FI77169C - SAMLARREAGENS FOER FLOTATION AV SULFIDMINERALIER OCH FOERFARANDE FOER ANRIKNING AV SULFIDMINERALIER. - Google Patents

Description

1 77169

Collector reagent for flotation of sulphide minerals and methods for enrichment of sulphide minerals

The present invention relates to flotation processes for recovering metals from base metal sulfide ores. In particular, it relates to new, improved sulfide collectors containing certain hydrocarboxycarbonylthionocarbamate compounds which have excellent metallurgical performance over a wide pH range.

Flotation is one of the most widely used methods for the enrichment of ores containing valuable minerals. It is used in particular to separate finely ground valuable minerals from side rock or to separate valuable minerals from each other. Method 15 is based on the affinity of appropriately prepared mineral surfaces for air bubbles. In flotation, foam is formed by introducing air into a stirred, finely ground aqueous slurry of ore containing flotation agents. The main advantage of flotation separation is that it:; 20 is a relatively efficient procedure compared to other methods at a substantially lower cost.

According to current theory and practice, the success of sulfide foaming depends to a large extent on the reagent or reagents, called aggregators, which selectively make the sulfide mineral separated from other minerals hydrophobic. The flotation separation of one mineral species from another thus depends on the relative wettability of the mineral surfaces by water. Typically, the free surface energy is intentionally reduced by causing the surface to adsorb heteropolar aggregators. According to this description, the hydrophobic coating thus obtained acts as a bridge so that the mineral particles may adhere to the air bubble. However, the practice of this invention is not related to this or other flotation theories.

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In addition to the collector, several other reagents are required. Of these, foaming agents are used to form a stable foaming foam which is sufficiently stable that it cannot be decomposed to perform the following process steps. The most commonly used blowing agents are tall oil, creosote and cresylic acid, as well as alcohols such as 4-methyl-2-pentanol, propylene glycols and ethers, etc.

In addition, the success of the flotation separation of sulfide minerals is highly dependent on certain other important reagents, such as modifiers. Modifiers include all reagents whose primary function is not to act as collectors or blowing agents, but to modify the surface of a mineral so that the collector either adsorbs or does not adsorb it. Modifiers can thus be depressants, activators, pH adjusters, dispersants, deactivators, etc. Often, the modifier performs several functions simultaneously. According to current theory and practice of sulfide flotation, the efficiency of all flocculant classes depends to a large extent on the degree of alkalinity or acidity of the ore slurry. As a result, modifiers that adjust the pH are very important. The most commonly used pH adjusters are lime, calcined soda and, to a lesser extent, sodium hydroxide. However, lime is by far the most used in Sulfi diva suggestion. For example, copper sulfide flotation, which is the major sulfide flotation industry, uses lime to maintain a pH above 10.5, more usually above 11.0 30, and often as high as 12 or 12.5. In known sulfide flotation methods, the pH of the slurry must be adjusted in advance to 11.0 or higher not only to control known sidestone ferrous mineral sulfides such as pyrite and pyrrotite, but also to improve the performance of most conventional sulfide collectors such as xanthates. The cost of adding lime becomes quite high, and plant management is interested in flotation methods that require little or no lime addition, i.e., flotation methods that can be performed at slightly temporal, neutral, or even acidic pH values. Neutral and acidic circulating foaming methods are particularly desirable because sludges can be easily acidified with li-10 by adjusting sulfuric acid, and sulfuric acid is obtained in many plants as a by-product of the smelter. Thus, flotation methods in which the pH does not need to be adjusted in advance or in which the pH can be pre-adjusted to neutral or acidic using less expensive sulfuric acid are more advantageous than current flotation methods because in current flotation methods adjusting the pH to at least high, about ll, to 0, using lime, is more expensive.

In order to better clarify the current problems, it should be noted that in 1980 the US copper and molybdenum mining industry used almost 25 million kg of lime. This amount of industrial lime accounted for almost 92.5% by weight of the total reagents used, and the monetary value of the lime used was about 51.4% of the total reagent costs in this industry, i.e., more than $ 28 million.

As mentioned above, the lime consumption of different plants can vary from about 0.5 kg per tonne of treated ore to as much as 10 kg. In certain areas, such as South America, lime is poorly available and transport and / or import of lime has significantly increased costs in recent years. Yet another application of the known strongly alkaline methods is that when large amounts of lime are added to obtain a sufficiently high pH in the plant's flotation equipment, scaling is formed, the frequent removal of which requires the plant to be stopped for cleaning and thus increases costs.

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It is thus obvious that it is highly desirable that the need for lime addition in sulfide flotation processes could be reduced or eliminated altogether, resulting in significant savings in reagent costs. The additional reduction or omission of lime in the sulfide ore treatment may have other advantages, as it also facilitates non-flotation operations, such as sludge treatment.

In the past, xanthates and thiophosphates have been used as sulfide collectors in the flotation of base metal sulfide ores. The main problem with these conventional sulfide collectors is that at pH values below 11.0, pyrite and pyrrotite can be poorly controlled. In addition, as the pH is lowered, the collecting capacity of these sulfide collectors also decreases, making them unsuitable for a slightly alkaline, neutral or acidic environment. This decrease in the aggregating capacity when the pH decreases, e.g. below less than about 11, Q, means that the dose of the collector has to be increased several times, which is less economically desirable. The decrease in the activity of the collector as the pH decreases may be due to many factors. The collector may affect different sulfide minerals differently at a given pH. On the other hand, low pH at low pH, such as poor stability of xanthate and trithiocarbonate solutions, may explain the poor aggregation activity observed.

In an effort to overcome the above shortcomings, neutral derivatives of xanthates, alkyl xanthene alkyl formates, were developed, which can be represented by the following general formula: S 0 RO-C-S-C-OR '

Alkylxanthene alkyl formates are patented as sulfide collectors in U.S. Patent No. 2,412,500. Other structural modifications of this general structure were disclosed later.

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For example, in U.S. Patent 2,608,572, alkyl formate substituents contain unsaturated groups. The alkyl formate substituents described in U.S. Patent 2,608,573 contain halogen, nitrile and nitro groups. U.S. Patent 2,608,814 describes bis-alkyl xanthogen formates for use as sulfide builders. These modified structures have not achieved the same commercial use as non-modified structures. For example, alkyl xanthene alkyl formate is currently commercially available under the tradename MINERAC A from Minerec Corporation. MINERAC A, which is ethylxa-togene ethyl formate, as well as its higher homologues, do not work completely satisfactorily at pH values below 11.0, but their aggregation and pyrite control leave room for action.

One class of sulfide collectors that has to some extent achieved commercial success in flotation includes oily sulfide collectors containing dialkyl thionocarbamate or diurethane compounds of the general

The formula for 20 is: S

RO-CNHR1 Γ: There are several disadvantages associated with the preparation and use of these compounds. U.S. Patent No. 2,691,635 discloses a process for preparing dialkylthionocarbamates. The three steps of the described reaction sequence are cumbersome to perform, and the final by-product is methyl mercaptan, which is an air contaminant that is expensive to handle. U.S. Patent 3,907,854 discloses an improved process for preparing dialkylthionocarbamates. Although good yields and high purity are presented as new features of the process, it should be noted that the by-product of the reaction is sodium hydrosulfide, which is also an environmental toxin and requires special treatment. U.S. Pat. No. 3,590,998 discloses a thionocarbamate-6,777,169 sulfide scavenger compound in which the N-alkyl substituent is attached via alkoxycarbonyl groups. The method described must use expensive amino acid esters for the thioester substitution reaction of xanthates. The by-products of this method are either methyl mercaptan or sodium thio-glycolate. In addition, this type of structurally modified thionocarbamate has not achieved much commercial success. As will be apparent from the description of the present invention below, dialkylthionocarbamates are weak accumulators when the pH drops below certain values.

It is therefore an object of the present invention to provide a new, improved sulphide collector and flotation process for enriching sulphide minerals using a flotation process in which it is not necessary to pre-adjust the pH to strongly alkaline.

Another object of the invention is to provide a novel, improved sulfide collector and flotation process for enriching sulfide minerals so as to selectively recover value metal sulfides and selectively control pyrite, pyrrotite and other side rock sulfides.

The invention further relates to a new, improved sulphide collector and flotation process for enriching sulphide minerals by a flotation process using a new class of sulphide sizing reagents which can be prepared and used without the formation of harmful by-products or environmental toxins.

The invention further relates to a flotation process for enriching sulfide ores at a pH of 10.0 or less using certain novel collectors containing new donor atom combinations and specifically designed for low pH flotation.

The invention further relates to a new, improved process for the selective foaming of sulphides in an acid-35 cycle in which cheap sulfuric acid is used to adjust the pH.

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More particularly, the invention relates to a collecting reagent for flotation of sulfide minerals. The collection reagent according to the invention is characterized in that it comprises at least one hydrocarboxycarbonylthio-5-carbarbamate which is N-ethoxycarbonyl-O-isobutylthioocarbamate? N-ethoxycarbonyl-O-sec-butyylitionokarba-carbamate; N-ethoxycarbonyl-O-n-amyylitionokarbamaatti; N-ethoxycarbonyl-O-isoamyylitionokarbamaatti; N-ethoxycarbonyl-O-phenylthionocarbamate N-phenoxycarbonyl-10-ethylthionocarbamate N-phenoxycarbonyl-O-isopropylthionocarbamate; N-phenoxycarbonyl-O-n-butyylitiono-carbamate; N-phenoxycarbonyl-O-isobutylthionocarbamate; N-phenoxycarbonyl-O-sec-butyylitionokarbamaatti; N-phenoxycarbonyl-O-n-amyylitionokarbamaatti; or N-phen-15-oxycarbonyl-O-isoamylthionocarbamate.

