,200 ° C.

After 1/5 hour of feeding, the run was stopped and the su-5 list of products was sampled. According to the analysis, they contained in weight percentages:

Cu Ni Fe S SiO 2 O 2 O 3 MgO Fe 2 O 2

Crude rubber 95.7 2.67 0.82 0.22

Depreciation slag 0.47 0.03 49.3 0.06 28.0 2.65 3.81 6.2 10

Numerous other runs were performed under similar conditions using different granular carbonaceous materials such as petroleum coke, coal, anthracite coke and varying the proportion of siliceous flux, as well as using 15 calcareous fluxes. Copper calcine with up to 1.9 wt% S was also used. At all these times, remarkably uniform results were obtained with a copper content of 0.4 to 0.9% in the slag. Consumption of carbonaceous materials (including coke breeze) increased from 5.6 to 8.0% by weight of calcine and the reduction potential of the waste gases ranged from 0.07 to 0.22. The latter values are many times lower than the equilibrium ratio CO / CO 2 n. 2, which is • required to calculate the copper content of the slag to n.

0.6%. This fact ensures that in the process of the present invention, the conditions in the reduction zone (thin coke bed) and on the back edge of the furnace (combustion zone) differ very substantially and favorably from that, so that the reduction and combustion can take place in the most efficient ways.

The following example illustrates the application of this invention. 30 to the reduction smelting of nickel calcine, which is a product containing a partial oxidation roasting of a sulfide-dinickel concentrate. Unlike copper calcine, nickel calcine contains significant amounts of sulfur, usually n.

5-15 wt%. Most of this sulfur is from sulfide, but 35 small amounts of sulfate sulfur may also be present in calcine. The majority of calcine iron is in the form of a solid solution of hematite and magnetite-trevorite, while a substantial portion of nickel still remains bound to sulfide-sulfur. The main goal of the reduction smelting-5 process applied to nickel calcine is to reduce and enrich nickel and other non-ferrous metals such as copper and especially cobalt in metal rock and to reduce iron oxides to wustite as required to produce discardable ferrous silicate slag. This object is best achieved by producing a sulfide-metal alloy containing nickel, copper, and iron (a low-sulfur metal rock) as recommended in SU Patent No. 383,753 and U.S. Patent No. 4,344,792.

Example 4 A mixture of nickel calcine and silicon-containing flux was continuously melted with natural gas and oxygen using the same equipment and procedure as described in the previous examples. Calcine had the following composition in weight percent:

20 Cu Ni Co Fe Stokona.sS042- S102 Al203 CaO MaO

3.10 15.8 0.51 40.6 12.2 1.16 6.0 1.5 0.97 0.99

About 75% of the iron in the calcine was in the trivalent oxidation state. The composition of the flux was the same as shown in Example 3 ': 25 and the addition amount was 10.5% by weight of calcine. To protect calcine sulfur from oxidation, the experiment was performed at a fuel aeration ratio of about 98% using natural. . non-gas and oxygen feed rates of 9 and 33 wt% calcine, respectively. Solid supply 2. .30 speed was 8.9 kg / h or 19 t / m per day. The same solid reducing agent as in the previous examples was used and its layer was kept 2-3 cm thick.

During operation, the temperature of the waste gas was about 1 360 ° C, the temperature above the receiving crucible was about 1 375 ° C and. ‘* .35 the temperature inside the bottom of the crucible was about 1,200 ° C.

(33 85878 After 1.5 hours of feeding, the run was stopped and the molten metal rock and slag were sampled. According to the analysis, they contained in weight percent:

Cu Ni Co Fe S SiO- Al-, 0 ·, Mgo Fe, 0. 5 Metal stone 7.77 42.1 1.0 25.3 22.9 -

Slag 0.17 0.19 0.10 45.0 1.47 28.7 3.78 3.65 6.5 These analyzes show excellent reduction, enrichment and recovery of non-ferrous metals in very high grade low sulfur metal (Cu + Ni + Co = 51 wt%). Coke consumption was only 2.4 wt% of the solid feed and the waste gas reduction potential was only 0.19.

