ok molibdenita
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2008:012
M A S T E R ' S T H E S I S
Purification of Molybdenite Concentrates
Ikumapayi Fatai Kolawole
Luleå University of Technology
Master Thesis, Continuation Courses Minerals and Metallurgical Engineering
Department of Chemical Engineering and GeosciencesDivision of Process Metallurgy
2008:012 - ISSN: 1653-0187 - ISRN: LTU-PB-EX--08/012--SE
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Abstract
A molybdenite flotation concentrate was refined by selective removal of copper impurity with
minimum dissolution of molybdenum in the concentrate. Copper is present in the concentrate
mainly as copper sulphide, chalcopyrite. Investigations were carried out on the removal of the
sulphide by their selective leaching in sodium cyanide, ferric chloride and ferric sulphate
solutions.
Approximately 80 to 100% of the copper concentration was removed under optimum
conditions with sodium cyanide and ferric chloride solutions with little dissolution of
molybdenum while ferric sulphate solution was ineffective due to a number of factors such as
passivation of chalcopyrite, temperature and redox potential.
Leaching with sodium cyanide was carried out with stoichiometric concentration of the salt,
20% solids, at ambient temperature and pH above11 with oxygen as oxidative gas, for 53 and
74 hours. Ferric chloride leaching was carried out at 35% solids with 10% ferric chloride, 1%
cupric chloride, 20% calcium chloride all by weight of solution at an average temperature of
80oC and pH about zero for 4 hours. Ferric sulphate leaching was carried out with
stoichiometric concentration of ferrous sulphate as the starting solution in which ferric
sulphate solution was generated by oxidation with addition of oxygen and sulphur dioxide in
ratio 10:1, 10% solids and pH below 0.5 at 65oC for 5 hours.
The intermediate pregnant solution samples and final leach solutions from the tests were
analyzed for dissolved copper with the aid of atomic absorption spectrophotometer, (AAS)
and the purified concentrate (solid residue) was analyzed for molybdenum, copper and other
important elements with different analytical techniques such as: (i) AAS for molybdenum,
calcium, bismuth and low level copper, iron and lead. (ii) Solution X-ray for high-level
copper, iron and lead. (iii) Fire assay for gold and silver. (iv) Leco-owen for sulphur. (v) Field
ionization mass spectroscopy, (FIMS) instrument for mercury and (vi) Selective electrode for
chloride ion.
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Table of content
Abstract ................................................................................................................................2
Acknowledgement ................................................................................................................6
1 Introduction .....................................................................................................................7 1.1 History..........................................................................................................................7 1.2 Characteristics ..............................................................................................................9 1.3 Applications................................................................................................................10
1.3.1 Use as parts and alloys .........................................................................................10 1.3.2 Use as lubricant....................................................................................................10 1.3.3 Use in petrol-chemistry ........................................................................................11 1.3.4 Other uses ............................................................................................................11
1.4 Occurence...................................................................................................................11
2 Recovery of molybdenite .................................................................................................13
3 Molybdenite production in Boliden ................................................................................15 3.1 Background ................................................................................................................15 3.2 Aim ............................................................................................................................16
4 Literature review.............................................................................................................17 4.1 Purification of molybdenite concentrate ......................................................................17
4.1.1 Microorganisms ...................................................................................................17 4.1.3 Sulphuric acid ......................................................................................................17 4.1.4 Sulphuric acid with sodium dichromate solution, Na2Cr2O7..................................18 4.1.5 Leaching with oxygen and halide e.g. CuCl2 ........................................................18 4.1.6 Conventional leaching with chloride salts.............................................................18 4.1.7 Hydrochloric acid and hydrofluoric acid...............................................................19 4.1.8 SO2/O2 as a strong oxidant ...................................................................................19
4.2 Molybdenum-water chemistry ....................................................................................19 4.3 Equilibrium diagram and its interpretation ..................................................................21
4.3.1 Stability of the compounds of trivalent molydenum..............................................25 4.3.2 Stability of the compounds of tetravalent molybdenum ........................................25 4.3.3 Stability of the compounds of pentavalent molybdenum, molybdenum blue.........26 4.3.4 Stability of the compounds of hexavalent molybdenum........................................26
4.4 Cu-CN chemistry........................................................................................................28 4.5 Leaching theory ..........................................................................................................34
4.5.1 Leaching under pressure.......................................................................................34 4.5.2 Leaching theory ...................................................................................................34
5 Experimental ...................................................................................................................35 5.1 Material ......................................................................................................................35 5.2 Reactors......................................................................................................................36 5.3 Methodology ..............................................................................................................37 5.4 Test performance ........................................................................................................37
5.4.1 Cyanide leaching..................................................................................................37 5.4.2 Ferric chloride (FeCl3) leaching ...........................................................................40 5.4.3 Ferric sulphate Fe2(SO4)3 leaching .......................................................................41
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6 Results and discussions....................................................................................................45 6.1 Cyanide tests...............................................................................................................45
6.1.1 Mo-losses to solution ...........................................................................................49 6.1.2 Concentrate weight changes .................................................................................49 6.1.3 Oxidising gases ....................................................................................................49
6.2 Ferric chloride (FeCl3) leaching ..................................................................................50 6.3 Ferric sulphate Fe2(SO4)3 leaching ..............................................................................51
7 Conclusion........................................................................................................................55
8 Suggested further work ...................................................................................................56
References...........................................................................................................................57
Appendix.............................................................................................................................60 Appendix I....................................................................................................................60
Properties of molybdenum ....................................................................................60 Appendix II ..................................................................................................................61
Estimation of reagent for cyanide leaching............................................................61 Procedure for cyanide leaching .............................................................................62 Table I-II Protocol and reagent consumption NaCN ..............................................63 Table I-III: Protocol and reagent consumption NaCN............................................64 Table I-IV: Leaching profile and material balance NaCN ......................................64 Table I-V: Profile of residual elemental analysis NaCN.........................................67 Table I-VI: Protocol and reagent consumption NaCN + CaCl2...............................67 Table I-VII: Protocol and reagent consumption NaCN + CaCl2 .............................68 Table I-VIII: Leaching profile and material balance...............................................68 Table I-IX: Profile of residual elemental analysis NaCN + CaCl2 ..........................71
Appendix III .................................................................................................................71 Procedure for ferric chloride leaching....................................................................71 Table I-X: Protocol and reagent consumption FeCl3 ..............................................72 Table I-XI: Protocol and reagent consumption FeCl3 .............................................73 Table I-XII: Leaching profile and material balance FeCl3 ......................................73 Table I-XIII: Profile of residual elemental analysis FeCl3 ......................................76
Appendix IV.................................................................................................................76 Procedure for ferric sulphate leaching ...................................................................76 Table I-XIV: Leaching profile and reagent consumption Fe2(SO4)3 ........................77 Table I-XV: Leaching profile and reagent consumption Fe2(SO4)3 .........................77 Table I-XVI: Molybdenum dissolution sodium cyanide leaching ...........................78 Table I-XVII: Molybdenum dissolution sodium cyanide + calcium chloride leaching.............................................................................................................................78
Appendix V ..................................................................................................................78 Atomic absorption spectroscopy, AAS ..................................................................78 Copper analysis with AAS ....................................................................................79 Procedure for copper analysis with AAS ...............................................................79 Table II-I: Copper concentration in standard solution and corresponding AAS signals...................................................................................................................80 XRD Equipment ...................................................................................................81
Appendix VI.................................................................................................................82 Composition of commercial molybdenite concentrate brands ................................82 Table II-II Molybdenum Concentrate Brands & Chemical Compositions ..............82
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Composition of commercial molybdenum oxide brands ........................................83 Table II-III :Technical Grade Molybdenum Oxide.Sadaci .....................................83 Table II-IV :Purified concentrates analysis. ...........................................................83 Table II-V: Cyanide leaching test assays................................................................83 Table II-VI: Cyanide + Chloride leaching test assays.............................................83 Table II-VII: Ferric chloride leaching test assays ...................................................84
Appendix VII................................................................................................................85 SCN formation......................................................................................................85
Appendix VIII ..............................................................................................................87 Mo-CN stability constant ......................................................................................87
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Acknowledgement
This research work is completed in partial fulfillment of the requirements for the award of
Masters of Science, Msc in minerals and metallurgical engineering, Luleå University of
Technology, (Ltu). Boliden Mineral AB assigned the project and it entails experiments that
involve handling of various materials and equipment such as chemicals, analytical
instruments, laboratory space and other logistics. All experiments and most analysis were
carried out in the mineral-processing laboratory at Boliden Mineral AB. XRD analysis was
carried out at the Department of Chemical Engineering, Ltu.
Hence my profound gratitude goes to the management and staff of Boliden Mineral AB for
providing all logistics and for giving me free access to use all facilities.
My appreciation goes to my supervisors: Prof. Åke Sandström (Ltu) and Jan-Eric Sundkvist
(Boliden) who has guided me through the research work; they are the brains behind the
success of this project. Many thanks also go to the laboratory staff: Rolf Danielsson, Amang
Saleh, Maria Lagerhielm, Johan Hansson and Mikael Widman who was ever ready to assist
me with all laboratory logistics and Chandra Sekhar Gahan who assisted with XRD analysis
at Ltu. The moral supports of all staff at the mineral processing department in Boliden are
greatly appreciated.
Finally my greatest thanks goes to my family members and particularly my late Dad; Ramon
Ikumapayi, my mother; Fausat Ikumapayi, my brother; Nurudeen Ikumapayi, and Anna
Bauer, all of whom have greatly contributed spiritually, morally and financially towards
successful completion of my study. It would have been impossible without them.
Boliden, August 2007
Fatai Ikumapayi
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1 Introduction
1.1 History
Molybdenite, originate from the Greek word molybdos, meaning lead, it is the principal
mineral from which molybdenum metal is now extracted. It was previously known as
molybdena. Molybdena was often confused with and implemented as graphite even after the
two ores were distinguishable; molybdena was still thought to be a lead mineral. In 1754,
Bengt Qvist examined the mineral and determined that it did not contain lead. It was in 1778
that a Swedish chemist named Carl Wilhelm Scheele put a clear light to the identification of
the metal. He discovered that molybdena was neither graphite nor lead. He and other chemists
then correctly assumed that it was the mineral of a distinct new element, named molybdenum
for the mineral in which it was discovered. In 1781 Peter Jacob Hjelm used linseed oil and
carbon to successfully isolate molybdenum [15].
The first major use of molybdenum came during World War I when its addition produced
steel with excellent toughness and strength at high temperatures for use as tank armor and in
aircraft engines. The major source through 19th century was Knaben mine in Norway in which
the molybdenite was concentrated by hand sorting. In 1918 Climax molybdenum company
(Colorado) uses the froth flotation process to concentrate the ore. Although the Climax mine
was shut down after the war due to decreased demand for the metal, it was however reopened
in 1924 when better peacetime applications was developed largely in the automotive industry.
In 1933 molybdenum production as a by-product of copper began when Anaconda Company
developed a method of selective flotation of molybdenite from porphyry copper ores at its
subsidiary Cananea in Mexico. The Kennecott Copper Company mine in Utah, El Teniente
mine in Chile and Anaconda’s Chuquicamata mine follows suite.
The United States supplied about 90% of the world molybdenum demand during the World
War II, most coming from Climax with balance from Kennecott’s Utah mines and Molycorp’s
Questa mine. Chile, Mexico and Norway remain the largest of the other western -world
producing countries.
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Climax Molybdenum Company continues to be the largest western-world producer into the
1980s while production from Mexico and Norway remained small, less expensive by-product
from Chile and United states continue to grow. Endako and other mines in Canada also began
molybdenum production. In 1977 the total world production of molybdenum exceeded 200
million pounds (approx. 90 000t) for the first time.
Late 1970s marks the opening of several molybdenum mines and expansion of many by-
products facilities; this was stimulated due to high price of molybdenum. However price
decline in the early 1980s made the by-products molybdenum a dominant economic force in
the market. The underground mine in Climax was closed by Climax molybdenum company
but the company still remained as a major producer with Henderson mine, and the open pit
mine at Climax Cyprus Minerals. Other major producers at the end of 1980s were Codelco’s
Chuquicamata mine in Chile, Placer’s Endako mine in Canada, the LaCaridad mine in
Mexico, and the Cuajone and Toquepala mines in Peru. The world annual production and
consumption averaged (180-200) million pounds (80 000- 90 000 t) of molybdenum during
the late 1980s [11].
Molybdenum has a value of approximately $65,000 per tonne as of 4 May 2007. It has
maintained a price at or near $10,000 per tonne from 1997 through 2002, and reached a peak
of $103,000 per tonne in June 2005. The world's largest producers of molybdenum materials
remained the United States, Canada, Chile, Russia, and China. In 2005, USA remains the top
producer of molybdenum with about 30% world share followed by Chile and China, figure1-1
[15].
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Figure1-1: Molybdenum output in 2005 [16]
The figure shows the world’s largest producer of molybdenum in 2005 as a percentage of the
top producer. The green band depicts 100% and corresponds to the United States as the top
producer with annual production of 56,900 tonnes, the yellow band depicts 10%, and the red
band depicts 1% respectively. The yellow bands correspond to Canada, Chile, China, and
other South American producers, while the red bands correspond to Russia, Mexico and other
Asian producer.
1.2 Characteristics
Molybdenum is a transition metal, silvery white in its pure metal form and very hard, it is
somewhat more ductile than tungsten. It has a melting point of 2623°C, and only tantalum,
rhenium and tungsten have higher melting points. Molybdenum burns only at temperatures
above 600°C. Its expansion during heating is the lowest compared to other commercially used
metals [15].
Molybdenum disulphide, also known as molybdenum sulphide or molybdenum (IV) sulphide,
with the molecular formula MoS2, molar mass 160.07 g/mol, and density 5.06 g/cm3 is a
black crystalline sulphide of molybdenum. It occurs as the mineral molybdenite and it is the
main commercial source of molybdenum. It is insoluble in all solvents and un-reactive toward
dilute acids. Its melting point is 1185°C, but it starts oxidizing in air from 315°C, limiting the
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range of its use as a lubricant in the presence of air between the temperatures of -185 and
+350°C; in non-oxidizing environments it is stable up to 1100°C [8]. Structurally MoS2 is
trigonal prismatic at Mo, and pyramidal at S. Molybdenum disulphide contains approximately
60% Mo, and 40% S. Detailed properties of molybdenum can be found in appendix I.
1.3 Applications
1.3.1 Use as parts and alloys
Molybdenum can withstand excessive heat at extreme temperatures without softening or
expanding, this unique property makes it very useful in the manufacture of aircraft parts,
filaments, amor tanks, electrical contacts and other applications that involve intence heat.
Molybdenum have very high weldability and it is also highly resistant to corrosion, hence
used in making high strength steel alloys: More than 43 000 tonnes of molybdenum is used
annually as alloying agent in high temperature superalloys stainless steels and tool steels,
although molybdenum contributes only about 8 to 25% composition of the alloys. It is being
implemented in place of tungsten due to its low density. It can be implemented both as an
alloying agent and as a flame-resistant coating for other metals. Molybdenum is better suited
for use in vacuum environments because it oxidises rapidly at temperatures above 760°C
although its melting point is 2623°C [15].
1.3.2 Use as lubricant
Molybdenum disulphide (MoS2) can form strong films on metal surface, highly resistant to
both extreme temperatures and high pressures; hence it is used as a lubricant and anti-
corrosion agent [15]. The structure, texture and appearance of molybdenum disulphide are
very similar to that of graphite; it is composed of sandwiched layers of molybdenum atoms
between the layers of sulphur atoms. The interactions between the sulphide atoms sheets are
weak and this is why MoS2 has a lubricating effect. Powdered MoS2 with particle sizes in the
range of 1-100 µm is a common dry lubricant. It is also often mixed into various oils and
greases, which keep the lubricated mechanisms running for a while longer, even in cases of
almost complete oil loss this is an important factor that makes it very relevant to aircraft
engines. It is often used also in motorcycle engines, especially in two-stroke engines, which
are otherwise not well lubricated. MoS2 grease is recommended for constant velocity joint
(CV) and universal joints. It is also used as a lubricating additive to special plastics, notably
nylon and teflon. During the Vietnam war, a commercial molybdenum disulphide product,
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"Dri-Slide", was used for lubricating troops' weapons; the military refused to supply it, as it
was "not in the manual", so it was sent to soldiers by their parents and friends privately.