In general, the novel, improved hydrocarboxycarbonylthionocarbamate collectors of the invention can be used per tonne of ore of about 0.0025-0.25 kg, preferably about 0.005-0.05 kg (but not limited to these values) to efficiently recover metals and minerals. and selectively from metal sulfide ores while selectively pressing pyrite and other side rock sulfides. The novel, improved sulfide collectors of the invention can generally be used regardless of the pH of the ore slurry. Again, without limitation, these collectors can be used by Dana at pH values of about 3.5-11.0, preferably about 4.0-10.0.

The invention also relates to a new, improved method for selectively enriching sulphide mineral-containing ore by pressing pyrite and pyrrotite, in which the ore is ground to flotation size particles, the particles are slurried in an aqueous medium, the slurry is prepared with fluxes and - 8,771 69 by a process wherein said metal collector contains at least one of the hydrocarboxycarbonylthionocarbamates described above.

A particularly preferred embodiment comprises a new, improved method for improving the recovery of copper sulphide minerals from ore containing various sulphide minerals by a flotation process carried out at a controlled pH of up to 10.0 and adding a collector to the flotation cell.

The present invention therefore relates to new sulphide collectors and to new flotation methods for base metal sulphide ores. The hydrocarboxycarbonylthionocarbamate collectors of the present invention surprisingly provide better metallurgical recovery in flotation separation than conventional sulfide collectors, even at lower collector doses, and are effective under acidic, neutral, and slightly temporal pH conditions. The invention provides a sulphide ore flotation process which simultaneously provides very good enrichment of sulphide minerals and considerable savings in lime consumption.

The hydrocarboxycarbonylthionocarbamate compounds of the invention can be conveniently prepared without the formation of environmentally hazardous by-products by first reacting the corresponding chloroformate with ammonium, sodium or potassium thiocyanate to give the isothiocyanate intermediate according to the following equation (1):

O O

R 10 -C-Cl + XSCN -> R 1 -O-C-NCS + XCl (1) 30 wherein R 1 is as defined above and X is NH 4 +, Na + or K +.

The hydrocarboxycarbonyl isothiocyanate intermediate is then reacted with the active hydroxyl compound according to Equation (2): 9 77169

0 OHS

R1-0-C-NCS + R2OH -> R ^ D-C-N-C-OR2 (2)

By active hydroxyl compound is meant any compound having a hydroxyl group that readily reacts with an isothiocyanate to form the corresponding thionocarbamate. Examples of active hydroxyl compounds are aliphatic alcohols, cyclic and acyclic, saturated and unsaturated alcohols which are unsubstituted or substituted by polar groups such as halogen, e.g. chlorine, bro-. . with iodine, nitrile or nitro groups, aryl alcohols such as benzyl alcohol, ethoxylated and / or propoxylated alcohols and phenols.

In the reaction according to Equation (1) above between the chloroformate and ammonium, sodium or potassium thiocyanate, the required chloroformates can be prepared with the corresponding aliphatic or aromatic alcohol V; in the reaction between lie and phosgene according to the following equation 20 (3): 'K 0 0 R40H + C1-C-C1-> R40-C-C1 + HCl (3) 4 25 wherein R is one of the active hydroxyl compounds defined above . The preparation of chloroformates from ethoxylated or propoxylated alcohols by this method can be further prepared by the following reaction scheme: R 0 (OCH „CH„) OH + C1-C-C1 -> R5 (0CHoCH ») 0C-C1 + HCl (4) 2 2 n 22 n wherein R 5 is C 1 -C 6 alkyl, and n is 1-4, the preparation of aromatic alcohols such as phenols, cresols 35 and xylenols is shown in the following reaction scheme: 10 771 69 ί :: · O (5) 2 6 ^ -OC- Cl + HCl

5 Cl-C-Cl-> R

where = H or CH 2, and R 2 = H, CH 2, Cl, Br, I, -NC 2 or -C = N.

Equations (1) and (2) showing the preparation of new, improved hydrocarboxycarbonylthionocarbamate sulfide collectors according to the invention show that only harmless sodium chloride is formed as a by-product in the reaction according to equation (1). In addition, the condensation of the isocyanate and the active hydroxyl compound according to Equation 15 (2) is rapid and complete, and no environmentally hazardous by-products are formed.

In accordance with the present invention, the hydrocarboxycarbonylthionocarbamates described above are used as sulfide collectors in a new, improved process for enriching sulfide minerals from base metal sulfide ores over a wide pH range and especially under acidic, neutral, slightly temporal and strongly temporal conditions.

According to the present invention, a new, improved, substantially pH-independent process for the enrichment of minerals from base metal sulphide ores comprises, as a first step, a reduction in the particle size of the ore to suit flotation. As will be appreciated by those skilled in the art, the particle size into which the ore is to be ground in order to release the desired minerals from the associated side rock or other minerals, i. release size, varies with different ores depending on several factors, such as the geometry of the mineral layers in the ore, e.g. beverages, agglomerates, common matrices, etc. In any case, as is common in the art, microscopy can be performed to determine the particle size reduction. In general, a suitable particle size ranges from about 50 mesh to about 400 mesh, however, these values are not limiting. The ores are preferably comminuted into flotation-sized particles ranging from about +65 mehs to about -200 mesh. Particularly preferably, for the purposes of the present invention, the base metal sulfide ore is comminuted so that about 14-30% by weight of the particles are +100 mesh and about 45-75% by weight of the particles are -200 mesh.

The ores can be comminuted by any method known in the art. For example, the ore may be crushed to a size of -10 mesh and then wet milled to a mesh size specified in a steel ball mill, or autogenous or semi-autogenous milling or milling may be used.

. . The method used to grind the ore is not critical; ; for the purposes of the invention, as long as particles suitable for efficient flotation are obtained.

The pre-adjustment of the pH is suitably carried out by adjusting the adjusting agent during the grinding step to the powdered ore.

. The pH of the ore slurry can be adjusted to any desired value by the addition of either acid or base, typically sulfuric acid or lime is used for this purpose. The process of the invention is clearly advantageous in that new, improved hydrocarboxycarbonylthionocarbamate sulfide collectors can be used in the process of the invention without prior pH adjustment, and flotation can generally be performed at a natural pH of 30 ml, which simplifies the process, reduces costs and reduces lime consumption and consequent downtime. Good enrichment has been obtained, for example, by the process according to the invention at pH values of 3.5-11.0, and a particularly good enrichment result has been observed at pH values of about 4.0-10.0.

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The comminuted ore, which contains e.g. particles comminuted to a release size, is then slurried in an aqueous medium as an easy-to-flow slurry. The aqueous slurry containing flotation-sized ore particles is typically adjusted in the flotation device to contain about 10-60% by weight of slurry solids, preferably 25-50% by weight and particularly preferably about 30-40% by weight solids.

In a preferred embodiment of the process according to the invention, the flotation of copper, zinc and lead sulphide is carried out at a pH of at most 10.0, preferably below 10.0. It has been found that by performing flotation at this pH, the new, improved hydrocarboxycarbonylthionocarbamate collectors according to the invention have a particularly good collector efficiency and at the same time very good collector selectivity even at reduced collector doses. Thus, in this preferred method, sulfuric acid is used, if necessary, to adjust the pH of the slurry to a maximum of 10.0.

In all cases and for any reason, the pH of the slurry can then be adjusted in advance by any known method, if desired.

Once prepared, the slurry is prepared by adding effective amounts of a blowing agent and a collector containing at least one of the hydrocarboxycarbonylthionocarbamate compounds described above. By "effective amount" is meant an amount of a component that provides the desired level of enrichment of the desired metal.

Any blowing agent could be used in the process of the present invention. Some suitable blowing agents may be mentioned as examples; straight or branched chain low molecular weight hydrocarbon alcohols such as C8-C8 alkanols, 2-ethylhexanol and 4-methyl-2-pentanol, also known as methyl isobutylcarbinol (MIBC), as well as tall oils, polyglycol or cresylic acid, monoethers and alcohol ethoxylates of polyglycols. In general, the blowing agent or agents are used in conventional amounts, and amounts of about 0.005 to 0.1 kg / ton of ore to be treated are suitable, however, these amounts can also be deviated from.

It has also surprisingly been found that, contrary to popular belief, a neutral, oily collector is most effective when added to a grinding rather than a flotation cell, as the new, improved hydrocarboxycarbonylthionocarbamate collectors of the invention are more effective when added to a flotation cell and not to grinding. Although the new collectors according to the invention are practically insoluble in water, they have the clear advantage of easy dispersibility. When new collectors are added to the flotation-15 cell, they provide better copper recovery in the first flotation step as well as improved overall copper recovery, suggesting improved flotation kinetics, which will be fully described below. Of course, new, improved collectors can also be added to the grinding process by a conventional method, and even then improved recovery of base metals is obtained.

The novel hydrocarboxycarbonylthionocarbamate collectors of the invention and methods of selectively concentrating or collecting certain metal sulfides, usually for the separation of copper, nickel, lead and zinc sulfides from side rock sulfides, e.g. , e.g. silicates, carbonates, etc. However, in some cases it is desirable to collect not only copper sulfide but also all sulfides contained in the ore, including sphalerite (ZnS) and iron sulfides, i. pyrite and pyrrotite.

In particular, this applies to certain massive or complex sulphide ores which contain large amounts of ferrous sulphide minerals, such as pyrite or pyrrotite. From these complex sulfide ores, it is often desired to foam ferrous sulfide minerals to recover sulfide from minerals from which further processing yields sulfur and sulfur reagents. Under these conditions, co-flotation is preferably used, i.e. composite flotation, i. foaming, where all sulfide minerals are foamed and assembled. Composite flotation is also preferably used to enrich precious metals from pyrite and pyrrotite minerals containing precious metals.

