CA2101517A1 - Method of treating a base metal bearing material - Google Patents
Method of treating a base metal bearing materialInfo
- Publication number
- CA2101517A1 CA2101517A1 CA002101517A CA2101517A CA2101517A1 CA 2101517 A1 CA2101517 A1 CA 2101517A1 CA 002101517 A CA002101517 A CA 002101517A CA 2101517 A CA2101517 A CA 2101517A CA 2101517 A1 CA2101517 A1 CA 2101517A1
- Authority
- CA
- Canada
- Prior art keywords
- base metal
- slurry
- concentrate
- tailings
- producing
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Abandoned
Links
Classifications
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
- B03B1/00—Conditioning for facilitating separation by altering physical properties of the matter to be treated
- B03B1/04—Conditioning for facilitating separation by altering physical properties of the matter to be treated by additives
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/002—Inorganic compounds
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/004—Organic compounds
- B03D1/012—Organic compounds containing sulfur
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/004—Organic compounds
- B03D1/014—Organic compounds containing phosphorus
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/02—Froth-flotation processes
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/007—Modifying reagents for adjusting pH or conductivity
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/02—Collectors
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/06—Depressants
Landscapes
- Chemical & Material Sciences (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
- Processing Of Solid Wastes (AREA)
Abstract
The invention relates to a method of treating a base metal bearing material for recovering a metal concentrate, by conditioning the base metal bearing concentrate with a slurry having a pulp density of at least 20 % solids in the presence of water whilst maintaining the slurry at a pH of at least 7 for a predetermined period of time and thereafter recovering the metal concentrate.
Description
W O 92~13640 PC~r/AU92/00043 1. ~101~17 METHOD OF T~EATING A BASE METAL BEARING MATERIAL
Fleld of the Inventlon The invention relates to a method of treating a base metal bearing material for recovering a metal concentrate. Typically this method is 5 applied to treat a base metal tailing.
BackR~round of the Inventlon Whilst the followlng descrlptlon of the invention is with reference to trestment of a talllng, the Invention is not so limited.
Typically a base metal talling is produced from mineral dressing 10 operations located on an ore slte. Such a talllng may contain commercially significant amounts of base metals, such as copper, lead, zinc and nickel. In these cases, mine operators have wanted to recover these base metals from the tailing in an economically viable way.
In thls case a tailing is recovered such as by dredging or sluicing and 15 then is subjected to a concentrntion process, which may include flotation and/or other techniques such as grflvity, to produce base metal concentrates. These tailing flotation processes can, although not exclusively, be applied to the concentration of zinc sulphide minerals from a tailing.
20 However, minerals in a tailing dam generfllly respond only poorly to a flotation process. This is due to the chemlcal environment in which they have been stored subsequent to their previous treatment.
Furthermore, because of their genernily low grade a high upgrade ratio is required to achieve a saleable col-centrate.
25 Consequently, flotation trentments ol suc}- a tai}ing have had only limited success in producing reliable flnd s~leable bnse metal concentrates.
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WO 92/13640 PCI'/AU92/00043 ~o~7 Summary Embodlments of the Invention Accordingly, investlgatlons have been directed to the effects of changes to the condltlonlng o the base metal bearing materlal and to the flotation conditions of the subsequent treatment.
5 Surprisingly, lt has been found that improvements in the level of recovery of base metal concentrate coupled wlth low pyrite recovery during the flotation step can be achleved by subJecting the base metal bearing material (e. g . tailing) to n particular conditioning treatment.
This can resull in substantial upgrading of feed values.
10 More particularly, there ls provided a method of condltioning a base metal bearing materlal for subsequent recovery of base metal concentrate comprises forming a slurry having a pulp denslty of at least 20% solids by the addition of water to the base metal material, and maintaining the slurry at a pH of at least 7 for a predetermined perlod of time. If 15 necessary the pH may be malntained at the 'desired pH by addition of alkali (e. g. lime or caustic soda) .
Preferably the base metal is copper, lead, zinc and/or nickel.
The pH of the slurry is maintained irl t}-e preferred r ange of 7.0 to 8.5.
The preferred range of pulp density o~ the slurry is from 30-60% solids.
20 The preferred predetermined perlod Or conditioning time is about 1 hour or more and more preferab]y from ~bout l hour to about 2 hours.
To optimise the conditioning Or the sllrf~ce of the base metal material it is desirable to agitate the slurry. In one preferred form of the invention the conditioning tre.stmenl comprises forming a ~lurry having a 25 pulp density of at least 20~6 solids l-y the additioll of water to the base metal bearing material and maintaining tllat slurry at a pH of at least 7 for a period of greater than F~bout I llour whilst agitating the slurry.
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W O 92/13640 PC~r/AU92/00043 ~ 3, ~ 7 The regulation of a mlnimum pulp density and preferably agltation of the slurry has found to allow conditioning times less than that previously expected. The higher the pulp density or the more intense the agitation the shorter the conditioning times.
5 The agitation may be by any suitAI)le means. However, preferably the means imparts shear to the slurry whllst maintaining the slurry in suspension.
Typically, where the base metal bearing material ls a tailing it may lnclude sphalerlte, pyrlte and other base metal sulphide minerals mixed 10 with non-sulphide gangue materlals (e.g. talc, chlorite and quartz).
Investlgatlons have slso found that the temperature of the slurry has little effect on the condltlonlng step though elevated temperature may subsequently affect the flotation reagents. Similarly during this condltioning there ls no need to add the flotatlon reagents. However, 15 pH modlflers may be added. In fact lt ls preferable that those reactants (other than pH modlflers) are added after the condltloning stage.
The refinement in operating practice of the process of the invention has potentially important commercial implications for enhanced profitability and reliability of recovery. The conclitioning has facllltated reliable, 20 repeatable recoverles of base metal concentrate.
According to another preferred form oî this invention there is also provided a method of proclucing n bnse metal concentrate from a base meta] bearing talllng which comprises:
(a) recovering a base metal bearing tailing and placing it in one or 25 more vessels;
(b) adding water to the tniling lo form a slurry having a density of at least 20~ solids;
(c) maintaining the slurry nl a p]l of flt least 7 for a period of about I hour or more;
30 (d) adding at least one fJotn~ion reagent to the slurly;
(e) subjecting the sh~rry lo f1 i~tiO~l to recover the base metal concentrnte; and .
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W O 92/13640 . ` PC~r/AU92/00043 ~o~l7 "
(f) dewaterlng (e.g. thickening and filtering) the base metal concentrate.