Another application is for coating bullets, giving them easier passage through the rifle barrel
with less deformation and better ballistic accuracy. Self-lubricating composite coatings for
high temperature applications were developed at the Oak Ridge National laboratory. A
composite coating of molybdenum disulphide and titanium nitride was created on the surface
of parts by chemical vapor deposition”.
1.3.3 Use in petrol-chemistry
Artificial MoS2 can also be used as a catalyst in petroleum refineries, specifically for
desulphurization of crude oil e.g. hydrodesulphurization. It has also been discovered that
doping with small amounts of cobalt and alumina can enhance its effectiveness as a catalyst.
This type of catalyst can be generated in-situ by treating molybdate/cobalt-impregnated
alumina with H2S or an equivalent reagent [8].
1.3.4 Other uses
Molybdenum trioxide (MoO3) is used as an adhesive to bind enamels and metals.
Molybdenum powder is also used in agriculture as a fertilizer for some vegetable plants, such
as cauliflower [15].
1.4 Occurence
Molybdenum commonly occurs in nature as the mineral molybdenite, MoS2, and can also be
found as veins in quartz rock [11]. Porphyry copper ores are characterized by sulphide
fractions containing traces of molybdenite which is usually separated as a flotation
concentrate containing up to 90% MoS2. This concentrate is normally characterized by high
rhenium concentration (about 700 ppm Re) and is a major source of this metal. It may also be
a source of uranium [1]. Though molybdenum is also found in such minerals as wulfenite
(PbMoO4) and powellite (CaMoO4), the main commercial source of molybdenum is
molybdenite (MoS2). Molybdenite is mined as a principal ore, and is also recovered as a by-
product of copper and tungsten mining such as in the porphyry copper ore aforementioned.
Molybdenum is the 42nd most abundant element in the universe, and the 25th most abundant
element in Earth's oceans, with an average of 10.8 mt/km³. The Russian Luna 24 mission
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discovered a single molybdenum-bearing grain (1 × 0.6 µm) in a pyroxene fragment taken
from Mare Crisium on the Moon [15].
Figure 1-2a: Molybdenite sample [17] Figure 1-2b: Molybdenite sample from Aitik.
(From a primary ore mine) (From a secondary ore mine)
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2 Recovery of molybdenite
Molybdenum ore can be mined either by underground or open-pit method as a primary ore
body and can also be produced as a by-product or a co-product of copper or scheelite
production. Typical primary ore body can contain 0.05-0.25% Mo and secondary ore bodies
from a copper porphyry ore with average of 0.3-1.6 % Cu can contain 0.01-0.05% Mo. The
preferred method for concentrating the mineral has been flotation. The flotation process starts
after grinding the ore to liberate the molybdenite singly or in combination with copper
sulphide or other sulphide minerals from the host rock, and then agitating the grounded ore
with water, a collector and other special chemicals to cause preferential wetting, settling
and/or suspension of the host rock particles, while the hydrophobic or un-wetted particles of
molybdenum and copper minerals are carried by air bubbles to the surface where they can be
subsequently recovered by a frothing agent [11].
Molybdenite belongs to the minerals group of easy flotation, which is related to its crystal
structure. Grounded particles of molybdenite present laminar structure that favors natural
hydrophobicity. This elevated hydrophobic capacity allows recovery of molybdenum
successfully from ores with low grades by flotation. The separation of molybdenum-copper
has been practiced for a long time using flotation technique. The predominant process is
normally to depress copper sulphides, capitalizing on the surface property of the easy
floatability of molybdenite. Several reagents have been used as depressant for copper
minerals, for example sodium cyanide, mainly used if copper is present as chalcopyrite. Other
reagents are certain types of sulphides such as sodium sulphide, sodium hydrogen sulphide or
phosphorous pentasulphide when there is a mixture of copper sulphides: bornite, chalcocite,
and chalcopyrite. There are other methods, which are not very common, this include the use
of sodium peroxide or sodium hypochlorite. The depressant agent mostly employed is sodium
hydrogen sulphide (NaHS), which is chosen considering environmental aspects. When the
reagent is added, the potential of copper sulphides is reduced to the point where they are un-
floatable (eqns. 2-1 and 2-2).
2CuX(s) + HS- → Cu2S + 2X- + H+ (2-1)
2CuFeS2 + HS- → S22- + H+ +Cu2S(s) + 2FeS(s) (2-2)
2HS- + 3/2O2 → S2O32- + H2 (2-3)
HS- + 3/2O2 → SO32- + H+ (2-4)
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HS- + 2O2 → SO42- + H+ (2-5)
Depression starts at redox potential of -250 mV and when the flotation is stable, the redox
potential is -450 to -500 mV for NaHS. Addition of NaHS allows for obtaining sulphides with
a fresh surface because it desorbs the collector, making its depression easy (eqn. 2-1). With
these conditions, molybdenite is ready for its recovery by flotation. Sometimes, it is useful to
add a little fuel oil, mainly when molybdenite can be associated with certain iron minerals.
It is very important to observe the slurry and try to understand what action is causing the
depression. Oxygen is not desirable in the slurry although with small or big dosages of NaHS,
there is certain amount of oxygen in the slurry, which reduces the sodium hydrogen sulphide
efficiency. The presence of dissolved oxygen in the slurry affect the flotation process because
it leads to an increase in redox potential and formation of sulphoxy species such as
thiosulphate ions (S2O32-), sulphite ions (SO3
2-), and sulphate ions (SO42-) (eqns. 2-3 to 2-5).
As a result, HS- ions are consumed by oxygen instead of copper sulphide, which give rise to
requirement of additional quantity of NaHS.
In summary, air addition promotes oxidation of sulphide ions to less active compounds. For
this reason nitrogen is commonly employed as flotation gas during rougher and cleaning
flotation stages. An option is to re-cycle used-air, if air is used as a flotation gas because as
the used air is depleted of oxygen, the nitrogen concentration increases. CO2 can be employed
as pH regulator, which forms a weak acid in the slurry; it can also be used to modify froth
texture by the formation of polysulphides in solution (S22-) through dissolution of So from
chalcopyrite surface [7].
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3 Molybdenite production in Boliden
“Boliden is one of the leading mining and smelting companies in Europe with operations in
Sweden, Finland, Norway and Ireland. Boliden’s main products are copper, zinc, lead, gold
and silver. Exploration and recycling of metals are also important within the company. The
number of employees is approximately 4500 and the turnover amounts to approximately � 3.8
billion annually. Its shares are listed on Stockholmsbörsen’s Large Cap list and on the
Toronto Stock Exchange in Canada” [28].
A molybdenite concentrate has been produced in Boliden Mineral AB in a pilot test as a by-
product from copper concentrate. The copper ore at Boliden Aitik mine contains
approximately 0.004 to 0.008% Mo, initial studies of molybdenum flotation from the copper
concentrate was previously carried out in lab scale during 1981 to 1982, which lead to a grade
of about 30% Mo. However the investigation did not proceed in a pilot scale due to falling
price of molybdenum at that time.
Recent dramatic increase in price of molybdenum during 2004 and 2005 with a peak value of
US$ 103,000 per tonne in June 2005 [15] and to approximately US$ 65,000 per tonne as of 4
May 2007 has motivated renewed interests in the molybdenum production. A long-term price
is hard to estimate, but it is reasonable to forecast a price higher than before. A price of US$
13,000 per tonne has been used in some calculations. The forecast for the nearest coming
years is higher with decreasing level down to US$ 22,000 per tonne in year around 2010 [12].
The price recently is approximately US$ 70,000 per tonne according to Info Mine [13].
Pilot tests with molybdenum recovery in 2006 showed that the process could give good grade
and recovery with reasonable consumption of reagents. A good grade of averagely 53.8% Mo
and 1.65% Cu have been achieved at the pilot plant in Aitik with the use of NaHS as a
depressant agent and air as flotation gas [12].
3.1 Background
In every manufacturing or production set-up, quality of products and services are the most
important factor. When discussing the quality of molybdenum concentrates, the copper
concentration is an undesired impurity and is the most important aspect. Lead, phosphorus
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and arsenic are also undesirable. There may also be issues concerning potassium, sodium, and
magnesium [12], but concerning the molybdenite concentrate production in Aitik, the most
important impurity is the copper concentration in the concentrate. A flotation circuit to extract
molybdenite from the copper by-products will be built in Aitik. The value of the molybdenite
concentrate will depend to a large degree on the extent of impurities in the molybdenite
concentrate. A number of leaching methods have been applied for the purification of
molybdenite concentrates.
3.2 Aim
This work aim to investigate different alternative leaching methods that can be used for
purification of the molybdenite concentrate from Aitik depending on the type of contaminant
to be leached and to study how well different methods are working on different contaminants
and how molybdenite is affected. The specific aim is to remove the copper concentration in
the molybdenite concentrate to a commercially acceptable minimum using different leaching
methods and selecting the optimum method that effectively reduce the copper concentration
with minimum dissolution of molybdenum. The major copper component of the Aitik
molybdenite concentrate is present as chalcopyrite.
The goal is to find the best leaching method for Aitik molybdenite concentrate that would be
employed in the construction of the plant that will be put into operation approximately in
2010.
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4 Literature review
4.1 Purification of molybdenite concentrate
Generally, molybdenite concentrates are complex as they contain substantial quantities of
base metal sulphides and numerous quantities of gangue in form of oxides and silicates, but
the specific sulphide to be removed from the molybdenite concentrate considered under this
study is chalcopyrite. A number of attempts have been made to purify molybdenite
concentrates based on the premise that it is possible to selectively dissolve the associated
sulphides, oxides and gangue by leaching. These include leaching with:
1. Microorganisms with ferric sulphate [2]
2. Sulphuric acid [9]
3. Sulphuric acid with sodium dichromate solution [3]
4. Oxygen and halide e.g. CuCl2 [4]
5. Hydrochloric acid [18].
6. Hydrochloric acid with hydrofluoric acid [18].
4.1.1 Microorganisms
Bioleaching tests carried out in an orbital shaker at 150 rpm, 20 g/l pulp density with 5 %
(v/v) Sulfolobus BC extreme thermopile previously adapted to molybdenite at 68oC claimed
100% dissolution of copper after 15 days. Chemical leaching of the same sample with 10 g/l
Fe2(SO4)3 under the same condition indicate 100% dissolution of copper after 15 days but at a
slower rate compared to bioleaching [2]. However assays of the leach solution and solid
residue are not presented; this could have given better insight to whether Mo is affected in
anyway and also the complexity of the leach solution as regards to how it could be treated
and/or disposed afterwards with regard to the environment.
4.1.3 Sulphuric acid
It has also been claimed that more than 98% Cu and most other impurities could be extracted
with less than 0.5% Mo dissolution. In a pulp containing concentrated sulphuric acid (more
than 93%) and molybdenite concentrate in ratio 1:1 in a sulphation reactor at 160-190oC for 2-
10 hours. The hot, acidic pulp discharging from the reactor is water leached, thickened,
filtered and washed to remove the soluble sulphates formed. Based on this claim a 250 kg/day
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pilot plant was operated continuously for 12 000 hours using several different molybdenite
concentrates with Cu concentrations ranging from 1-9.2 % Cu and less than 0.1% Cu was
consistently achieved in the purified molybdenite concentrate. However operational variables,
reactor designs and material selections are critical to the success of the continuous process.
Two configurations of reactors were developed using glass and teflon linings. Heat transfer
oil jackets provided the necessary heat to maintain the reacting pulp at desired temperature. A
80 t/day commercial plant based on this process was projected [9].
4.1.4 Sulphuric acid with sodium dichromate solution, Na2Cr2O7
Leaching with 0.3 M H2SO4 and 0.12 M Na2Cr2O7 for 90 min at 100oC and 560 rpm claimed
to dissolve 95% copper with little dissolution of molybdenum [3]. The presence of chromium
in the leach solution is critical to environmental requirements.
4.1.5 Leaching with oxygen and halide e.g. CuCl2
A US patent have claimed effective reduction of copper concentration from a molybdenum
concentrate by subjecting the molybdenum concentrate to pressure oxidation in the presence
of oxygen and a feed solution containing copper (e.g. CuSO4) and halide (e.g. CuCl2) to
produce a solution containing copper and a solid residue containing molybdenum. The
solution may be combined with the feed solution to a second pressure oxidation in which a
copper concentrate is treated for the recovery of copper. [4]
4.1.6 Conventional leaching with chloride salts
Another US patent have claimed removal of copper, iron, and lead impurities from
molybdenum flotation concentrates by mixing the feed concentrates with a non-volatile
chloride salt, heating the mixture to a temperature of from about 200o to 350oC for a time
sufficient to activate the lead impurities in the concentrates so that they can be leached during
the subsequent leach step, and leaching copper, iron, and lead impurities from the heat-treated
concentrates with a mildly oxidizing leach solution containing chloride ions and having a pH
of 4 or less. Good homogenization of the chloride salt and the concentrates can be achieved
by thoroughly mixing the feed concentrate with an aqueous solution of the chloride salt. The
patent also stated that, it is advantageous to use an aqueous ferric chloride solution to leach
the heat-treated concentrates as the lead values leached from the concentrates can be
crystallized from the pregnant leach solution and the resulting spent solution can be recycled
without further treatment, by mixing it with the feed concentrates, or it can be treated to
19
oxidize its ferrous constituent to ferric and then recycled for repeated use as regenerated leach
solution [5]. Finally a United States Patent claimed impurity removal from molybdenite
concentrates by leaching the concentrates at a temperature of about 70oC with an aqueous
solution containing an alkali metal or alkaline earth metal chloride and an oxidizing chloride,
for example cupric and ferric chlorides [29].
4.1.7 Hydrochloric acid and hydrofluoric acid
The aforementioned attempts have recorded greater success in removal of sulphide minerals
but only the attempts with hydrochloric acid and hydrofluoric acid administered singly,
sequentially or in mixed mode have proved to remove both the oxides and silicates gangue as
well as the metallic sulphides effectively [18]. However operational variables, reactor designs
and material selections seem to be critical for the success of this process.
4.1.8 SO2/O2 as a strong oxidant
It has also been demonstrated that SO2/O2 can function as a strong oxidant in acidic medium,
which can oxidise Fe2+ to Fe3+. The Fe3+ can be used to leach copper sulphides minerals and
uranium oxides, the Fe3+ and H2SO4 can also be regenerated back as oxidants to leach more
copper sulphides and uranium oxides. [10]. Also SO2/O2 is useful for the oxidative
precipitation of Mn(II) as MnO2, from Co(II) and Ni(II) leach liquors at around pH 3-4 [6].
4.2 Molybdenum-water chemistry
“Two dissolved substances
Relative stability of the dissolved substances
Oxidation number, Z = +6
1. HMnO4- � MoO4
2- + H+
log HMoOMoO
−
−
4
2
4 = -6.00 + pH
Z = +3 to +6
2. Mo3+ + 4H2O � HMoO4- + 7H+ + 3e-
E� = 0.39 – 0.1379 pH + 0.0197 logMo
HMoO+
−
34
3. Mo3+ + 4H2O � MoO42- + 8H+ + 3e-
20
E� = 0.508 – 0.1576 pH + 0.0197 logMo
MoO+
−
3
2
4
Limits of the domains of relative predominance of the dissolved substances
1´. HMoO4-/MoO4
2- pH = 6.00
2´. Mo3+/HMoO4- E� = 0.300-0.1379pH
3´. Mo3+/MoO42- E� = 0.508-0.1576pH
Two solid substances
Limits of domains of relative stability of the solid substances
Z = 0 to +4
4. Mo + 2H2O � MoO2 + 4H+ + 4e- E� = -0.072-0.0591pH
Z = +4 to +6
5. MoO2 + H2O � MoO3 + 2H+ + 2e- a. E� = -1.091-0.0591pH
b. E� = 0.320-0.0591pH
One dissolved substance and one solid substance
Solubility of solid substances
Z = +6
6. MoO3 + H2O � HMoO4- + H+ a. log[HMoO4
-] = -51.42 + pH
b. log[HMoO4-] = -3.70 + pH
Z = 0 to +3
7. Mo � Mo3+ + 3e- E� = -0.200 + 0.0197log[Mo3+]
Z = 0 to +6
8. Mo + 4H2O � MoO42- + 8H+ + 6e-
E� = 0.154 – 0.0788pH + 0.0098log[MoO42-]
Z= +3 to +4
21
9. Mo3+ +2H2O � MoO2 + 4H+ + e-
E� = 0.311 – 0.2364pH – 0.0591log[Mo3+]
Z= +3 to +6
10. Mo3+ + 3H2O � MoO3 + 6H+ + 3e-
a. E� = -0.623-0.1182pH - 0.0197log[Mo3+]
b. E� = 0.317-0.1182pH - 0.0197log[Mo3+]
Z = +4 to +6
11. MoO2 + 2H2O � HMoO4- + 3H+ + 2e-
E� = 0.429 – 0.0886pH + 0.0295log[HMoO4-]
12. MoO2 + 2H2O � MoO42- + 2e-
E� = 0.606 – 0.1182pH + 0.0295log[MoO42-]
4.3 Equilibrium diagram and its interpretation
The relations above have been used by Pourbaix M et.al to draw the equilibrium diagram in
figure 4-1a which represents the conditions of thermodynamic equilibrium of the system
molybdenum-water at 25oC, in the absence of complexing substances and substances forming
insoluble salts.
