FI78403C - AVSKILJNING AV PENTLANDIT GENOM FLOTATION AV PYRROTIT MEDELST ANVAENDNING AV SVAVELDIOXID-LUFTKONDITIONERING. - Google Patents

Description

78403

This invention relates generally to the enrichment of complex sulfide ores, and more particularly to a process which primarily defoams pentlandite while leaving the chalcopyrite and pyrrhotite to foam for subsequent separation.

In the Sudbury area of Canada, as elsewhere in the world, 10 masses of nickel are found in finely divided pyrotechnic complex ore along with other precious metals such as copper. The sulphide main minerals consist of pentlandite (Ni, Fe) gSg, chalcopyrite (CuFeS2) and pyrrotite (Fen_ ^, sn) and the undesired pyrrotite (which itself contains only a small part of nickel in solid solution) is present in much larger amounts than pentlandite.

The ratio of pyrrotite to pentlandite can be, for example, about 5: 1. These ores and their extraction have been under investigation for many years and as a result 20 much has been discovered.

The methods adopted for the enrichment of these complex ores separate the valuable parts into a nickel stream, a copper stream, a pyrrotite stream and a wreck stream, and are essentially compromises in which the overall recovery and degree of enrichment of the desired precious metals is balanced. In ore flotation, a practice has been adopted to form an unsorted nickel-copper concentrate and a pyrrotite and side rock wreck, followed by further flotation to achieve a separation of nickel and copper. When performing nickel-copper separation, chalcopyrite is preferably foamed than pentlandite. It is necessary to foam a large amount of pyrrotite to recover the slowly foaming fine pentlandite and intermediate particles.

35 Consequently, considerable amounts of pyrrhotite remain in the nickel concentrate, with the result that the nickel concentrate contains approximately 78% nickel, according to analysis. In contrast, the copper concentrate obtained contains about 30% copper. The nickel concentrate still contains copper and the copper concentrate still contains nickel, factors that further complicate processing in the 5 smelters. Pyrotite concentrate discarded from the cycle should contain as little nickel and copper as possible.

The fact that, in order to maximize the recovery of nickel (or copper), the nickel concentrate currently obtained is of relatively poor quality means 10 that the total sulfur content of the nickel concentrate fed to the smelter is higher than desired, taking into account operating costs, nickel throughput, nickel loss, cobalt losses and sulfur dioxide emissions. For example, increasing the grade of concentrate from about 1% nickel silicon-15 to about 17% would reduce sulfur dioxide emissions by about 25%, which is a highly desirable result and other economic benefits can be achieved. Although the goal of providing nickel sulfide ore to obtain a better quality nickel concentrate has long been recognized as a goal, no practical means have been developed to do so without facing serious economic obstacles.

In connection with the above, it has become increasingly important to reduce sulfur dioxide (SC 2) emissions. Sulfur dioxide is a undesirable by-product of sulfide ore refining.

Current gas treatment and sulfuric acid production are hugely expensive. Thus, emphasis has been placed on removing as much pyrotite as possible before the treated ore is subjected to pyrometallurgical processes. It is clear that the more sulfur that can be removed in the early stages of ore treatment, the lower the amount of sulfur dioxide must be combated during the final purification steps.

Flotation is one of the most useful mineral irrigation techniques available to an engineer to concentrate valuable minerals contained in ores. Many special flotation techniques have been developed for the treatment of various ores and a wide variety of chemicals have been used to achieve many advantageous results in the application of these techniques.

5 However, as far as can be deduced, there is no mass commercial flotation method that separates pent-landite from pyrrotite. The only other flotation method that succeeds in doing this is the subject of CA Patent 1156380, but this method has not been commercialized. This method uses a combination of sodium carbonate and cyanide to foam pentlandite and chalcopyrite while pressing pyrrotite. Unfortunately, this method consumes a large amount of cyanide and leaves stable complex metal cyanides in the waste. At present, other flotation methods currently in use are only moderately effective. For example, a process referred to by some as the pyrrotite rejection cycle at Inco Limited's Copper Cliff plant (located at 20 Sudbury, Ontario) recovers only 60% of nickel and 25% of pyrrotite.

The present invention relates to a process for the selective enrichment of sulphide ore containing pentlandite, calcopyrite and pyrrotite by co-foaming sulphide ore using xanthate as a collection reagent. The invention is characterized in that the ore is formed into a slurry; adding a reagent to the slurry to desorb the xanthate present in the slurry after the previous operations; xanthate and liquid are removed; adding a sufficient amount of xanthate to the ore to foam the sulfides to give a sulfide co-concentrate as a float product while leaving the side rock in the zinc stream; the sulfide co-concentrate is diluted to a second slurry; adjusting the pH of the second slurry; adding sulfur dioxide and air; and a second slurry is foamed to recover chalcopyrite and pyrrotite as a float product and pentlandite printing.