For comparison, the experiment was performed by flame reduction melting of nickel calcine according to the prior art. This test is described as follows.

Comparative Test

The same mixture of nickel calcine and silicon-containing flux as in Example 4 was melted with natural gas and oxygen using the same equipment as in the previous pre-20 fractions, but without using a coke bed. The experiment was performed at a fuel aeration ratio of 62% using natural gas and oxygen feeds of 20 and 45 wt% calcine, respectively, and no solid reducing agent was used.

The solids feed rate and temperatures were similar to those reported in Example 4.

After 1.5 hours of feeding, the run was stopped and the molten metal rock and slag were sampled. According to the analysis, they contained in weight percentages: · ·: 30

Cu Ni Co Fe S Si ° 2 -2-3 Mg0 Fe3 ° 4

Metal stone 9.65 42.5 1.1 22.9 22.4 - -

Slag 0.25 0.4 0.2 47.7 1.35 27.0 4.0 3.31 9.9 In this experiment, the fuel and oxygen feeds were 35 2.2 and 1.4 times higher than in Example 4. 34 85878, the reduction potential of the waste gas was 0.77, i.e. 4 times higher than in Example 4. In any case, the reduction and recovery of non-ferrous metals were significantly worse because the nickel and cobalt contents of the slag were 5 2 times higher than in Example 4.

The present invention has so far been described in this application with respect to materials containing non-ferrous parent metals in the chemical form of oxides (a mixture of copper and nickel oxides and copper paricalcine) or oxides plus sulfides (nickel-10 calcine) in the presence of iron. mainly in oxide form (copper and nickel calcins). Reduction smelting of these types of materials can only be performed using external fuel.

On the opposite side of the scope of the present invention, the reduction smelting process is applied to finely divided sulfide materials, such as sulfide ores, tailings, metal rocks, etc., which can be autogenously smelted with oxygen, i.e. without using any external fuel. With these materials, the formation of metal oxides, which are to be reduced in the thin layer of granular coke in the reduction zone, results from the oxidation (combustion) of sulfide materials with oxygen. The highly exothermic oxidation of the sulfides in the combustion zone generates virtually all the heat required for the entire melting process. If necessary, some additional heat may be generated by post-combustion of carbon monoxide rising from the reduction zone. The concept of fuel aeration ratio is no longer suitable for use in the combustion of sulphides, as the desulphurisation metallurgically desired is to be achieved. 30 degrees and oxygen utilization efficiency determine the weight fraction of oxygen used relative to the weight of sulfide materials in a given mineral mixture.

Those skilled in the art of extraction metallurgy are aware that the desired degree of desulfurization depends on the nature of the sulphide materials to be smelted and the compositional requirements of the final reduced products. Particularly in the case of copper sulphide concentrates, metal stones and white metal (Cv 2 S), it is metallurgically possible and desirable to oxidize and eliminate virtually all sulfur in the form of SC 1 so that the only copper product subsequently obtained in the reduction zone is copper metal. A high degree of desulfurization is also metallurgically possible and desirable in accordance with this invention in the case of processing a sulfidine-nickel concentrate obtained from the separation of copper-10 nickel converter metal by slow cooling and flotation techniques, as well as nickel or nickel cobalt sulfide concentrates precipitated from aqueous salts.

On the other hand, in the case of nickel sulfide-rich, which contains significant amounts of iron, it is desirable to produce a metal rock and discard as much iron as possible. Therefore, the proportion of sulfide sulfur to be oxidized and removed is controlled by the desired metal rock quality and the desired metal-rock sulfur content. The quality and sulfur content of the metal rock are the 20 most important parameters that affect the losses of non-ferrous metals, especially cobalt, with the effluent slags generated in the reduction zone, as well as the liquefaction temperatures of the metal rock. In addition, there may be a number of metallurgical situations involving the processing of complex polymetallic sulphide materials or monometallic materials containing copper and nickel or copper and lead or copper, lead and zinc, etc., which situations may require either complete or partial desulphurisation. autogenic process. 30 goals. For example, a copper-nickel alloy may be desired as the only non-ferrous metal product in the reduction smelting.