Figure 4-2 represents the theoretical conditions of corrosion, immunity and passivation of
molybdenum in the presence of solutions free from substances with which this metal can form
soluble complexes or insoluble salts.
Figures 4-1a and 4-2 depict molybdenum as a base metal, as its domain of stability lies
completely below that of water. It is not found in nature in the native state.
In alkaline solutions, it has tendency to decompose water with the evolution of hydrogen gas,
dissolving in the hexavalent state as molybdate ions MoO42-. In the presence of non-
complexing acid solutions, it tends to dissolve in the trivalent state with the formation of Mo3+
22
ions and also evolution of hydrogen gas. In neutral or slightly acidic or alkaline solutions it
tends to cover itself with tetravalent dioxide MoO2.
In practice molybdenum is noticed to be attacked only slightly by dilute non-complexing acid,
only dilute nitric acid acting as an oxidizing agent attacks it appreciably, while concentrated
nitric acid covers it with a layer of MoO3 which protects the metal from further attack.
In hydrochloric acid the metal is slightly attacked and passivated, it is probable that a film of
insoluble chloride is formed which passivates the metal. Molybdenum in powdered form can
be oxidized by tap water and distilled water (free from CO2). Water turns blue when in
contact with molybdenum.
Figure 4-1a: Potential pH equilibrium diagram for the system molybdenum-water, at 25oC by
M.Pourbaix et al.
23
Figure 4-1b: Potential pH equilibrium diagram for the system molybdenum-water, at 25oC
drawn from thermodynamic software, Fact-Sage.
Figure 4-2: Theoretical conditions of corrosion, immunity and passivation of molybdenum, at
25oC
24
Figure 4-3: Influence of pH on the solubility of MoO3, at 25oC
Molybdenum is distinguished by a peculiar behavior with regard to passivation: depending on
the pretreatment of the metal it is either active or passive; chemically passivating reagents are
not only oxidizing agents such as nitric acid concentrated chromic acid and ferric chloride but
also dilute hydrochloric acid and sulphuric acids; the highest degree of passivation is however
obtained by anodic polarization under certain conditions of current density. Activation of the
metal is produced chemically, by KOH or NH3 and reducing agents, or electrochemically by
cathodic polarization in a solution of KOH. This behavior peculiar to molybdenum seems to
be in agreement with the equilibrium diagram, according to which, in alkaline media the
metal is constantly active and dissolves as molybdate ions MoO42- by oxidation, while in acid
media it is liable to dissolve at first as Mo3+ at relatively low electrode potentials and cover
itself with a passivating layer of MoO2 or MoO3 [lines (9) and (10) of Fig.4-1a] at higher
potentials. The passivation of molybdenum in acid media may also be due to the formation of
a film of insoluble salt.
According to fig.4-1a molybdenum can be obtained electrolytically from acid solutions of
molybdenum salts or alkaline solutions of molybdates. In practice, the electrolytic reduction
of the molybdenum chlorides gives a positive result only in non-aqueous solutions, for
25
instance solutions of MoCl2 and anhydrous HCl in absolute alcohol. The usual processes for
the electrolytic separation of molybdenum are however based on the electrolysis of the molten
salts (a mixture of calcium molybdenate and molybdenum carbide in bauxite, or a molten
mixture of sodium and molybdenum chlorides).
4.3.1 Stability of the compounds of trivalent molydenum
The molybdenous ion Mo3+ is stable only in strongly acidic reducing media, on being
oxidized, the Mo3+ ions are slowly converted into a red compound whose composition is
unknown, but according to fig.4-1a the principal oxidation product should be the purple-
brown oxide MoO2.
The dehydration of Mo(OH)3 by warming always give rise to oxidation to MoO3, even in a
current of hydrogen, it is most probable that the water contained in the hydroxide acts as an
oxidizing agent at the temperature necessary for the dehydration.
4.3.2 Stability of the compounds of tetravalent molybdenum
Tetravalent molybdenum is known only in solution in the form of complex ions, which are
not considered in figure 4-1a.
By reducing ammonium molybdate solutions with hydrogen at ordinary temperature and
pressure in the presence of colloidal palladium, a brown-black hydroxide Mo(OH)4 (or
MoO2.2H2O) is formed; by drying this hydroxide carefully in the cold the monohydrate
MoO2.H2O [or MoO(OH)2] is obtained; if the hydroxide is dried by warming the brown-
purple anhydrous oxide MoO2 is obtained, which, of these three forms of oxide, is the only
one considered in figure 4-1a. This oxide MoO2 can be obtained by other means, for example
by heating molybdenum in air, or in water vapour, by reducing solutions of MoO3 with metals
such as Zn, Cd, Mg, and by the electrolytic reduction of molten MoO3.
From fig.4-1a strong non-oxidising acid should cause MoO2 to split up into molybdenous ions
Mo3+ and the acid molybdate ions HMoO4- [according to the family of lines (9) and (11)] with
the possible formation of molybdenum blue and MoO3. Non-oxidising alkali should cause
MoO2 to split up into molybdate MoO42- and metallic molybdenum [according to the family
of lines (12) and line (4)], which would react with water to form molybdate and hydrogen;
oxidizing alkalis should convert it into molybdate.
In actual fact, MoO2 is oxidized by nitric acid to MoO3, but it is, in general, insoluble in non-
oxidizing acids. In the present of aerated water Mo(OH)4 is easily oxidized by forming a
26
solution of “molybdenum blue”; it is soluble in concentrated acids, forming solutions which
are red to purplish brown and are feebly reducing; it was not clear if the reducing property
was due to the formation of trivalent Mo3+, or to the existence of complexes of tetravalent
molybdenum.
4.3.3 Stability of the compounds of pentavalent molybdenum, molybdenum blue
Pentavalent molybdenum compounds are also known only in solution in form of complexes
just like tetravalent molybdenum. The hydroxide MoO(OH) or pentavalent Mo2O5.3H2O,
molybdenyl hydrate, is obtained as a dark-brown precipitate by treating a solution of
ammonium molybdenum oxychloride (NH4)2 (MoOCl5) with NH3. Due to lack of availability
of thermodynamic data, no derivative of pentavalent molybdenum could be considered
quantitatively in the figure. It was pointed out that Mo2O5.3H2O is easily oxidized by air, and
seems liable to decompose into compounds of tri- and hexavalent molybdenum: when it is
treated with KOH in an atmosphere of hydrogen, it is partially converted into Mo(OH)3 and
the solution contain molybdenum in the hexavalent state.
The dehydration product of MoO(OH)3 is the oxide Mo2O5, which can also be obtained by
reduction of MoO3; it is a purplish-black powder which is very sparingly soluble in acids.
Together with the compounds of pentavalent molybdenum, molybdenum blue can be
considered a compound containing oxygen and molybdenum whose percentages of oxygen
and molybdenum fits approximately the formula Mo3O8, an oxide in which molybdenum
would have valency between +5 and +6; however, it is usually considered to be a molybdenyl
molybdate Mo2O5.xMoO2. It is a blue compound which is very soluble in water and easily
forms colloidal solutions; it is obtained either by reduction of molybdate solutions or by the
oxidation of lower oxide such as MoO2.
4.3.4 Stability of the compounds of hexavalent molybdenum
The compounds of the hexavalent molybdenum are the most important ones; they are
illustrated by molybdenum trioxide or molybdic anhydride MoO3, by its hydrates (molybdic
acids) and by its dissolved forms, principally the molybdate ion MoO42- obtained by the
action of alkalis on MoO3. By varying the relative quantities of MoO3 and alkalis a whole
series of salts can be obtained: the di-, trimolybdates, etc; to these various salts there should
be various condensed ions. Studies on molybdate solutions by diffusion measurements and
27
conductometric titrations have shown that the nature of the ions varies with the pH in such a
way that a domain of stability can be attributed to each of them; for example:
6 < pH < 14 MoO42-
4.5 < pH < 6 Mo3O114-
1.5 < pH 4.5 Mo6O216-
0.9 < pH 1.5 Mo12O4110-
pH = approx. 0.9 Mo24O7812-
This complex range of substances has been symbolized by the ion HMoO4- in establishing the
equilibrium diagram.
The oxide MoO3, prepared by roasting ammonium molybdate, is a white powder; its
solubility in water is about 2 g/l i.e 10-1.85 mole/l, which is in agreement with figure 4-3. It
does not combine directly with water to give hydrates; these are obtained only from
molybdates; there exists two well-defined hydrates MoO3.H2O and MoO3.2H2O.
The fairly close solubility values for MoO3 and its two hydrates show that these three
compounds have stabilities which are practically equal; the domain of stability attributed to
MoO3 in the equilibrium diagram in fig.4-1 can therefore also be attributed to one of its
hydrates. The influence of pH on the solubility of MoO3 is represented in fig.4-3; the
characteristics of the solution obtained by dissolving this oxide in pure water until a saturated
solution is obtained are given in this figure by the coordinates of the point of intersection of
this solubility line with the line showing the values of [(H+)-10-7]; these characteristics are:
pH = 1.85, [HMoO4-] = 10-1.85, i.e. 2 g MoO3/l.
Molybdic solutions treated with hydrogen peroxide give rise to the formation of the so-called
permolybdates, which are yellow-orange in acid media and intense red in alkaline media. On
account of lack of precise data relating to these compounds, in which molybdenum would
have a valency of +7, it is not taken into account in establishing the equilibrium diagram in
which they appear purely as a guide [22].
28
4.4 Cu-CN chemistry Copper can dissolve in cyanide solution to form different copper-cyanide aqueous complexes, known as cyanocuprate ions, this includes, Cu(CN)2
-, Cu(CN)32- and Cu(CN)4
3- The complexes have been confirmed to undergo the following successive equilibrium steps in reaction with free and un-dissociated hydrocyanic acid: CuCN� Cu+ + CN- Ksp (4.3-1) CuCN + CN- � Cu(CN)2
- K2 (4.3-2) Cu(CN) + 2CN- � Cu(CN)2
- �2 (4.3-3) Cu(CN)2
- +CN- � Cu(CN)32- K3 (4.3-4)
Cu(CN)3
2- + CN- � Cu(CN)43- K4 (4.3-5)
HCN � H+ + CN- Ka (4.3-6)
“The equilibrium constants for copper cyanide complexes differ among different authors, due
to different methods of measurement and the processing of data” [32]. However the calculated
constants listed in table 4-1 have a common agreement.
Table 4-1: Equilibrium constants for copper cyanide system
Temperature
oC
log Ka log Ksp log �2 log K3 log K4
25 -9.21 -20 24 5.3 1.5
40 -8.84 -19.1 22.98 4.91 1.11
50 -8.60 -18.33 22.35 4.67 0.86
60 -8.41 -17.6 21.75 4.45 0.64
Cupric ions react with CN- to form cupric complexes, which are unstable and decompose
rapidly. The distribution and equilibrium potential of copper cyanide species has been found
to depend on the mole ratio of cyanide to copper, total cyanide concentration, pH and
temperature. With increasing CN:Cu mole ratio, the distribution of copper cyanide species
shifts more completely to the highly coordinated complex (Cu(CN)43-). The equilibrium
potential for Cu(I)/Cu decreases with increasing CN:Cu mole ratio. Increasing pH is directly
proportional to increasing the free cyanide concentration. Increasing temperature gave rise to
reduced stability constants. Therefore the distribution of copper cyanide shifts to the lowly
coordinated complexes. The potential measurements have been carried out to confirm the
validity of the calculated results in table 4-1. The potential-pH diagrams shows that the copper
29
complexes, CuCN, Cu(CN)2-, Cu(CN)3
2- and Cu(CN)43- can predominate in different pH
regions. The free energy data for all species are listed in table 4-2 all the potential-pH
diagrams were calculated based on the data in table 4-2
Table 4-2: Gibbs energy data for copper and cyanide species (J mol-1) [32]
Figure 4-4: CN-H2O potential-pH diagram at all solute species activities of 1 and P(CN)2 = 1
atm and 25oC assuming HCNO and CNO- are stable [32].
Figures 4-4 and 4-5 show the potential-pH diagram for the CN-H2O system assuming that
CN-, CNO-, HCN, HCNO and (CN)2 are stable, although not all of them are stable in
practise. CN- and HCN are oxidised at high potential range and are not stable, they are
however metastable at low potential range shown in figures 4-4 and 4-5 as the potentials for
their oxidation are are much higher (1.0 - 1.2V). Hence they are considered stable in Cu-CN-
H2O potential-pH diagram.
30
Figure 4-5: CN-H2O potential-pH diagram at all solute species activities of 1 and P(CN)2 = 1
atm and 25oC assuming (CN)2 is stable [32].
31
Figure 4-6: Potential-pH diagrams for Cu-CN-H2O system at 25oC at all solute species
activities of 1, assuming Cu(OH)2 as a stable species, HCNO, CNO- and (CN)2 are not
considered [32].
Figure 4-6 depicts the potential-pH diagram at the activities of all species equal 1, CuCN,
Cu(CN)32-, and Cu(CN)4
2- are shown to be stable in the three pH regions; -2 to 4.5, 4.5 to 8
and 8 to16 respectively.
Figure.4-7: Potential-pH diagrams for Cu-CN-H2O system at 25oC and all solute species
activities of 10-2, assuming Cu(OH)2 as a stable species, HCNO, CNO- and (CN)2 are not
considered [32].
Figure 4-7 depicts the potential-pH diagram at the activities of all species equal 10-2, CuCN,
Cu(CN)2-, and Cu(CN)3
2- are shown to be stable in the three pH regions; -2 to 4.5, 4.5 to 6 and
6 to 16 respectively.
32
Figure 4-8: Potential-pH diagrams for Cu-CN-H2O system at 25oC and all solute species
activities of 10-6, assuming Cu(OH)2 as a stable species, HCNO, CNO- and (CN)2 are not
considered [32].
Figure 4-8 depicts the potential-pH diagram at the activities of all species equal 10-6, only
CuCN, and Cu(CN)2-, are shown to be stable in the three pH regions; -2 to 4.5 and 4.5 to 16
respectively.
33
Figure 4-9: Potential pH diagram for Cu-CN-H2O system at 25oC and solute copper species
activities of 0.01 and cyanide species activities of 0.1 considering Cu(OH)2 as a stable
species. HCNO, CNO- and (CN)2 are not considered [32].
Figure 4-9 shows the potential-pH diagram at the activities of copper species equal 0.01 and
activities of cyanide species equal to 0.1 it can be seen that all copper-cyanide species are
stable in their correponding pH regions [32].
“Generally copper cyanide species are stable in certain potential and pH regions, with
increasing potential copper cyanide can be oxidised to Cu2+, Cu(OH)2 or CuO and CuO2-.
Cyanide can also be oxidised to cyanate based on thermodynamics data” [32].
34
4.5 Leaching theory 4.5.1 Leaching under pressure
The leaching method suitable and commonly adopted for the purification of molybdenite
concentrate is pressure leaching.
“Pressure leaching can be broadly divided into two types namely:
(i) In absent of oxygen: In this case the ore is heated with the leaching agent above the boiling
point of the solution to achieve a high reaction rate. Therefore the process must be carried out
in a closed vessel that can withstand the vapor pressure of solution at that temperature. An
example is leaching of bauxite with caustic soda solution
(ii) In present of oxygen: Here the pressure in the autoclave (reactor) is due to the solution
pressure as well as oxygen pressure (or air pressure if air is used instead of oxygen). This
method is used mainly for leaching sulphide ores or uranium oxide ores [1]
The method adopted in this study is a form of pressure leaching in the present of oxygen or air
with some modification with addition of sulphur (iv) oxide, SO2 in some tests.
