FI77792C - Neutral sulfide collectors and foaming procedures. - Google Patents

Description

77792

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

Flotation is one of the most widely used methods in the enrichment of ores containing valuable minerals. It is used in particular for separating finely ground valuable minerals from side rock or for separating valuable minerals from each other. The method is based on the affinity of appropriately treated mineral surfaces for air bubbles. In flotation, foam is formed by introducing air into a mixed aqueous slurry of finely ground ore containing a flotation agent. The main advantage of flotation separation is that it is a relatively efficient operation 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 reagent 25, or reagents called aggregators, which selectively render the sulfide mineral separated from other minerals hydrophobic. The flotation separation of one type of mineral from another thus depends on the relative wettability of the mineral surfaces by means of water 30. Typically, the free surface energy is intentionally reduced by causing the surface to adsorb heteropolar collectors. 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.

2 77792

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 but not so 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 contraceptives, activators, pH adjusters, dispersants, deactivators, etc. Often, the modifier performs several functions simultaneously. According to the current theory and practice of sulfide flotation, the efficiency of all flocculant categories 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 soo-25 da and, to a lesser extent, sodium hydroxide. However, lime is by far the most widely used in sulphide foaming. For example, copper sulfide flotation, which is the major sulfide flotation industry, uses lime to keep the pH above 10.5, more usually above 11.0, and often as high as 12 or 12.5. In known sulfide flotation processes, the pH of the slurry must be adjusted in advance to 11.0 or higher not only to press known side iron ferrous mineral sulfides such as pyrite and pyrrotite, but also to improve the dithiochosphates of most conventional sulfide collectors such as xanthates, xanthocarbates. performance. The cost of adding lime becomes quite high, and the management of the plants is interested in flotation methods which 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 flotation circuit methods are particularly desirable because sludges can be easily acidified by the addition of sulfuric acid, and sulfuric acid is obtained in many plants as a by-product of the smelter. Thus, flotation methods that do not require prior pH adjustment or that can be pre-adjusted to neutral or acidic using less expensive sulfuric acid are more advantageous than current flotation methods because current flotation methods favorably adjust the pH to high time values of at least 15 times. to about 11.0, using lime, is more expensive.

To better clarify the current problems, it should be noted that in the 1980s, the U.S. copper and molybdenum mining industry used nearly 250 million kg of lime. This amount of industrial lime represented almost 92.5% by weight of the total amount of reagents used, and the monetary value of the lime used was about 51.4% of the total cost of reagents in this industry, i.e. more than 28 million US dollars.

As mentioned above, the lime consumption of different plants can vary from about 0.5 kg per tonne of treated ore to as high as 10 kg. In certain regions, such as South America, lime is poorly available and the transport and / or import of lime has significantly increased costs in recent years. Another problem with the known strongly alkaline methods is that the addition of large amounts of lime to obtain a sufficiently high pH in the flotation plant of the plant produces scum, the frequent removal of which requires the plant to be shut down for cleaning and thus increases costs.

It is thus obvious that it is highly desirable that the need to add lime to the sulfide flotation process could be reduced or eliminated altogether, resulting in significant savings in reagent costs. In addition, the reduction or omission of lime in the sulphide ore treatment may have other advantages, since this 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 biggest problem with these conventional sulfide collectors is that at pH values below 11.0, pyrite and pyrrotite can be poorly printed. In addition, lowering the pH also reduces the collection capacity of these sulfide collectors, making them unsuitable for mildly alkaline, neutral, or acidic environments. This decrease in the collecting capacity 15 when the pH value drops below e.g. about 11.0 results in the dose of the collector having to be increased several times, which is less economically pleasant. The decrease in the activity of the collector as the pH decreases may be due to many factors. The collector may affect 20 different sulfide minerals at a given pH in 20 different ways. On the other hand, at low pH, the poor stability of the solution, such as the poor stability of the xanthate and trithiocarbonate solutions, may explain the poor size and activity observed.

25 In an effort to remedy the above shortcomings, neutral derivatives of xanthates, alkyl xanthene alkyl formates, were developed, which can be represented by the following general formula:

S O Il II

30 RO-C-S-C-OR '

Alkylxanthene alkyl formates are patented as sulfide assemblies in U.S. Patent No. 2,412,500. Other structural modifications of this general structure were disclosed later. 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 include halogen, nitrile and nitro groups. U.S. Patent 2,608,814 describes bis-alkylxanthogen formates for use as sulfide builders. These modified structures have not achieved the same commercial use as unmodified structures. For example, alkyl xanthene alkyl formate is currently commercially available under the tradename "MINERAC A" from Minerec Corporation. "MINERAC A", which is ethylxanthene ethyl formate, as well as its higher homologues, do not perform completely satisfactorily at pH values below 11.0, but their aggregating force and pyrite printing leave much to be desired.

One class of sulfide collectors that has to some extent achieved commercial success in flotation comprises oily sulfide collectors containing di-alkylthionocarbamate or diurethane compounds of the general formula:

S

II

RO-CNHR '20 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 25-mercaptan, which is an air contaminant that is expensive to handle. U.S. Patent 3,907,854 describes 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. Patent 3,590,998 discloses a thionocarbamate sulfide scavenger compound in which the N-alkyl substituent is attached via alkoxycarbonyl groups. In the method described, expensive amino acid esters must be used for the thioester substitution reaction of xanthate pathways. By-products of this process are either methyl mercaptan or sodium thio-6,77792 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 falls below certain values.