Oxygen utilization efficiency (not oxygen efficiency) is the ratio of the amount of oxygen required stoichiometrically to achieve the desired degree of desulfurization, 35 to the amount actually used to achieve this desulfurization. Usually, the oxygen efficiency is quite close to 100% when producing a relatively low quality metal rock, i.e. when the desired desulfurization is well below 100%. In this case, the fuel gases consist almost exclusively of pigs and do not contain significant concentrations of free oxygen.

When the desired degree of desulfurization approaches 100%, the oxygen efficiency ratio can be expected to be less than 100% and therefore the combustion gases may contain some concentration of free oxygen.

In this case, it is obvious that the amount of heat generated in the combustion zone during the combustion of the given sulfide material depends on the degree of desulfurization and therefore there is a fixed dependence between the heat generated and the sulfur-15 injection achieved. In addition, the amount of heat generated at a given degree of desulfurization also depends on the nature of the sulfide materials to be combusted and the nature of the combustion products containing the metals produced.

When treating sulfide materials that are substantially free of iron, the combustion process is controlled in such a way that the metal-containing combustion products produced in the flame are mainly represented by metals, exemplified by the following chemical equations:

Cu2S + O2 -> 2 Cu + SO2 ...... (1) ;; 25 Ni3S2 + 2 ° 2 -> 3 Ni + 2 SO2 ...... (2)

However, it was found that the use of stoichiometric amounts of oxygen according to these reactions does not necessarily lead to the obtaining of metals with a sufficiently low sulfur content. : 30 contents.

--- In order to obtain a sufficiently low sulfur content in metals obtained from the combustion of sulphide materials with oxygen, it is preferred in accordance with the present invention to use oxygen in quantities slightly above the stoichiometric requirements, such as those corresponding to equations 1 and 2.

37 8 5 8 7 8 In this case, relatively small amounts of metal oxides are formed. Ideally, these amounts would be close to or slightly above those corresponding to the solubility limits of the given metal oxide (oxygen).

5 Under these conditions, the metals obtained from the combustion of sulphide materials with oxygen are either saturated or slightly over-saturated with oxygen and therefore contain little sulfur. In practice, however, it is difficult to control the process so that the oxygen content of the metal is exactly 10, which corresponds to the oxygen saturation. In addition, the oxygen saturation of the metals at high flame temperatures corresponds to a substantially higher oxygen content of the metals than at lower temperatures in the reduction zone, and this results in the separation of oxygen from the metals in the form of metal oxides. Therefore, it is quite advantageous to use such an excess of oxygen over the stoichiometric requirements of the reactants (1-2) that the metals produced in the flame are saturated with oxygen and a small part of them is obtained in the form of metal oxides. In addition, in the combustion of sulfide materials, oxide is always present in the material entering the furnace because of the economic and environmental requirement that oxide-containing smoke dust be recycled to the furnace. Dissolved oxygen and oxygen from metal oxides are then eliminated in the reduction zone, resulting in final metals with low concentrations of both sulfur and oxygen: -'25. This application of the present invention is particularly advantageous when the copper sulfide material contains a significant amount of nickel, as nickel tends to oxidize primarily to produce a soft mass of nickel oxide.

When treating sulfide materials containing-. Non-ferrous base metals and iron, the amount of heat that can be generated by oxygen combustion to achieve the desired degree of metallurgical desulphurisation usually far exceeds the process heat balance requirements and the excess heat can become even higher due to the afterburning of carbon monoxide ... 35 in the reduction zone. 7 8 gases. Thus, the thermal equilibrium requirements conflict with the metallurgical objectives of producing final reduced products with the desired composition.