According to another preferred form of this Invention there is also provided a method of producing a base metal concentrate from a ba.se 5 metal bearing tailing which comprises:
(a) recovering a base metal tailing and placing it In one or more vessels;
(b) addlng water which has a pH of at least 5 to the tailing to provide a slurry having a denslty Or at least 20%;
10 (c) agitating the slurry whilst maintaining the slurry at a pH of at least 7 for a perlod of up to about 2 hours;
(d) addlng at least one flotation reagent to the slurry;
(e) subjecting the slurry to flotation to produce the base metal concentrate; and 15 (f ) dewatering the base metal c oncentrate .
When necessary the water has been treated to maintain the pH by the addition of alkali reagents such as caustic soda or lime.
Preferably a number of holding vossels are used to provide a surge capacity to ensure continuous supply and the necessary conditioning for 2 o successful subsequent flotation .
Generally, flotation will take place in a number of stages (e.g. four), comprising a rougher stage followecl by a number of cleaning stages (e . g . three) . After the final cleaning stage the base metal zinc concentrate is de-watered.
25 Preferab]y, flotation reagents are used after the slurry has been preconditioned to render the desirecl mineral selectively smenable to the flotation process. The reAgent P~ddition is tailored to suit the mineral or minerals from which it is desired ~o r2cover the bnse metal.
In the case of recovery Or zinc beariTlg material, a number OT'` reagents 30 have been found to be prefers~blc. Tht? reagents added can be classified into three grour)s, namel v:
:
: ' W O 92/13640 PC~r/AU92/0~043 ~`- 5 ~ t (1) activators (such as CuSO,I);
(2) depressants (such as sodium metabisulphite (MBS)); and (3) collectors (such as potassium amyl xanthate (PAX), sodium isobutyl xanthate (SIBX) or a dithiophosphate thioncarbamate formulation ( e . g . AERO 4037 ) ) .
The reagents used in the rougher stage are typically either CuSO4, MBS, an alkali and SIBX or CuSO4, SBS, an alkali and SIBX. A
preferred reagent composition for the flotation is 1OOO g/t MBS, 500 g/t CuSO4, 100 g/t SIBX at a pH oE between 9 and 9. 5 .
The reagents used In the flrst and second cleaning stages are typically alkali and SIBX, preferably in the following amounts:
300 g/t NaOH and 0-10 g/t SIBX.
In the first cleanlng stage the pH of the slurry is preferably 10 to 11. 5 and more preferably 10 . 2 to 10 . 4, whereas in the second cleaning stage the pH is preferably 11 to 12.
The third cleaning stage generally uses an alkall, typically at an addition rate sufficient to give a pTI of least 11Ø
By way of example, the slurry density to the roughing stage should be about 25% ~ 40% solids and the slurry density in the cleaning stages should be in the range of about 20n ~ 50% solids and preferably 20 ~ 35%
solids .
The tailing from the rougller stAge m~y be pumped directly back to a disposal site. The water may be re-!overed for re-use.
It is far from clenr why condltloning ~ccordillg to the invention causes the significant improvement in the recovery of base metal concentrate.
However, it is thought to arise from an attrition which results in cleaner particle surfaces and/or some dis;lggregatioll of the particles which increases the liberation of minerals. In either case the particles are rendered more suseptible to subsequen~ ~lotation techiques.
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2101~17 6. ~
Brlef Descrlptlon of the Drawln~s The invention is now further illustrated wlth reference to the -~
accompanying clrawings and examples in which:
~IGURE 1 shows graphlcal]y a Zn grade versus Zn reeovery curve for 5 the roughening stage for a tailing whieh was eondltioned and one which was unconditioned in early test work.
FIGURE 2 shows graphieal]y a Zn recovery versus Zn grade eurve for the process of thls Invention.
FIGURE 3 ls a flow chart of the f]oation circuit used in the subsequent 10 investigation o the process of the invention.
FIGURE 4 shows graphica]ly a Zn grade versus Zn recovery curve for a series of samples under dlfferlng conditloning conditions.
In preliminary lnvestigstlons the first attempt to recover Zn from a tailing was a two stage process where the tai]ing was aerated in a slurry 15 and then a bulk concentrate wa.s floated from the slurry. The conditions for the aeration stage (2 - 6 hours) were 40% solids mixed in a pH 3 solution that contained a minimum quantity of copper (approximately 3 - 500 ppm Cu). Under these conditions up to 2096 of the contained zinc ( 10 - 20 g/l Zn ) was ]eached and the structure of the 2 o partic]es was modlfied . Aeration was followed by flotation at the natural pH of the aeration solution using a fAtty acid flotation reagent. A bulk zinc/]ead concentrate was produced.
When these parameters were carriecl out on a tailing from Wood]awn Mines New South Wflles Australia the results were not reproduced.
25 The process was then modifiecl to float a low grade zinc coneentrate (8 - 10% zinc at 80~ reeovery) using pH 3 wat~r contaminated with metal ions. The flotation step also recovered the pyrite from the tailing .
' . ' . ~ ' ' . ' ' ' ' ~ . '' ' :' ., . .,, : - . , : . .:
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W O 92tl3640 PCT/AU92/00043 ` 7.. ~ 7 While the process wns belng deflned In a pilot plsnt, a flotation concentrate was produced which floated less pyrite and contained a high grade ~inc rougher concentrnte with low weight recovery - typically 40% zinc at 409~ recovery.
5 The flotation conditions were:
1000 g/t MBS
500 g/t CuSO4 75 g/t PAX
pH 7.0 - 7.5 uslng tap water 10 Laboratory tests were then conducted to identlfy the conditions for achieving these results. The same reagent reglme in the laboratory did not give the same results as those from the Pllot Plant, i . e .
substantially more pyrite was recovered in the concentrate.
The Pllot Plant practice was then investlgated. The procedure for 15 mlxing a batch of tailing for feed was recognised as a maJor variation.
There was a delay of at least 2 hours between when the slurry was prepared and the flotation test carried out.
This procedure was simulated in the laboratory by conditioning the slurry sample in the float cell for 2 hours prior to flotation with water 20 at pH 5-7. Simllar results to those for the Pilot Plant resulted. An example of the results showing this conditioning effect is shown in Figure 1.
A flotation reagent scheme of:
1000 g/t MBS
25 500 g/t CuSO4 75 g/t PAX
pH 7.0 - 7.5 was used.