It is therefore an object of the present invention to provide a new, improved sulphide collector and a 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 new, improved sulfide collector and flotation process for enriching sulfide minerals by selectively recovering precious metal sulfides and selectively pressing 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 collector 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 specially designed for low pH flotation.

The invention further relates to a new, improved process for the selective foaming of sulphides in an acid-30 cycle in which cheap sulfuric acid is used to adjust the pH.

One object of the invention is to provide a novel, improved collector composition for the selective enrichment of ore containing sulfide minerals by pressing pyrite and other siliceous sulfides or non-sulfides, which collector composition contains at least one hydrocarboxycarbonylthiourea compound of formula 7,77792

OHS

3 II 1 11 2

R O-C-N-C-NHR

2 3 wherein R and R independently of one another denote a saturated or unsaturated alkyl group. Particularly preferred hydrocarboxycarbonylthioureasulfide scavengers for use in the process of the invention are those compounds of the above formula wherein R is C 3 -C 8 alkyl or allyl and R is C 1 -C 6 alkyl.

In general (without limitation), the novel, improved hydrocarboxycarbonylthiourea collectors of the invention can be used per tonne of ore of about 0.0025 to 0.25 kg, preferably 0.005 to 0.15 kg, to efficiently and selectively recover metals and minerals from metal sulfide ores while selectively by pressing pyrite and other side rock sulfides or non-sulfides. The new, 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 at pH values of about 3.5 to 11.0, preferably about 4.0 to 10.0.

According to one embodiment of the invention, there is provided a new, improved method for selectively enriching an ore containing sulfide minerals by pressing pyrite and other side rock sulfides and non-sulfides, wherein preferably flotation on pyrite and other side rock sulfide minerals or non-sulfides by a flotation process, wherein the metal collector contains at least one hydrocarboxycarbonylthiourea compound of the above formula.

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

Other objects and advantages of the invention will become apparent from the following detailed description and exemplary embodiments illustrating the invention.

According to the present invention, metal sulfide minerals are recovered by a flotation process in the presence of a new sulfide collector containing at least one hydrocarboxycarbonylthiourea compound having the formula

10 0 H S

2 H I II p

R O-C-N-C-NHR

2 3 wherein R and R independently of one another denote a saturated or unsaturated alkyl group. Alkyl group 15 denotes straight-chain or branched, saturated or unsaturated, cyclic or acyclic hydrocarbon radicals containing hydrogen and carbon atoms.

In a preferred embodiment of the invention, the hydrocarboxycarbonylthiourea collectors of the above formula are preferably compounds in which R is C 1 -C 6 alkyl, for example methyl, ethyl, n-propyl, isopropyl, n-butyl, sec-butyl, isobutyl, n- amyl, isoamyl, n-hexyl, isohexyl, heptyl, n-octyl or 3 2-ethylhexyl, and R is C 1 -C 6 alkyl.

Some examples of compounds of the above formula for use as sulfide collectors of the invention are given below: N-ethoxycarbonyl-N'-isopropylthiourea, N-ethoxycarbonyl-N'-isobutylthiourea, N-ethoxycarbonyl-N'-allylthiourea.

Used in the flotation method according to the invention. hydrocarboxycarbonylthiourea compounds can be conveniently prepared without the formation of environmentally hazardous by-products. By first reacting the corresponding chloro-35 formate compound with ammonium, sodium or potassium thiocyanate to form an isothiocyanate intermediate according to the following equation (1): 9 77792 o o

3 H 3 II

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

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

0 O H S

3 I · 3 H I H 2 10 R -O-C-NCS + H2NR2 - »R O-C-N-C-NHR (2)

The corresponding chloroformates for the reaction with ammonium, sodium or potassium thiocyanate according to Equation (1) above can be prepared from the corresponding aliphatic or aromatic alcohols by reaction with phosgene according to the following Equation (3):

O O

3 * 3 11 R OH + Cl-C-Cl-> · R O-C-Cl + HCl (3) 3 20 where R OH is the active hydroxyl compound. By active hydroxyl compound is meant a compound having a hydroxyl group that readily reacts with phosgene to form the corresponding chloroformate compound.

Ethoxylated or propoxylated alcoholic chloroformates can be prepared by this method, e.g., according to the following reaction scheme:

Ö O

5 II 5 II

R (OCH 2 CH 2) n -OH + Cl-C-Cl-> R (OCH 2 CH 2) n> O-C-Cl + HCl (4) wherein R is C 1 -C 6 alkyl, and n is 1 to 4.

According to Equation (1) above, in the preparation of the new, improved hydrocarboxycarbonylthiourea collectors according to the invention, it can be seen that the only by-product in the reaction according to Equation (1) is harmless sodium chloride.

According to the present invention, the hydrocarboxycarbonylthioureas described above are used as sulfide assemblies in a new, improved process for the enrichment of 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 method for enriching minerals from base metal sulfide ores comprises, as a first step, reducing 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. the release size varies with different ores depending on several factors, such as the geometry of the mineral deposits in the ore, e.g. beverages, agglomerates, common matrices, etc. In any case, as is common in the art, the particle size reduction can be determined by microscopic examination. In general, a suitable particle size of 20 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 mesh 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 autogenic or semi-autogenous milling or milling may be used. The method used to grind the ore is not critical to the invention as long as particles suitable for efficient foaming are obtained.