However, these massive or complex sulfide minerals 10 often contain, in addition to various base metal sulfides such as copper, zinc, lead, nickel, cobalt sulfides, closely related side rock such as carbonates as well as silicates and siliceous materials.

Such massive or complex sulfide ores are not uncommon and represent a unique problem in flotation enrichment. Composite sulfide flotation of these ores cannot be successfully performed under conventional flotation conditions, i.e., at pH 10.0, 20 because pyrite and pyrrotite are controlled at high pH values. At pH values of 3.0-5.0, the composite flotation result is good when using conventional collectors such as xanthates, and sulfuric acid is used as a modifier to lower the pH of the slurry to these values. The carbonate side minerals contained in these complex ores are acid-soluble, so large amounts of sulfuric acid are required, e.g., about 5-6 kg / ton of ore, which is not economically attractive, and the use of sulfuric acid in alkaline earth metal carbonates such as calcite, dolomite, etc. Ores form large amounts of insoluble alkaline earth metal sulphates, which cause severe scale formation in the plant, which in turn requires frequent and costly downtime. At slurry pHs of about 6.0-9.0, the composition of the composite sulfide foam using conventional collectors such as xanthates is below optimum.

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Surprisingly, it has been found that the novel, improved hydrocarboxycarbonylthionocarbamate collectors according to the invention give, under well-defined conditions, an optimum flotation result of co-sulphides from sulphide-containing ores. According to this aspect of the present invention, the optimum result of composite sulfide flotation is obtained by flotation at neutral or slightly temporal pH values, especially pH 6.0-9.0, and using higher amounts of hydrocarboxy-10 carbonylthionocarbamate collectors according to the invention, namely at a level of about 0.05 -0.5 kg / ton or in other words at a level of at least about 0.3 mol / ton of ore.

Once the sulfide co-concentrate has been prepared by foaming at these pH conditions and using the specified collection doses, the copper, lead and zinc sulphides are separated from the large amounts of ferrous sulphides in the co-concentrate in a second flotation step at a higher pH, i.e. above 9.0 pH. the sulfides are assembled and the ferrous sulfides are selectively pressed.

Previously, xanthate collectors at pH 3.0-5.0 were used in composite flotation, and the second flotation step, in which iron sulfides were selectively pressed, had to be performed at a pH of about 11.0 because pyrite prevention for xanthate collectors is poor at pH values below 11.0. It can be seen that significant amounts of lime had to be added to adjust the pH for this second flotation step. According to this aspect of the present invention, using a hydrocarboxycarbonylthionocarbamate collector, co-sulfide flotation can be performed at a higher pH of 6.0-9.0, and the amount of lime required in the second flotation step, i.e., separation of more valuable metal sulfides from ferrous sulfides, is lower. In addition, the hydrocarboxycarbonylthionocarbamate collectors of the invention are much more efficient at collecting copper, lead and zinc at pH values of 9.0-11.0, so that the second flotation step can be performed at pH values just sufficient to control iron sulfides. , in which case it is not necessary to increase the pH value by more than 11.0, and thus additional savings in lime consumption are obtained.

Other objects and advantages achieved by the new, improved assemblers and method of the invention will become apparent from the following working examples, which are intended only to further illustrate the invention so that those skilled in the art can better understand and practice the present invention.

Preparation 1

Synthesis of Ethoxycarbonyl Isothiocyanate A 2-liter three-necked round bottom flask equipped with a vertical condenser with a tube containing 15 anhydrous calcium sulfate as a moisture shield, a dropping funnel and a mechanical stirrer was placed in a heating mantle. To the flask were added 700 ml of anhydrous acetonitrile and 194 g of potassium thiocyanate. The mixture was heated to 70 ° C with stirring, then external heating was stopped. 217 g of ethyl chloroformate were added to the mixture from a dropping funnel over 40 minutes with stirring. The reaction was exothermic. The mixture thickened and turned yellow. At the end of the addition, the temperature of the reaction mixture was 77 ° C. The reaction mixture was stirred without external heating for 3 hours. The reaction mixture was then cooled to room temperature and the precipitate was removed by filtration. The precipitate was washed with anhydrous acetonitrile. The filtrate and washings were combined and concentrated by evaporation under reduced pressure. The residual liquid was distilled using a fraction column. 86.9 g of ethoxycarbonyl isothiocyanate were obtained as a colorless liquid having a mp of 45 ° C / 11 mm Hg or 48 ° C / 12 mm Hg.

Preparation 2 Synthesis of N-ethoxycarbonyl-O-isopropylthionocarbamate 35 To 10 g of ethoxycarbonyl isothiocyanate (V-17,771 69 m 1) was added 40 ml of isopropyl alcohol and the solution was stirred well. After completion of the exothermic reaction, the reaction solution was allowed to stand overnight, and the progress of the reaction was monitored by the infrared spectrum of the reaction solution.

Completion of the reaction was indicated by the disappearance of the N = C = S group absorption belt at 1960-1990 cm. The excess isopropyl alcohol was evaporated under reduced pressure to give an oily residue. Crystallization from petroleum ether (b.p. 35-60 ° C) gave 13.1 g of colorless crystalline N-ethoxy-10 carbonyl-O-isopropylthiocarbamate, mp 32-33 ° C.

Preparation 3 Synthesis of N-ethoxycarbonyl-O-isobutylthionocarbamate To 10 g of ethoxycarbonyl isothiocyanate (Val-15 mist 1) was added 40 ml of isobutyl alcohol. After completion of the reaction, the excess isobutyl alcohol was evaporated under reduced pressure. 15 g of N-ethoxycarbonyl-O-isobutylthionocarbamate were obtained as a colorless oil.

• - Preparation 4 Synthesis of N-phenoxycarbonyl-O-ethylthionocarbamate

Complete reaction series A <2> ~ »S- 25 30 250 ml round bottom three neck flask was equipped with a condenser, thermometer, dropping funnel and mechanical stirrer. To the flask were added 100 ml of ethyl acetate and 9.7 g of potassium thiocyanate. The mixture was stirred and heated. 15.7 g of phenyl chloroformate were added dropwise to the mixture over 30 minutes from a dropping funnel. After the exotherm ceased, the reaction mixture was stirred for 1.5 hours (gas chromatography showed complete reaction of phenyl chloroformate). 10 ml of anhydrous ethyl alcohol was added. The reaction mixture was stirred and the reaction was monitored by infrared until the phenoxycarbonyl isothiocyanate was completely reacted. To dissolve the solid, 50 ml of water was added. The reaction mixture was transferred to a 250 mL separatory funnel. The organic layer was separated. It was dried over MgSO 4 and filtered. The filtrate was concentrated by evaporation of the volatiles. 2Q, 4 g of solid were obtained. It was recrystallized from hexane. Pure product mp 81-83 ° C.

The hydrocarboxycarbonylthionocarbamates synthesized above were used as collectors in the enrichment of various sulfide ores, their enrichment properties were tested at different pH values and compared to known sulfide collector compounds. Other homologous and / or analogous hydrocarboxycarbonylthionocarbamates used in the following examples may be prepared by substantially identical preparation methods using the appropriate corresponding active hydroxyl compound to give the indicated R group in the compound.

25 In all of the following examples, the following general preparation and test methods were used:

Sulfide ores were crushed to a size of -10 mesh. About 500-20 g of crushed ore was wet milled in a steel ball mill / with a number of steel balls of 5.3-10.7 kg and a solids content of 50-75%; the grinding time was 6-14 minutes, i.e. until the slurry had obtained the reported size distribution, usually about 10-20% +65 mesh, 14-30% +100 mesh, and 43-80% -200 mesh. Lime and sulfuric acid were used as modifiers to adjust the pH to the required range. These modifiers 35 were generally added to the milling. The foam used in some experiments was added to the grinding and in others it was added to the flotation cell. In certain experiments, 50% of the collector was added to the grinding, otherwise the collector was added to the flotation-5 cell in the first and second stages of conditioning.

The comminuted slurry, possibly containing foaming and collector additives, was transferred to a rectangular Denver D12 flotation cell. The volume of the slurry was adjusted to 1200-2650 ml by adding water so that the ti-10 of the slurry became about 20-45% solids and the level of the slurry in the cell was about 2 cm below the edge.

A collector and / or foam was added to the slurry by stirring the slurry at a speed of about 1100-1400 rpm. The slurry was conditioned for 2 minutes, during which time temperature measurements were made. At the end of the 2 minute conditioning, about 5-7 liters / min of air was drawn from the compressed air bottle. Flotation was continued for about 3 minutes during which time the first stage concentrate was collected. The il-Mahana was then sealed and an additional collector 20 was added to the slurry and the frother and slurry were conditioned for an additional 2 minutes. After a second 2 minute conditioning step, the air tap was opened and the concentrate from the second step was collected. The foaming times were predetermined so that the foam obtained at the end of the foaming did not contain minerals.

The concentrates and concentrates of the first and second stages were filtered, dried, and copper, iron and sulfur were determined from the samples taken from them. Tap water of the required temperature was used in all experiments.

Example 1 Foaming at Natural pH In this series of experiments, copper ore with a copper content of 0.3% and a pyrite (FeS2) content of 1.7% was used in the southwestern part of the USA. In this and all the following examples of side iron iron minerals such as pyrite, pyrrotite, etc., for the sake of simplicity, a pyrite is used. The main copper minerals were, chalcosite and chalcopyrite.