This problem of excess heat is solved simply and efficiently by roasting some of such sulfide materials and mixing the resulting low-sulfur calcines with unroasted sulfide materials to obtain a mixed feed for autogenous smelting, as previously described in U.S. Patent No. 4,415,356.

This experienced technique provides for the control of the thermal power of the mixed feed to any required value by autogenous thermal equilibrium and therefore makes any desulfurization achievable. The use of this technique within the scope of this invention provides the autogenic process with some additional benefits, as it improves heat transfer from the combustion zone to the reduction zone and reduces the amount of reduction work to be performed.

The improvement in heat transfer is mainly due to the fact that the superheated condensed particles formed in the flames as a result of the combustion of the sulfide portion of the mixed feed and the mixed calcine particles heated in the flame are all rare excipients of some rare nickel- and cobalt-containing oxides. All of these molten particles, when superheated in a flame of about 200 to 400 ° C above the required temperature of the final products, can carry all the heat that must be made available in the reduction zone and therefore the need for heat transfer by radiation. 30 decreases to a minimum. The reduction in the amount of reduction work carried out in the reduction zone compared to, for example, the case of copper-calcine reduction smelting (Example 3 above) * ···: occurs mainly due to two factors. Another of them is that the majority of the pairs of the sulphide part of the mixed feed ... 35 are converted to copper metal back in the flame during combustion. Another factor is that under high temperature and well below 10 kPa oxygen partial pressure, iron oxides with a lower oxidation state are formed in the sulphide part of the mixed feed, while the ferric iron and copper copper oxides of the uncharged calcine part are partially dissociated because they are not thermodynamically stable under these conditions. As a result, less oxygen enters the reduction zone with the superheated particles of the products of the combustion process, and therefore less reduction work is required in the reduction zone to produce crude copper and slag. Similar considerations apply equally to nickel, copper-nickel, and other sulfide materials.

As previously shown, convection of flue gases into the reduction zone is undesirable because it would lead to excessive gasification of the coke and poor conditions of the reduction process. When the autogenous process of the present invention is present, this requirement to prevent flue gas and intrusion into the reduction zone becomes increasingly important because the infiltration of SO 2 -containing gases leads not only to excessive gasification of coke but also to sulfidation of non-ferrous metals as well as iron. Indeed, the conditions in the reduction zone are thermodynamically and kinetically very favorable for the occurrence of sulfidation with SO 2 as the sulfur source. This image is followed by an example.

Example 5

The copper sulfide concentrate and copper calcine used in this experiment had the following compositions at 30% by weight:

Cu Ni Fe S.. , SiO_ - - - —total -2

The rich 28.1 0.66 31.4 33.4 1.45

Calcine 34.3 1.14 35.0 0.45 2.70 ίο 85878

The concentrate and calcine were mixed together and the siliceous flux in a weight ratio of 100:20: 15 in the same order, and the mixture was continuously injected with oxygen into the same adiabatic furnace used above as used in the previous examples. The solids feed rate was 3.5 kg / h and the oxygen feed was 46.7 wt% feed.

During the run, the temperature of the waste gas was about 1,420 ° C and the temperature above the receiving crucible was about 1,480 ° C.

Anthracite carbon was charged to the receiving crucible in an amount of 11% by weight of the copper feed (copper concentrate plus calcine). About 64% of this amount was consumed during the experiment, with the remainder accumulating in the receiving crucible on the surface of the molten bath. Despite the presence of a reducing agent, the waste gas contained 5-7% by volume of free oxygen in addition to sulfur and carbon dioxide-15.