The pre-adjustment of the pH is suitably carried out by adding the adjusting agent during the comminution step to the ore π 77792.

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 hydrocarboxycarbonylthioureasulfide collectors can be used in the process of the invention without prior pH adjustment, and flotation can generally be performed at the natural pH of the ore slurry, simplifying the process, reducing costs and reducing 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 to 11.0, and a particularly good enrichment result has been observed at pH values of about 4.0 to 10.0.

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 to 60% by weight of slurry solids, preferably 25 to 50% by weight and particularly preferably about 30 to 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 hydrocarboxycarbonylthiourea collectors according to the invention have a particularly good collector efficiency and at the same time a 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 method known in the art, if desired.

12 77792

Once prepared, the slurry is prepared by the addition of effective amounts of a blowing agent and a collector containing at least one of the hydrocarboxycarbonylthiourea 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 can be used in the method of the present invention. Some suitable blowing agents include Examples 10; straight-chain or branched lower 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, cresyl, and alcohol ethoxylates.

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 may also be deviated from.

The novel, improved hydrocarboxycarbonylthioureasulfide collectors used in the process of the invention are generally added in an amount of about 0.0025 to 0.25 kg / ton of ore, preferably about 0.005 to 0.15 kg / ton of ore to be treated. In flotation where pyrite and other side rock sulfides must be selectively prevented from being exposed to copper sulfide, the collector amounts used are generally 0.005 kg / ton to 0.025 kg / ton.

The slurry prepared according to the method of the invention, comprising an effective amount of a blowing agent and an effective amount of a collector containing at least one hydrocarboxycarbonylthiourea compound, is foamed by a conventional flotation method to obtain the desired sulfide mineral selectively on the surface foam concentrate and pyrite.

It has also surprisingly been found that, contrary to popular belief, a neutral, oily collector is effective when added to a grinding rather than a flotation cell, as the new, improved hydrocarboxycarbonylthiourea collectors of the invention are more efficient when added to a flotation cell and not to a flotation cell. 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 cell, they provide a better copper recovery result in the first flotation step as well as an improved copper recovery of 1 ° overall, suggesting improved flotation kinetics, which are fully described below. Of course, new, improved collectors can also be added to the grinding process by a conventional method, and even then an improved recovery of base metals is obtained.

The novel hydrocarboxycarbonylthiourea collectors of the invention and methods for 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 sulphide but also all sulphides contained in the ore, including sphalerite (ZnS) and iron sulphides, i.e. 25 pyrite and pyrrotite.

This applies in particular to certain massive or complex sulphide ores containing large amounts of ferrous sulphide minerals, such as pyrite and pyrrotite. Of 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. flotation, in which all sulfide minerals are foamed and assembled. Co-flotation is also preferably used to enrich precious metals from pyrite and pyrrotite minerals containing precious metals.

77792 14 However, these massive or complex sulfide minerals 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. Co-sulphide flotation of these ores cannot be successfully performed under conventional flotation conditions, i. at pH 10.0 because pyrite and pyrrotite become printed at high pH values. At pH values of 3.0 to 5.0, the co-flotation result is good when using conventional collectors such as xanthates-15, 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 when sulfuric acid is used alkaline earth metal carbonates such as calcite, dolomite and others. ores contain 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 pH values of about 6.0 to 9.0, the total sulfide flotation result using conventional collectors such as xanthates is below optimum.

Preparation I

Synthesis of Ethoxycarbonyl Isothiocyanate A 2-liter 3-neck round bottom flask equipped with a vertical condenser with a drying tube containing anhydrous calcium sulfate, 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 with stirring from a dropping funnel over 40 minutes. An exothermic reaction began. The mixture turned thick and yellow. At the end of the addition, the temperature of the reaction mixture was 77 ° C. The reaction mixture was stirred for 3 hours without external heating. The reaction mixture was then cooled to room temperature and the precipitate was filtered. The precipitate was washed with anhydrous acetonitrile. The filtrate and washings were combined and evaporated under reduced pressure. The liquid obtained as residue-10 was distilled using a fraction column. 86.9 g of ethoxycarbonyl isothiocyanate were obtained as a colorless liquid, b.p. 45 ° C / 11 mmHg or 48 ° C / 12 mmHg.

Preparation II

Synthesis of N-ethoxycarbonyl-N'-isopropylthiourea A solution of isopropylamine (7.1 g) in 40 ml of anhydrous ethyl ether was added dropwise over 30 minutes with stirring to a solution of ethoxycarbonyl isothiocyanate (Preparation I, 15.5 g) in 10 ml of anhydrous ethyl ether. . The reaction vessel was cooled in an ice-water bath. The reaction mixture 20 was allowed to stand at room temperature. When the reaction was complete, the solution was concentrated by evaporating most of the solvent under reduced pressure. The crystals were filtered and washed with hexane. The first batch weighed 8.1 g, m.p. 52.5-54 ° C. The second batch weighed 7 g, m.p. 52-54 ° C.