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Figure 1 is a general flow diagram of the invention.

Figure 2 is a modification of the invention.

Figure 3 is a graph showing the resolution power.

Figure 1 shows a flow chart of the process.

5 The feed in the following description is of poor quality material obtained from Inco Limited's Copper Cliff plant. The stream has been developed after removing the easily foamable chalcopyrite and pentlandite and discarding by magnetic separation most of the monoclinic-10 pyrrhotite. However, it is to be understood that the method of the invention can be used with appropriate materials regardless of the source.

In this case, according to the analysis, the feed contains about 3% chalcopyrite (CuFeS2>, 5% pentlandite 15 (Ni4 yFe ^ 2S8 ^ · ^% pyrrotite (Fe7Sg) and 50% side-by-side. The aim is to recover chalcopyrite and pentlandite as separate sulfur all pyrrotite and side stone are discarded.

Simply for ease of description, the following treatment relates to a one kilogram sample of the feed. Larger amounts are treated in the same proportion. In all cases, the feed was slurried to two liters of ore slurry.

* · The process begins with an operation called a sulfide wash. The purpose of this step is to provide a means of removing excess xanthate that may follow the feed material as a result of previous processing. After the addition of a sulfide-containing reagent (e.g., sodium sulfide, sodium hydrogen sulfide, ammonium sulfide, or ammonium hydrogen sulfide), the slurry is prepared for a few minutes and the slurry is then thickened, during which time the slurry is continuously scraped. In fact, the sulfide reagent may be selected from alkali metal or alkaline earth metal sulfides or hydrogen sulfides. It is also within the scope of this invention to use lime washing instead of sulphide washing.

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0.5 g / kg of sodium sulfide was added, after which the slurry was prepared for five minutes. The purpose of the scraping is to ensure that the slurry and the solution are mixed so that the xanthate desorbed from the sulfides is evenly distributed throughout the solution. It is known that the conditions required for desorption of xanthate are very negative redox potential and low oxygen conditions. After thickening for a long time (17 hours), the solution is decanted and discarded to remove desorbed xanthan-10.

The thickened solids are then ground at a solids content of 50% by weight to obtain a particle size distribution where 35% remains on a 37 micron sieve. After grinding, fresh xanthate (usually 0.07 g / kg feed) is added and the sulphides are foamed off the side rock. The foam time was 16 minutes.

The sulfide co-concentrate obtained in the previous step contains substantially all of the chalcopyrite, pentlandite and pyrrotite that were in the feed. Typically 20 to about 40% by weight of the initial amount of side rock enrichment pellets containing about 0.1 Cu, 0.25 Ni and about 4% of the total copper and nickel.

The sulphide co-concentrate, diluted to 1.7 l of slurry, is heated to 40 ° C, the pH is adjusted downwards to 8.1, if necessary, and sulfur dioxide is then added at a rate of 9 ml / min for 50 minutes. The total SC 2 addition is therefore 450 ml or 1.2 g (1.2 g / 1.7 l or 0.71 g / l). Air is added at a rate of 590 ml / min simultaneously with the SO 2 blowing to ensure that the level of dissolved ha-30 pen remains above about 5.4 ppm. During the addition of sulfur dioxide and air, the slurry is stirred vigorously to create a vortex that prevents foam from accumulating on the surface of the slurry. After the addition of sulfur dioxide for 50 minutes, the gas flow is stopped and sufficient lime is added to the slurry at 40 ° C to obtain a pH of 11.5. The slurry is prepared for another 20 minutes with vigorous stirring 78403 6 again to create a vortex. Air is added again to keep the dissolved oxygen above about 5.4 ppm.

After this SC / air / lime training, the total sulfide concentrate is subjected to flotation for about 10 minutes.

5 Chalcopyrite and pyrrotite foam, while pent Landite is pressed. In a successful experiment, 95% of chalcopyrite and more than 90% of pyrrotite is in a float product with less than 25% pentlandite. The foamed material is returned to the cell for the cleaning flotation step.

10 Lime is added again to obtain a pH of 11.5, after which the slurry is subjected to flotation at ambient temperature.

The non-foamed portion of the first and second (post-flotation) flotation steps (called the pyrrotite-15 post-flotation enrichment pellet) is combined into a nickel grit. In the post-foaming step, the foamed material contains essentially all of the chalcopyrite and most of the pyrrotite.