460 g of ore was ground at 60% solids for 8.5 minutes to give a slurry with the following size distribution: 17.5% +65 mesh, 35.2% +100 mesh and 41% -200 mesh. The natural pH value of the slurry, i.e. the pH value without the addition of lime or sulfuric acid, was 5.5. The slurry was conditioned at a solids content of 30% using the indicated collector and a frother containing 50/50 MIBC / tall oil at a weight ratio of 0.04 kg / ton of ore, and the first and second flotation were performed by the methods described above. The collectors used and the analyzes of concentrates and concentrates are shown in Table 1: 15 Table 1

Natural pH 5.5; frother - 1: 1 MIBC / tall oil 0.04 kg / ton. Elemental analysis: CU - 0.3%, FeS2 - 1.7%

Eg Collector Dose Recovered Cu content Recovered% taken _Cu% _FeS ^,% 20 A Sodium ethyl xanthate 0.054 18.6 0.7 2.1 B Sodium ethyl xanthate 0.200 82.8 4.0 89.7 C Sodium diethyldithio-0.100 66.6 3.3 64.4 phosphate D Sodium N-isobutyl-0.054 69.3 2.3 11.2 dithiophosphinate 25 E Ethyl xanthogen-ethyl formate

Panos 1 0.054 84.1 2.3 74.2

Panos 4 0.054 79.2 1.9 51.2

Batch 5 0.054 86.4 3.9 91.1 30 F (isopropyl N-ethyl 0.054 73.2 2.7 57.1 thionecarbamate G O-isobutyl N-ethyl 0.054 78.0 3.2 51 , 1 thionecarbamate 1 N-ethoxycarbonyl-O-iso-0.054 90.8 9.6 67.3 propylthionocarbamate 35 aMINEREC®A, Minsrec Corporation, Baltimore, Md USA.

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It is obvious that the conventional collectors of Examples A-G are much weaker at this pH than the collector of Example 1 according to the invention. In addition to the maximum recovery of copper from the hydro-carboxycarbonylthionocarbamate of Example 1, the accumulators tested also had a maximum copper content while the pyrite control was acceptable.

Examples 2-3

The following examples used copper-molybdenum ore from the southwestern United States with an Elemental Analysis of 0.458% copper and 2.2% pyrite. Malini contained chalcopyrite, chalcosite and covellite as the main copper minerals. The ore was ground at a solids content of 63% in a steel ball mill to a slurry with a particle size distribution of about 16.4% +65 mesh, 30% +100 mesh and 43.9% -200 mesh. The natural pH of the ore slurry was 5.0. 50 g / ton of 1: 1 MIBC / tall oil was used as the frother. To make comparisons more accurate and significant, the collectors were dosed in equimolar amounts of 0.03 mol / ton, corresponding to approximately 0.005 kg / ton. In addition, a selectivity / performance factor was calculated for each collector tested.

The selectivity / performance factor was calculated according to the following equation: T _ (100 -% recovered pyrite) 25 1cu ~ ”“ '2 (100 -% copper recovered)

The selectivity coefficient of copper is a suitable measure of not only the recovery of copper but also the selectivity to combat pyrite. For example, if in the ta-3Q case of this ore, 90% copper recovery and 50% pyrite recovery are accepted as optimal, then the selectivity coefficient for copper is: = (100-50) CU (100-90) 2 = 0.5 35 22 771 6 9

The collectors tested and the results obtained with them are shown in Table 2 below:

Table 2 5 Natural pH 5.0 (no lime); frother - 1: 1 MIBC / tall oil 50 g / ton; collectors 0.03 mol / ton (approx. 0.005 kg / ton)

Eg Collector Recovered - Cu content Recovered

taken% FeS ~ I

___ Cu, t_2 01 10 H Sodium isobutyl 33.2 4.3 9.6 0.02 Xanthate I O-isobutyl N-ethyl 76.8 8.2 46.6 0.100 thionocarbamate J O-isopropyl N-methyl - 67.7 5.8 38.8 0.059 thionocarbamate 15 K Ethyl xanthene ethyl ethyl 84.6 9.2 50.0 0.211 formate, charge 1,88.2 7.1 55.5 0.319

Ethyl xanthene ethyl 86.2 6.3 52.7 0.248 lyformate, batch 2

Ethyl xanthene ethyl 85.7 6.4 56.9 0.212 20 lyformate, batch 3, pure L Sodium trin-butyl tri-58.8 6.4 16.9 0.049 tick carbonate M Isobutyl xanthene 85.6 7.7 38.2 0.297 ethyl lyformate 25 N Isopropylxanthogen 86.2 6.5 65.5 0.180 ethyl lyformate O Isopropyl xanthogen 88.7 6.1 64.6 0.277 nibutyl formate P Ethyl xanthogen 83.3 8.4 45.4 0.196 phenylformate 30 2 N-ethoxycarbonyl-90.8 9.6 67.3 0.389 O-isopropylthionocarbamate 3 N-ethoxycarbonyl-91.1 6.7 58.3 0.525 O-anrylthionocarbamate 35 X This single run yielded unusually high copper recovery. All other figures represent reproducible results.

23 771 69

The results of Table 2 clearly show the superiority of the collectors according to the invention, i.e. the better results of Examples 2-3 compared to the results obtained with conventional collectors in Examples H-P. In Examples 2 and 3, the recovery of cup-5 was high while the control of pyrite was satisfactory. Only the collectors of Examples 2 and 3 gave I values close to the optimum of 0.5. cu

Examples 4-7

In the following examples, the experiments were performed using the same ore as in Examples 2-3 and using different collectors to determine whether the collectors of the invention were better even at higher dose levels or whether their superiority was limited to just one dose level. The collectors tested, the doses and the results obtained are summarized in Table 3: 771 69 24

Table 3

Eg Collector Dose Recovery- Cu Content- Recovery-

irol / t taken by mouth,% taken I

_Cu,% _Fes2,% Q Ethyl xanthogen-5 pharmaceutical (batch 1) 0.04 86.0 8.8 55.6 0.227 (batch 3), silicon 0.04 88.6 6.0 65.7 0.261 4 N- tecycarbonyl-O-0,04 91,2 7,1 78,5 0,277 isopropylthionocarbamate 5 N-ethoxycarbonyl-O-0,04 89,5 6,6 63,9 0,330 10 butylthionecarbamate R Sodium ethyl-0, 14 41.3 5.0 38.3 0.018 Tate 5 Sodium diisobutyl 0.14 59.0 7.7 24.0 0.045 Dithiophosphinate T Ethyl xanthogen 0.14 86.0 8.1 91.5 0.043 15 e style formate (batch 1) U Sodium butyl trithio-0.14 83.8 6.6 43.5 0.216 carbonate V Diallyl trithiocarbo- 0.14 79.4 9.2 50.1 0.117 natt 20 W Amyl allyl xanthate 0.14 75 , 4. 9.8 40.4 0.099 Tiester 6 N-Ethoxycarbonyl-O 0.03 90.8 9.6 67.3 0.389 Isopropylthicycarbamate 7 N-Ethoxycarbonyl-O-0.014 89.4 8.4 48.6 0.456 Isopropylthionocarbamate 25 naatti

As the figures in Table 3 show, the new, improved collectors of both Examples 4 and 5 provided better copper recovery and I-value at a dose of 0.04 w / v 3Q mol / ton than the conventional collectors of Example Q. In addition, the conventional collectors R-W were inferior to the collectors of Examples 4, 5, 6 and 7 also at a dose of 0.14 mol / ton. These figures show that at pH 5.0, the new collectors according to the invention give significantly better results than the conventional collectors, even at greatly reduced dose levels, e.g. 1/5 in Example 6 and 1/10 in Example 7. .

Examples 8-14 Using the same ore, the performance of conventional collectors 5 and collectors of the invention was compared, comparing the effect of hydrocarbon chain length and structure using known dialkylxanthogen formates for comparison. The collectors tested and the results obtained are shown in the following Table 4: ^ Table 4

Natural pH 5.0 (no lime)? frother - 1: 1 MIBC / tall oil 50 g / ton; collectors 0.03 mol / ton (approx. 0.005 kg / ton)

Eg Collector Recovered Cu Concentration, As Art I

15 taken% taken 011

Cu,% FeS2 '%

Alkylxanthogenalkyl (phenyl) formate S 0 tsp tsp

R1-O-C-S-C-O Ro 20

R1 = C2H5, R2 = C2H5, 50 0 0 211

Batch 1 84 | 6 9> 2 U'Zi

Batch 2 86> 2 6> 3 52.7 ° '248

Batch 3 8V 6 »4 56'9 ° '212 25 Y Ri = i-C3H7, R2 = C2H5 86> 2 6 * 5 65» 5 ° »180 Z Ri = i-C4H9, R2 = C2H5 85» 6 7I7 38 '2 0.297 AA Ri = i-C3H7, R2 = n-C4H9 88t7 6 »1 64» 6 0.277 3Q BB R1-1-C3H7, R2 = CeH5 89 »7 6.7 71.1 0.273 CC Ri * C2H5, R2 -C6H5 83'3 8.4 45f4 0.196 26 77169

Eg Collector Recover- CU Hold- Recover- I

taken favor,% taken 011

Cu, S ^ e ^ 2 '*

Ethoxy (phenoxy) karbonyylialkyylitionokarbamaatti

5 0 S

R10-C-NH-C-R2 8 R1 = C2H5, R2 = C2H5 83 ^ 2 10.1 38.2 0.219 9 R1 = C2H5, R2 = i-C3H7 90.8 9.6 67.3 0.389 10 ' 10 R1-C2H5, R2 = i-C4H9 90.6 7.7 81.3 0.212 11 Ri = C2H5, R2 = n-C4H9 88.2 10.2 49.6 0.359 12 Ri = C2H5, R2 = C5Hh 91, 1 6.7 58f3 0.525 15 13 Ri = C6H5, R2 = C2H5 90, 1 6.8 71.9 0.284 14 Ri = C6H5, R2 = i-C3H7 90.9 6.5 68.1 0.384

The results in Table 4 show that, with the exception of diethyl homolog, all the collectors of Examples 8-14 of the invention were superior in copper recovery than the corresponding conventional collectors having the same R 1 and R 2 substituents. I values were also correspondingly better. The amyl homologue of Example 12 had the best 25 copper recovery and the best Icu value.