After the run was stopped, the following molten products were found in the crucible: metal, metal rock, and slag. According to the analysis, they contained in weight percentages:

Cu Ni Fe S SiO 2 -3-4 20 Metal 80.2 3.3 10.9 3.9

Metal stone 51.5 1.7 23.0 20.5

Slag 1.1 0.1 34.5 0.65 32.0 4.1

The copper composite material containing metal: 25 plus metal rock contained in weight percentages: 6.20 Cu, 2.3 Ni, 18.6 Fe and 14.4 S. Approximately 25% of the sulfur in the feed appeared in this composite and therefore the degree of desulfurization was n. 75%. However, under exactly the same conditions, but without the use of a reducing agent only crude copper and in abundance. , 30 silicate slags containing copper and iron, impregnated with a magnet, were routinely produced in the same furnace with a degree of desulfurization close to 100%.

* · * ·: It is therefore obvious that the above car; in a genetic smelting run using a reducing agent, about 25% of the 35 sulfur fed penetrated downstream from the combustion zone into the reduction zone in the form of sulfur dioxide. This sulfur dioxide was reduced by the action of carbon monoxide and carbon to elemental sulfur, which in turn reacted with the copper metal formed in the combustion zone. The overall result of these side reactions is the formation of copper sulfide, i.

back sulfidation. In addition, the presence of sulfur and / or sulfide sulfur in the coke bed improved iron reduction and back sulfation. As a result, sulfur-saturated metal and metal-10 stone were produced instead of crude copper alone, as shown in the last table above.

On the other hand, Example 5 clearly shows that even in the used furnace, which was not particularly suitable for carrying out the autogenous process of this invention, excellent reduction of iron oxides occurred, resulting in iron silicate slag with very low magnetite and copper content despite free oxygen in the waste gas. . At the same time, the carbon consumption was substantially higher than what is stoichiometrically required to produce a highly reduced slag having the above composition. All these facts show that n.

25% of the volume of combustion gases penetrated the reduction zone, causing back sulphation and excessive wear of the carbon (Gasification).

-. In the autogenous process of the present invention, the volume of reaction gases rising from the reduction zone is small. For example, in the case of a sulfide-copper concentrate, it is only about 10% of the volume of the combustion gases as opposed to about 30 and 60% in the case of the reduction smelting of copper calcine and non-ferrous metal oxides in the same order using external fuel. Reaction gases V; the relatively small volume is an important reason for the penetration of sulfur-containing combustion gases into the reduction zone, but. ···. the combination of the relatively small volume of reaction gases, the current geometric shape of the furnace and the trajectory of the feed jets (flames) must be addressed / to prevent re-sulphidation of non-ferrous metals and possible iron as well as excessive carbonification of the carbon.

This is achieved in accordance with the present invention by providing a non-reducing and substantially sulfur-free intermediate zone between the reduction zone and the combustion zone. From below, this intermediate zone gradually changes to a reduction zone, while from above it joins the combustion zone.

In this way, the atmospheric reduction / oxidation potential distribution of the furnace in the direction of the furnace height is realized.

As the intermediate zone gradually changes to a reduction zone, the reduction potential gradually changes from a value close to zero to a value approaching one (Fig. 1).

On the other hand, as the intermediate zone gradually becomes a combustion zone, virtually no gaseous reducing agent is present. Instead, free oxygen may be available in the furnace atmosphere in addition to sulfur and carbon dioxide, as was the case in Example 5.

20 One or more combinations of practical means may be used to form an intermediate zone. First, an oxygen-containing gas such as air or oxygen-enriched air, cold or preheated, can be injected into this intermediate zone to burn off carbon monoxide rising from the reduction zone * •• 25 as well as to increase the volume of the reaction gases. ·. promoting the prevention of intrusion into the flue gas reduction zone. In addition, a solid granular carbonaceous material intended to replace that consumed in the reduction. .30 tu coke, can be injected together with an oxygen-containing gas. In this case, only a small portion of the coarse carbonaceous material has time to burn off * ·· *: with oxygen before it falls into the reduction zone.

♦ "* · The total amount of oxygen fed to the intermediate zone must be * · # 35 at least equal to that required for any Λ • · * · · i! 43 85878 carbon monoxide afterburning whether it rises from the reduction zone or is the result of partial combustion of coarse carbonaceous material.