25 Preparation III

Synthesis of N-ethoxycarbonyl-N'-isobutylthiourea

A solution of ethoxycarbonyl isothiocyanate (Preparation I, 5.3 g) in 100 ml of petroleum ether (b.p. 35-60 ° C) was stirred while cooling in an ice-water bath. A solution of isobutylamine (3.9 g) in 50 ml of petroleum ether was added dropwise to the solution over 20 minutes. The reaction vessel was cooled during the addition in an ice-water bath. At the end of the addition, the vessel was removed from the ice-water bath and left at room temperature overnight. The solution was concentrated by evaporating most of the solvent. The concentrated solution was cooled in an ice-water bath. The crystals were filtered and washed with hexane. The product weighed 7.5 g, m.p. 50-52 ° C.

i6 11192

Preparation IV

Synthesis of a liquid product containing N-ethoxycarbonyl-N'-isopropylthiourea and N-ethoxycarbonyl-N'-isobutylthiourea To a 250 ml round bottom flask were added 11.86 g of n-octane and 11.86 g of ethoxycarbonyl isothiocyanate (Preparation I). The flask was immersed in an ice-water bath, and the mixture was stirred with a magnetic stirrer for 5 minutes. A solution of isopropylamine 10 (2.63 g) and isobutylamine (3.57 g) in n-octane was added dropwise from a dropping funnel to the solution.

The reaction vessel was immersed in an ice-water bath, and the reaction mixture was stirred during the addition of the amine solution. The reaction vessel was then removed from the ice-water bath and the reaction mixture was stirred at room temperature until the reaction was complete. The reaction solution was concentrated by evaporation of volatiles containing mainly n-octane to give 18.34 g of a liquid product. It contained N-ethoxy-N'-isopropylthiourea and N-ethoxycarbonyl-N'-isobutylthiourea in a molar ratio of 1: 1, and the content of these two thioureas was 87.4%.

The hydrocarboxycarbonylthioureas synthesized above were used as aggregators in the enrichment of several different sulfide ores, whereby their enrichment properties were tested at different pH values and compared with the corresponding properties of known sulfide-ores. In the following examples, other homologous and / or analogous hydrocarboxycarbonylthioureas may be used which can be readily prepared by substantially identical methods of preparation using the appropriate corresponding active amine compounds to give the desired R groups in the compound.

In all of the following examples, the following general manufacturing and testing methods were used:

Sulfide ores were crushed to a size of -10 mesh. Approximately 500-200 g of crushed ore was wet milled at a solids-35 content of 63% in a steel ball mill with a steel ball count of 10.7 kg for about 8 minutes, i.e. until the sludge size distribution had reached the indicated level, usually

II

17,77792 about 10-20% +65 mesh, 14-30% +100 mesh and 40-80% -200 mesh. Lime and sulfuric acid were used as modifiers to adjust the pH to the desired level. The blowing agent used was added to the grinding in some experiments and to the flotation cell in others. In certain experiments, 50% of the collector was added to the grinding, otherwise the collector was added to the flotation cell in the first and second stages of training.

The comminuted slurry, possibly containing blowing and collecting additives, was transferred to a rectangular Denver D12 flotation cell. The volume of the slurry was adjusted to 2650 ml by adding water so that the slurry contained about 30-35% solids and the surface of the slurry in the cell was about 2 cm from the edge.

A collector and / or blowing agent was added to the slurry with stirring at about 1400 rpm. The slurry was cured for 2 minutes, at which time pH and temperature measurements were performed. After 2 minutes of training, about 7 liters / min of air was drawn from the compressed air bottle. Foaming was continued for about 3 minutes, during which time the first concentrate was collected. The air tap was then closed, and more collector and blowing agent were added to the slurry, and the slurry was prepared for an additional 2 minutes. After a second 2 minute training step, the air tap was closed and the concentrate from the second step was collected. The flotation times had been predetermined so that the foam obtained at the end of the flotation did not contain minerals.

The concentrates and concentrates from the first and second stages were filtered, dried and copper, iron and sulfur were determined from 30 samples taken from them. Tap water of the required temperature was used in all experiments.

Examples 1-2

Acid flotation circuit

The following examples used South American copper-molybdenum ore containing 1.65% copper, 2.5% pyrite and 0.025% molybdenum. The ore contained chalcosite, chalcopyrite, is 77792 covellite, bornite, and some oxide minerals such as malachite and cuprite as copper minerals. Although the ore contained a large amount of chalcopyrite, a significant portion of it contained chalcosite and covellite as the edge mineral.

About 500 g of -10 mesh ore was wet milled at 63% solids for about 13 minutes in a steel ball mill with 5.3 kg of steel balls to give a slurry with a size distribution of 14% + 100 mesh and 62% -200 mesh. The natural pH of the ore slurry was about 4.0. In each of the 10 examples, 10.5 g / ton of diesel oil was also added. Experimented collectors were added to the flotation cell in the first and second stages of coaching. The flotation method described above was used in all flotation experiments.

The standard collectors for this ore are a mixture of ethyl 15 xanthene ethyl formate / diesel oil / MIBC 60:30:10 and 2.5 g / ton sodium diethyldithiophosphate. For further comparisons, experiments were also performed using diethylxanthogen formate in pure form and another standard collector, dialkylthionocarbamate. The first and second flotations were performed on Stan-20 standard collectors as well as the new, improved hydrocarboxycarbonylthiourea collectors according to the invention. The copper content, recovered copper and recovered pyrite were determined from the foam concentrate and enrichment slurry of each flotation step. In addition, a selectivity / performance factor was calculated for each collector tested.