Examples 15-17

In the following examples, South American copper molybdenum ore was used. This ore contained 1.65% copper, 2.5% pyrite and 0.025% molybdenum. Copper 30 was in the form of chalcosite, chalcopyrite, covellite, bornite, and the ore also contained some copper oxidine minerals such as malachite and cupritite. Although the ore contained a large amount of chalcopyrite, a substantial portion of it contained chalcosite and covellite as curbs.

35 About 500 g of -10 mesh ore was wet milled for 13 mi-27,771 69 minutes in a steel ball mill at 63% solids using 5.3 kg of steel balls to give a slurry with a size distribution of 14% + 1QQ mesh and 62% -200 mesh. In all experiments, 10.5 g / ton of diesel oil was also added.

5 The natural pH of the ore slurry was 5.5. The standard collector for this ore is a neutral alkyl xanthogen-alkyl formate mixture, e.g. a mixture containing ethyl

O

xanthene ethyl formate (MINEREC A), gasoline and 4-methyl-2-pentanol (MIBC) in a ratio of 60:30:10. Polypropylene glycol monoalkyl ether, such as polypropylene glycol monomethyl ether, is added in an amount of 60 g / ton. The standard collector was compared in mixed and unmixed form to the collectors of the invention at different dose levels; the collectors were added to the flotation cell in the first and second steps according to the flotation test method described above. The test results are shown in the following Table 5:

Table 5

Natural pH 5.5 (no lime or sulfuric acid); 2Q frother - 60 g / ton

Eg Collector Dose Recover- Cu content- Recover- I

g / t taken by mouth,% taken Cu,% * ED Ethyl xanthene-20 66.1 6.7 69.5 0.026 ethyl formate / -2 5 petrol / MlBC 60 / 3Q / 1Q- mixture EE - "- 40 70.9 6 .6 72.0 0.033 FF 58 72.9 6.2 76.8 0.032 15 N-ethoxycarbonyl-O-20 73.8 9.4 77.2 0.033 isopropylthionocarbamate 3 40 76.4 8.3 80, 3 0.034 17 - "- 58 79.5 7.5 GG Ethyl xanthogen 20 67.4 6.9 75.5 0.023 ethyl formate (batch 1) HH 40 72.3 6.7 75.8 0.032 35 II 58 68.4 7.2 70.5 0.030 28 771 6 9

As the results in Table 5 show, the unmixed hydrocarboxycarbonylthionocarbamate collectors of Examples 15-17 of the invention were superior to conventional neutral xanthene formate collectors, either in pure form or in mixtures, in terms of both copper recovery and copper content.

Examples 18-25

In the following examples, the collectors according to the invention in mixed and unmixed form were compared to neutral xanthogen formate collectors for the same South American copper molybdenum ore and by the same test method, but this time the pH of the ore slurry was adjusted to 5.0 by adding 5.0 kg / ton of sulfuric acid. conditioning and flotation test. Again, the collectors li-15 were only set in the flotation cell during the first and second conditioning steps. The collectors used and the results obtained are shown in the following Table 6: 29 7 7 1 69

Table 6

Cu = 1.65%, FeS2 = 2.5%; sulfuric acid 5.0 kg / ton; pH: 4.0; frother Dow 1012 = 60 g / ton

Eg Collector Dose Recover- Cu content- Recover- I

g / t taken by mouth,% taken 5 Cu,% FeS2 '% JJ Ethylxanthene ethyl 5 33.4 3.4 15.8 0.019 formate / petrol / -' MIBC standard mixture 60/30/10 KK 10 46.7 4.5 21 .1 0.028 10 LL 20 80.4 6.7 79.4 0.054 m 30 89.6 7.2 91.5 0.078 NN 40 90.1 7.2 92.2 0.080 00 Pure ethylxanthone 5 61.7 6, 6 44.5 0.038 15 gene style formate (batch 3); PP 15 88.5 8.8 88.2 0.090 // - qq 20 90.6 8.4 93.4 0.075 18 Ethoxycarbonyl isoprop-5 68.7 6.6 52.3 0.049 yl thionecarbamate 19 10 89, 7 7.9 92.1 0.074 20 20 93.2 7.4 91.7 0, 180 21 N-ethoxycarbonyl-O-20 92.4 7.9 90.5 0, 163 isopropylthionecarba- 2 5 countries ti / benz iin / talBC mixture 60/30/10 22 N-acetylcarbonyl-O-20 91.6 9.1 90.3 0.136 isopropylthionocarbamate / lencin / MrBC mixture 60/30/10 30 0-3 v ^ ·. 5 67.5 7.2 45.0 0.052 23 N-tecyccarbanyl-O- * thisobutylthiancarba- 24 ^ ttL 10 89.8 9.0 87.7 0, 119 25 20 93f5 8.2 97.0 0.072 30 771 69

Table 6 shows that the collectors of the invention in pure form (Examples 18-20 and 23-25) or in mixed form (Examples 21 and 22) are collectors with stronger activity than standard xanthogen formate collectors, either as mixtures or pure at all dose levels tested. In addition to the fact that the copper recovery in Examples 18-25 was about 3% higher without a decrease in the copper content, an increased% recovery was obtained at a much lower dose than with the corresponding standard collector 10. The dosing advantage of the hydrocarboxycarbonyl thionocarbamate collectors according to the invention makes them economically advantageous, i.e. a better recovery result is obtained with lower reagent costs.

It should be noted that in enriching this particular ore, the recovery of pyrite was remarkably high and appeared to closely follow the% recovery of copper. Microscopic analysis showed that in this particular ore, the pyrite in the degree of grinding used was in close contact and / or as an edge mineral with the copper minerals, so that a high% recovery of copper necessarily produced a high% recovery of pyrite. Although all the collectors tested in Table 6 showed high pyrite recovery, only the collectors of Examples 18-25 obtained the highest I values of this ore at pH 4.0. In addition, cu 25 as shown in Example 22 of Table 6, a mixture containing only 36% hydrocarboxycarbonylthionocarbamate collector resulted in higher copper recovery than standard collectors.

Examples 26-27 The following examples were performed using the same South American ore as in Examples 15-25 to study the pH sensitivity of the collectors of the invention and their effectiveness under strongly acidic conditions. The flotation conditions and reagents of Examples 18-25 were used in the following experiments. The collector dose was 5 g / ton.

31 7716 9

Sulfuric acid was used to adjust the pH of the slurry to the indicated pH. Collectors were tested at pH 2.75 and 3.70 and the results obtained are shown in the following Table 7: 5 Table 7

Eg Collector pH HjSO ^ Recover- Cu-pi- Recover- k. taken up taken g / Cu,%% FeS2,% 26 N-ethylcarbonyl-O-2.75 8.0 79.9 7.5 59.3 isobutylthionocarbamate 27 3.70 5.2 80.1 8.4 81.0 RR Ethyl xanthene ethyl 2.75 8.0 24.5 2.6 11.8 Liformate (pure) SS - "- 3.70 5.2 19.1 2.5 8.5 15

The figures in Table 7 clearly show that the collectors of the present invention beat conventional neutral xanthogen formate collectors even under strongly acidic conditions by a clear margin, and that the hydrocarboxycarbonylthionocarbamate collectors of the invention are generally not pH sensitive. As can be seen from Table 7, under identical conditions, only 20-25% copper recovery was obtained with standard collectors, while the new 25 collectors of Examples 26 and 27 according to the invention had 80% copper recovery. This result probably depends on the significantly better hydrolytic stability of the collectors according to the invention compared to standard collectors. Examples 28-30 In these flotation experiments, the same U.S. Lou-30 female ore containing 0.458% copper and 2.2% pyrite was used as in Examples 2-14. As the frother, a tall oil / MIBC mixture of 50 g / ton was used. Sulfuric acid was used to adjust the pH to the indicated level; To obtain a pH of 4.0, about 1.7 kg / ton of sulfuric acid was added.