The above measures are improved if the geometry of the furnace 5 is somewhat stretched vertically, which provides better conditions for the distribution of the reduction / oxidation potential of the furnace atmosphere with respect to the height of the furnace. In addition, the auxiliary burners for injecting oxygen-containing gas into the intermediate zone are inclined upwardly, about 10 to 15 ° from the horizontal, and the main burners, as shown in Fig. 2, are positioned so that the jet momentum becomes negligible before the jets reach the intermediate zone.

The present invention is exemplified by two opposite extreme cases in its field of application, namely the case where all the thermal requirements of the process are met using external fuel and the autogenic process where no external fuel is required. There are numerous technological 20 situations between these two extreme cases.

An example of such a situation concerns a high grade copper concentrate containing the predominant proportion of copper in the form of chalcosite (C ^ S) and also containing a significant proportion of refractory rock material which may contain varying concentrations of silica, alumina, potassium oxide, magnesium oxide and iron oxides. in various mineralogical forms. One goal for processing such a concentrate is to produce crude copper and slag. Usually this requires. .30 addition of flux, as it is rare for a rock material to be converted to slag with acceptable physical properties without being properly treated with flux. Under these conditions, even complete desulfurization of the concentrate by oxygen combustion may not be able to generate as much heat as required.

44 85878 According to the present invention, there are two alternative ways to achieve the goal of producing crude copper and slag from a concentrate having the above properties. Another way is to mix this concentrate with another sulfide material so that the mixture can be melted autogenously, as shown by the examples above. An alternative way is to supplement the sulphides by oxidizing the heat generated by (combustion) with the heat of combustion of the fuel added to the feed. In the latter case, the amount of oxygen relative to the added fuel is always much higher than that corresponding to 100% of the fuel aeration ratio and the excess oxygen is used to burn the sulfides.

Another example of a situation where not all of the required heat can be generated by burning metal sulfides alone is in the processing of materials consisting of a mixture of finely divided non-ferrous elemental and sulfide forms plus recycled dust when only partial desulfurization is desired. This is the case with the intermediate separation product obtained by slow cooling, grinding, flotation and magnetic separation of a copper-nickel converter-20 termetal stone. This intermediate typically contains by weight: 10 to 20 Cu, 55 to 65 Ni, 1 to 3 Co, 1 to 3 Fe, 14 to 18 S, and significant amounts of noble metals. Mineralogically, this material: ‘: 25 is mainly represented by nickel-copper alloy metals and copper and nickel sulfides. The processing of this intermediate. . ·. metallurgical goal is to convert it into the feed most suitable for carbonylation. The feed is most suitable for carbonylation when it contains significantly less sulfur and .30 is melted and then quickly quenched by granulation. During carbonylation, most of the nickel in the feed is separated from copper, cobalt, sulfur and precious metals. Today, this feed is produced in a top-blown rotary converter in several successive stages, such as .35 smelting, partial desulphurisation (top-down

II

45 85878 blown oxidation) in the reduction of nickel oxide formed during desulphurisation and finally in granulation.

According to the present invention, the above intermediate and fuel additive are injected through the burner into the preheated combustion zone 5 with oxygen at a weight ratio of oxygen to fuel substantially greater than that corresponding to 100% aeration, adjusting the excess oxygen sufficient to provide the desired degree of desulfurization. The combustion process in this case involves three main features: (a) complete combustion of the fuel, (b) partial desulfurization, and (c) partial oxidation of the nickel-copper alloy metal.

The superheated condensed products 15 from the combustion process are separated from the combustion gases and fall down into the reduction zone to a thin layer of granular coke floating on the surface of the molten bath of the final product produced or previously preloaded. All phenomena occurring in the reduction zone have been described previously. The final 20 molten product is then granulated. In this variation of the present invention, the solid granular carbonaceous material is fed to the reduction zone separately, i.e., not through a feed burner.