To separate the chalcopyrite from the pyrrotite, the foamed material is prepared in a saturated lime solution (prepared with 1.6 g of lime), a small amount (0.1 g) of cyanide is added, then prepared by aeration until the redox potential (Au / saturated calomel is at least increased) -225 mV, followed by foaming of chalcopyrite-25 for four minutes. The non-foaming part is the final pyrrotite concentrate. Typically, its analysis is about 90% pyrrotite, 1.4% Ni (1.5% Ni / Ρο) and contains n.

75% pyrrithite and 10% pentlandite.

Analysis of the 30 copper concentrates from this single flotation step shows about 16% copper and the concentrate contains 80% feed copper. Analysis of the concentrate also shows 1.2% Ni, 31% pyrrotite, corresponding to 2.7% of total nickel and 4.4% of total pyrrotite.

The results are presented in more detail below: 35 7 78403

Table I

Pentlandite / pyrrotite separation results Analysis (%) Distribution (%) ^ Product Cu Ni S Po Wt Cu Ni Pn Po

Cu-rich 16.8 1.2 30.3 31.0 5.9 80.3 2.7 2, A 4.4

Rich concentrate 0.3 1.3 35.9 87.8 35.2 9.4 17.4 9.1 74.6

Ni-rich 0.6 12.0 22.2 27.3 16.5 16.5 75.4 84.8 10.9

Rk sterns 0.1 0.28 4.2 9.9 10 Feed 1.2 2.6 19.9 41.4

Rk - side rock Po pyrrotite Pn = pentlandite A. pH and training time

The effects of final pH and training time during the addition of SO 2 and air are shown in Figure 3. The procedure used in these 15 experiments was shown in Figure 2.

This procedure differs from the previous one in several ratios, some of which depend on the sample. No sulfide wash (thickening step described in the previous procedure) was performed on this sample because it did not contain excess xanthate. 20 However, Figure 3 is relevant to the previous procedure.

Figure 3 shows the efficiency of pentlandite / pyrrotite separation (S.E.) as a function of training time and final pH. In this series of experiments, SO 2 was added as a 1.5 wt-25% mixture in air. The actual rate of SO 2 addition was 20 ml / min until the pH reached the desired value, at which point the additional rate of gas was reduced so that the pH was kept constant until the entire training time had elapsed. All training was performed at room temperature (25 ° C).

30 Separation efficiency is a numerical value used to evaluate the efficiency of a physical separation process. It is simply calculated from the difference between pentlandite and pyrrotite in flotation. In this case, the foamed fraction is calculated from the sulphide co-concentrate. Losses of pentlandite and pyrrotite 35 in the side rock enrichment are not taken into account. The separation efficiency measures the fraction of the feed that was completely separated.

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Figure 3 was obtained using a stepwise multiple regression program to obtain an equation that relates the difference efficiency to two independent variables. The difference efficiency curves were then developed from the regression equations. The results show an optimum condition with a pH of about 6.0 and a training time of about 110 minutes. Figures outside the change curves show the actual 10 experimental results.

B. Lime washing

During a previous experiment, it was found that the feed contained many times more xanthate than the SO 2 / air procedure could survive at room temperature for 60 ml to 15 minutes. Since the amount of xanthate associated with the feed of the ripper was also variable, a method was sought to desorb the excess xanthate and stabilize the feed to the separation process. The method adopted was lime washing. Lime washing was performed by adding about 5 g of 20 lime to one kg of feed in a 2.27 liter container filled with water. The slurry was gently stirred for 30 minutes, after which the slurry was allowed to settle for 45 minutes, after which the solution was decanted and discarded. The thickened slurry was then ground as needed, the pH was adjusted to 9.5 with sulfuric acid, xanthate was added, followed by SO 2 / air training.

C. Temperature Time Sample A was subjected to a series of experiments by this procedure. The varied parameters were temperature SO2 / air-30 during training and duration of the conditioning period.

The optimum range was found to be at 40 ° C with a training time of at least 60 minutes. In the four experiments performed under these conditions, the difference in efficiency averaged 0.77. It was determined that SO 2 / air training causes inhibition of pentlandite flotation, whereas it has only a minor effect on pyrrotite. The best nickel 9 78403 in pyrrotite corresponds to the optimum range of the separation efficiency.

D. pH Adjustment with Acid In these experiments, others were performed in which the pH was adjusted to the desired value (typically pH-5 to 6.5) before the addition of SC> 2. None of these experiments succeeded in obtaining a significant difference, confirming that the separation observed after SO 2 / air treatment was not simply due to pH adjustment. In addition, it was found that the use of sulfur dioxide alone to lower the pH to 6.5 after washing with lime resulted in the pressing of too much pyrrotite. Needless to say, the SO2 addition was too large.

