Note: Descriptions are shown in the official language in which they were submitted.
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.
'~' - ' ~ ,' .
. ::
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.
.
~,
. , .
WO 92/13640 PCI/AU92/00043
Table 2 - Varlstlon of resu]ts wlth Conditionin~ Time
TIME HRS ZINC CONCENTRATl~ GRADE ZINC RECOVERY
% %
0 12 69
O. 5 14 49
19 79
2 24 75
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'
~ .' . ~ : ' . . ' '