The selectivity / performance factor was calculated according to the following equation: 30 i (100% recovered pyrite) cu (100% recovered copper) ^

The selectivity coefficient of copper suitably indicates not only the recovery of copper but also the selectivity for the printing of side rock sulfides such as pyrite and pyrrotite. For example, if, in the case of this ore, 90% copper recovery and 92% pyrite recovery are accepted as optimal, then the collector optimum selectivity coefficient for copper using this X, 77792 ore is 0.08. The collectors tested and the foaming results are shown in Table 1 below.

table 1

Acid flotation circuit 5 Ore analysis: Cu = 1.65%, FeS2 = 2.5%, pH 4.0; blowing agent = polypropylene glycol monomethyl ether 60 g / ton; sulfuric acid 5.0 kg / ton to adjust the pH to 4.0

Eg Collector Dose Recover- Cu content- Recover- I

10 g / t taken orally,% taken

Cu,% FeS2 '% A'. Standard part 10 46.7 4.5 21.1 0.028 B '. "20 78.9 7.0 80.9 0.043 C." 30 89.6 7.2 91.5 0.078 15 D '. "40 90.1 7.2 92.2 0.080 E '. Sodium diethyldithiophosphate 20 65.1 6.2 45.4 0.045 F1. Diethyl xanthogen formate 15 88.5 8.8 88.2 0.09 G'. "20 90.6 8.4 93.4 0.075 20 h". Isopropylethylthionocarbamate 15 76.3 7.8 83.0 0.030 1. N-ethoxycarbonyl-N1-isopropylthiourea 9.5 91.3 8.6 92.1 0.104 2. N-ethoxycarbonyl-N'- isopropylthiourea 10.2 90.7 8.3 93.5 0.075

a ^ 60:30: 10 mixture of ethyl xar.togenate formate / diesel oil / MIBC

From the results in Table 1, it can be seen that the new, improved hydrocarboxycarbonyl lithiourea collectors of the invention 30 (Examples 1-2) gave better metallurgical results at lower doses than using a conventional standard collector mixture (Examples A'-D1) or the sodium diethyldithiophosphate of Example E1. In Examples 1 and 2, the performance of the collector was better than that of pure diethylxanthogen formate (Examples F 'and G') or isopropylethylthionocarbamate (Example H '). Table 1 shows that the hydrocarboxycarbonylthiourea collector according to the invention provides better copper recovery at a lower dose than known collectors.

5 Only the I-values of the new collectors were at the required cu level.

Examples 3-6

Foaming at a Mildly Alkaline pH In these examples, 10 ores from the southwestern United States containing 0.867% copper and 7.0% pyrite were used. The main copper mineral was chalcopyrite, although the ore also contained some chalcosite, covellite, and bornite.

510 g of ore was ground at 65% solids for 8.5 minutes in a steel ball mill to give a size distribution of 5.8% +65 mesh, 19% +100 mesh and 53.3% -200 mesh. Lime was used to adjust the pH to slightly alkaline. A 70:30 polypropylene glycol / polypropylene glycol monomethyl ether mixture of 91 g / 20 tons was used as blowing agent. To make the comparison more significant, the collector doses are reported in equimolar amounts in mol / ton. The standard collector for this ore is sodium methyl xanthate, which has an optimum performance at pH 11.5. Collectors were tested at different doses and at different pH values, and the results are shown in Table 2 below.

21 777 92 (N ^ vsmco m «h ro m o o m m m com 2 rHCMincsip 10 mm u ^ ***.« * * ·> · · «· * *»

H OOOOO OO OO

+ J

-p

<u oj o cr> o o h ro 1 — I

4-> * · · * 0 rnvom <N <r> id m C <#> cn cn m cn (N cn m cn <u Φ *

+ J CN I — I CO

m o> H fc w • H § C3 w m m m ^ r- x> co ¢ 3 -H · * I * · I * * · ** ** O o VD CD OiCOOCncn

Jj -P «H» -H

a

^ I

3 ι-H U <#> crt <D 5

£ + J

• rH Q)

(rt + j o o i £> m ** ι-ι ro id rH

O V »» ** - 'li ¢ 3 oo ro co ο σ> σ> σ> o oo £ a) t'- oo oo co oo ooco co oo + J a) dp

0 + J

S 3 3 «id H °

<0 M

> 0) O O

O o o m o m oo oo “^ K OI O rH O rH O 'ο σ> o

^ ^ ft rH rH t — I rH rH rH

o 3

„(Rt Cfl-P" d * ^ Hf IN CN <ί N H1 CN

O '·' - IN IN CN l £> l £> CM VO CN lO

'+1 C3 rH rH τ-ί I — i O O rHO rHO

II 3 <Cg ooooo oo oo + J (N +1 cn a) £ H A) + J a! + J m (n m r ~ in r ~ -

MCh-H i-H 'v ΓΟ CN σ »Γ0 CN Γ0 CN

-I-I 0 ιβ O '* »' '

rH 'E * 4 M O rH CO O rH Ο rH

ft OP rH

W -H I I

d i "- o o hj io c to to

0 oo o t! V V

-b · * JJ J J

Λ ox S zzd ft id ii tö n -μ -μ · η> ιο S Tl Tl H .pk fO> 1 - r, CO>, <13>, CU x cm cm <U -P -hoho ω C ”Q ) rH = ϊ ϊ ϊ Λ P = Λ M ϊ • h · η n> ό μ omd

> i id -μ <d O

1? ϊ g m μ m -h Cö> 1 O (3 -H · μ -μ 4->> 1 C Id E tO rH to-μ

:: rH rH c -o 3 j *>, a <rH

"* 2 10 o '0 5 δ s £ o -h« c -g s r 2 y s Λί + JrtSX Z 2 ft 2 Λ ^ M C · η

d: <0 -H O

h> e o ^ .....

d m h λ .g ..