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W O 92/13640 PC~r/A U92tO0043 21~1~17 However, this conditioning effect wns found to be variable with the procedure being unsuccessful In some sample locations.
The variables which were investigAted included time, temperature and pulp density. Included in this work WAS the optimisation of the flotation 5 reagents.
Temperature Initial indications were that by increasing the temperature the conditioning time could be decreased. This did not prove to be the case during these tests and a conditioning temperature of 20 to 40C was 10 best. The effect of tempersture is shown in Table 1.
Table 1 - Variation of Results with Conditioning Temperature TEMPERATURE ZINC CONCENTRATE GRADE ZINC RECOVERY
C % %
15.4 62 14. 1 65 3.4 33 1.8 6 At the high temperatures ( l40C) tl-e flotation reagents were destroyed.
2 o Time A minimum conditioning time of ni~ol~ 2 hours WFI.q estnblished and the variation with conditloning time is showll in Table 2.
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Table 2 - Varlstlon of resu]ts wlth Conditionin~ Time TIME HRS ZINC CONCENTRATl~ GRADE ZINC RECOVERY
% %
O. 5 14 49 Pulp Denslt,y Tests showed that there was no variAtlon in float results when the pulp 10 density during conditionlng was variecl between 30 - 60% solids.
~lotation Reagents A series of reagents for flotation were evaluated. The end result was the choice of PAX and Cyanamid Aero 4037 for the flotation.
Results were still variable dependin g on sample location . Testwork to 15 date had been restricted to near surfnce samples.
At this point, +/-30% engineering estimste was carried out for the processlng of 3 milllon tonnes of tPIillng.
At thls stage the p~I reglme wAS increAsed to a p~ of 9-9. 5 .
20 The deslgn basis ror the flotation wns derine(3 as:
1000 g/t MBS
500 g/t CuSO4 50 g/t PAX
50 g/t 4037 P~-~ 9.0 - 9.5 , -: . . - : ~, W092/13~40 Z101Sl7 IO. PCT/~L192/00043 In conjunction with this study n core sAmpling programme of the tai3ing dam at Woodlawn Mines was completed and the preconditioning and flotation condltlons were tested on core sections.
The previous variability in results with locntion was not present. -5 Analysis of the results indicated this mny have been due to the fact that the natural pH of the slurry was always greater than 7.
Also during the 2 hour conditioning period the pH remained above 7.
The importance of maintaining the pH of at least 7 during the conditioning period was then confirmed when It was shown that samples 10 that had not previously responded to the technique did so when the slurry pH was adjusted to above 7 by addition of lime for the condltioning period.
Zinc rougher recoveries for the core samples averaged ~3 - 85%
compared to previous 7596. Feed grnde varied between 2 - 7% Zn.
15 The flotation sequence was defined ns:
ConditioninEt 2 hours p~l >7.0 Flotation 1000 g/t MBS
500 g/t CuSO"
50 g/t PAX
50 g/t 4037 p~l 9 - 9.5 The ~ /-309~ study indicatecl that tlle project should proceed to more detailed evaluation. In particular flnt~tion conditions were addressed.
25 Previous MBS, CuSO4 and pH level~ were found to be satisfactory.
The combined collector of PAX nnd ~10'17 was foul-d to be unsatisfactory during locked cycle testwork. The collector wa~ chnnge~3 to SIBX only , . . ' :
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W O 92/13640 PC~/AU92/00043 which corrected this and gave the ndditional benefit of using one collector site wide.
The alkali used for pll modlicatlon was also varied with lime and caustic soda being investigated. Each gnve similar results.
5 The conditions for upgradlng the talling were deflned as:
Conditionin~ 2 hours - minimum pH >7.0 Flotation 1000 g/ t MB S
500 g/t CuS04 100 g/t SIBX
pH 9 - 9.5 Extended eondltlonlng times for a perlod up to 24 hours have also been investigated and with no real change occurring after 2 hours.
Typieal results using these condltions on the tailing are a rougher 15 eoneentrate grade of 20 - 25% zinc at a zinc recovery of 80 - 85% in 1096 of the weight. Typlcal Iron recovery into the rougher concentrate is around 10%. Cleaning glves a final concentrate of 4796 zinc at 61%
recovery .
The optimised Zn recovery for A snmple wlth average composltion of the 20 dam is shown in Flgure 2. These results were attained wlth an average conditioning time oî 1 hour or more.
Further investigations took place to r ed~lce the conclitioning times whilst at least maintainlng recovery Integrit y .
A slurry was formecd by introducing tailings and water from the 25 Woodlawn Mine into two large holding tanks. Each tank held the slurrv for approximately lO hours. ~he.sr? tanks were filled cluring the day and eontinuously operatecd. rlle reecl rlowed throu~h these tanks and into the conditioning plant. l }~e ~shlrrv in the holding tanks was - . . .. .
:. , . . ' .: ' ' .. '- ' ~ . . : ,: -W O 92/13640 PC~r/AU92/00043 ~ 17 12. ~ , subjected to continuous mild agitation to ensure a slurry of predictable pulp density was delivered to the conditioning plant.
In the conditioning plant, the slurry w~s held for approximately 1 hour at a pH of about 7 . 7 . During this time the slurry hnd a pulp density o~
5 approximately 35%. The s]urry was also agitated during this period by an agitator capable of impartlng high int.enslty shear to the slurry.
Details of the agitator were as follows:
Type of Agitator Twin Level l~xial Flow/Radial Turbine Agitator Speed 6.2 m/ s Impeller Power Number (Np) 0.37 Axial Plow 5 . 0 Rsdial Turbine Impeller Pumping Number (Nq) 0.62 Axlal Flow 0.72 Radial Turbine 15 Installed Power/Unlt Volume 2.2 kw/m3 Torque/Unit Volume 206 Nm/m3 Ratio Impeller Diameter/
Tank Diameter I . 2m/4 . 25m = 0. 28 The conditioned slurry was then subjected to A flotation circuit under 2 0 the following conditions:
Flotation 600 g/t SO2 ndded SBS or MBS
500 g/t CuSO4 100 g/t SIBX
p~ 9 - 9.5 2 5 The flotation circuit was in severnl sepnrate stages . Each stage begAn with a feed which was separnted into A concentrate and A tail.
rougher stage was initially produced followed by three cleaning stages.
Feed for the rougher stage was from the condltioning plant and product from each stage provided the feed for ench succeeding stnge and the tail 30 returned to ench preceding stage. The circuit is illustrated in Figure 3.