4.5.2 Leaching theory
Factors influencing the rate of leaching process can be summarized in the following points:
1. Rate of leaching increases with decreasing particle size of the ore since the smaller the
particles the larger is the surface area per unit weight.
2. If a leaching process is diffusion controlled then it will be greatly influenced by the
speed of agitation. On the other hand if it is chemically controlled then it will not be
influenced by agitation, provided that enough agitation is done to prevent the solids
from settling.
3. Leaching rate increases with increasing temperature. However, this increase is much
less remarkable for a diffusion-controlled process than for a chemically controlled
process.
4. Rate of leaching increases with increasing concentration of the leaching agent.
5. Rate of leaching increases with decreasing pulp density, i.e. when a large volume of
leaching agent is added to a small volume of solids.
6. If an insoluble reaction product is formed during leaching, then the rate will depend on
the nature of this product. If it forms a nonporous layer, then the rate of leaching will
greatly decrease. If however, the solid product is porous, it will not affect the
rate’’[14].
The particle size distribution of the material used in this test is shown in table 5-1.
35
5 Experimental
5.1 Material
The feed material for purification used in the tests is a molybdenite flotation concentrate from
Aitik whose chemical composition is given in table 5-2, particle size analysis in table 5-1 and
can also be viewed in figure 5-1. The mineral phases identified by XRD in the concentrate are
mainly molybdenite (MoS2) and chalcopyrite (CuFeS2) as shown in figure 5-2.
Table 5-1: Particle size distribution of feed ore
wt % 4.4 10.3 20.2 41.6 9.8 13.7 Apparent particle size, �m +90 -90, +63 -63, +45 -45, +20 -20, +10 -10
The particle size distribution of the ore indicated that it was fine enough with 65% below 45
�m (350 mesh) and 4% over 90 �m (170 mesh). This ensures enough surface area per unit
weight of the ore. The leaching process in this case is chemically controlled and stirrers kept
the material evenly suspended in the pulp as shown in the experimental set-up, figure 5-5.
Table 5-2: Head assays: chemical analysis of the molybdenite concentrate used in this work Mo %
Cu %
Pb %
Fe %
S %
Au g/t
Ag g/t
Bi %
Hg %
Ca %
Cl %
46 2.13 0.019 2.83 37.4 1.6 98 0.011 0.0016 0.11 <0.1
Figure 5-1: Feed molybdenite concentrate to be purified
36
Figure5-2: XRD analysis of the feed molybdenite concentrate
The XRD analysis, figure5-2 shows the main mineral phases present in the feed material. The
high peaks corresponds to molybdenite as the dominant mineral and the low peaks
corresponds to traces of chalcopyrite being the main copper mineral present as impurity in the
concentrate. Details of XRD analysis equipment and procedure are presented in appendix II.
5.2 Reactors
Only one type of experimental set up was used for all the experiments. The experiments were
carried out in round bottom glass reactors of varying capacities (1000-3000 ml) with five
openings and capacity to withstand heating temperature above 100oC. The central opening of
the reactor was used to connect and adjust the shaft of the stirrer, while the other openings
were fitted with a programmable thermometer to automatically regulate the heat supply at the
programmed temperature, a water condenser to prevent and/or reduce evaporation during
heating and leaching, one or two hose(s) as the case may be to supply the appropriate gas(s)
into the pulp. Heat is supplied by an electro-mantle heater, which is automatically regulated
by the programmable thermometer. The stirrer speed was 500 rpm. The experimental set-up is
shown in figure 5-3 below.
37
Figure 5-3: Experimental set-up
5.3 Methodology Leaching was carried out at both high and low temperatures; the cyanide leaching was carried
out at 25oC, the ferric chloride leaching was carried out at temperature ranging between 70oC
and 100oC, while the optimum temperature for the ferric sulphate leaching tests was 65oC.
The leaching reagent for cyanide leaching was added approximately based on stoichiometric
requirements for the “cyanide only” test and was in excess of stoichiometric requirement for
the “cyanide + calcium chloride” test, the pH was constantly maintained above 11 with
addition of sodium hydroxide pellets. The leaching reagent for ferric chloride test was
stoichiometric concentration of ferric chloride with appropriate estimation of copper chloride
and calcium chloride, the pH was kept at about zero with addition of hydrochloric acid. The
leaching reagent for ferric sulphate leaching was stream of Fe3+ produced by SO2/O2
oxidation; the pH was kept below 0.5 initially with addition of sulphuric acid.
5.4 Test performance 5.4.1 Cyanide leaching
Sodium cyanide (NaCN) was used as the leaching agent. Sodium hydroxide (NaOH) was used
as a pH regulator and to eliminate any contamination of the concentrate by gypsum
precipitation. Calcium chloride (CaCl2) was added to one of two set-ups as inhibitor of
molybdenum dissolution, oxygen and air was used separately as oxidizing agents. Copper
38
analysis in leach solution was carried out with the aid of atomic absorption
spectrophotometer, AAS. Detail of AAS and procedure of copper analysis is presented in
appendix II.
Seven experiments were planned to be carried out according to table 5-3 with temperature and
NaOH as variable factors and with leaching time, %solid, NaCN and CaCl2 as constant
factors. It was however reduced based on former experience and logistics reasons.
Table 5-3: Design of cyanide experiment with modde software
Experiment
No
Leaching
time
Hours
Solid
concentration
%Solids
Temperature oC
NaOH
kg/t
NaCN
kg/t
CaCl2
kg/t
1 72 20 20 90 90 150
2 72 20 80 90 90 150
3 72 20 20 300 90 150
4 72 20 80 300 90 150
5 72 20 50 195 90 150
6 72 20 50 195 90 150
7 72 20 50 195 90 150
The design of cyanide leaching conditions was based on Boliden in-house experience with
respect to leachability of chalcopyrite from pyrite by cyanide solution based on the in-house
experience of the 4:1 rule of thumb i.e. four moles of cyanide to one mole of gold; this
normally is for gold leaching but it also works to some extent for copper leaching because the
leaching solution from gold leaching is found to always contain some copper-cyanide
complexes in form of either Cu(CN)2-, Cu(CN)3
2-, or Cu(CN)43-.
In cyanide leaching of chalcopyrite, possible stoichiometric relations are expressed in
equations 5-1, 5-2 and 5-3 (also in section 4.4) and the overall relation in equation 5-4. The
equations depict cyanide-consuming reactions. However it has been confirmed that formation
of CNO- is very negligible. Hence only (5-2) and (5-3) are valid i.e. one mole of copper
requires three moles of cyanide for dissolution and formation of one mole of Cu(CN)32-
complex, and also three moles of cyanide is required for the formation of three moles of
thiocyanide complex,(SCN-)[30].
39
2CN- + O2 � 2CNO- (5-1)
3CN- +3/8 S8 � 3SCN- (5-2)
Cu + 3CN- � Cu(CN)32- (5-3)
CuFeS2 + 6CN- + H2O + 1/8S8 + O2 + H+ � Cu(CN)32- + 3SCN- + Fe(OH)3(s) (5-4)
The mole and mass ratio of Cu:SCN and CN:Cu was found to vary between 2 and 3 from
analytical results of cyanide leaching tests on Aitik final cleaner tailing, see appendix VII.
Hence the total cyanide requirement for this test was calculated based on the 3 moles
requirement for dissolution of Cu and S as shown in appendix II.
The leaching tests were carried out batch-wise at two different leaching conditions in
oxidative environment. The first experiment was carried out using only NaCN as the leaching
agent, and the other was carried out using NaCN with addition of CaCl2 in order to confirm an
earlier observation that it inhibits the dissolution of Mo [12]. The two experiments were
carried out at ambient temperature 25oC and oxygen was used as oxidative gas. Although the
effect of air as an oxidative gas was previously observed in two preliminary experiments; the
use of air gave rise to flotation in the pulp, this may be as a result of the small volume of the
laboratory reactor, however this effect could be negligible in a pilot or a large scale reactor,
hence air could be a cheaper alternative oxidative gas to oxygen. The duration of the
experiments was 48 for the test with only NaCN and 72 hours for NaCN+CaCl2 test.
The leaching conditions and total amounts of additives are summarized in table 5-4.
Table 5-4: Leaching conditions and additives for cyanide leaching.
Test Leaching
time
Hours
Pulp solid
concentration
%-Solids
Temp oC
NaOH
kg/t
NaCN
kg/t
CaCl2
kg/t Oxidizing
gas
NaCN +
CaCl2
72 20 25 100 150 600 O2
NaCN 48 20 25 25 100 - O2
The redox potential measured throughout the experiment varied between –0.1V to –0.2V. The
detailed procedure, protocol and material balance can be found in appendix I.
40
5.4.2 Ferric chloride (FeCl3) leaching Ferric chloride, (FeCl3) and cupric chloride (CuCl2) were used as oxidizing reagents and
calcium chloride (CaCl2) was added in order to elevate the boiling point of the solution as
well as to enhance the extraction rate of copper from the concentrate by complex formation
[29]. Hydrochloric acid was added as a pH regulator, air was supplied for aeration. Copper
analysis in leach solution was carried out with the aid of AAS.
Five experiments were planned to be carried out with only temperature as variable factor
while leaching time, %solid, FeCl3, CuCl2, HCl and CaCl2 were constant factors. It was
however limited to one experiment because the temperature was observed to vary between 70
and 110oC during the experiment, due to the exothermic reactions.
The design of the ferric chloride leaching conditions was based on examples found in the US
patent, No 3674424. The patent claimed removal of sulphide impurities from molybdenite
concentrates with aqueous solution containing an alkaline metal or an alkaline earth metal
chloride and an oxidizing chloride,for example ferric chloride and cupric chloride. A number
of examples in the patent showed that the process was more effective with solution
concentration containing 20% CaCl2 as the alkaline earth metal chloride, 10% FeCl3 and 1%
CuCl2 as oxidizing chlorides [29]. The concentration of the oxidizing metals was estimated in
such a way that it is sufficient for the desired extraction of copper from the molybdenite
concentrate with regard to the stoichiometric concentration as can be observed in equation (5-
7) and (5-8) respectively. The stoichiometric concentration required at least 4 moles of FeCl3
and 3 moles of CuCl2 to leach one of Cu from the chalcopyrite in the concentrate.
CuFeS2 + 4FeCl3 � CuCl2 + 5FeCl2 + 2S (5-7)
CuFeS2 + 3CuCl2 � 4CuCl + FeCl2 + 2S (5-8)
Although it was stated that cupric and ferric chloride could be deployed singly, but deploying
them in a mixed mode was observed to have a synergistic effect [29]. Hence the mixed mode
approach was employed in this test. The leaching test was carried out batch-wise oxidative
environment. The experiment was done using, 20% CaCl2, 10% FeCl3 and 1% CuCl2 as
leaching reagent at a temperature between 70 and 110oC and air was also used as oxidative
gas for 4 hours. The leaching conditions and total amounts of additives are summarized in
table 5-5.
41
Table 5-5: Leaching conditions and additives for ferric chloride leaching.
Test Leaching
time
Hours
Pulp solid
concentration
%-Solids
Temp oC
FeCl3
kg/t
CuCl2
kg/t
CaCl2
kg/t Oxidizing
gas
Ferric
chloride
4 35 70-110 186 18.6 372.1 Air
The detailed procedure, protocol and material balance can be found in appendix I.
5.4.3 Ferric sulphate Fe2(SO4)3 leaching
Iron II sulphate hexahydrate (FeSO4.6H2O) was used as initial source of Fe2+. A mixture of
SO2 and O2 gases was used as oxidizing agent to oxidize Fe2+ to Fe3+; the regenerated Fe3+
served as leaching reagent. Fe3+ and Fe-total analysis in the solution was done with the aid of
absorption spectroscopy. Analysis of copper concentration in the leach solution was carried
out with the aid of AAS. The oxygen gas and sulphur dioxide were metered separately with a
5 mm diameter hose as shown in figure 5-3 before feeding directly into the pulp.
Concentrated sulphuric acid was carefully added to maintain the pH below 0.5 during the
tests.
The design of ferric (Fe3+) sulphate leaching tests was based on the claim by [6], [10], and
[21] that Fe3+ can serve as an oxidant to leach some metal sulphides and oxides including
copper sulphides and uranium oxides in acidic medium. The claims demonstrated leaching of
naturally occurring chalcocite and chalcocite concentrates [10], precipitation of manganese
[6] and leaching of uranium [21] with Fe3+ and the regeneration of Fe3+ using SO2/O2.
Preliminary tests were carried out to determine the kinetics of Fe2+ conversion to Fe3+ using
SO2 and O2 combined in a ratio of 1:10, at the temperatures 25, 45, 65 and 80oC as shown in
figure 5-6 and a comparison test using only O2 at the temperature with fastest rate; 65oC
figure 5-7.
42
0
500
1000
1500
2000
2500
3000
3500
4000
4500
5000
0 20 40 60 80 100 120 140 160 180 200
Time in minutes
Fe3+
con
tent
in m
g/l
0
10
20
30
40
50
60
70
80
% o
f tot
al F
e co
nver
ted
to F
e3+
Figure 5-6: Rate of Fe2+ conversion to Fe3+ as a function of time at 25oC (×), 45oC (�) 65oC
(�) and 80oC (�) using a mixture of 10% SO2 and 90% O2 as oxidizing gases and the
percentage of total Fe converted to Fe3+ at 65oC (�).
The figure shows plots of concentration of Fe3+ generated in solution as a function of reaction
time. It can be seen that the fastest reaction rate was obtained at 65oC.
43
0
500
1000
1500
2000
2500
3000
3500
4000
4500
5000
0 50 100 150 200
Time in min
Fe3+
con
tent
in p
pm
0
10
20
30
40
50
60
70
80
% o
f tot
al F
e co
nver
ted
to F
e3+
Fig
ure 5-7: Fe3+ formation as a function of time at 65oC using SO2/O2 (�)and O2 (�)as oxidizing
gases and percentage of total Fe converted to Fe3+ using SO2/O2 (�).
The figure shows a comparison between the rates of Fe3+ formation using SO2/O2 as well as
O2 separately. It can be seen that the SO2 have strong effect as an oxidizing agent when
combined with O2. It has greatly enhanced the conversion of Fe2+ to Fe3+ as can be seen in
figure 5-7. A kind of autocatalytic phenomenon can also be observed to have taken place after
60 minutes of reaction when about 470 ppm Fe3+ is present in solution with SO2/O2, a rapid
conversion rate can be seen in figure 5-7.
Stoichiometry relations of Fe2+ conversion to Fe3+ based on the mechanism proposed by
Zhang et al are shown in equations 5-9 to 5-16 [10] and leaching of chalcopyrite by Fe3+ in
equation 5-17.
SO2.H2O ⇔ HSO3- + H+ (5-9)
HSO3- ⇔ SO3
2- + H+ (5-10)
Fe3+ + SO32- ⇔ FeSO3
+ (5-11)
FeSO3+ → Fe2+ + SO3
- Slow (5-12)
SO3- + O2 → SO5
- Fast (5-13)
44
Fe2+ + SO5- + H+ → Fe3+ + HSO5
- Fast (5-14)
2Fe2+ + HSO5- + H+ → 2Fe3+ + SO4
2- + H2O Fast (5-15)
2HSO5- → SO4
2- + O2 + 2H+ (5-16)
CuFeS2 + 4Fe3+ → Cu2+ + 5Fe2+ +2So (5-17)
“It involves the slow initial formation of a ferric sulphite complex and decomposition to
produce sulphite radical SO3-. This is followed by a fast reaction with O2 to form a peroxo-
monosulphate species SO5-, and subsequently HSO5
-which is responsible for the oxidation of
Fe(II) and sulphite species”[10]. From (5-17) four moles of Fe3+ is required for dissolution of
one mole of chalcopyrite to produce one mole of Cu2+ five moles of Fe2+ and two moles of
elemental sulphur. Hence it can be concluded that to start up the reaction the mole ratio of Cu
to Fe3+ can be taken as 1:4. The detailed procedure, protocol and material balance can be
found in appendix I.