.: cÖ-HtOO tn HOWi-lSro »s · in ιο

Eh J S «M

22 77792

The results in Table 2 show that the hydrocarboxycarbonylthiourea collectors of the invention provide an equally good metallurgical result at pH 9.0 and 10.0 with lime consumption of only 10-30% of the amount required to use a standard collector, sodium methyl xanthate (Examples 1 'to M '). The results show that the collectors according to the invention (Examples 3-6) provide higher copper recovery and selectivity for pyrite with lower lime consumption than when using 10 known standard collectors. This is also seen in the higher I value of these new collectors. It is important to note that standard collectors give a very poor metallurgical result at pH 9 and 10, as seen in Examples I ', J * and L'.

15 Examples 7-8

Flotation at a slightly alkaline pH The following examples used South American copper-molybdenum ore containing 1.844% copper and 4.2% pyrite. The copper minerals were mainly 20 chalcosite, chalcopyrite, covellite and bornite.

510 g of ore was wet milled at 68% solids for 7.5 minutes in a steel ball mill to give an ore slurry with a total distribution of 24.7% +65 mesh, 38.3% +100 mesh and 44% -200 mesh. In all experiments, 2.5 g / ton of di-sec-butyl dithiophosphate was added to the 25 millings. To obtain the desired pH in the flotation, lime was also added to the grinding. The slurry was transferred to a flotation cell and cured at 1100 rpm at 32% solids. The blowing agent used was a 1: 1: 1 mixture of polypropylene-30 niglycol monomethyl ether / MIBC / tall oil 20 g / ton. The collectors of the invention were compared to several standard collectors, and the results of the experiments are shown in Table 3 below.

23 77792

Γ ~ iH H CO H CM

m m <i n co cm co

3 O O O O O H O

υ 1 M ooooooo <d o 5 £ 5 O P cm o h ro ro ro m · X o ~ G <#> o co ro o o cm a) co co oo cm · «* o oo J“ a) -

O + J CM

- Η ω σ> <0 cd H a

S

Λ U) H 3 H 3 5 H 00 Ο O CM Οι Ο ΓΟ C O '

0 -P H CM O CM no si · (N

ι) Ή rH rH rH rH rH rH rH

tP ^

O dP

o ^ 3 0) £ β 0) T? Γ- ro o ^ * H -Μ · * · *> - * ··· * * m O ^ <p <p cn ro m rn C co co r- co co 2 Φ

p 0) dP

P 4J

Ο Ή * frt c0 3 S ^ u ίΰ id m lD o o o o o> .......

S o O CP CP (P CP CP

iH rH

(#p

•B

<N ^

*. X m m CN CN CN CN CN

^ 03 CP OOOOOOO

ώ X

II

cm + j inooL / ntnuiin

rn \ CM Ο 01 CM CM CM CM

, rH rHrHrHrHrHrHrH

• HQ) 0.

Pi Cn g OOOOOOO

•B

• H *.

Λ dP H

2 ΰ II

3 ^ 03 OO

4J o3 co cn

O 00 -P Ή H

^ G H I I

H, (0 P - -

H Ή w Q) Z Z

<0 X -MI i I

(δ II -H -H M O -H H

S ö rH (1) -H rH rH

r * - * H + J> i> <i -fis> i -M Ή> i 03> i

CU O CL) + J Ό C 0) C rQ

0) -P O s s (0 -H O P O 0)

C ·· ">. P fflH.Q3.QP

H -H H ft— P>, -H PO P 3

JJmn O H G>, + J rfl H (fl O

'jr.y wtj nj + J4 -> ^; - p ^: - H

• o r * 'c HP M 3 to -p H H + J

Ä> i rfl 6 <fl Λ rfl 0} H (0 H

rHHin -η 3Ό h o c. *> i .¾ h COIÖIÖCM O H G H tfl H O> i O> i £ O pro>, H H P a -P> i

n _i h · * P P> ι I u> <1> O 0) P

O Ή «0 * O id ID E H O I P I 3 Λί4-> O ^ Z ~ <Q H Z a Z Λ

M to G

3 «0 -H m <-»> E O e .....

3 0) H c .5. .

flJ-H (0C w z o a OI 05 r ~ co h J 20 «24 77792

The results in Table 3 show that with the new hydrocarboxycarbonylthioureas of the invention (Examples 7-8) the copper recovery at pH 9.0 is substantially the same as that obtained with the sodium isopropyl xanthate standard-5 collector at pH 10.5 (Examples N'-O '). With a standard collector, poor copper recovery at pH 9.0 was obtained even at a dose level of 0.19 mol / ton (Example P '). The use of new hydrocarboxycarbonylthiourea collectors reduces lime consumption by more than 50% as can be seen by comparing Examples 7 and 8 with standard collector examples N '- P'. The collectors of Examples 7 and 8 gave a satisfactory copper content in the concentrate and better selectivity for pyrite than the standard collectors. It can also be noted that other conventional collectors (Examples 15 Q 'and R') gave very poor copper recovery at pH 9.0.