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The results of the testing are sllown in Figure 4. Figure 4 shows graphically Zn grade versus Zn recovery curves for the samples. In particu]ar the four dlfferent samples tested were as follows:
TA-201 SAMPLE - Thls sample dld not undergo any conditionlng prior 5 to flotatlon.
TA-201 SAMPLE - This sample underwent conditioning for 60 minutes in a laboratory flotation cell.
TA-201 SAMPLE - Thls sample underwent condltloning for 60 minutes whilst being agitated wlth an agltator hflving a speed of 3.8 m/s. The 10 agltator was a laboratory scale twin level axial flow/radial turblne.
TA-201 SAMPLE - This sample unclerwent conditioning for 60 minutes whilst being agitated with the same agitator but havlng a speed of 6 . 4 m/s These examples showed that conditioning for tlmes below 2 hours, in 15 particular about 1 hour, whllst Imparting agitatlon to a slurry, improved flotation recovery of a base metal from the base metal bearing material.
Accordingly by selecting a particulnr pl~ range and minimum residence tlme it is possible to improve flotation recovery of A base meta] from a base metal bearing material.
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" t '., ~ r' ~ .' . ~ : ' . . ' '
Fleld of the Inventlon The invention relates to a method of treating a base metal bearing material for recovering a metal concentrate. Typically this method is 5 applied to treat a base metal tailing.
BackR~round of the Inventlon Whilst the followlng descrlptlon of the invention is with reference to trestment of a talllng, the Invention is not so limited.
Typically a base metal talling is produced from mineral dressing 10 operations located on an ore slte. Such a talllng may contain commercially significant amounts of base metals, such as copper, lead, zinc and nickel. In these cases, mine operators have wanted to recover these base metals from the tailing in an economically viable way.
In thls case a tailing is recovered such as by dredging or sluicing and 15 then is subjected to a concentrntion process, which may include flotation and/or other techniques such as grflvity, to produce base metal concentrates. These tailing flotation processes can, although not exclusively, be applied to the concentration of zinc sulphide minerals from a tailing.
20 However, minerals in a tailing dam generfllly respond only poorly to a flotation process. This is due to the chemlcal environment in which they have been stored subsequent to their previous treatment.
Furthermore, because of their genernily low grade a high upgrade ratio is required to achieve a saleable col-centrate.
25 Consequently, flotation trentments ol suc}- a tai}ing have had only limited success in producing reliable flnd s~leable bnse metal concentrates.
'~' - ' ~ ,' .
. ::
WO 92/13640 PCI'/AU92/00043 ~o~7 Summary Embodlments of the Invention Accordingly, investlgatlons have been directed to the effects of changes to the condltlonlng o the base metal bearing materlal and to the flotation conditions of the subsequent treatment.
5 Surprisingly, lt has been found that improvements in the level of recovery of base metal concentrate coupled wlth low pyrite recovery during the flotation step can be achleved by subJecting the base metal bearing material (e. g . tailing) to n particular conditioning treatment.
This can resull in substantial upgrading of feed values.
10 More particularly, there ls provided a method of condltioning a base metal bearing materlal for subsequent recovery of base metal concentrate comprises forming a slurry having a pulp denslty of at least 20% solids by the addition of water to the base metal material, and maintaining the slurry at a pH of at least 7 for a predetermined perlod of time. If 15 necessary the pH may be malntained at the 'desired pH by addition of alkali (e. g. lime or caustic soda) .
Preferably the base metal is copper, lead, zinc and/or nickel.
The pH of the slurry is maintained irl t}-e preferred r ange of 7.0 to 8.5.
The preferred range of pulp density o~ the slurry is from 30-60% solids.
20 The preferred predetermined perlod Or conditioning time is about 1 hour or more and more preferab]y from ~bout l hour to about 2 hours.
To optimise the conditioning Or the sllrf~ce of the base metal material it is desirable to agitate the slurry. In one preferred form of the invention the conditioning tre.stmenl comprises forming a ~lurry having a 25 pulp density of at least 20~6 solids l-y the additioll of water to the base metal bearing material and maintaining tllat slurry at a pH of at least 7 for a period of greater than F~bout I llour whilst agitating the slurry.
:, - . : , .:, :' . . ::, ,, , -- . :, : . . , . .
W O 92/13640 PC~r/AU92/00043 ~ 3, ~ 7 The regulation of a mlnimum pulp density and preferably agltation of the slurry has found to allow conditioning times less than that previously expected. The higher the pulp density or the more intense the agitation the shorter the conditioning times.
5 The agitation may be by any suitAI)le means. However, preferably the means imparts shear to the slurry whllst maintaining the slurry in suspension.
Typically, where the base metal bearing material ls a tailing it may lnclude sphalerlte, pyrlte and other base metal sulphide minerals mixed 10 with non-sulphide gangue materlals (e.g. talc, chlorite and quartz).
Investlgatlons have slso found that the temperature of the slurry has little effect on the condltlonlng step though elevated temperature may subsequently affect the flotation reagents. Similarly during this condltioning there ls no need to add the flotatlon reagents. However, 15 pH modlflers may be added. In fact lt ls preferable that those reactants (other than pH modlflers) are added after the condltloning stage.
The refinement in operating practice of the process of the invention has potentially important commercial implications for enhanced profitability and reliability of recovery. The conclitioning has facllltated reliable, 20 repeatable recoverles of base metal concentrate.
According to another preferred form oî this invention there is also provided a method of proclucing n bnse metal concentrate from a base meta] bearing talllng which comprises:
(a) recovering a base metal bearing tailing and placing it in one or 25 more vessels;
(b) adding water to the tniling lo form a slurry having a density of at least 20~ solids;
(c) maintaining the slurry nl a p]l of flt least 7 for a period of about I hour or more;
30 (d) adding at least one fJotn~ion reagent to the slurly;
(e) subjecting the sh~rry lo f1 i~tiO~l to recover the base metal concentrnte; and .
, .
W O 92/13640 . ` PC~r/AU92/00043 ~o~l7 "
(f) dewaterlng (e.g. thickening and filtering) the base metal concentrate.
According to another preferred form of this Invention there is also provided a method of producing a base metal concentrate from a ba.se 5 metal bearing tailing which comprises:
(a) recovering a base metal tailing and placing it In one or more vessels;
(b) addlng water which has a pH of at least 5 to the tailing to provide a slurry having a denslty Or at least 20%;
10 (c) agitating the slurry whilst maintaining the slurry at a pH of at least 7 for a perlod of up to about 2 hours;
(d) addlng at least one flotation reagent to the slurry;
(e) subjecting the slurry to flotation to produce the base metal concentrate; and 15 (f ) dewatering the base metal c oncentrate .