In numerous experiments with an intermediate product having the above composition, it was found that the degree of desulfurization can be easily adjusted by the value of the excess oxygen in the combustion gases. For example, when the free oxygen content of the combustion gases (after complete combustion of the required fuel addition) was increased from 9% to 35% by volume, the desulfurization rate increased from 33% to 62%. It was also found that when a solid reducing agent was not used, it was not possible to achieve both the desired degree of desulfurization and a homogeneous melt due to the formation of significant amounts of soft pulp. This soft pulp typically contained 35 weight percent: 7 to 16 Cu, 50 to 65 Ni, 1 to 3 Co, 3 to 10 Fe, 4 to 6 S, and 8 to 15 oxygen.

46 85878

An alternative way to prepare the feed for carbonylation is to mix nickel oxide / readily available to a copper-nickel converter metal producer with the above intermediate and melt the mixture in the same manner as in the nickel calcine reduction smelting as described in Example 4. If some desulfurization is still required, ratio can be adjusted accordingly.

Although a specific embodiment of the invention has been illustrated and illustrated herein in accordance with the terms of the Regulations, those skilled in the art will appreciate that changes may be made to the form of the invention covered by the claims and that certain features of the invention may sometimes be utilized without others.

• I t il

Claims (10)

A process for reducing melting of finely divided material containing at least one base metal from the group consisting of copper, nickel and cobalt, at least partially in oxide form, characterized in that a) the material is sprayed together with fuel and oxygen and possibly flux into a confined space while the fuel is combusted to produce a high temperature non-reducing flame; b) superheating finely divided particles containing base metal in the substantially non-reducing flame to a temperature higher than the highest melting point of the reduced product to be produced; c) bringing the superheated particles substantially uniformly onto a thin layer of particulate coke flowing on the surface of a molten bath of the reduced products, the thin coke layer and the nearby atmosphere forming a reduction zone; d) reduces oxides of the base metals in the reduction zone, whereby the reduced products are obtained in liquid form, while all required heat is conducted into the reduction zone only in the form of free heat of the superheated particles and by radiation from the non-reducing flame; e) filtering the reduced products through the coking layer into the molten bath; F) removing the reduced products from the molten bath, and g) feeding solid carbonaceous granular material to the thin granular coke layer to complete the coke consumed in the reduction. 35 2. A process according to claim 1, characterized in that the fuel is at least partly made up of finely divided sulphide material and one of the reduced products is metal or pebble. 3. Process according to claim 2, characterized in that the sulphide material contains a base metal from the group consisting of copper, nickel and cobalt. 4. A process according to claim 1, characterized in that the base metal of the finely divided material is essentially in oxide form, carbonaceous fuel is used to provide the non-reducing flame with an aeration ratio close to 100% and the reduced product is essentially metallic. Process according to claim 1, characterized in that the finely divided material contains a substantial amount of a material to be treated with the flux, the flux is heated in the flame and a molten slag layer which can be discarded is between the thin layer of particulate coke and the molten reduced product. 20 6. A process according to claim 1, characterized in that the fuel is a liquid substance selected from the group consisting of hydrocarbon gases, hydrogen, hydrocarbon liquids, finely powdered carbonaceous solids and elemental sulfur. 25 7. A process according to claim 1, characterized in that finely divided pebbles are present in the finely divided material except the oxide. 8. A method according to claim 1, characterized in that atomized material, fuel, oxygen and flux are injected into the confined space, where the flame is maintained with at least one burner. 9. A method according to claim 8, characterized in that at least one burner is disposed to provide at least one flame which generally extends in road planes over the limited space, wherein 52 85878 gaseous products and non-gaseous products of at least a flame is separated by gravity, the non-gaseous products falling on and through the thin layer of particulate coke, which is on the lower side of the flame. 10. A process according to claim 2, characterized in that the flame and the layer of particulate coke are separated by a non-reducing, substantially sulfur-free gas between them. 1st