Foaming at a slightly alkaline pH

In the following examples, copper-molybdenum ore in the southwestern United States was used, containing about 0.778% copper and about 5.7% pyrite. This ore was one of the most difficult ores tested in terms of complex mineralogy, low copper recovery, high lime consumption and flotation problems. The ore contained mainly chalcosite, however, the ore pyrite was rich in chalcosite and covellite as 25 edge and scattering minerals. Separation of pyrite in pre-flotation, i.e. the first flotation, was thus not possible and was not attempted. 80 g of ore was prepared with ammonium sulfide (500 g / ton) and ground at 55.5% solids for 6 minutes in a steel ball mill to give a size distribution of 17.4% +65 mesh, 33% +100 mesh and 47.4% -200 mesh. The slurry was prepared at a rpm of 1500 rpm with a solids content of 20.4%.

In the standard treatment of this ore, the pH is 11.4-35 11.5 and N-ethyl-O-isopropylthionocarbamate is used as the collector. The consumption of lime to obtain the required pH of 11.4 to 11.5 is about 3.07 kg / ton. Standardivaahdotus-

II

25 7 7 7 9 2 The substance is cresylic acid, about 150 g / tonne.

Collectors were tested at the indicated doses and at the indicated pH values. The results are shown in Table 4 below.

26 77792 p r- o LO cn> £> ^ <τ »tj» isO o r ** O CNCOO ^ CN i— ♦ rH (N ιΗ CO Γ0

M O O rH O rH O O O O O O

OOOOO OOOO O O

3 4->

P

d)

P

O m o ^ I1 co iH ^ vocn m <T> G dP - * · * * 0 »n m I1 ^ r- O ro rH ro 0) *> * r- i £> vO cn in cn" sf ro ro m ro

P 04 rH CO

ro <U E- * Pm

•B

c -rH

5 o dP

O -P ΓΟΓ'Η'3'kO νΟΓ-ΟΟ CO CO., LTD

-P -H- - - \ o »w co cT) o m m r-coojo σ» <-h

13 «H t — 1 rH t — I rH rH

KJ p p O w o in - 5

φ S

G p lOCMvOCOO ro CN O- CO rH O

• H 0 - - (Ti G 00 CO t-H O 'H iDO'iOin o O' rn <D cO O CO UI CD CN ro i £> in 3 jS - P H - O ms

XJ HU

x:

2 -H

m * ·> xi + J m * 3 · o r * n co on co o i — i \ cm r- ~ in o o cm cm r- cn cm t ~~ men '· _ | s2 Λ! ΗΓΟΓΟιΗΓΟΓΟ OOOO o o

OP

Γ "ooro ^ j · ^ OOOO o o ^ \ N ·· S S ·. ·» · »·» · »·.

, r, CC COCTiOrHrHCOCOCTCOCOCT

U) 04 rH rH rH

II

:: p o m o m o m o m o o m

(VJ \ rH O rH O rH o rH O rH rH O

^ rH (NrHCNrHCM H CM rH CN (N rH

-HO) g ooooo oooo o o P Pn

. * * * H

• H * \ H

Qj OP I * "* to o 0)

2 _ C -M I

3 oo o en o i + J 00 -rH I 0) -rH o o t '' P O -H -M Ή

, Π Ή -H P -H P

3 rH P P -O -H

-p °>, -hm -hr — i <0> 1 P m rH> 1 (0 II ft PP> 1> i> O -HC> 1 4-> -oe .hmp 3 f, rH ft> 1 tn ρ λ

S M O> 1 U. Λ O

0) tn p -h o tn

e ·· -H s s = r m = »—t r tn -H

• H -H I -Ή Λ Ή> i -H I

rH m OP I p> 1 I -H

j ί ”I P CP g -H -H Ό -H

H -nm im m ό p 6 p .¾> i m rnm g ro -h b p m p f — i r — i -n>, g cc * h mm -mm o> i m * ho> 1 -h ro e ro

X 0 P XI C XJ 5h PM-IOP

. O M <1) C PM rH p tn g tn o -h ro o im mm h ro o e o

X -V X 2 M Z X <2 P <P

X tn ö

3; ro -H

Ή> £ X ...........

3 0) rH .g ...........