When necessary the water has been treated to maintain the pH by the addition of alkali reagents such as caustic soda or lime.
Preferably a number of holding vossels are used to provide a surge capacity to ensure continuous supply and the necessary conditioning for 2 o successful subsequent flotation .
Generally, flotation will take place in a number of stages (e.g. four), comprising a rougher stage followecl by a number of cleaning stages (e . g . three) . After the final cleaning stage the base metal zinc concentrate is de-watered.
25 Preferab]y, flotation reagents are used after the slurry has been preconditioned to render the desirecl mineral selectively smenable to the flotation process. The reAgent P~ddition is tailored to suit the mineral or minerals from which it is desired ~o r2cover the bnse metal.
In the case of recovery Or zinc beariTlg material, a number OT'` reagents 30 have been found to be prefers~blc. Tht? reagents added can be classified into three grour)s, namel v:
:
: ' W O 92/13640 PC~r/AU92/0~043 ~`- 5 ~ t (1) activators (such as CuSO,I);
(2) depressants (such as sodium metabisulphite (MBS)); and (3) collectors (such as potassium amyl xanthate (PAX), sodium isobutyl xanthate (SIBX) or a dithiophosphate thioncarbamate formulation ( e . g . AERO 4037 ) ) .
The reagents used in the rougher stage are typically either CuSO4, MBS, an alkali and SIBX or CuSO4, SBS, an alkali and SIBX. A
preferred reagent composition for the flotation is 1OOO g/t MBS, 500 g/t CuSO4, 100 g/t SIBX at a pH oE between 9 and 9. 5 .
The reagents used In the flrst and second cleaning stages are typically alkali and SIBX, preferably in the following amounts:
300 g/t NaOH and 0-10 g/t SIBX.
In the first cleanlng stage the pH of the slurry is preferably 10 to 11. 5 and more preferably 10 . 2 to 10 . 4, whereas in the second cleaning stage the pH is preferably 11 to 12.
The third cleaning stage generally uses an alkall, typically at an addition rate sufficient to give a pTI of least 11Ø
By way of example, the slurry density to the roughing stage should be about 25% ~ 40% solids and the slurry density in the cleaning stages should be in the range of about 20n ~ 50% solids and preferably 20 ~ 35%
solids .
The tailing from the rougller stAge m~y be pumped directly back to a disposal site. The water may be re-!overed for re-use.
It is far from clenr why condltloning ~ccordillg to the invention causes the significant improvement in the recovery of base metal concentrate.
However, it is thought to arise from an attrition which results in cleaner particle surfaces and/or some dis;lggregatioll of the particles which increases the liberation of minerals. In either case the particles are rendered more suseptible to subsequen~ ~lotation techiques.
,. .
t , .
" ' , ' ' -. :, W O 92/13640 PC~r/AU92/00043 ~ .
2101~17 6. ~
Brlef Descrlptlon of the Drawln~s The invention is now further illustrated wlth reference to the -~
accompanying clrawings and examples in which:
~IGURE 1 shows graphlcal]y a Zn grade versus Zn reeovery curve for 5 the roughening stage for a tailing whieh was eondltioned and one which was unconditioned in early test work.
FIGURE 2 shows graphieal]y a Zn recovery versus Zn grade eurve for the process of thls Invention.
FIGURE 3 ls a flow chart of the f]oation circuit used in the subsequent 10 investigation o the process of the invention.
FIGURE 4 shows graphica]ly a Zn grade versus Zn recovery curve for a series of samples under dlfferlng conditloning conditions.
In preliminary lnvestigstlons the first attempt to recover Zn from a tailing was a two stage process where the tai]ing was aerated in a slurry 15 and then a bulk concentrate wa.s floated from the slurry. The conditions for the aeration stage (2 - 6 hours) were 40% solids mixed in a pH 3 solution that contained a minimum quantity of copper (approximately 3 - 500 ppm Cu). Under these conditions up to 2096 of the contained zinc ( 10 - 20 g/l Zn ) was ]eached and the structure of the 2 o partic]es was modlfied . Aeration was followed by flotation at the natural pH of the aeration solution using a fAtty acid flotation reagent. A bulk zinc/]ead concentrate was produced.
When these parameters were carriecl out on a tailing from Wood]awn Mines New South Wflles Australia the results were not reproduced.
25 The process was then modifiecl to float a low grade zinc coneentrate (8 - 10% zinc at 80~ reeovery) using pH 3 wat~r contaminated with metal ions. The flotation step also recovered the pyrite from the tailing .
' . ' . ~ ' ' . ' ' ' ' ~ . '' ' :' ., . .,, : - . , : . .:
:, . : ~
W O 92tl3640 PCT/AU92/00043 ` 7.. ~ 7 While the process wns belng deflned In a pilot plsnt, a flotation concentrate was produced which floated less pyrite and contained a high grade ~inc rougher concentrnte with low weight recovery - typically 40% zinc at 409~ recovery.
5 The flotation conditions were:
1000 g/t MBS
500 g/t CuSO4 75 g/t PAX
pH 7.0 - 7.5 uslng tap water 10 Laboratory tests were then conducted to identlfy the conditions for achieving these results. The same reagent reglme in the laboratory did not give the same results as those from the Pllot Plant, i . e .
substantially more pyrite was recovered in the concentrate.
The Pllot Plant practice was then investlgated. The procedure for 15 mlxing a batch of tailing for feed was recognised as a maJor variation.
There was a delay of at least 2 hours between when the slurry was prepared and the flotation test carried out.
This procedure was simulated in the laboratory by conditioning the slurry sample in the float cell for 2 hours prior to flotation with water 20 at pH 5-7. Simllar results to those for the Pilot Plant resulted. An example of the results showing this conditioning effect is shown in Figure 1.
A flotation reagent scheme of:
1000 g/t MBS
25 500 g/t CuSO4 75 g/t PAX
pH 7.0 - 7.5 was used.
- : :.-' . ' :-~
W O 92/13640 PC~r/A U92tO0043 21~1~17 However, this conditioning effect wns found to be variable with the procedure being unsuccessful In some sample locations.
The variables which were investigAted included time, temperature and pulp density. Included in this work WAS the optimisation of the flotation 5 reagents.