45
6 Results and discussions
6.1 Cyanide tests
The results obtained from AAS analysis of copper concentration in leach solution was used to
calculate the copper recovery in the solution as well as the amount of copper remaining in the
concentrate, the same concentrates was sent for analysis. The copper recovery, the calculated
residual copper concentration remaining in concentrate and the analyzed copper concentration
were plotted as a function of time as shown in figure 6-1 for the sodium cyanide only test and
in figure 6-3 for the sodium cyanide + calcium chloride test.
0
10
20
30
40
50
60
70
80
90
100
0 10 20 30 40 50 60
Leaching time hrs
Cu
reco
very
%C
u
0
0.5
1
1.5
2
2.5
3
Res
idua
l Cu
grad
e in
con
c.%
Cu
Figure 6-1: Copper recovery and residual grade in concentrates as a function of time using
sodium cyanide as a leaching reagent at ambient temperature. Cu recovery based on
calculated head (�). Calculated residual Cu grade in concentrate (�). Analyzed residual Cu
grade in concentrate (�).
The recovery curve shows three different zones: first zone between 0 and 10 hours
characterized by very fast kinetics and follows almost a linear relationship, second zone
between 20 and 40 hours characterized by a slower kinetic and follows a separate linear
relationship from first zone and the third zone which is very slow and also follows a
46
relationship. It can be summarized as having a rapid leach rate, which slows down. It can be
observed from the diagram that approximately 91% recovery of copper in the solution was
achieved after 53 hours. The calculated and analyzed copper grade curves are reasonably
inversely proportional to the recovery curve as expected, although there is little deviation
between calculated and analyzed results; probably due to homogenization problems in the
pulp.
Handling of the final pulp was a bit complex in this test; it was difficult to filter the pulp and
to wash the solid residues because the pulp was thick which lead to laminar flow. This may be
due to the fact that the duration of leaching was long which gave rise to too fine particles due
to prolonged attrition. The color of the dry residue seems to be closer to that of commercial
molybdenite as shown in figure 6-2 below.
Figure 6-2: Purified molybdenite concentrate with sodium cyanide
Table 6-1: Head and Purified concentrate assays sodium cyanide leaching
Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag
% % % % % % % % % g/t g/t
Head 46 2.6* 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate 49 0.55 0.02 35.9 2.53 0.023 0.16 0.1 0.0014 0.6 18
In addition to copper notable amounts of gold and silver dissolved in the experiment.
* The copper head assay was calculated from the analyzed copper concentration in the
pregnant solutions and the purified residues. It is a bit higher than the initial head analysis
(2.13%Cu) this could be due to sampling and homogeneity errors.
47
0
10
20
30
40
50
60
70
80
90
100
0 10 20 30 40 50 60 70 80
Leaching time
Cu
reco
very
in so
lutio
n %
0
0.5
1
1.5
2
2.5
3
Cu
grad
e in
con
c. %
Cu
Figure 6-3: Copper recovery and grade in concentrates as a function of time using sodium
cyanide and calcium chloride as leaching reagents at ambient temperature.
Cu recovery (�). Calculated Cu grade in concentrate (�). Analyzed Cu grade in concentrate
(�).
The curve also shows three different zones as the curve in figure 6-1: first zone between 0 and
20 hours which is also characterized by very fast kinetics and following almost a linear
relationship, second zone between 20 and 40 hours characterized by a slower kinetic and
following a separate linear relationship from first zone and the third zone which is very slow.
It can be summarized as having a rapid leach rate, which slows down with time.
It can be observed in the diagram that approximately 86% recovery of copper in the solution
was achieved around 48 hours, after which the recovery starts to decline although the
analyzed copper grade seems not to support this. The decline of recovery after reaching its
peak was also observed in a number of preliminary tests with cyanide. This was confirmed to
have resulted from reduced concentration of NaCN because the recovery normally rises again
after addition of more NaCN. Hence it can be concluded that based on the conditions of this
experiment, maximum recovery of the copper concentration into solution could also be
achieved within 48 hours.
48
Handling of the final pulp from this test was much more complex than that of the sodium
cyanide only test it was difficult to filter the pulp and to wash the solid residues because the
thickened pulp was much more laminar due to too fine particle sizes. The colour of the dry
residue was not similar to commercial molybdenite; it is ash like in colour as can be seen in
figure 6-4 and the particles are too fine; it is actually dusty. This may be due to the fact that
the duration of leaching was longer; 72 hours, which gave, rise to extensive and prolonged
attrition.
Figure 6-4: Purified molybdenite concentrate with sodium cyanide and calcium chloride
Table 6-2: Head and Purified concentrate assays sodium cyanide + calcium chloride leaching
Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag
% % % % % % % % % g/t g/t
Head 46 2.13 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate 43 0.47 0.02 32.4 2.44 0.023 5.50 0.47 0.0012 0.4 7.0 In addition to copper, it can also be observed that substantial amount of gold and silver also
dissolved in the experiment.
49
6.1.1 Mo-losses to solution The dissolution of molybdenum in cyanide leaching tests was negligible as can be seen in
figure 6-5. However, the addition of calcium chloride strongly lowered the dissolution.
0
0.05
0.1
0.15
0.2
0.25
0.3
0.35
0.4
0 20 40 60 80
Leaching time in hours
Mol
ybde
num
rec
over
y in
%
Figure 6-5: Mo-losses to solution as a function of time. Mo losses in CN test (�). Mo losses in
CN+Cl test (�).
6.1.2 Concentrate weight changes
An increase in weight of the concentrate (approximately 7%) was observed in sodium cyanide
+ calcium chloride leaching which also was observed in the preliminary experiments, while
all other leaching experiments gave weight losses. This could be due to precipitation of
compounds like for example gypsum which could have been induced by increased calcium in
residue. Purified concentrates samples was subjected to XRD analysis in order to identify the
precipitated substance but none of the possibly precipitated compound could be found.
Therefore the precipitated compound could be a non-crystalline substance. This may have
contributed to the slow filtration rate observed in the experiment.
6.1.3 Oxidising gases
Oxygen was mainly used in the cyanide leaching experiments; however air was also used in a
number of preliminary tests. The major difference observed between the use of air and pure
oxygen as oxidising agents was that, the pulp was floating after two hours with air as
oxidising agent (which could have lead to incomplete reaction in the pulp because large
50
amounts of material stocked to the wall of the reactor cover. When pure oxygen was used, the
pulp was more stable. The problem with floating is most probably smaller in large scale
reactors. Hence the use of air as oxidising agent could be more reasonable in large scale.
6.2 Ferric chloride (FeCl3) leaching The results of the copper concentration in leach solution obtained from AAS analysis was
used to calculate the copper recovery in the solution as well as the copper concentration
remaining in the concentrate, the same concentrates was sent for analysis. The copper
recovery, the calculated copper concentration remaining in concentrate and the analyzed
copper concentration were plotted as a function of time as shown in figure 6-5 below.
Figure 6-5: Copper recovery and grade in concentrate as a function of time using ferric chloride at high temperature above 70oC. Cu recovery (�). Calculated Cu grade in concentrate (�). Analyzed Cu grade in concentrate (�).
51
The curve shows that the leaching rate is fast during the initial two hours and thereafter
becomes slower. It can be observed from the figure that approximately 76% copper recovery
into the solution was achieved after 4 hours leaching time.
Handling of the final pulp from this test was easier than handling the final pulps in the
cyanide test; it was easier to filter the pulp and to wash the dry wet residue, the average
particle size is more granular than the particle sizes of the residues from cyanidation tests and
the dry residue gave a brighter gray color with closer resemblance to normal commercial
molybdenite as shown in figure 6-6; this may be due to the fact that the duration of leaching
was shorter in this test than the cyanidation tests leading to less extensive attrition.
Figure 6-6: Purified molybdenite concentrate after ferric chloride leaching
Table 6-2: Head and Purified concentrate assays ferric chloride leaching
Mo Cu Pb S Fe Bi Ca Cl Hg Au Ag
% % % % % % % % % g/t g/t
Head 46 3.06* 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate 49 0.65 0.003 37.8 1.40 0.004 - - 0.0002 3.29 20 The copper head assay was calculated from the analyzed copper concentration in the pregnant
solutions and purified residues. It was also found to be higher than the initial head analysis
(2.13% Cu) this is probably due to sampling and homogeneity problems. It can also be
observed that a substantial amount of lead, iron, bismuth, mercury, and silver was leached.
6.3 Ferric sulphate Fe2(SO4)3 leaching
The fastest conversion rate of Fe2+ to Fe3+ using an SO2/O2 ratio at 10% SO2 was obtained at
65oC as shown in figure 5-6.
52
0.0
0.5
1.0
1.5
2.0
2.5
3.0
3.5
4.0
4.5
5.0
0 0.5 1 1.5 2 2.5 3 3.5 4 4.5
leaching time in hours
Cu
reco
very
in %
495
500
505
510
515
520
Red
ox p
oten
tial
mV
Figure 6-7: Cu recovery and redox potential as a function of time using a ferric sulphate
solution generated from a ferrous sulphate solution by oxidation with a mixture of sulphur
dioxide and oxygen combined at 10% SO2 and 90% O2 as oxidising gases at 65oC.
Cu recovery (�). Redox potential (�).
The leaching result at the optimum conditions for ferric generation is also shown in figure 6-
7. It could however be seen that the result does not look really good but it shows that it works
to some extent with little extraction of copper, which probably could be improved upon based
on explanations by various previous works [23], [24], [33]: The major reason for the
ineffective leaching of copper is probably due to passivation of chalcopyrite at high redox
potential and lower temperature. The extent of copper extraction can however be increased at
elevated temperature and at lower potentials [23] [24]. It has also been demonstrated that
controlling the thermal and redox potential of the medium can counteract the passivation of
chalcopyrite [24]. Lowering the pH is another possibility [25].
The fact that chalcopyrite is the most refractory to leaching, among the copper sulphides,
makes the chemical leaching of chalcopyrite by an acidified solution of ferric sulphate
proceeds at a very slow rate. The rate of this reaction in the temperature range 50–110 oC is
also very low and the decrease in the rate is due to the formation of a film (passivation),
53
which builds up on the surface of the mineral and opposes the electron transfer from
chalcopyrite to the ferric medium which is necessary for the redox reaction. Jarosite
(H(Fe)3(SO4)2(OH)6) formation on the surface of chalcopyrite is a possibility as shown in
equation 6-1 and figure 6-8 [33].
H+ + 3Fe3+ + 2SO42- + 7H2O � H3O(Fe)3(SO4)2(OH)6 + 6H+ (6-1)
Figure 6-8: Passivation of chalcopyrite by jarosite
In order to improve the chalcopyrite-leaching rate, much effort has been made and several
catalysts have been proposed, silver ion being the most effective one. It is well known that
low concentrations of silver ions greatly accelerate the chalcopyrite leaching. Therefore, it is
necessary to employ a catalyst such as Ag (I) that kinetically activates the reaction between
chalcopyrite and Fe (III). It has been shown that there is a sharp increase in copper extraction
with addition of 0.2-3.0 mg Ag/g concentrate, further increases of the amount of catalyst have
little effect on the copper extraction [26]. Silver has been effectively recovered from residues
by leaching them with acidic brine medium with 200 g/L of NaCl and 0.5 M sulphuric acid
provided that elemental sulphur had been previously removed. High silver extractions (above
98 wt. %) were obtained in 1 h at 70oC for both concentrates. It is possible to obtain total
recovery of the silver added as a catalyst plus some of the silver originally present in the
concentrate by increasing the temperature to 90oC provided that the acid was not limiting
[26].
Handling of the final pulp from this test was easier; it was easy to filter the pulp and to wash
the dry wet residue, the average particle size was granular and the dry residue gave a gray
color with closer resemblance to the original molybdenite concentrate as shown in figure 6-9;
Jarosite CuFeS2 H3O(Fe)3(SO4)2(OH)6
Passivation
54
this may be due to the fact that the duration of leaching was short in this test leading to less
extensive attrition.
Figure 6-9: Treated concentrate from ferric sulphate leaching
The treated concentrate was not sent for analysis because the copper concentration in pregnant
solution was not substantial enough.
55
7 Conclusion
��91% of the copper concentration in the concentrate was removed with sodium cyanide
in oxidative environment within 53 hours. The molybdenite in the purified concentrate
is about 85% (49% Mo and 36% S) and dissolution of molybdenum was less than
0.4%; however there is concentrate weight loss of about 7%.
��74% of the copper was leached with a mixture of sodium cyanide and calcium
chloride solution in oxidative environment within 72 hours. The molybdenite
concentration in the purified concentrate had about 75% (43% Mo and 32% S).and
dissolution of molybdenum was only 0.1%; but there is a concentrate weight increase
of 7% which reduced the molybdenite concentration in the concentrate.
��76% of the copper concentration was removed with a solution of ferric, copper and
calcium chloride in oxidative environment within four hours. The molybdenite
concentration in the purified concentrate is about 87% (49% Mo and 38% S) the
molybdenum dissolution has not been analyzed but the concentrate weight loss was
5%. However the kinetics of copper recovery indicated increased copper recovery
with time.
��Ferric sulphate solution is not effective with only 4.5% copper removal based on the
pregnant solution analysis. The solid residue was not analyzed.
��It is possible to reduce the material loss in pilot scale experiments, because most
losses in the lab scale experiments are due to sticking on the equipment used.
56
8 Suggested further work
• Treatment of leach solution; It can be observed from the assays that a substantial
amount of Au and Ag also dissolved during cyanide leaching; hence the solution
can be routed through the gold leaching line and Au and other precious metals can
be recovered. The solution will therefore be treated at the final stage of former
cyanide destruction, hence requiring no separate treatment.
• Effect of the concentration of cyanide should be closely studied by taken closely
range samples; this would be more practicable in a pilot scale because the quantity
of concentrate used in this lab scale was small.
• The use of air as oxidising agent is more reasonable and cost effective and should
be tested also in the pilot scale.
• Detailed reagent consuming reactions (cyanide consuming reaction etc) should be
determined in more detail in order to optimize the process.
• Scanning electron microscope, SEM analysis should be carried out on the purified
concentrates in order to verify the precipitated compound, as it may be a non-
crystalline compound that could not be detected by XRD analysis.
• Recovery of free cyanide with addition of NaHS should be carried out and this
increases the possibility of copper recovery by complex formation and
precipitation of Cu2S. Cu(CN)2 + NaHS,(S2-) → Cu2S + CN-
Cu(CN)2- + HS → CuS + S
• It could be reasonable to test counter-current method of leaching with possible
bleeding periodically along the cycle in a pilot scale in order to speed up the
leaching rate.
57
References
[1] Fathi Habashi, Text book of Hydrometallurgy, Laval University, Quebec City, Canada,
second Edition
[2] P. Romano, M.L. Blazquez, F.J. Alguacil. J.A. Munoz, A. Ballester, F. Gonzalez (2001)
Comparative study on selective chalcopyrite bioleaching of a molybdenite concentrate with
mesophilic and thermophylic bacterial FEMS Microbiology Letters 196 (1), 71-75 (Volume
196 Issue 1 Page 71 – March 2001) doi: 10. 1111/j.1574-6968.2001.tb10543
[3] Ruiz M.C. and Padilla R, Department of Metallurgical Engineering, University of
Concepción Casilla 53-C, Concepción, CHILI, Copper removal from molybdenite concentrate
by sodium dichromate leaching, Hydrometallurgy (Hydrometallurgy) ISSN 0304-386X
CODEN HYDRDA, 1998, vol. 48, n 3, pp. 313-325 (13 ref.)
[4] Process for the treatment of molybdenum concentrate also containing copper, European
Patent EP1451380, http://freepatentsonline.com/EP1451380.html, Abstract of correspondent:
US2003124040.
[5] Purifying molybdenum flotation concentratesUnited States Patent 4083921
http://www.freepatentsonline.com/4083921.html
[6] W. Zhang, P. Singh and D. Muir., 2001. Oxidative preparation of manganese with
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[7] http://www.startprospecting.com/html/molybdenum-copper_separation_b.html
[8] Topsøe, H.; Clausen, B. S.; Massoth, F. E. "Hydrotreating Catalysis, Science and
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Arvena, J Copper 91 (Cobre 91); Ottawa; Ontario; Canada; 18-21 Aug. 1991. 1992
http://md1.csa.com/partners/viewrecord.php?requester=gs&collection=TRD&recid=1993094
40081MD&q=refining+of+molybdenite+concentrate&uid=790088489&setcookie=yes
[10] W. Zhang, .P .Singh, and D .Muir., 2000. SO2/O2 as an oxidant in hydrometallurgy.