35 The efficiency and selectivity of the new collectors under acidic 32 771 69 conditions for this ore were evaluated using various standard collectors for comparison. The collectors tested and the results obtained are shown in the following Table 8: 5 Table 8 pH 4.0; sulfuric acid 1.7 kg / Storm; frother 1: 1 tall oil / KlBC 50 g / ton. Collector dose 0.01 mmol / ton (approx. 2 g / ton), unless otherwise stated

mention Dose TalteeJXJfcettu Cu-pi- Recovery I

Eg Collector no ^ t **% content taken% F * SO,% iO -------------- * -r ~ .- ΤΓ Sodium diisobutyl-j 0.01 4 .0 26.2 2.6 11 * 9 0.016 dithiophosphate w 0.03 4.0 53.1 4.5 34.5 0.030 W - "- 0.10 4.0 95.0 5.7 95.5 0.180 lii · # WW Sodium isobutyl 0.01 4.0 20.0 1.9 10.6 0.014 xanthate r> '; XX 0.03 4.0 51.4 3.5 33.1 0.028 Ή 0.10 4.0 94.9 5.1 99.5 0.020 20 ZZ Ethyl xanthogen 0.01 4.0 92.3 6.7 93.1 0.117 ethyl formate (batch 1) AAA Ethyl xanthene 0.01 3.7 65.2 5, 6 70.1 0.025 Ethyl formate (batch 1) pH 3.7 25 BBB Ethyl xanthene ethyl acetate, 01 4.0 91.0 6% 0 96.4 0.045 Liformate (batch 2) '' 1 '' COC Ethyl xanthene ethyl acetate * - 0.01 3 .9 54.1 4.4 59.0 0.019 liformate (batch 2) pH 3.9 30 DDD Ethyl xanthene ethylene, 01 4.0 94.8 5.3 93.4 0.249 lyformate (batch 3), r '' 'pure 28 N-ethoxycarbonyl-0.01 4.0 96.3 6.4 91.5 0.621 O-isopropylthione.

carbamate „_, Q Aoin? QQ

25 29 N-ethoxycarbonyl-O-0.01 3.9 94.9 5.8, »isopropylthionocarbamate, pH 3.90 30 N-ethoxycarbonyl-O-isobutylthionocarbamate'0.01 4; 0 97.5 5.9 97.4 0.426 33 7716 9

The results in Table 8 clearly show the superiority of the new collectors of the invention (Examples 28-30) over conventional collectors (Examples TT-DDD) when compared to copper recovery, copper content and selectivity factor. From the results in Table 8, it can be seen that the doses of water-soluble ionic collectors, dithiophosphate and xanthate had to be 10 times higher than the new collectors in order to achieve a high copper recovery rate (about 95%), which is still lower than that obtained with the new collectors ( 97%). Also, with a neutral collector, ethylxanthene ethyl formate, which is considered suitable for acid circulating flotation, the% recovery of copper is only 91-95% (Examples ZZ-DDD). In addition, the performance of this collector appears to be very sensitive to small pH variations; a slight decrease in pH from 4.0 to 3.9 or 3.7 significantly reduced the copper recovery from 92.3 to 65.2% (compare Examples ZZ and AAA) and 91% to 54% (Examples BBB and CCC). New collectors do not feel this way (compare pre-2Q signs 28 and 29). This unusual stability to pH fluctuations is a clear advantage for new collectors over conventional collectors, especially in factory conditions where pH fluctuations cannot be avoided.

Examples 31-32 It is generally believed that a neutral, oily collector is most effective when added to grinding instead of a flotation cell. This is usually true if the collector is very sparingly soluble and insoluble in water. Although the hydrocarboxycarbonylthio-30-carbamate collectors of the invention are virtually insoluble in water, they are readily dispersible. The experiments were designed to assess whether the dispersibility of these collectors would provide additional benefits for their use. In this connection, two flotations were performed, one by adding 50% of the collector to grinding and the other by adding 50% of the collector during flotation, and the other by adding 50% of the collector to the cell in the first flotation step and the other 50 * in the second flotation step. The same ore and collectors were used as in Examples 4-14. Flotations were performed and copper was determined from the concentrates and concentrates. The results obtained are shown in Table 9 below:

Table 9 1Q Eg Collector / addition - Dose Step 1 Step 2 Cu-tal- Cu-dark as rack mol / t Recovery- Recovery other taken taken tot. % (size),

Cu,% Cu,% _% 31 N-tecycarbonyl-O- 0.01 22.1 74.1 96.3 5.6 1 sobutyl thionacarbamate added 50% to grinding + 50% to step 2 flotation 32 N- etcxicarboryl 0.01 84.6 12.9 97.5 5.9 O-isobutylthionecarbamate added 50% to step 1 flotation and 50% to step 2 foaming

The results in Table 9 unexpectedly show that adding the collectors of the invention to the flotation cell 22 instead of grinding gives better results. Comparing Examples 31 and 32, it can be seen that when the collectors according to the invention are added to the cell alone, the total copper recovery increases, i.e. the result of Example 32 to 97.5% recovery from 96% recovery of Example 31. This difference may be explained by the fact that although the new collectors according to the invention are selective for iron, some collector power for copper may be lost in the grinding step due to adsorption of the collector in an iron steel ball mill or 35 grinder.

It is interesting to note that the use of the new collectors according to the invention when the addition to the flotation cell and not to the grinding results in an unexpected and somewhat dramatic improvement in the flotation kinetics. In particular, the improved kinetics are manifested in better recovery of 5 copper in the flotation step. As will be readily appreciated by those skilled in the art, the ore is foamed to obtain a preconcentrate and concentrates. The tailings are usually discarded and the preconcentrate is re-ground, conditioned and then subjected to more accurate foaming. This results in a cleaner concentrate and cleaner concentrates. The purer concentrate is usually dried and sent to a melting step or other further enrichment step. The purer enrichment slurry is then subjected to rinsing foam after re-coding. The rinse concentrate can then be combined with the purer concentrate. The rinsed tailings can be combined with the main feed for pre-flotation. In the reconditioning steps between pre-, purification and rinsing flotation, the pH of the slurry is generally increased to provide better selectivity and better copper recovery.

As can be seen from the figures in Table 9, the new collectors according to the invention obtained a much higher and faster collector activity when added to the cell than in the case where 50% of the collector was added to the grinding.

When the collector was added only to the cell (Example 22), about 84% of the copper was recovered in the flotation step 1, whereas in Example 31 only 22.1% of the copper was recovered in the flotation step 1. Thanks to the improved flotation kinetics, less pure concentrate containing more copper is obtained in step 1, followed by the practice that reagent consumption can be reduced by appropriately controlling the supply of reagents and possibly reducing the number of cells in the flotation cell. The factory's production can also be increased 35.

36 771 69

Examples 33-34

The following flotation examples were performed using copper-molybdenum ore from the U.S. redemption containing 0.458% copper and 2.2% pyrite. A 50:50 tall oil / MIBC mixture of 50 g / ton was used as the frother. Collectors were added at 0.01 mol / ton (about 2 g / ton). Lime was added to adjust the pH of the slurry to about 8.3 to 1.76 kg / ton. Conventional flotation of this ore requires about 4.412 kg / ton of lime to adjust the pH to 11.2-11.3. 1Q The collectors tested and the results obtained are shown in the following Table 10:

Table 10 pH 8.3, lime 1.76 kg / ton; frother tall oil / MIBC 15 1.1, 50 g / ton; aggregator dose 0.01 mol / ton (about 2 g / ton ni)

Eg Collector Recover- Cu Contain- Recover- I

taken by mouth,% taken Cu,% FeS2 '% EEE Ethyl xanthene 84.2 5.4 28.6 0.287 Ethyl formate (batch 3, pure) FFF Ethyl xanthene 85.9 6.7 29.6 0.352 ethyl formate (batch 2) 33 N-ethoxycarbonyl-90.6 9.3 38.9 0.688 O-isopropylthiocarbamate 34 N-ethoxycarbonyl-O-90.8 7.7 46.2 0.633 isobutylthionecarbamate

The results in Table 10 show that the hydrocarboxycarbonylthionocarbamates of the invention 30 are clearly superior to a conventional ethylxanthene ethyl formate collector. Examples 33 and 34 obtained about 5% higher copper recovery than the conventional collector in Examples EEE and FFF, and also copper-35 concentrations and Icu values were significantly higher.

37 7 716 9 These results were obtained with a 60% lower consumption of lime, ie 1,76 kg / tonne, when the amount of lime required for the use of conventional collectors of this ore was 4,412 kg / tonne.

Examples 35-36 5 Similar foaming experiments were performed using the same ore, the same foaming and collector dosage as in Examples 33-34, except for the pH of 7.2 used in the experiments. To achieve this pH, about 1.18 kg / tonne of lime was added, corresponding to a 73% reduction in lime consumption to 10 standard lime doses, compared to 4.412 kg / tonne, which is the pH value used in the previous flotation of this ore 11, 2-11.3 to achieve. The collectors tested and the results obtained are shown in the following Table 11:

Table 11

pH 7.2, lime 1.18 kg / ton; frother tall oil / MIBC

1: 1 50 g / ton; aggregator dose 0.01 mol / ton (approximately 2 g / ton)

Eg Collector Recover- Cu Contain- Recover- I

taken by mouth,% taken _Cu,% _FeS2,% _ 20 QQG Sodium diisobutyl 61.9 6.1 20.6 0.055 Lithithiophosphate HHH Sodium isobutyl 45.6 4.9 14.2 0.029 Xanthate III Ethyl xanthogen 83.3 9.8 24.6 0.270 Ethyl formate „(batch 1) JJJ Ethyl xanthene- 86.0 7.7 30.8 0.351 ethyl formate (batch 2) KKK Ethyl xanthogen- 86.0 7.6 33.5 0.339 ethyl formate (batch 1) 3, pure) 30 35 N-ethoxycarbonyl-90.9 7.4 47.6 0.632 O-isopropylthicocarbamate 36 N-ethoxycarbonyl-89.9 7.4 54.8 0.444

Oisobutylcycno carbamate 35 As the figures in Table 11 show, even at pH 38,771 69 7.2, the new, improved hydrocarboxycarbonylthionocarbamate collectors of the invention provide better metallurgical performance than conventional collectors (Examples GGG-KKK) when compared to better copper 5 recovery 4-45% higher than that obtained with conventional collectors, higher Cu content of the concentrate and higher I, values, cu

Example 37

Similar experiments were performed using the same 1Q ore, the same frother and collector dosage as in Examples 33-36, except for the pH of 10.0 used in the experiments. Lime was added at 2.75 kg / tonne, corresponding to a 38% reduction in lime consumption compared to the usual amount previously used, 4.412 kg / tonne. The results of the experiments performed at pH 10.0 are shown in the following Table 12:

Table 12

pH 10.0, lime 2.75 kg / ton; foam - tall oil / MIBC

1: 1 50 g / ton; aggregator dose 0.01 mol / ton (approximately 2 g / ton)

2Q Eg Collector Recover- Cu Contain- Recover- I

taken by mouth,% taken 01

Cu,% FeS2,% r.TJ. Ethyl xanthene 90.8 6.5 76.0 0.284 ethyl formate (pure, batch 3) 25 MM Ethyl xanthene 88.7 7.3 52.6 0.373 ethyl formate (batch 2) 37 N-ethoxycarbonyl 89.7 8.2 31 .8 0.640

Oisdbutylthianocarbamate 30 The results in Table 12 show that at pH 10.0, the hydrocarboxycarbonyl collector of Example 37 obtained about 1% lower copper recovery than the conventional collector of Example LLL, however, the selectivity to pyrite was clearly better, i.e. in Example 35. 37 Recovery of 32% pyrite compared to 76% 39 7 7 1 69 in Example LLL, which difference is also reflected in the higher I value, cu

Examples 38-43

In the following examples, copper-molybdenum ore from the southwestern part of the USA containing about 0/778% copper and about 5.7% pyrite was used. This ore was one of the most difficult ores tested for its complex mineralogy, low total copper recovery, high. . lime consumption and flotation problems. The ore contained mainly chalcosite, however, the pyrite contained in the ore was rich in chalcosite and covellite as peripheral and pyrotal ores. Separation of pyrite by crude foaming was thus not possible and was not attempted.

880 g of ore was conditioned with 500 g / ton of ammonium-15 sulfide, and the ore was ground for 6 minutes at a solids content of 55.5% to give a slurry with a size distribution of 17.4% + 65 mesh, 33% + 100 mesh and 47.4% -200 mesh. The slurry was codified at 1500 rpm and a solids content of 20.4%.

In a standard treatment using an N-ethyl-O-isopropylthionocarbamate collector, the pH of this ore is 11.4-11.5. Approximately 3.07 kg / tonne of lime is required to reach a treatment pH of 11.4-11.5. The standard frother is cresylic acid, which is added at about 150 g / tonne.

25 The collectors tested, their doses and conditions, and the results of the experiments are shown in the following table 13: 40 7 71 6 9

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I I 10 k «» # · i> ^ · »« m k ^ ^ ^

O Sh Ai O O O O co CO O OOO OO

Ό s co cd uo m

H> m O 000- ^ 0 OOO OO

* Ä ·. ··, ·· »f“ H l — M «. mm 0tm ^ ** O · »^ oo oo o os · —ii— · oo oo σ 'oo oo σ'

Ai <#> cd

^ r- S

, —1 * 2 ι / 'O SO O uo O uo Oinuo O Ό »

3 in 3 O r-I o ^ O <—1 O rH o O t — IO

(Q r-l I — 1 CM r-H CM rH CM rH CM rH ι — I CM rH

FH II O (0 -P * ► K f fc, N r r λ · * s · · m QS o oooooo ooo oo cm -p Bq

w <E

03 O

Ph M

-P I

- S § s 1 i fe f ^ O · § 8 * -8 o> -p »to

ϋ · Η Ή * H

? f Φ2 Φ d ® g -h m -η t_> a p d g r-j

•• · m ?. § · gj i'S

s "8 5 p 11 f 2 M 11 11« s * S 3 5 $ yav Ϊ a = = 1 = 1 = = 1 2 g = = 1 I -δ = 1 = 'dOj ZAil IIII Zftl I a -p II 6 0 6 s * wi § 6 § ISSS § 3 513 41 77169

The results in Table 13 show that the new collector according to the invention (Examples 38-40 and 42-43) was metallurgically superior at pH 8.0 and 9.0 than the standard collector (Examples NNN-SSS) at pH 11.5. When using new collectors at pH 8.0 and 9.0, an acceptable metallurgical effect is obtained with significantly reduced lime consumption (7.5% of total lime consumption "to obtain pH 8.0 and 25% to obtain pH 9.0). and at a lower collector dose (0.105 mol / ton when a standard 10 di-collector at pH 11.5 is required at 0.210 mol / ton.

Examples 44-47

The following experiments used South American Cu-Mo ore containing 1.844% Cu and 4.2% pyrite. The copper minerals were mainly chalcosite, cal-15 copyrite, covellite, and bornite.

510 g of ore was wet milled for 7.5 minutes at 68% solids to a slurry with a size distribution of 24.7% +65 mesh, 38.3% +100 mesh and 44% -200 mesh. In all experiments, 2.5 g / ton of di-20 sec-butyl dithiophosphate was added to the milling. Lime was also added to the grinding to obtain the desired pH in the flotation. The slurry was transferred to a flotation cell and conditioned at 1100 rpm and a solids content of 32%.

The same ore was used for the second flotation tests in a slightly alkaline cycle with new collectors according to the invention. The standard collection program is: about 30-40 g / ton of sodium isopropyl xanthate and 2.5 g / ton of di (sec-butyl) dithiophosphate, and the standard flotation pH is 10.5. The standard foam is 1: 1: 1 poly-30 propylene glycol monomethyl ether / MIBC / tall oil 20-25 g / ton. The collectors tested and the results obtained are shown in the following Table 14: 42. 77169 r -'- oo I — I r- ~ O '3-0- in in oo <r oo o

O O © O CM i-H i-H

ij ^ W · 1. ^ o o o o o o o lii +5 Jj tlO'-Hf-ICMO 1- · CN (Tv H QJ CO I »►» · 1.

(N co oo i £> oo in m oo HQin ^ oooooor 'in inoo 1 «h Oiio oovooon is mo <r C fg - - - - - - - d Jj5 vovomm ^ f r-> vo <?

Q O 0} f “H i-H r-H rH r-H r-H r-H r-H

-P

"So rgao on in <r cni on in O 'm ^ ^ 1. 2-. 1- - v. -P -po © m <j- <r vo si-σ' in (e'dWJvor ^ oooooo oo oo oo ™ 0 eh 0 0 3 'cm η' Π I — I σ 'on is in cm cm <r “· · ϊ — 1 3 2+) r — I cm m in cm ih> -h cm <f ιΗ (0 Π \ - - »: 0 4->" 3 Cn ooooo o oo

II> 1 K M

c o · 1 cm .a in

rH CO g +> o O in in O O OO

0) --N, fQ · ^ 1 1 »1 · 1» · -

O CO On o O ON 00 COON

\ J flQ ^ f — 1 r-H

** HH I

3 H 3G

H m ^ o ^ n oo O> in in cm in m

ig »in L O 'O' O 'CM cm or CM CM

£ h r-H csl -¾ nH rH r-H rH r-H O rH rH

I M s “K- <H M S, K K Λ il> b g \ ooooo o oo _ o o. 5 ¾ => q 5 8 a to 1 · 63

^ · Η I

I -j ^ to:: = Ϊ1 ^ · ^ C 5 o -p1 in C § Si & Λ I h - .¾ 1 = = = I I = =

«J T3 O V o Q

§ 'S, 2 ^ 3 Tj 3 c £ 2 g = = = IS ixl = = • d a 3 1 e -p h gi 111 X +3 v d <p 6 & I = = = U L · -. -.

2 6 -H H O> X <r in voc1 tn H o> x -J- <r <r mt

H H O> X

43 7 7 1 69

The results in Table 14 show that the new collectors according to the invention are excellent in both copper recovery (no losses in Cu content) and pyrite control, while the consumption and dose of collector are lower than when using standard collectors. It is important to note that the standard collector provided low, unacceptable copper recovery at pH 8.0 even at the higher collector dose.

10 Example 48

The previous examples have shown that the novel, improved hydrocarboxycarbonylthioocarbamate collectors of the invention have better performance with low lime consumption or no lime addition than a large number of conventional collectors when used for pre-foaming or first stage flotation of various ores. In practice, the preconcentrate is purified in one or more steps to obtain a high degree of copper mineral or copper-molybdenum mineral concentrate for further processing to produce metal.

The following examples illustrate the use of new improved ... hydrocarboxycarbonylthionocarbamate collectors in cleaner flotation systems to provide high grade copper concentrates for use in smelters and the like. In the following examples, the same ore was used as in Examples 44-47. The first stage flotation, i.e. pre-flotation, was performed according to the methods of Examples 1-47. The concentrate was filtered, dried and then ground to a slurry with a solids content of about 30%. The pH of the re-ground slurry was adjusted with lime and, if necessary, more collector and frother were added. The re-ground slurry was conditioned and foamed as above in the preparation of the pre-concentrate to give a purer concentrate and cleaner concentrates. The purer tailings were gradually rinsed at 77169 44 higher pH values, possibly with the addition of an additional collector and a frother, and finally rinsed at a pH above 11.0 with the addition of an additional collector to recover all remaining copper minerals, and the product of each step was analyzed separately. .