E. Oxygen level during SO2 / air training

An attempt was made to repeat the above procedure with Sample A. The test results are shown in Table II along with dissolved oxygen levels during SO 2 / air training.

Table II

Results of repeated experiments with sample A

20

Experiment 02 (ppm) Separation efficiency 1 6-6.5 0.78 2 4.9 - 6.3 0.27 3 5.1 - 6.1 0.67 4 6.1 - 6.4 0.81 25 5 6.1 to 6.4 0.80 These results were undoubtedly surprising at first. In Experiment 1, an excellent difference was obtained (difference efficiency 0.78), but in Experiment 2, this result was not successful for another 30 countries. Similarly, Experiment 3, although better than Experiment 2, did not have a high level of Experiment 1. Examination of the experiment revealed that in experiments 2 and 3, there was little dissolved oxygen at the beginning of the SO 2 addition. A change in procedure was made: the ore slurry 35 was saturated with oxygen during the heating of the slurry in Experiments 4 and 5 so that oxygen was present when the SO 2 addition was started. There was only a slight observed change in the measured values of clearly dissolved oxygen, but the results of the separation improved enormously. The procedure is then determined by oxygen saturation during heating and at least 5.4 ppm oxygen during SO 2 addition.

F. SO ^ addition rate

Experiments were performed on sample B using three different sulfur dioxide flow rates. The total addition time was set at 60 minutes, so the total volume of 10 SO 2 added was also changed as the flow rate was changed. The results show that with this procedure and this sample, the optimum xanthate addition was 0.05 g / kg. Moreover, it turns out that in the tested range (480-600 ml in 60 minutes) the rate of SO 2 addition did not have much effect on the result. Because it is much easier to maintain high levels of dissolved oxygen at low SO 2 addition rates, the procedure was changed from 20 ml / min to 9 ml / min. A provision was later added that the final pH could not be allowed to fall below 6. Thus, the total addition time of SC> 2 was shortened to 50 minutes.

G. Lime aeration before post-foaming of pyrrhotite

The standard procedure included a post-flotation step with pyrrotite pre-concentrate. Initially, the pyrrotite pre-concentrate was prepared with SO 2 / air 15 minutes before post-25 foaming, but this was diluted to a dilution with water previously acidified to pH 6 with bogs.

Because the lime aeration process with Cu / Ni separation is so effective at pressing pentlandite while foaming chalcopyrite and pyrrotite, this technique was evaluated. After 60 minutes of lime aeration, the post-foaming of the pyrrithite concentrate was greatly improved, as the results in Table III show.

and 78403

Table III

Effect of pyrrithite pre-concentrate lime / aeration 5 Sample Best separation efficiency Best Ni% / Po SO ^ / air Lime / air SO ^ / air Lime / air B 0.73 0.76 1.69 1.71 C 0.61 0.70 1 .99 1.36 D 0.46 0.58 3.20 2.24 10 E 0.43 0.71 2.20 1.31 F 0.55 0.69 1.89 1.52 G 0.68 0 .70 2.09 1.80 H 0.53 0.59 1.98 1.39

In all cases, the efficiency of pentlandite / pyrrotite separation was better with the lime aeration process than with standard post-foaming in water with a pH of 6.

A previous pilot plant experiment included the addition of sodium hydrogen sulfide followed by thickening to remove desorbed xanthate. Very early in the program, it was observed that very little xanthate was decanted, but there was a very high xanthate concentration in the thickener bottom stream solution. It has long been known that a reducing and oxygen-poor environment was required for the desorption of xanthan-25 tin, but apparently this information has not been fully utilized. To remove xanthate, the thickener bottom stream was pumped back to the thickener feed in a sealed tube. Leakage current was taken from this recycle loop to be fed to the process. This solved the problem. As a result of these 30 findings, the laboratory procedure was re-examined. It was found that the static thickener had a heterogeneous xanthate distribution with the highest concentration associated with the thickened solids. To prevent this, solids were scraped with a slowly rotating scraper grade-35 timer (8 rpm). Samples taken from different levels of thickener 12 78403 after 17 hours of scraping showed even distribution of xanthate.

The previous circulation process involved SO 2 / air coaching of the heated, sulfide co-concentrate, followed by a single flotation step. Prior to the pilot plant project, the Penk tests had included a post-flotation step for pyrrhotite, in which the dilution water was acidified with SC 2 to pH 6. The run actually provided a ready source for the pyrrotite pre-concentrate. This opportunity was taken to perform a series of batch-like bench-scale post-foaming tests with lime after coaching and aeration. Pyrotite flotation improved only slightly with the addition of aeration in lime, but pentlandite flotation deteriorated substantially as the aeration time was extended; as a result, the post-flotation separation efficiency (S.E.) improved from about 0.25 without aeration to about 0.7 after 60 minutes of aeration.