Temperature Initial indications were that by increasing the temperature the conditioning time could be decreased. This did not prove to be the case during these tests and a conditioning temperature of 20 to 40C was 10 best. The effect of tempersture is shown in Table 1.
Table 1 - Variation of Results with Conditioning Temperature TEMPERATURE ZINC CONCENTRATE GRADE ZINC RECOVERY
C % %
15.4 62 14. 1 65 3.4 33 1.8 6 At the high temperatures ( l40C) tl-e flotation reagents were destroyed.
2 o Time A minimum conditioning time of ni~ol~ 2 hours WFI.q estnblished and the variation with conditloning time is showll in Table 2.
.
~, . , .
Table 2 - Varlstlon of resu]ts wlth Conditionin~ Time TIME HRS ZINC CONCENTRATl~ GRADE ZINC RECOVERY
% %
O. 5 14 49 Pulp Denslt,y Tests showed that there was no variAtlon in float results when the pulp 10 density during conditionlng was variecl between 30 - 60% solids.
~lotation Reagents A series of reagents for flotation were evaluated. The end result was the choice of PAX and Cyanamid Aero 4037 for the flotation.
Results were still variable dependin g on sample location . Testwork to 15 date had been restricted to near surfnce samples.
At this point, +/-30% engineering estimste was carried out for the processlng of 3 milllon tonnes of tPIillng.
At thls stage the p~I reglme wAS increAsed to a p~ of 9-9. 5 .
20 The deslgn basis ror the flotation wns derine(3 as:
1000 g/t MBS
500 g/t CuSO4 50 g/t PAX
50 g/t 4037 P~-~ 9.0 - 9.5 , -: . . - : ~, W092/13~40 Z101Sl7 IO. PCT/~L192/00043 In conjunction with this study n core sAmpling programme of the tai3ing dam at Woodlawn Mines was completed and the preconditioning and flotation condltlons were tested on core sections.
The previous variability in results with locntion was not present. -5 Analysis of the results indicated this mny have been due to the fact that the natural pH of the slurry was always greater than 7.
Also during the 2 hour conditioning period the pH remained above 7.
The importance of maintaining the pH of at least 7 during the conditioning period was then confirmed when It was shown that samples 10 that had not previously responded to the technique did so when the slurry pH was adjusted to above 7 by addition of lime for the condltioning period.
Zinc rougher recoveries for the core samples averaged ~3 - 85%
compared to previous 7596. Feed grnde varied between 2 - 7% Zn.
15 The flotation sequence was defined ns:
ConditioninEt 2 hours p~l >7.0 Flotation 1000 g/t MBS
500 g/t CuSO"
50 g/t PAX
50 g/t 4037 p~l 9 - 9.5 The ~ /-309~ study indicatecl that tlle project should proceed to more detailed evaluation. In particular flnt~tion conditions were addressed.
25 Previous MBS, CuSO4 and pH level~ were found to be satisfactory.
The combined collector of PAX nnd ~10'17 was foul-d to be unsatisfactory during locked cycle testwork. The collector wa~ chnnge~3 to SIBX only , . . ' :
" ' , ' . ' . '.
, .. . .
W O 92/13640 PC~/AU92/00043 which corrected this and gave the ndditional benefit of using one collector site wide.
The alkali used for pll modlicatlon was also varied with lime and caustic soda being investigated. Each gnve similar results.
5 The conditions for upgradlng the talling were deflned as:
Conditionin~ 2 hours - minimum pH >7.0 Flotation 1000 g/ t MB S
500 g/t CuS04 100 g/t SIBX
pH 9 - 9.5 Extended eondltlonlng times for a perlod up to 24 hours have also been investigated and with no real change occurring after 2 hours.
Typieal results using these condltions on the tailing are a rougher 15 eoneentrate grade of 20 - 25% zinc at a zinc recovery of 80 - 85% in 1096 of the weight. Typlcal Iron recovery into the rougher concentrate is around 10%. Cleaning glves a final concentrate of 4796 zinc at 61%
recovery .
The optimised Zn recovery for A snmple wlth average composltion of the 20 dam is shown in Flgure 2. These results were attained wlth an average conditioning time oî 1 hour or more.
Further investigations took place to r ed~lce the conclitioning times whilst at least maintainlng recovery Integrit y .
A slurry was formecd by introducing tailings and water from the 25 Woodlawn Mine into two large holding tanks. Each tank held the slurrv for approximately lO hours. ~he.sr? tanks were filled cluring the day and eontinuously operatecd. rlle reecl rlowed throu~h these tanks and into the conditioning plant. l }~e ~shlrrv in the holding tanks was - . . .. .
:. , . . ' .: ' ' .. '- ' ~ . . : ,: -W O 92/13640 PC~r/AU92/00043 ~ 17 12. ~ , subjected to continuous mild agitation to ensure a slurry of predictable pulp density was delivered to the conditioning plant.
In the conditioning plant, the slurry w~s held for approximately 1 hour at a pH of about 7 . 7 . During this time the slurry hnd a pulp density o~
5 approximately 35%. The s]urry was also agitated during this period by an agitator capable of impartlng high int.enslty shear to the slurry.
Details of the agitator were as follows:
Type of Agitator Twin Level l~xial Flow/Radial Turbine Agitator Speed 6.2 m/ s Impeller Power Number (Np) 0.37 Axial Plow 5 . 0 Rsdial Turbine Impeller Pumping Number (Nq) 0.62 Axlal Flow 0.72 Radial Turbine 15 Installed Power/Unlt Volume 2.2 kw/m3 Torque/Unit Volume 206 Nm/m3 Ratio Impeller Diameter/
Tank Diameter I . 2m/4 . 25m = 0. 28 The conditioned slurry was then subjected to A flotation circuit under 2 0 the following conditions:
Flotation 600 g/t SO2 ndded SBS or MBS
500 g/t CuSO4 100 g/t SIBX
p~ 9 - 9.5 2 5 The flotation circuit was in severnl sepnrate stages . Each stage begAn with a feed which was separnted into A concentrate and A tail.
rougher stage was initially produced followed by three cleaning stages.
Feed for the rougher stage was from the condltioning plant and product from each stage provided the feed for ench succeeding stnge and the tail 30 returned to ench preceding stage. The circuit is illustrated in Figure 3.
. : . . . :
.
.
, . : ~' , ~ : .
' . . -. ,, . : ... ~ , : . ~ , ,, WO 92/13640 ,ci ~ /AU92/00043 13.