(ö -H ro tn en HD> SX> HN <c (n u

Eh ^ g h <3 m u

II

27 77792 2 m o ^ m r * - u m co r- n cd M O O O o o o o o o o o o 3 P P 0)

P

0 o «H 00 m ro m G <#> *“ - *

0) O Ό CO vO (N

Q) * CO VO v £> OOOvO

P (N i — I W 03 Q) E-ι Ui

•B

O <#> P (N in O CO ΓΟ Γ '' • H v * »<*» * V ·. · »O ^ tn co o o r ^ cr \ <T *

13 P P

3 3 u w 3

P

P

CL)

p o r * - CO LO CM (N

O - ““ * * * C 00 r- 00 P P CTv 0) r- r-- oo co r * <u <*>

P

ip ** rtj 3 tn o

B

M P LOO ^ J * ιο O Ό

P \ (N P <N P

dj O '*. * »» * · »«. «>

te M OOP OOP

o o r- o o o

X CD (Τ 'CO T O

O * P

p o lti m o LO lo \ p o o P o o

P ΓΜ P P (N P P

o - * · S o o o o o o

I I

0 o tn m

H -H

1 I

z z i i

• H * H

i-H Ή Ä> 1> 1 tri ^ 7 ^ ^ 7 S CO) C Ifl

O OM O <U

M X) 3 = = Λ M Ϊ - JJ MO M3

, rt m -H <β O

M JJ M -H

π · Μ -H -M + J

(0 tn H (0 -H

• n, * 5n, * Ή • O 0> 1 O> 1 O -P CU -P> 1 «. * <U o ω -p

2 o IM IP

X * Z CU Z Λ λ; 3 3 e 3 * r1 ......

id W OOPCNrOTt

Eh ΰ P P P P P

28 77792

The figures in Table 4 show that the hydrocarboxycarbonylthiourea collectors according to the invention (Examples 9 to 14) at pH 8.0 and 9.0 give essentially the same copper recovery as the standard collectors at pH 5 10.3 or 11.4 (Examples U ', V and W). Copper concentrations were also comparable. It is important to note that by using the new collectors according to the invention, lime consumption can be reduced by more than 50-70% of the standard consumption. In fact, using the collector according to the invention at a dose of 0.210 mol / ton at pH 8.0, lime consumption could be reduced by 92% (Examples 12-14) and at a collector dose of 0.105 mol / ton at pH 9.0, lime consumption could be reduced 78 % (Examples 12-14). At pH 9.0, selectivity for pyrite is acceptable for this ore, as higher pyrite recovery is unavoidable, as explained above. It should be noted that when several other conventional collectors were used (Examples X1 to CC '), copper recovery was very poor. In addition, it should be noted that the standard-20 collectors of Examples S '- W gave a very poor metallurgical result at pH 8.0 and 9.0.

The above examples demonstrate the significant improvements and advantages achieved with the novel, improved hydrocarboxycarbonylthiourea collectors of the invention over a number of conventional collectors known in the art.

Although the method of the invention has been described in terms of certain preferred embodiments, modifications and alterations may be made to those skilled in the art. For example, instead of N-ethoxycarbonyl-N'-alkylthioureas of the above formula, other hydrocarboxycarbonyl lithioureas such as N- (3-butene) -1-oxycarbonyl-N'-alkylthioureas can be used as sulfide scavengers according to the invention. As mentioned above, method 35 can also be applied using mixtures of two or more hydrocarboxycarbonylthiourea collectors with another known collector selected, for example, from the following: a) xanthates or xanthate esters of the formulas respectively

S S

r fi H - +. n II q

5 R O-C-S M Da R O-C-SR

b) dithiophosphates

S

8 H - +

(R 0) -P-S M

C) thionocarbamates S

8 11 9

R O-C-NHR

d) dithiophosphinate 15 g ® +

(R) 2P-S-M

(e) dithiocarbamates and their derivatives of the formula

R11 S R11 S

R8N-C “S“ M + and R8N-C-S-R1 f) trithiocarbonates and their derivatives of the formulas respectively

25 s S

8 ·· - + 8 U 9

R S-C-S M and R S-C-S-R

(g) mercaptans

30 R2SH

8 wherein in the above formulas (a) to (e) R is x δ 11

alkyl, R is C 1 -C 4 alkyl, aryl or benzyl, and R

2 * 8 "10 is hydroxy or R and in formula (f) R is C1-C12" alkyl · 35 Instead of copper minerals, the process according to the invention can be used to enrich other sulphide minerals and metals from sulphide ores, for example lead, 77792 30 zinc, nickel, cobalt, molybdenum , for the enrichment of iron as well as for the enrichment of precious metals such as gold, silver, platinum, palladium, rhodium, iridium and osmium 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.

Il

Claims (8)

Collecting reagents for flotation of sulphide minerals of base metals, characterized in that it contains at least one hydrocarbon carboxycarbonylthiourea compound formed in OHS 3. II in II 2 R OCNC-NHR I 2 and R are independently saturated or unsaturated alkyl group. A collector reagent according to claim 1, characterized in that R is C 1 -C 6 alkyl and R is C 1 -C 6 alkyl. 0 A collector reagent according to claim 1, characterized in that R is isopropyl and R is ethyl. 4. A reagent according to claim 1, characterized in that R is isobutyl and R is ethyl. The collector reagent of claim 1, characterized in that it comprises a liquid 2 mixture of a compound of formula I, wherein R is isopropyl and R is ethyl and a compound of formula I wherein R is isobutyl and R is ethyl, in an aliphatic or aromatic hydrocarbon solvent. 25 6. A process for enriching copper sulphide minerals from non-ferrous metal mime by printing iron yarn minerals and other yarn sulphides at a pH not exceeding 10.0, characterized in that a) is prepared from finely divided ore particles having a grain size suitable for release, an aqueous slurry having a pH not exceeding 10.0; b) conditioning the ore sludge by an effective amount of a flotation agent and a metal collector comprising at least one hydrocarbon carboxycarbonylthiourea compound of formula I