Minerals Engineering, Vol. 13. No.13 pp.1319-1328. 2000
[11] Fathi Habashi, Handbook of Extractive Metallurgy, Vol. III, WILEY-VCH, 1997
[12] Pilot tests with Mo recovery in Aitik, Boliden in-house report, process technology,
September 2006.
[13] http://www.infomine.com/investment/metalschart.asp?c=Molybdenum&r=7d
58
[14] Fathi Habashi, Principles of Extractive metallurgy, Volume 2, Hydrometallurgy,
1980,pp. 18-19, Laval University, Quebec City,Canada
[15] http://www.dayah.com/periodic/?lang=en, under Mo
[16]http://upload.wikimedia.org/wikipedia/en/5/54/2005molybdenum_%28mined%29.PNG
[17] http://commons.wikimedia.org/wiki/Image:Molybdenite.jpg
[18] Manoj Kumar, T.R. Mankhand, D.S.R. Murthy, R. Mukhopadhyay, P.M. Prasad. 2006.
Refining of low-grade molybdenite concentrate. ScienceDirect Hydrometallurgy 86 (2007)
56-62.
[19] Website of China Tungsten Online (Xiamen) Manufacturing & Sales Corp.,
http://chinatungsten.com/ctop2.htm
[20] SADACI, websites, http://www.sadaci.be/molybdenum.html
[21] Elizabeth M. Ho, Clifford H. Quan ANSTO Minerals, PMB 1, Menai NSW 2234,
Australia Iron (II) oxidation by SO2/O2 for use in uranium leaching SciencDeirect
Hydrometallurgy 85 (2007) 183–192
[22] Marcel Pourbaix, atlas of electrochemical equilibria in aqueous solutions Pergamon
press, (1966) pp 272-279
[23] G. J. Olson · J. A. Brierley · C. L. Brierley Bioleaching review part B: Progress in
bioleaching: applications of microbial processes by the minerals industries Appl Microbiol
Biotechnol (2003) 63:249–257
[24] A.F. Tshilombo, J. Petersen, D.G. Dixon The influence of applied potentials and
temperature on the electrochemical response of chalcopyrite during bacterial leaching
Minerals Engineering 15 (2002) 809–813
[25] D.B. Johnson, Importance of microbial ecology in the development of new mineral
technologies. Hydrometallurgy 59_2001.147–157
[26] R. Romero, A. Mazuelos, I. Palencia, F. Carranza Copper recovery from chalcopyrite
concentrates by the BRISA process Hydrometallurgy 70 (2003) 205–215
[27] “cyanide” Wikipedia, Wikipedia 2007. Answers.com 25 Jun. 2007
http://www.answers.com/topic/cyanide
[28] www.boliden.com
[29] Process for purifying molybdenite concentrates. United States patent.3673424
http://www.freepatentsonline.com/3674424.pdf
[30] Boliden Internal report.
[31] Atomic absorption spectroscopy. (2007, May 22). In Wikipedia, The Free Encyclopedia
Retrieved 21:55, July 4, 2007, from
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http://en.wikipedia.org/w/index.php?title=Atomic_absorption_spectroscopy&oldid=13265717
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[32] Jianming Lu, D.B. Dreisinger, W.C Cooper Thermodynamics of the aqueous copper-
cyanide system Hydrometallurgy 66 (2002) 23-36
[33] Åke Sandström, Andrei Shchukarev, Jan Paul XPS characterization of chalcopyrite
chemically and bio-leached at high and low redox potential Minerals Engineering 18 (2004)
505-515
60
Appendix Appendix I Properties of molybdenum Physical properties Molybdenum is the second member of group 6 of the periodic table, with electronic
configuration [Kr] 4d6 5s1. It may have valency of 2, 3, 4, 5, or 6. It posses typical metallic
properties and lustrous silver-white colour in the massive state, and a dull grey colour in
powdery state. It has a body-centered cubic lattice with ao = 0.31472 nm. Accepted main
physical properties of molybdenum are given in table I-I below.
Table I-I Physical properties of molybdenum
Melting point, mp 2617-2623oC
Boiling point (101.3 kPa),bp 4612oC
Latent heat of fusion at mp 35.6 kJ/mol
Mean specific heat (0-100oC) 251 Jkg-1K-1
Density (20oC) 10.22g/cm3
Thermal conductivity (0-100oC) 137 Wm-1K-1
Electrical resistivity (20oC) 5.7µ�.cm
Temperature coefficient (0-100oC) 4.35×10-3K-1
Elastic constants of poly crystalline metal
(20oC)
Young’s modulus 324.8 Gpa
Rigidity modulus 125.6 Gpa
Bulk modulus 261.2 Gpa
Poisson’s ratio 0.293
Linear coefficient of thermal expansion (0-
100oC)
5.1 x 10-6K-1
Standard electrode potential EoMo3+,Mo -0.200V
Chemical properties. The lustre property of molybdenum can be retained indefinitely especially when it has been
drawn into wires. Its oxide, molybdenum trioxide, MoO3 through electrolytic oxidation at
prolonged temperature below 600oC, passivates it. The oxide sublime at 600oC and rapid
oxidation occurs thereafter. The metal also burns in oxygen at 500 to 600oC and it gets
61
oxidised slowly by steam and also attacked by fluorine when it is cold and by chlorine and
bromine when it is hot. It is slightly affected by dilute acids and concentrated hydrochloric
acid. It can be dissolved by moderately concentrated nitric acids but gets passivated in
concentrated nitric acid. A mixture of concentrated nitric and hydrofluoric acid dissolves
molybdenum effectively. Molybdenum is practically unaffected by alkaline solutions and
fused alkali-metal hydroxides, but can be dissolved rapidly by fused oxidising salts such as
sodium peroxide, sodium or potassium nitrate or perchlorate. It can react with carbon, boron,
silicon and nitrogen on heating and it forms many alloys. It finds application in a variety of
catalyst as mentioned earlier (section 1.3.3) especially in combination with cobalt in the
desulphurisation of petroleum. Biologically; molybdenum enhances the performance of
enzymes in the reduction of nitrogen to ammonia and also in the reduction of nitrates [11].
Appendix II Estimation of reagent for cyanide leaching
Total cyanide requirement for the test was calculated based on the 3 moles requirement for
dissolution of Cu and S as stated in equations (II-I) and (II-II) below:
MCu(CN)3- = nCN/Cu × nCu × MNaCN (II-I)
MSCN = nSCN/Cu × nCu × MNaCN (II-II)
MCu(CN)3- = Quantity of cyanide required for dissolution of Cu and formation of Cu(CN)32-
complex in gram.
MSCN = Quantity of cyanide required for formation of SCN- complex in gram.
nCN/Cu = mole ratio of CN: Cu
nSCN/Cu = mole ratio of SCN : Cu
nCu = number of mole of Copper in the concentrate in mol
MNaCN = molar mass of sodium cyanide in g per mole
Hence the total sodium cyanide requirement for the reaction, MCN is
MCN = MCu(CN)3- + MSCN
62
MCN was calculated to be approximately 20 g for this experiment.
This was calculated thus; nCN/Cu = 3 and nSCN/Cu = 3
The copper concentration in the concentrate is 2.13% Cu i.e. 4.26 g of Cu in 200 g
concentrate. Hence nCu = 4.26 g/63.55 gmol-1 = 0.0670 mol
MNaCN = 23 + 14 + 12 = 49 gmol-1
Therefore, MCu(CN)3- = 3 × 0.0670 mol × 49 gmol-1 = 9.85 g
Also MSCN = 3 × 0.0670 mol × 49 gmol-1 = 9.85 g
Hence MCN = MCu(CN)3- + MSCN = 19.71 g ~ 20 g NaCN required for complete dissolution of
Cu with formation of Cu(CN)32- and formation of SCN- complexes.
Procedure for cyanide leaching
The experimental set-up was as shown in figure 5-3, the concentrate and distilled water was
initially placed in the reactor and conditioned by heating the set-up to the specified
temperature, 25oC. The pulp solid concentration was initially 20%. Thereafter estimated
quantity of NaOH pellets (also CaCl2 in the second set-up) and NaCN were added along with
the supply of oxidative gas. The pH was constantly maintained above 11 throughout the
experiment by addition of NaOH pellets when required.
Pulp samples (pregnant solution) were taken periodically with the aid of a pipette; the sample
weight is always about 20 to 25 g (and it is assumed to be a true representative pulp sample).
The pulp sample is filtered to separate the solution from the solid residue. The density of the
solution is determined immediately after filtration (The weight of 5 ml of the solution was
measured and the density is calculated). The solution samples were analyzed for dissolved
copper concentration at Boliden Bioleaching Lab., with the aid of atomic absorption
spectrophotometer, AAS. The wet solid residues are oven-dried at about 70oC for about 72
hours. The dry weight of residue was measured accordingly and it was sent for analysis at
Rönskär’s central laboratory.
At the end of the experiment, the final pulp was weighted filtered and washed respectively.
The density of the final solution was determined (also by measuring the weight of 5 ml of the
solution). The weight and density of wash water was also measured. The wet residue was also
oven-dried at about 70oC for about 72 hours. The dry weight of residue was measured
afterwards and it was sent for analysis.
63
Table I-II Protocol and reagent consumption NaCN Leaching of Aititk Molybdenite concentrate to remove copper concentration by means of sodium cyanide Protocol and reagent consumption Date of Experiment 2007-07-03/05 Reference number of sample 31334 Material Aitik Mo-conc Grinding time Screen analysis, k80
Test description High Cyanide dosage Leaching
Test temperature 25 C Tare 0.00 kg
Copper concentration 2.60 %Cu Wt. of Copper 5.2 g Cu
Weight of material 0.20 kg 6.5 g/l Weight of solution 0.80 kg 852 weight of pulp 1.00 kg 0.52 %solid 19.51 % Gross weight 1.00 kg Solid density 4.30 kg/dm3 assumed 4.192 g/ml measured solution density 1.03 kg/dm3 Volume of pulp 0.8232 litre Pulp density 1.2148 kg/dm3
Reactor+Pulp wt 1085.80 g Empty reactor wt 450.01 g Final pulp weight 635.79 g Calc final solution concentration wt. 511.73 g
Density of final filtrate 1.04 g/ml, kg/dm3
Volume of final filtrate 481.66 ml Cu conc. in filtrate 4942.14 mgCu/l Cu concentration. in filtrate 2380.44 mg Weight of wash water 1322.09 g Volume of wash water 1307.55 ml
Density of wash water 1.01 g/ml, kg/dm3
Cu cont. in wash water 179.10 mgCu/l Cu concentration in wash water 234.19 mg Ini.wt.diswater+bot 578.00 g Fin.wt.diswater+bot 362.57 g Water added 215.43 g
64
Calc dry cake wt. 124.06 g Dry cake wt 137.00 g Cu concentration in dry cake 0.55 %Cu
Table I-III: Protocol and reagent consumption NaCN
Table I-IV: Leaching profile and material balance NaCN Leaching profile and material balance Date: 2007-07-03/05 Reference number: 31334 Material: Aitik Mo-conc Grinding time: Leaching temperature: 25 C Screen analysis:
Test description: High Cyanide dosage Leaching
Performed by: Fatai Ikumapayi Feed: Tare: 0.000 kg
Final weight of material: 0.185 kg Weight in 0.200 kg
Final weight of solution 0.726 kg Weight in 0.800 kg
Final weight of pulp: 0.912 kg
%solid: 20.334 % Solid losses 0.015 kg
Gross weight 0.912 kg 7.315 % Solid density: 4.300 kg/dm3 assumed Solution density: 1.035 kg/dm3 Volume of pulp: 0.745 liter Pulp density 1.224 kg/dm3 Cu concentration 2.600 %Cu Inventory of solution concentration Total final pulp weight 0.912 kg Total weight of NaCN added 0.020 kg Total weight of NaOH added 0.005 kg Total weight of CaCl2 added 0.000 kg Water addition 0.215 kg Total weights in 1.040 kg Total water losses 0.314 kg
65
Cu cont. In all soln. Samples 771.653 mg specific gravity of filtrate 1.036 Volume of filtrate 481.662 ml Copper conc. in filtrate 4,942.140 mg/l Copper concentration in separated filtrate 2,380.442 mg Weight of wash water 1.322 kg Copper conc. in wash water 179.104 mg/l Specific gravity of wash water 1.011 Copper concentration in wash water 234.187 mg Total copper concentration in solution 3,386.281 mg Weight of final leach residue 0.137 kg Copper conc. in leach residue 0.550 % Copper concentration in leach residue 753.500 mg Total weight of copper found 4,139.781 mg Copper concentration in the material 5,200.000 mg Cu mol in material 0.082 mol Copper concentration loss 1,060.219 mg 20.389 % % Copper leached 85.510 % Recovery in solution 65.121 % Recovery in residue 14.490 % Molybdenum concentration Conc. (HEAD) 46.000 % Molybdenum concentration purified conc. 49.000 % Mo weight in conc. (Head) 92.000 g Mo weight in purified conc. 67.130 g Molybdenum loss/leached 24.870 g Molybdenum recovery 72.967 % Mo-mol in Head 0.959 mol Mo-mol in purified conc 0.700 mol Sulphur concentration Conc (Head) 37.400 % Sulphur concentration purified Conc 35.900 %
66
S weight in conc. (Head) 74.800 g S weight in purified conc. 49.183 g Sulphur loss/leach 25.617 g Sulphur recovery 65.753 % S-mol in Head = 2*Mo-mol 2.332 mol S-mol in purif.con = 2*Mo-mol 1.534 mol Lead concentration Conc (Head) 0.019 % Lead concentration purified Conc 0.020 % Pb weight in conc. (Head) 0.038 g Pb weight in purified conc. 0.027 g Lead loss/leach 0.011 g Lead recovery 72.105 % Pb-mol in Head 0.000 mol Pb-mol in purif.con 0.000 mol Iron concentration Conc (Head) 2.830 % Iron concentration purified Conc 2.530 % Fe weight in conc. (Head) 5.660 g Fe weight in purified conc. 3.466 g Iron loss/leach 2.194 g Iron recovery 61.239 % Fe-mol in Head 0.101 mol Fe-mol in purif.con 0.062 mol Gold concentration Conc (Head) 0.000 % Gold concentration purified Conc 0.000 % Au weight in conc. (Head) 0.000 g Au weight in purified conc. 0.000 g Gold loss/leach 0.000 g Gold recovery 25.69 % Au-mol in Head 0.000 mol Au-mol in purif.con 0.000 mol
67
Silver concentration Conc (Head) 0.010 % Silver concentration purified Conc 0.002 % Ag weight in conc. (Head) 0.020 g Ag weight in purified conc. 0.002 g Silver loss/leach 0.017 g Silver recovery 12.582 % Ag-mol in Head 0.000 mol Ag-mol in purif.con 0.000 mol
Table I-V: Profile of residual elemental analysis NaCN
Purified analysis Leach time Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag hrs % % % % % % % % % g/t g/t 1 46 1.82 0.02 36.9 2.66 0.019 0.20 <0.1 0.0015 1.1 60 2 45 1.61 0.02 37.1 2.67 0.019 0.20 <0.1 0.0015 1.3 55 6 45 1.30 0.02 36.4 2.59 0.019 0.20 <0.1 0.0015 1.1 44 24 48 0.85 0.02 35.9 2.54 0.019 0.19 <0.1 0.0015 1.1 46 26 47 0.89 0.02 35.9 2.72 0.019 0.20 <0.1 0.0016 1.0 45 30 48 0.84 0.04 35.8 2.68 0.018 0.19 <0.1 0.0016 1.0 36 48 48 0.67 0.02 35.8 2.55 0.019 0.20 <0.1 0.0015 1.0 33 51 47 0.74 0.02 35.7 2.70 0.023 0.20 <0.1 0.0016 1.0 35 53 49 0.55 0.02 35.9 2.53 0.023 0.16 0.1 0.0014 0.6 18
Table I-VI: Protocol and reagent consumption NaCN + CaCl2 Leaching of Aititk Molybdenite concentrate to remove copper concentration by means of sodium cyanide Protocol and reagent consumption Date of Experiment 070710/13 Reference number of sample 31334 Material Aitik Mo-conc Grinding time Screen analysis, k80
Test description High Cyanide dosage Leaching
Test temperature 25 C Tare 0.00 kg
Copper concentration 2.13 %Cu Weight of Copper 4.26 g Cu
Weight of material 0.20 kg 5.325 g Cu/l Weight of solution 0.80 kg 705.3969953 weight of pulp 1.00 kg 0.426 %solid 17.09 % Gross weight 1.17 kg Solid density 4.30 kg/dm3 assumed 4.192 g/ml measured solution density 1.03 kg/dm3 Volume of pulp 0.8232 litre Pulp density 1.2148 kg/dm3
68
Reactor+Pulp wt 1337 g Empty reactor wt 448.35 g Final pulp weight 888.65 g Calc final solution concentration wt. 736.744 g