The results of the testing are sllown in Figure 4. Figure 4 shows graphically Zn grade versus Zn recovery curves for the samples. In particu]ar the four dlfferent samples tested were as follows:
TA-201 SAMPLE - Thls sample dld not undergo any conditionlng prior 5 to flotatlon.
TA-201 SAMPLE - This sample underwent conditioning for 60 minutes in a laboratory flotation cell.
TA-201 SAMPLE - Thls sample underwent condltloning for 60 minutes whilst being agitated wlth an agltator hflving a speed of 3.8 m/s. The 10 agltator was a laboratory scale twin level axial flow/radial turblne.
TA-201 SAMPLE - This sample unclerwent conditioning for 60 minutes whilst being agitated with the same agitator but havlng a speed of 6 . 4 m/s These examples showed that conditioning for tlmes below 2 hours, in 15 particular about 1 hour, whllst Imparting agitatlon to a slurry, improved flotation recovery of a base metal from the base metal bearing material.
Accordingly by selecting a particulnr pl~ range and minimum residence tlme it is possible to improve flotation recovery of A base meta] from a base metal bearing material.
.
" t '., ~ r' ~ .' . ~ : ' . . ' '
Claims (28)
THE CLAIMS DEFINING THE INVENTION ARE AS FOLLOWS:
1. A method of conditioning base metal tailings for subsequent production of a base metal concentrate which comprises forming a slurry having a pulp density of at least 20% solids by the addition of water to the base metal tailings, and maintaining the slurry at a pH of at least 7 for a predetermined period of time.
2. A method of conditioning a base metal bearing material for subsequent production of a base metal concentrate prior to the addition of flotation reagents or the extraction of the base metal or metals which comprises forming a slurry having a pulp density of at least 20% solids by the addition of water to the base metal bearing material, and maintaining the slurry at a pH of at least 7 for a predetermined period of time.
3. A method of conditioning according to either claim 1 or 2 wherein the pH of the slurry is maintained by the addition of alkali.
4. A method of conditioning according to any one of claims 1, 2 or 3 wherein the slurry is maintained at a pH in the range of 7.0 to 8.5.
5. A method of conditioning according to any one of claims 1 to 4 wherein the pulp density of the slurry is from at least 30% solids.
6. A method of conditioning according to any one of claims 1 to 4 wherein the pulp density of the slurry is from 30% - 60% solids.
7. A method of conditioning according to any one of claims 1 to 6 wherein the predetermined period of time is about 1 hour or more.
8. A method of conditioning according to any one of claims 1 to 6 wherein the predetermined period of time is from about 1 hour to about 2 hours.
9. A method of conditioning base metal tailings for subsequent production of a base metal concentrate which comprises forming a slurry 15.
having a pulp density of at least 20% solids by the addition of water to the base metal tailings and maintaining that slurry at a pH of at least 7 for a period of greater than about 1 hour whilst agitating the slurry.
having a pulp density of at least 20% solids by the addition of water to the base metal tailings and maintaining that slurry at a pH of at least 7 for a period of greater than about 1 hour whilst agitating the slurry.
10. A method of conditioning a base metal bearing material for subsequent production of a base metal concentrate prior to the addition of flotation reagents or the extraction of the base metal or metals which comprises forming a slurry having a pulp density of at least 20% solids by the addition of water to the base metal bearing material and maintaining that slurry at a pH of at least 7 for a period of greater than about 1 hour whilst agitating the slurry.
11. A method of conditioning according to either claim 9 or 10 wherein the agitation imparts shear to the slurry whilst maintaining the slurry in suspension.
12. A method of conditioning according to any one of claims 1 to 11 wherein the base metal tailings include sphalerite, pyrite or other base metal sulphide minerals mixed with non-sulphide gangue materials.
13. A method of conditioning according to any one of claims 1 to 11 wherein the base metal bearing material includes sphalerite, pyrite or other base metal sulphide minerals mixed with non-sulphide gangue materials.
14. A method of producing a base metal concentrate from base metal tailings which comprises:
(a) recovering base metal tailings and placing them in one or more vessels;
(b) adding water to the base metal tailings to form a slurry having a pulp density of at least 20% solids;
(c) maintaining the slurry at a pH of at least 7 for a period of about 1 hour or more and optionally adding at least one flotation reagent to the slurry;
(d) where a flotation reagent has not been added in step (c), adding at least one flotation reagent to the slurry;
16.
(e) subjecting the slurry to flotation to recover the base metal concentrate; and (f) dewatering the base metal concentrate.
(a) recovering base metal tailings and placing them in one or more vessels;
(b) adding water to the base metal tailings to form a slurry having a pulp density of at least 20% solids;
(c) maintaining the slurry at a pH of at least 7 for a period of about 1 hour or more and optionally adding at least one flotation reagent to the slurry;
(d) where a flotation reagent has not been added in step (c), adding at least one flotation reagent to the slurry;
16.
(e) subjecting the slurry to flotation to recover the base metal concentrate; and (f) dewatering the base metal concentrate.
15. A method of producing a base metal concentrate from base metal tailings which comprises:
(a) recovering base metal base metal tailings and placing them in one or more vessels;
(b) adding water which has a pH of at least 5 to the base metal tailings to provide a slurry having a pulp density of at least 20%
solids;
(c) agitating the slurry whilst maintaining the slurry at a pH of at least 7 for a period of up to about 2 hours and optionally adding at least one flotation reagent to the slurry;
(d) where a flotation reagent has not been added in step (c), adding at least one flotation reagent to the slurry;
(e) subjecting the slurry to flotation to produce the base metal concentrate; and (f) dewatering the base metal concentrate.
(a) recovering base metal base metal tailings and placing them in one or more vessels;
(b) adding water which has a pH of at least 5 to the base metal tailings to provide a slurry having a pulp density of at least 20%
solids;
(c) agitating the slurry whilst maintaining the slurry at a pH of at least 7 for a period of up to about 2 hours and optionally adding at least one flotation reagent to the slurry;
(d) where a flotation reagent has not been added in step (c), adding at least one flotation reagent to the slurry;
(e) subjecting the slurry to flotation to produce the base metal concentrate; and (f) dewatering the base metal concentrate.
16. A method of producing a base metal concentrate from base metal tailings according to either claims 14 or 15 wherein in step (b) the slurry has a pulp density of at least 30% solids.
17. A method of producing a base metal concentrate from base metal tailings according to either claims 14 or 15 wherein the pH of the slurry is maintained by addition of alkali.