The following table shows the results of experiments in which Malmi was subjected to first stage flotation and second stage flotation, i.e. cleaner flotation, using a standard collector, sodium iso-10 propyl xanthate at pH 11.0, for comparison. In Example 48, additional collector was added in the cleaner flotation of the second stage because it was found that the amount added in the pre-flotation was not sufficient for the cleaner flotation. Sufficient transfer of the standard collector was made in the second stage flotation, so no collector was added to the second stage control. The results obtained are shown in Table 15:

Table 15

Cleaner flotation. Main analysis: Cu = 1.85%,

FeS2 = 4.2%; frother - 1: 1: 1 Dow 250 / MIBC / tall oil 2Q Eg FFFF 48

Sodium isopropyl N-ethoxycarbonyl-O-xanthate isobutylthionocarba- ___received Ά. First stage Pre-foaming 25 Collector dose, g / ton 30.0 12.8 pH 10.5 8.2 Lime used, 0.608 0.108 kg / ton

Recovered,%

Cu 86.9 88.1 30 FeS2 90.9 63.7

Md 64.0 55.6

Concentration,%

Cu 18.30 21.30

Fe 20.70 16.40 «77169 B. Second stage FFFF_48

Cleaner flotation

Collector dose g / tonne - 4.2 pH 11-11.6 8.7-9.6 5 Lime used, 0.343 0.118 kg / tonne

Content of purer concentrate,%

Cu 39.4 41.9

Fe 22.2 18.6 MO 0.56 0.58

Added collector total, 30.17 17.0 g / ton

Added lime to a total of 0.951 to 0.226 tea, kg / ton

Added foam 38.0 39.0 ^ total, g / ton

The results in Table 15 clearly show that the novel, improved hydrocarboxycarbonylthionocarbamate of the invention is clearly better in performance in both pre-foaming and cleaner flotation than the standard collector used as a reference. In particular, the Cu content of the purer concentrate was about 2.5 percentage points higher in Example 48 than in the comparative example (former 41.9%, the latter 39.4%) and the Cu content of the first 25 concentrate in Example 48 was similarly 3 percentage points higher. higher than in the control experiment. The total size and dose to obtain this Cu content was only 17 g / ton in Example 48, compared to 30 g / ton in the comparative experiment. Example 48 shows that better copper recovery and concentration is obtained by using collectors according to the invention with collector cost savings of about 45%. Example 48 shows that the collector according to the invention gives better copper recovery and Cu content in a cleaner flotation cycle using less lime, i.e. 0.226 kg / ton, than in the control experiment where lime 46 7 71 6 9 was used at 0/951 kg / ton. In lime consumption, this corresponds to a saving of 75%. The purer concentrate of Example 48 contained nearly 4 percentage points less iron than the concentrate obtained using the standard collector (former 18.6%, the latter 22.2% Fe), indicating that the collectors of the invention are clearly more selective for pyrite than the standard collector. Better selectivity of the collectors according to the invention was also found in the low pyrite recovery of the pre-foam (63.7%) compared to the pyrite recovery of the standard collector (90%). In addition, the amount of copper recovered by the collector of the invention in the pre-foaming of Example 48 was greater than that obtained by the standard collector, however, the dose of the collector of the invention was less than half the dose of the standard collector in the pre-foaming.

Example 49

Koostesulfidivaahdotus

The following flotation experiments used copper-zinc-pyrite-pyrrotite ore in the southeastern United States. It contained 0.5-0.7% copper as chalcopyrite, 0.9% zinc and 30-35% iron as pyrrotite and pyrite. The ore also contained a large amount of carbonate minerals as a side rock, such as calcite, dolomite, etc., in addition to the usual silicate or siliceous side rock types.

25 All experiments used a product ground in a ball mill at an industrial plant. The slurry contained ore particles, of which about 40% was -200 mesh. Approximately 4 liters of slurry was modified with 0.5-5 kg / ton of concentrated sulfuric acid at a solids content of 25% for 30 seconds at 3 rpm 1800 rpm. The collector and frother were then added and the slurry was conditioned for 2 minutes. Flotation was performed for 4 minutes with air and stirring at 1800 rpm, and the first stage concentrate was collected. The slurry was then conditioned for 30 seconds with the addition of more foam, and the second stage flotation concentrate was collected over 47 7 7 1 69 4 minutes. The first and second stage concentrates and concentrates were filtered, dried, and assayed for copper, iron, sulfur, and zinc.

The results are shown in Table 16. The usual size was sodium ethyl xanthate and the foam was polypropylene glycol (OP 515, Oreprep Inc).

48 771 69 and r- »r ** and CO H · - '· - r- m un m en co ro m O o O o

C <—i i — i r- O

N -

H r * H Q rH

(#> O CM <Γ Ό

Oi I »M ti to <r <r <r g and and un and • 3 S Q esi and m 00" H 3 CM CM rH o O 'S u -

t rH rH · l-H rH

II

00 m ^ CO r- CM rH (n CM - * h “· ^ O» vO Ό 00 fT) O 'O' O '

^ C rH Ό CO rH

11 -d ^ uO cm co and 4) g and and <r and

2 S B

7j _ ^ ti O lC <r rH CM

S JS 8 60! * o- rT oo * · nT

2 .m g o> oo oo θ '0 (0 rH 00 | 3>> "3 =“ i ° Z>' Σ * c e tn O Ό «O cm m ^ ^ O 'O' (0 * 4-1 o

M N | <r 4J

5 T9 oy) 00 σ '<r Γ "2 vo (8 co Q« om cm oo M o ai · 4 *' »<Γ <Γ -Γ 8 11 '* 3 UJ Ό CM O © CJ m ^ - * ~ a Ή. (h. rH o ΓΗ • HS oi o O Γη Γη

03 (3 —I rH

S 8 s § -d h H i 1 M = = & § S ta # § '3 8 s. a | g «f - - agio

Id lii I Is · · m

O X i-H ON

(Λ O X rn <j ·

W © X rH

O X rH

49 7 7 1 69

The results in Table 16 show that the hydrocarboxycarbonylthionocarbamate collectors of the invention in Example 49 provide substantially the same metallurgical result in composite sulfide flotation at about 25% lower dose and about 62% lower sulfuric acid utilization than using the conventional GGGG-IIII collector of Examples.

The above examples illustrate the significant improvements and advantages achieved with the novel, improved hydrocarboxycarbonylthionocarbamate collectors of the invention over many conventional collectors known in the art.

As mentioned above, the process can also be applied using mixtures of two or more hydrocarboxycarbonylthionocarbamate collectors and a mixture of at least one hydrocarboxycarbonylthionocarbamate collector with another known collector selected from, for example: a) xanthates or xanthate esters of formula:

O O

o "_ I o n q

R O-C-S M and R O-C-SR

b) dithiophosphates 25 ® (R 0) -P-S M; c) thionocarbamates

S

o ·· g 30 R O-C-NHR; d) dithiophosphinates

S

p n i (R6) 2P-S-M +; (E) dithiocarbamates and their derivatives, having the following formulas:

R11 S R11 S

R8N-C-S'M + and R8N-C-S-R9 (f) trithiocarbonates and their derivatives of the formulas

S S

R ° S-C-S M and R S-C-S-R

g) «arcapans 10 R10SH;

O

wherein in the above formulas (a) to (e) R is C 1 -C 6 alkyl, R is C 1 -C 6 alkyl, aryl or benzyl, and R 10 is hydroxy or R and in the formula ( f) R is C 1 -C 4 alkyl.

Instead of copper minerals, the process according to the invention can be used for the enrichment of other sulphide minerals and metals from sulphide ores, for example lithium-20, zinc, nickel, cobalt, molybdenum, iron, as well as precious metals such as gold, silver, platinum, palladium, rhodium, iridium and osmium enrichment. One skilled in the art can make all such modifications and alterations without departing from the scope of the invention as defined in the following claims.

Claims (7)

51 7716 9 A collector reagent for flotation of sulfide minerals, characterized in that it comprises at least one hydrocarboxycarbonylthionocarbamate which is N-ethoxycarbonyl-O-isobutylthionocarbamate; N-ethoxycarbonyl-O-sec-butylthionocarbamate? N-ethoxycarbonyl-O-n-amyylitionokarbamaatti; N-ethoxycarbonyl-O-isoamyylitio nokarbamaatti; N-ethoxycarbonyl-O-phenylthionocarbamate-10 ti; N-phenoxycarbonyl-O-etyylitionokarbamaatti; Fenok-N-ethoxycarbonyl-O-isopropyylitionokarbamaatti; N-fenoksikarbo-phenyl-O-n-butyylitionokarbamaatti? N-phenoxycarbonyl-O-isobutyylitionokarbamaatti; N-phenoxycarbonyl-O-sec-butyl-tyylitionokarbamaatti; N-phenoxycarbonyl-O-n-amylthiono-carbamate; or N-phenoxycarbonyl-O-isoamylthionocarbamate. 2. A process for the selective enrichment of base sulphide minerals from base metal sulphide ores by inhibiting side rock sulphide materials at a pH of less than 10.0, characterized in that a) the ore is made into a fine aqueous slurry with a pH of less than 10 containing minerals suitable for mineral release , b) the ore slurry is prepared by adding effective amounts of a blowing agent and a metal collector containing at least one hydrocarboxycarbonylthionocarbamate of the formula OHS η in «R OCNC-OR 30 1. wherein R is C 1 -C 6 alkyl or aryl and R is C C8-alkyl, and c) the sulfide minerals of the base metals are subjected to flotation. Process according to Claim 2, characterized in that the metal collector is added in an amount of 52 7 7 1 6 9 of about 0.0025 to 0.05 kg / ton of ore. A method according to claim 2, characterized in that the pH of the aqueous slurry is about 3.5-10.0. Process according to Claim 2, characterized in that the metal collector has the formula ethyl-2 and isopropyl. Process according to Claim 2, characterized in that the metal collector has the formula ethyl and R is isobutyl. 7. A process for the co-enrichment of sulphides from complex sulphide ores, characterized in that a) the ore is made into a fine aqueous slurry containing ore particles of a size suitable for the release of minerals and having a pH of about 6.0-9.0, b) the ore slurry is prepared by adding an effective amount a blowing agent and about 0.05-0.5 kg / ton of a metal collector containing at least one hydrocarboxycarbonyl lithium carbamate having the formula 20 OHS