Density of final filtrate 1.089 g/ml, kg/dm3
Volume of final filtrate 675.036 ml Cu conc. in filtrate 3396.148 mgCu/l Cu concentration. in filtrate 2292.522 mg Weight of wash water 748.380 g Volume of wash water 735.668 ml
Density of wash water 1.017 g/ml, kg/dm3
Cu conc. in wash water 451.710 mgCu/l Cu concentration in wash water 332.308 mg Ini.wt.diswater+bot 566.17 g Fin.wt.diswater+bot 117.32 g Water added 448.85 g Calc dry cake wt. 151.906 g Dry cake wt 153.725 g Cu concentration in dry cake 0.470 %Cu
Table I-VII: Protocol and reagent consumption NaCN + CaCl2
Table I-VIII: Leaching profile and material balance Leaching profile and material balance Date: 070710/13 Reference number: 31334 Material: Aitik Mo-conc Grinding time: Leaching temperature: 25 C Screen analysis: Test description: Performed by: Fatai Ikumapayi Feed: Tare: 0.000 kg
69
Final weight of material: 0.215 kg Weight in 0.2 kg
Final weight of solution 1.051 kg Weight in 0.8 kg
Final weight of pulp: 1.266 kg Solid losses -0.015 kg
%solid: 16.986 % -7.48888 %
Gross weight 1.266 kg Solid density: 4.300 kg/dm3 assumed Solution density: 1.088 kg/dm3 Volume of pulp: 1.015 liter Pulp density 1.247 kg/dm3 Cu concentration 2.13 %Cu
Inventory of solution concentration Total final pulp weight 1.26562 kg Total weight of NaCN added 0.03 kg Total weight of NaOH added 0.02 kg Total weight of CaCl2 added 0.12 kg Water addition 0.44885 kg Total weights in 1.41885 kg Total water losses 0.368 kg Copper concentration in all soln. samples 785.16719 mg specific gravity of filtrate 1.08872 Volume of filtrate 675.0358219 ml Copper conc. in filtrate 3396.148 mg/l Copper concentration in separated filtrate 2292.5216 mg Weight of wash water 0.74838 kg Copper conc. in wash water 451.71 mg/l Specific gravity of wash water 1.01728 Copper concentration in wash water 332.30844 mg Total copper concentration in solution 3409.9972 mg Weight of final leach residue 0.153725 kg Copper conc. in leach residue 0.47 % Copper concentration in leach residue 722.5075 mg Total weight of copper found 4132.5047 mg Copper concentration in the material 4260 mg Cu mol in material 0.06703 mol Copper concentration loss 127.4953106 mg
70
% Copper leached 3412.990037 % Recovery in solution 80.04688238 % Recovery in residue 16.96026995 %
Mo weight in conc. (Head) 92 g Mo weight in purified conc. 66.10175 g Molybdenum loss/leached 25.89825 g Molybdenum recovery 71.84972826 % Mo-mol in Head 0.95893 mol Mo-mol in purified conc 0.688990515 mol Sulphur concentration Conc (Head) 37.4 % Sulphur concentration purified Conc 32.4 % S weight in conc. (Head) 74.8 g S weight in purified conc. 49.8069 g Sulphur loss/leach 24.9931 g Sulphur recovery 66.58676471 % S-mol in Head = 2*Mo-mol 2.33239788 mol S-mol in purif.con = 2*Mo-mol 1.553068288 mol Lead concentration Conc (Head) 0.019 % Lead concentration purified Conc 0.02 % Pb weight in conc. (Head) 0.038 g Pb weight in purified conc. 0.030745 g Lead loss/leach 0.007255 g Lead recovery 80.90789474 % Pb-mol in Head 0.000183398 mol Pb-mol in purif.con 0.000148383 mol Iron concentration Conc (Head) 2.83 % Iron concentration purified Conc 2.44 % Fe weight in conc. (Head) 5.66 g Fe weight in purified conc. 3.75089 g Iron loss/leach 1.90911 g Iron recovery 66.27014134 % Fe-mol in Head 0.101342883 mol Fe-mol in purif.con 0.067160072 mol
71
Gold concentration Conc (Head) 0.00016 % Gold concentration purified Conc 0.00004 % Au weight in conc. (Head) 0.00032 g Au weight in purified conc. 0.00006149 g Gold loss/leach 0.00025851 g Gold recovery 19.215625 % Au-mol in Head 1.62461E-06 mol Au-mol in purif.con 3.1218E-07 mol Silver concentration Conc (Head) 0.0098 % Silver concentration purified Conc 0.0007 % Ag weight in conc. (Head) 0.0196 g Ag weight in purified conc. 0.001076075 g Silver loss/leach 0.018523925 g Silver recovery 5.490178571 % Ag-mol in Head 0.0001817 mol Ag-mol in purif.con 9.97567E-06 mol
Table I-IX: Profile of residual elemental analysis NaCN + CaCl2
Purified analysis Leach time Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag hrs % % % % % % % % % g/t g/t 1 42 1.67 0.02 31.6 2.6 0.019 4.52 1.59 0.0013 1.3 68 3 42 1.42 0.02 31.1 2.58 0.019 4.72 1.88 0.0012 1.3 49 5 42 1.18 0.02 32.2 2.52 0.019 4.65 1.86 0.0011 1.1 43 7 42 1.09 0.02 31.8 2.52 0.019 4.89 2.06 0.0014 1.1 41 24 42 0.90 <0.02 30.9 2.55 0.023 5.50 2.42 0.0014 1.0 42 27 42 0.85 0.02 31.4 2.47 0.019 5.00 2.31 0.0014 0.9 37 48 41 0.72 0.02 31.2 2.42 0.022 5.10 2.79 0.0014 1.2 28 52 40 0.73 0.02 30.7 2.52 0.019 5.40 2.86 0.0014 1.0 34 55 38 0.69 0.02 29.2 2.50 0.019 7.00 2.94 0.0012 1.0 41 72 40 0.59 <0.02 29.4 2.25 0.023 5.40 3.47 0.0011 0.8 41 74 43 0.47 0.02 32.4 2.44 0.023 5.50 0.47 0.0012 0.4 7
Appendix III Procedure for ferric chloride leaching
The experimental set-up was as shown in figure 5-3, the concentrate and distilled water was
initially placed in the reactor. The concentrate was conditioned by heating the set-up to the
specified temperature about 90oC. The pulp solid concentration was initially 35%. Thereafter
estimated quantity of FeCl3, CuCl2 and were added along with the air supply. The pH was
72
constantly maintained about zero throughout the experiment by addition of aqueous solution
of HCl when required.
Pulp sample of about 58 g was taken after 2 hours of leaching with the aid of a pipette. The
pulp sample was filtered to separate the solution from the solid residue. The density of the
solution was determined immediately after filtration. The sample solution was analyzed for
dissolved copper concentration with the aid of AAS. The wet solid residues are oven-dried at
about 70oC for about 72 hours. The dry weight of residue was measured accordingly and it
was sent for analysis.
At the end of the experiment, the final pulp was weighted filtered and washed respectively.
The density of the final solution was determined. The weight and density of wash water was
also measured. The wet residue was oven dried at 70oC for about 72 hours. The dry weight of
residue was measured and was subsequently sent for analysis.
Table I-X: Protocol and reagent consumption FeCl3 Leaching of Aititk Molybdenite concentrate to remove copper concentration by means of Ferric chloride and other ligands Protocol and reagent consumption Date of Experiment 7/3/2007 Reference number of sample 31334 Material Aitik Mo-conc Grinding time Screen analysis, k80
Test description Ferric Chloride leaching
Test temperature 110 C Maximum Tare 0.00 kg
Copper concentration 3.06 %Cu Weight of Copper 13.16316 g Cu
Weight of material 0.43 kg 23.93302 g/l Weight of solution 0.55 kg 0 weight of pulp 0.98 kg 0.61224 %solid 34.58 % Gross weight 0.98 kg Solid density 4.30 kg/dm3 assumed 4.192 g/ml measured
solution density 1.03 kg/dm3
Volume of pulp 0.6340 litre Pulp density 1.5458 kg/dm3
Reactor+Pulp wt 1707.00 g Empty reactor wt 450.05 g
73
Final pulp weight 1256.95 g Calc final solution concentration wt. 822.34 g
Density of final filtrate 1.19 g/ml, kg/dm3
Volume of final filtrate 726.49 ml Cu conc. in filtrate 12478.23 mgCu/l Cu concentration. in filtrate 9065.31 mg Weight of wash water 529.11 g Volume of wash water 516.34 ml
Density of wash water 1.02 g/ml, kg/dm3
Cu cont. in wash water 1426.89 mgCu/l Cu concentration in wash water 736.76 mg Ini.wt.diswater+bot 576.78 g Fin.wt.diswater+bot 300.59 g Water added 276.19 g Calc dry cake wt. 434.61 g Dry cake wt 391.70 g Cu concentration in dry cake 0.65 %Cu
Table I-XI: Protocol and reagent consumption FeCl3
Table I-XII: Leaching profile and material balance FeCl3 Leaching profile and material balance Date: 7/3/2007 Reference number: 31334 Material: Aitik Mo-conc Grinding time: Leaching temperature: 70-110 C Screen analysis:
Test description: Ferric Chloride leaching
Performed by: Fatai Ikumapayi Feed: Tare: 0.000 kg
74
Final weight of material: 0.408 kg
Weight in 0.430 kg
Final weight of solution 0.906 kg
Weight in 0.550 kg
Final weight of pulp: 1.315 kg
%solid: 31.052 % Solid losses 0.022 kg
Gross weight 1.315 kg 5.070 % Solid density: 4.300 kg/dm3 assumed Solution density: 1.190 kg/dm3 Volume of pulp: 0.856 liter Pulp density 1.535 kg/dm3 Cu concentration 3.061 %Cu
Inventory of solution concentration Total final pulp weight 1.315 kg Total weight of FeCl3 added 0.080 kg Total weight of CuCl2 added 0.008 kg Total weight of CaCl2 added 0.160 kg Water addition 0.276 kg Total weights in 1.074 kg Total water losses 0.168 kg specific gravity of filtrate 1.191 Volume of filtrate 726.490 ml Copper conc. in filtrate 12,478.230 mg/l Copper concentration in separated filtrate 9,065.314 mg Weight of wash water 0.529 kg Copper conc. in wash water 1,426.894 mg/l Specific gravity of wash water 1.025 Copper concentration in wash water 736.757 mg Total copper concentration in solution 9,802.070 mg Weight of final leach residue 0.392 kg Copper conc. in leach residue 0.650 % Copper concentration in leach residue 2,546.050 mg Total weight of copper found 12,348.120 mg Copper concentration in the material 13,163.160 mg Cu mol in material 0.207 mol Copper concentration loss 815.040 mg 6.192 % % Copper leached 9,808.262 %
75
Recovery in solution 74.466 % Recovery in residue 19.342 %
Molybdenum concentration Conc. (HEAD) 46.000 % Molybdenum concentration purified conc. 191.933 % Mo weight in conc. (Head) 197.800 g Mo weight in purified conc. 751.802 g Molybdenum loss/leached -554.002 g Molybdenum recovery 380.082 % Mo-mol in Head 2.062 mol Mo-mol in purified conc 7.836 mol Sulphur concentration Conc (Head) 37.400 % Sulphur concentration purified Conc 37.800 % S weight in conc. (Head) 160.820 g S weight in purified conc. 148.063 g Sulphur loss/leach 12.757 g Sulphur recovery 92.067 % S-mol in Head = 2*Mo-mol 5.015 mol S-mol in purif.con = 2*Mo-mol 4.617 mol Lead concentration Conc (Head) 0.019 % Lead concentration purified Conc 0.003 % Pb weight in conc. (Head) 0.082 g Pb weight in purified conc. 0.012 g Lead loss/leach 0.070 g Lead recovery 14.383 % Pb-mol in Head 0.000 mol Pb-mol in purif.con 0.000 mol Iron concentration Conc (Head) 2.830 % Iron concentration purified Conc 1.400 % Fe weight in conc. (Head) 12.169 g Fe weight in purified conc. 5.484 g Iron loss/leach 6.685 g Iron recovery 45.064 %
76
Fe-mol in Head 0.218 mol Fe-mol in purif.con 0.098 mol Gold concentration Conc (Head) 0.000 % Gold concentration purified Conc 0.000 % Au weight in conc. (Head) 0.001 g Au weight in purified conc. 0.001 g Gold loss/leach -0.001 g Gold recovery 187.310 % Au-mol in Head 0.000 mol Au-mol in purif.con 0.000 mol Silver concentration Conc (Head) 0.010 % Silver concentration purified Conc 0.002 % Ag weight in conc. (Head) 0.042 g Ag weight in purified conc. 0.008 g Silver loss/leach 0.034 g Silver recovery 18.590 % Ag-mol in Head 0.000 mol Ag-mol in purif.con 0.000 mol
Table I-XIII: Profile of residual elemental analysis FeCl3 Purified analysis Leach time Mo Cu Pb S Fe Bi Hg Au Ag hrs % % % % % % % g/t g/t 0 *3.0612 2 45 1.11 0.004 34.4 1.91 0.004 0.0004 3.22 27 4 49 0.65 0.003 37.8 1.40 0.004 0.0002 3.29 20
Appendix IV Procedure for ferric sulphate leaching
The experimental set-up was as shown in figure 5-3, 1.5 kg of 0.1 M solution of FeSO4.6H2O
was placed in the reactor, and the set-up was heated to the fastest generating temperature,
65oC. SO2/O2 was subsequently supplied to the solution for minimum of one hour (when
sufficient Fe3+ has been formed to start up the leaching). The concentrate was added to the
solution to make up 10% solid concentration of the pulp. The pH was constantly maintained
below 1.00 throughout the experiment by addition of aqueous solution of H2SO4 when
required. The experiments lasted for 3 to 5 hours.
77
Pulp samples were taken at every one-hour and the solution concentration of the intermediate
samples and the final pulp was analyzed for dissolved copper concentration with the aid of
AAS. The wet solid residue was oven dried at about 70oC for about 72 hours. The dry weight
of residue was measured accordingly and kept for analysis.
Table I-XIV: Leaching profile and reagent consumption Fe2(SO4)3 leaching of Aitik Moly-conc to separate Copper through Ferric sulphate leaching date of experiment 5/22/2007 Provnr (sample number) 0.004683 kg Cu in 250g conc Tare 0 kg Cu-halt (Cu-concentration) 2.81 % 7.37E-05 Kmol Cu in 250g conc
Antag (% solid) 10 % 0.016464 kg Fe3+ required stoich. To leach Cu in conc.
Godsmängd(amount or quantity of material) 0.167 kg Lösningsvolym start(initial solution volume) 1.500 kg Temp 65 C Pulp weight 1.667 kg
Gross weight 1.667 kg (Tare+pulp weight)
Total Fe3+ requred stoichiometrically to leach Cu concentration in Conc. 0.016 kg
FeSO4 concentration added 40.5 g
0.099 kg/kg conc
Cu concentration 4.68333 g
Final pulp+reator 2838.7 g Reactor empty 941.6 g Final pulp weight 1897.1 g wwbini 563.48 g final 183.28 g water addition 380.2 g wash water+bottleini 573.16 g wash wate+bottle final 50.98 g washwater 522.18 g Weight of dry purified residue 149.3 g
Table I-XV: Leaching profile and reagent consumption Fe2(SO4)3
78
Table I-XVI: Molybdenum dissolution sodium cyanide leaching
Leaching time
Mo concentration in leach solution
Cu concentration in solution Cal. Cu grade Mo rec. Cu rec.
hrs mgMo/l mgCu/l %Cu % Mo % Cu 30 210 4400 0.84 0.18832 69.8077 51 420 5800 0.36 0.37663 92.0192 53 390 5300 0.56 0.34973 84.0865 Table I-XVII: Molybdenum dissolution sodium cyanide + calcium chloride leaching
Leaching time
Mo concentration in leach solution
Cu concentration in solution Cal. Cu grade Mo rec. Cu rec.
hrs mgMo/l mgCu/l %Cu % Mo % Cu 27 47 3200 0.7 0.04955435 72.8638498 48 85 4000 0.31 0.08961957 91.0798122 74 95 3600 0.48 0.10016304 81.971831 Appendix V Atomic absorption spectroscopy, AAS
Atomic absorption spectroscopy, AAS is an analytical technique for determining the
concentration of a particular metal element in a sample. Atomic absorption spectroscopy can
be used to analyze the concentration of over 62 different metals concentration in different
solutions. The technique normally makes use of flame to atomize the sample and turning the
sample into an atomic gas. Three basic steps are involved in achieving the atomization:
1. Desolvation – the liquid solvent is evaporated, and the dry sample remains
2. Vaporisation – the solid sample vaporises to a gas
3. Volatilisation – the compounds making up the sample are broken into free atoms
The flame is arranged such that it is laterally long (usually 10 cm) and not deep. Controlling
the flow of the fuel mixture must also control the height of the flame. A beam of light is
focused through this flame at its longest axis (the lateral axis) onto a detector past the flame.