18. A method of producing a base metal concentrate from base metal tailings according to either claims 14 or 15 wherein the at least one flotation reagent is not added in step (c).
19. A method of producing a base metal concentrate from base metal tailings according to any one of claims 14 to 18 wherein the base metal is copper, lead, zinc or nickel.
17.
17.
20. A method of producing a zinc concentrate according to claim 19 wherein the flotation reagent is a combination of at least one compound selected from each of the following three groups:
(1) activators including CuSO4;
(2) depressants including sodium metabisulphite (MBS); and (3) collectors including potassium amyl xanthate (PAX), sodium isobutyl xanthate (SIBX) or a dithiophosphate thioncarbamate formulation (e.g. AERO 4037)).
(1) activators including CuSO4;
(2) depressants including sodium metabisulphite (MBS); and (3) collectors including potassium amyl xanthate (PAX), sodium isobutyl xanthate (SIBX) or a dithiophosphate thioncarbamate formulation (e.g. AERO 4037)).
21. A method of producing a base metal concentrate from base metal tailings according to any one of claims 14 to 20 wherein the flotation reagent is 1000 g/t MBS, 500 g/t CuSO4, 100 g/t SIBX at a pH of between 9 and 9.5.
22. A method of producing a base metal concentrate from base metal tailings according to either claims 14 or 15 wherein flotation occurs in a plurality of stages which comprises a rougher stage followed by at least a first cleaning stage and a second cleaning stage.
23. A method of producing a base metal concentrate from base metal tailings according to claim 22 wherein the flotation reagent used in the rougher stage is either CuSO4, MBS, an alkali and SIBX or CuSO4, SBS, an alkali and SIBX.
24. A method of producing a base metal concentrate from base metal tailings according to claim 22 wherein the flotation reagent used in the first cleaning stage and the second cleaning stage is an alkali and a collector respectively.
25. A method of producing a base metal concentrate from base metal tailings according to claim 22 wherein the first cleaning stage the pH of the slurry is between 10 to 11.5, and in the second cleaning stage the pH is between 11 to 12.
26. A method of producing a base metal concentrate from base metal tailings according to claim 22 further comprising a third cleaning stage 18.
wherein the flotation reagent used is an alkali added at a rate sufficient to maintain a pH of at least 11Ø
wherein the flotation reagent used is an alkali added at a rate sufficient to maintain a pH of at least 11Ø
27. A method of producing a base metal concentrate from base metal tailings according to claim 22 wherein the slurry density in the roughing stage is between about 25% - 40% solids and the slurry density in the cleaning stages is between about 20% - 50% solids.
28. A method of producing a base metal concentrate from base metal tailings according to any one of claims 22 to 27 wherein the base metal tailings from the rougher stage are pumped directly back to a disposal site and water is recovered for re-use.
Applications Claiming Priority (4)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| AUPK4510 | 1991-02-06 | ||
| AUPK451091 | 1991-02-06 | ||
| AUPK9628 | 1991-11-22 | ||
| AUPK962891 | 1991-11-22 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA2101517A1 true CA2101517A1 (en) | 1992-08-07 |
Family
ID=25644002
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA002101517A Abandoned CA2101517A1 (en) | 1991-02-06 | 1992-02-06 | Method of treating a base metal bearing material |
Country Status (5)
| Country | Link |
|---|---|
| CA (1) | CA2101517A1 (en) |
| MX (1) | MX9200479A (en) |
| PT (1) | PT100091A (en) |
| WO (1) | WO1992013640A1 (en) |
| YU (1) | YU11692A (en) |
Families Citing this family (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| AUPO590997A0 (en) * | 1997-03-26 | 1997-04-24 | Boc Gases Australia Limited | A process to improve mineral flotation separation by deoxygenating slurries and mineral surfaces |
| CN102849757B (en) * | 2012-09-28 | 2014-04-02 | 中国科学院青海盐湖研究所 | Floatation method for extracting aphthitalite and potassium chloride from carbonate type salt lake |
| CN105457760B (en) * | 2015-12-22 | 2017-10-31 | 广西中金岭南矿业有限责任公司 | A kind of preparation method of pyrite activator |
Family Cites Families (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US1448929A (en) * | 1920-03-06 | 1923-03-20 | Luckenbach Processes Inc | Concentration of ores by flotation |
| US1678259A (en) * | 1927-06-30 | 1928-07-24 | Harold S Martin | Process of concentrating mixed-sulphide ores |
| US2470150A (en) * | 1946-01-02 | 1949-05-17 | Erie Mining Co | Froth flotation of oxide iron ore |
| CA872730A (en) * | 1968-06-07 | 1971-06-08 | F. Sirianni Aurelio | Process for separation of siliceous and phosphatic material from iron bodies |
| AU5374173A (en) * | 1973-03-26 | 1974-09-26 | Weston D | Flotation of copper ores |
| CA1062819A (en) * | 1976-12-07 | 1979-09-18 | Stephen J. Thorndyke | Flotation separation of copper and nickel sulfides |
| US4132635A (en) * | 1977-01-13 | 1979-01-02 | Michigan Technological University | Beneficiation of iron ores by froth flotation |
| GB2086768B (en) * | 1980-03-21 | 1983-02-23 | Inco Ltd | Selective flotation of nickel sulphide ores |
| CA1147970A (en) * | 1980-12-23 | 1983-06-14 | Victor A. Ettel | Process for cobalt recovery from mixed sulfides |
| US4929344A (en) * | 1989-05-01 | 1990-05-29 | American Cyanamid | Metals recovery by flotation |
-
1992
- 1992-02-04 MX MX9200479A patent/MX9200479A/en not_active IP Right Cessation
- 1992-02-04 PT PT100091A patent/PT100091A/en not_active Application Discontinuation
- 1992-02-04 YU YU11692A patent/YU11692A/en unknown
- 1992-02-06 CA CA002101517A patent/CA2101517A1/en not_active Abandoned
- 1992-02-06 WO PCT/AU1992/000043 patent/WO1992013640A1/en active Application Filing
Also Published As
| Publication number | Publication date |
|---|---|
| WO1992013640A1 (en) | 1992-08-20 |
| PT100091A (en) | 1993-05-31 |
| MX9200479A (en) | 1992-08-01 |
| YU11692A (en) | 1994-06-24 |
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| Date | Code | Title | Description |
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| FZDE | Discontinued | ||
| FZDE | Discontinued |
Effective date: 19990208 |