H. Hot Lime Coaching

A series of experiments were performed to investigate the effect of elevated temperature 20 on the lime training step. This resulted in the addition of lime to the slurry at the end of the Si / air training carried out at 40 ° C. The effect of the training time following the addition of lime is shown in Table IV.

25 Table IV

Test results of hot lime training before pyrotite pre-foaming (sample I)

Lime hardening time Observed maximum Test (min.) Difference efficiency 30 2 0.62 1 5 0.68 2 10 0.60 3 15 0.74 4 15 0.70 5 35 20 0.69 6 20 0.68 7 13 78403

Table IV (continued)

Lime curing time Observed maximum Test (min.) _ Separation efficiency 5 30 0.75 8 25 0.75 9 30 0.73 10 A curing time of 25 minutes after the addition of lime to pH 11.5 at 40 ° C is considered quite satisfactory. Although specific embodiments of the present invention have been described and illustrated herein in accordance with the provisions of the Patent Act, those skilled in the art will appreciate that changes may be made to the form of the invention covered by the claims and certain features of the invention may be utilized without the need for other features.

Claims (19)

A process for the selective enrichment of sulphide ore containing pentlandite, chalcopyrite and pyrrotite by co-foaming sulphide ore using xanthate as a collecting reagent, characterized in that the ore is formed into a slurry; adding a reagent to the slurry to desorb the xanthate present in the slurry after the previous operations; xanthate and liquid 10 are removed; adding a sufficient amount of xanthate to the ore to foam the sulfides to give the sulfide co-concentrate as a float product while the side rock remains in the zinc stream; the sulfide co-concentrate is diluted to a second slurry; adjusting the pH of the second slurry; adding sulfur dioxide 15 and air; and the second slurry is foamed to recover the chalcopyrite and pyrrotite as a float product and to press the pentlanite. Process according to Claim 1, characterized in that the temperature of the second slurry is kept at a value of about 35 to 50 ° C. Process according to Claim 1, characterized in that the pH of the second slurry, after the addition of sulfur oxide and air, is raised to about 11.5 at a temperature of the slurry of about 40 ° C, after which it is prepared for some time. Process according to Claim 1, characterized in that the pH of the float product obtained from the second slurry is increased, cyanide is added to the float product obtained from the second slurry, the redox potential of the slurry is increased to at least -225 mV and the chalcopyrite is foamed. Process according to Claim 1, characterized in that the reagent is sodium sulphide, sodium hydrogen sulphide, ammonium sulphide or ammonium hydrogen sulphide. Process according to Claim 1, characterized in that the reagent is an alkali metal or alkaline earth metal sulphide or hydrogen sulphide. is 78403 A method according to claim 1, characterized in that the pH of the second slurry is adjusted to a value of about 8 before the sulfur dioxide is fed. The method of claim 1, characterized in that sulfur dioxide is added to the slurry at a rate of about 0.7 g / L of slurry while maintaining the oxygen content above about 5 ppm. A method according to claim 1, characterized in that the pH of the second slurry is adjusted to a value of about 6-6.5 by adding sulfur dioxide and air. A method according to claim 1, characterized in that the second slurry is mixed to create a vortex to reduce the formation of foam. The method of claim 1, wherein the reagent is added to the first slurry at a rate of about 0.5 g / kg of solids in the slurry. A method according to claim 1, characterized in that the pH of the second slurry is raised to about 11.5 after the addition of sulfur dioxide and air using lime at a slurry temperature of about 40 ° C, after which the slurry is further prepared. A method according to claim 1, characterized in that the pH of the pyrrithite float product obtained from the second slurry is raised to about 11.5 before training for about 20 minutes, followed by a second or post-foaming in which chalcopyrite and pyrrotite are obtained. as a float product, while pentlan dite is pressed. The method of claim 1, characterized in that xanthate is added to the ore in an amount of about 0.07 g per kg of ore. Process according to Claim 1, characterized in that the slurry containing the sulphide reagent is scraped in a thickener. 16 78403 A method according to claim 1, characterized in that the thickened ore is ground to ensure the release of minerals. The method of claim 1, wherein the pH of the thickened ore is adjusted to about 9.5. Process according to Claim 1, characterized in that lime milk is used to desorb the xanthate. A method according to claim 13, characterized in that lime is added in an amount of about 5 g / about 1 kg of ore. 78403