A hollow cathode lamp produces the light that is focused into the flame. The lamp contains an
anode and a cylindrical metal cathode that holds the metal for excitation. When a high voltage
is applied across the anode and cathode, the metal atoms in the cathode are excited into
producing light with a certain emission spectra. The type of hollow cathode tube depends on
the metal being analyzed. For analyzing the concentration of copper in an ore, a copper
cathode tube would be used, and likewise for any other metal being analyzed. The electrons of
the atoms in the flame can be promoted to higher orbitals for instant by absorbing a set of
79
quantum energy. The amount of energy is specific to a particular electron transition in a
particular element. As the quantity of energy put into the flame is known, and the quantity
remaining at the other side (of the detector) can be measured, it is possible to calculate how
many of these transitions took place, and thus get a signal that is proportional to the
concentration of the element being measured [31].
Copper analysis with AAS
The copper analysis in leach solution was carried out on fresh pregnant leach solution in the
Boliden Bioleach-Cyanide Lab., with the aid of Atomic Absorption Spectrometer, AAS
model PU 9100X made by Phillips, using acetylene as burner gas.
Procedure for copper analysis with AAS
A copper-in-cyanide standard solution was prepared by diluting 1.41 g CuCN and 4 g NaCN
to make 1liter of solution.
Standard solutions containing 1 ppm, 3 ppm, 5 ppm, 10 ppm and 20 ppm respectively was
prepared from the copper-in-cyanide standard solution.
1ppm standard was prepared by diluting 100 �l 100 times in a 100 ml standard volumetric
reagent bottle, 3 ppm by diluting 300 microns 100 times, 5 ppm by diluting 500 microns 100
times, 10 ppm by 1 ml 100 times and 20 ppm by diluting 2 ml 100 times. The figure shows
the linear curve as ppm as a function of AAS signals in each of the standard solutions. A line
of best fit was plotted such that the R2 value is 0.9-1.0. The copper quantity in solution is
calculated using the equation of the curve and the appropriate signals. The value of signals
obtained from experimental samples must lie within the range of signals obtained from the
standard samples therefore appropriate dilution of the experimental samples must be done in
order to achieve this. An example is shown in figure II-I.
80
Cu analysis
y = 4.7019x + 1.5251
R2 = 0.9999
0
20
40
60
80
100
120
0 5 10 15 20 25
Cu contents in ppm
AA
S si
gnal
s
Figure II-I
The figure shows the plots of AAS signals as a function of copper concentration of the
corresponding standard solution in table II-I.
Table II-I: Copper concentration in standard solution and corresponding AAS signals Cu concentration in ppm
AAS signals
1 6 3 15.5 5 25.5 10 48.5 20 95.5
If 200 µm of a leach solution is diluted to 100 ml i.e. 500 times and the AAS signals from this
solution is 37.7. The copper concentration in the solution can be calculated using the equation
of the curve since the signal lies within the standard solutions signals and the R2 of the curve
is 0.999: y = 4.7019χ + 1.5251
5007019.4
5251.1 ×−= yx
χ = 3825.57 ppm Cu i.e. the copper concentration in the solution.
81
Figure II-I: Atomic Absorption Spectrometer
The AAS equipment used in this study is shown in figure II-I.
XRD Equipment The material was placed in sample holders and pressed manually with a glass slide to achieve
a flat surface. The equipment utilized was a Siemens D5000 X-ray diffractometer, figure II-II,
using copper K� radiation with accelerating voltage of 20 kV. XRD patterns were recorded
from 30 to 50° and 30 to 45°. The phase identification was made by reference patterns in an
evaluating program known as EVA.
Figure II-II: X-ray diffractometer
82
Appendix VI Composition of commercial molybdenite concentrate brands Table II-II Molybdenum Concentrate Brands & Chemical Compositions
Chemical compositions
Impurities, %, Max
Brand
Mo
%
min SiO2 As Sn P Cu Pb CaO WO3 Bi
Kmo53-
A 53 6.5 0.01 0.01 0.01 0.015 0.15 1.50 0.05 0.05
Kmo53-
B 53 5.0 0.05 0.05 0.02 0.20 0.30 2.00 0.25 0.10
Kmo51-
A 51 8.0 0.02 0.02 0.02 0.20 0.18 1.80 0.06 0.06
Kmo51-
B 51 5.5 0.10 0.06 0.03 0.40 0.40 2.00 0.30 0.15
Kmo49-
A 49 9.0 0.03 0.03 0.03 0.22 0.20 2.20 -- --
Kmo49-
B 49 6.5 0.15 0.06 0.04 0.60 0.60 2.00 -- --
Kmo47-
A 47 11.0 0.04 0.04 0.04 0.25 0.25 2.70 -- --
Kmo47-
B 47 7.5 0.20 0.07 0.05 0.80 0.65 2.40 -- --
Kmo45-
A 45 13.0 0.05 0.05 0.05 0.20 0.30 3.00 -- --
Kmo45-
B 45 8.5 0.22 0.07 0.07 1.20 0.07 2.60 -- --
[19]
83
Composition of commercial molybdenum oxide brands Table II-III :Technical Grade Molybdenum Oxide.Sadaci Mo 57.00% min.
Cu 0.50% max.
S 0.10% max.
C 0.10% max.
P 0.05% max.
Pb 0.05% max.
[20]
Table II-IV :Purified concentrates analysis. Test Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag % % % % % % % % % g/t g/t CN 49 0.55 0.02 35.9 2.53 0.023 0.16 0.1 0.0014 0.6 18 CN+Cl 43 0.47 0.02 32.4 2.44 0.023 5.50 0.47 0.0012 0.4 7 FeCl3 49 0.65 0.003 37.8 1.40 0.004 - - 0.0002 3.29 20 Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag % % % % % % % % % g/t g/t Head 46 2.13 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Table II-V: Cyanide leaching test assays Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag % % % % % % % % % g/t g/t Calculated head assays
46 2.6* 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate
assays 49 0.55 0.02 35.9 2.53 0.023 0.16 0.1 0.0014 0.6 18
Table II-VI: Cyanide + Chloride leaching test assays Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag % % % % % % % % % g/t g/t Calculated head assays
46 2.13 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate
assays 43 0.47 0.02 32.4 2.44 0.023 5.50 0.47 0.0012 0.4 7
84
Table II-VII: Ferric chloride leaching test assays Mo Cu Pb S Fe Bi Ca Clx Hg Au Ag % % % % % % % % % g/t g/t Calculated head assays
46 3.06* 0.019 37.4 2.83 0.011 0.11 <0.1 0.0016 1.6 98
Purified
concentrate
assays 49 0.65 0.003 37.8 1.40 0.004 - - 0.0002 3.29 20 * Assays calculated from analyzed solutions and residues
85
Appendix VII SCN formation The major cyanide consuming reactions in the cyanidation tests is the formation of Cu(CN)3
-, SCN complexes and negligible CNO complex formation.
Table III-I: SCN formation from previous study
Time in hours Cu mg/l CNS mg/l mg/mg ratio
mol/mol ratio
48 124 370 2.983871 3.268882 48 127 290 2.283465 2.501575 48 819 1300 1.587302 1.738916 48 860 1310 1.523256 1.668753 2.094473 2.294532 - Average ~3 ~3
The table shows SCN formation as a ratio of copper complex formation from previous
experiments on cyanide leaching of Lakefield chalcopyrite. It gives estimation of cyanide
consumed by both sulphur and copper in a typical reaction containing chalcopyrite and other
metal sulphides. It can be seen that both the mass: mass and mol: mol ratio can be rounded up
to 3 [30]. This shows an indication of cyanide consuming species in the reaction involving
cyanide and metal sulphides, especially chalcopyrite. It can be seen that sulphur is always
consuming as much mole of cyanide as copper is consuming. It was also observed in the
study that CNO is also formed during such reaction, but the quantity is negligible compared to
CNS formation. Therefore the major consumption of cyanide is the formation of copper
complex and sulphur thiocyanite.
Table III-II: SCN formation this study NaCN
Time in hours SCN mg/l Cu mg/l mg/mg mol/mol 1 2085 1300.2 1.603599 1.757047 6 3987.5 2503.1 1.593025 1.745461 24 6460 3900 1.65641 1.814912 30 7210 4270 1.688525 1.850099 51 8830 5420.67 1.62895 1.784824
Average 1.634102 1.790468
Still approximately 1:1 but as 2 mol: 2 mol unlike the observation from the previous study
which is 3 mol: 3 mol.
The previous higher formation rate of SCN versus Cu dissolution seems to have come from
the high pyrite concentration or pyrrotite. However it seems not to have any significant
86
contribution of sulphur from MoS2 to the SCN formation in this case. Hence there seems to be
better utilization of CN in this test. Although it can be observed that the SCN formation rate is
increasing with time; this may be due to increased concentration of sulphur in the solution,
which seem to increase the formation rate of the complex.
Table III-III: SCN formation this study NaCN + CaCl2
Time in hours SCN mg/l Cu mg/l mg/mg mol/mol 1 2002 1175.095 1.703692 1.866718 5 3506 2011.236 1.743207 1.910014 27 5905 3009.314 1.962241 2.150007 48 6659 3756.239 1.772784 1.942421 74 6030 3396.148 1.775541 1.945442
Average 1.79149 1.96292
The formation rate of SCN is much reduced in the test with addition of CaCl2 as can be seen
in the table III-III, the formation is actually still approximately 1:1 i.e. mg/mg is ~2 with more
reduced mol/mol ~2 indicating that sulphur is not consuming much of the cyanide and a better
cyanide utilization.
87
Appendix VIII Mo-CN stability constant START Experiments recorded for Boliden Mineral AB, Boliden, Sweden from SC-Database on Monday, 20 July, 7-02 at 08:51:38 Software version = 5.4 Data version = 4.51 Experiment list contains 26 experiments for 4 ligands : Cyanide, Cyanate, Thiocyanate, Selenocyanate 5 metals: Mo(0), Mo(III), Mo(IV), Mo(V), Mo(VI) (no references specified) (no experimental details specified) ***************************************************************************** CN- HL Cyanide CAS 74-90-8 (230) Cyanide; ----------------------------------------------------------------------------- Metal Mtd Medium Temp Conc Cal Flags Lg K values Reference ExptNo ----------------------------------------------------------------------------- Mo(IV) nmr KNO3 25°C 0.10M C 1994RLa (2705) 1 *K(MoO(CN)4(H2O)=-9.88 Method: N.M.R. ----------------------------------------------------------------------------- Mo(IV) con oth/un 25°C dil U M 1974FIb (2706) 2 K(K+Mo(CN)8)=1.8 K(Me4N+Mo(CN)8)=2.5 K(Et4N+Mo(CN)8)=2.3 ----------------------------------------------------------------------------- Mo(IV) gl none 25°C 0.0 U T H 1973BKa (2707) 3 K(MoOOH(CN)4+H)=8.81 K=8.86(30 C). K=8.90(35 C). K=8.97(40 C). K=9.04(45 C). K=9.13(50 C). DH=23.4 kJ mol-1 ----------------------------------------------------------------------------- Mo(IV) sp NaClO4 25°C var U 1973MHa (2708) 4 K(Fe+Mo(CN)8)=2.6 ----------------------------------------------------------------------------- Mo(IV) sp NaClO4 25°C var U M 1971JSb (2709) 5 K(Fe+Mo(CN)8)=2.6 ----------------------------------------------------------------------------- Mo(IV) sp oth/un 25°C var U M 1969KBc (2710) 6 K(UO2+Mo(CN)4(OH)3(H2O))=3.71 ----------------------------------------------------------------------------- Mo(IV) sp oth/un 25°C var U M 1968DBb (2711) 7 K(VO+MoL4(OH)3H2O)=4.86 ----------------------------------------------------------------------------- Mo (IV) gl oth/un 25°C 0.0 U 1968PNb (2712) 8 K (H+MoO2L4) =12.62 K (H+MoOOHL4) =9.98 ----------------------------------------------------------------------------- Mo(IV) con oth/un 25?°C dil U M 1958SEa (2713) 9 Ks(KAg2Y(s))=-13.96
88
Ks(Ag3Y(s))=-13.83 Ks(Mn3Y2(s))=-12.35 Ks(Fe3Y2(s))=-16.28 Y=MoSOHL4 (H2O)2---. Ks(Co3Y2)=-13.92; Ks(Ni3Y2)=-18.23; Ks(Cu3Y2)=-18.46; Ks(Zn3Y2)=-13.62; Ks(Cd3L2)=-18.32; Ks(Hg3Y2)=-18.73; Ks(Pb3Y2)=-18.52 ***************************************************************************** SCN- HL Thiocyanate CAS 463-56-9 (106) Thiocyanate; ----------------------------------------------------------------------------- Metal Mtd Medium Temp Conc Cal Flags Lg K values Reference ExptNo ----------------------------------------------------------------------------- Mo(III) kin oth/un 25°C 2.00M U 1997NCa (14850) 10 K(Mo4S4(H2O)12+L)=3.11 K(Mo7S8(H2O)18+L)=2.94 Medium: Li-p-toluenesulfonate. ----------------------------------------------------------------------------- Mo(III) kin oth/un 25°C 2.00M U 1993HLa (14851) 11 K(Mo4S4+L)=3.11 Medium: Li toluene-p-sulfonic acid. For Mo(IV), K=3.72; for mixed Mo(III)/ Mo(IV) (Mo4S4+++++), K=3.48. ----------------------------------------------------------------------------- Mo(III) kin oth/un 25°C 1.0M U K1=5.0 1974SSd (14852) 12 medium:lithium p-toluenesulfonate ----------------------------------------------------------------------------- Mo(III) sp oth/un ? 1.0M U K1=0.6 1972KTa (14853) 13 Medium: p-toluenesulfonic acid ----------------------------------------------------------------------------- Mo(IV) kin NaClO4 25°C 2.00M U 1993LMb (14854) 14 K(Mo3Se4+NCS)=3.38 K(Mo3OSe3+NCS)=3.23 K(Mo3O2Se2+NCS)=3.66 K(Mo3O3Se+NCS)=3.18 K(Mo3O4+NCS)=2.99. Medium: 2.0 M HClO4. ----------------------------------------------------------------------------- Mo(IV) kin NaClO4 25°C 2.00M U 1993VSa (14855) 15 K(Mo3S4(H2O)9+L)=3.36 K(Mo2WS4(H2O)9+L)=3.48 K(MoW2S4(H2O)9+L)=3.68 Medium: 2.0 M HClO4. For mixed Mo/W species data refer to L binding to Mo. Metals are Mo(IV) and W(IV). ----------------------------------------------------------------------------- Mo(IV) kin oth/un 25°C 2.0M U T K1=2.54 1976OSa (14856) 16 Medium: LiClO4/HClO4, metal: MoO++. K1=2.89 (10 C); 2.73 (15 C); 2.61 (20 C) ----------------------------------------------------------------------------- Mo(V) sp mixed 20°C ? C 1986CZa (14857) 17 B(CuH-2L)=-7.88 B(CuH-3L)=-15.12 Medium: DMSO/acetone -----------------------------------------------------------------------------
89
Mo(V) kin NaClO4 25°C 1.00M U M 1976CSa (14858) 18 K(Mo2O4(C2O4)2+L)=0.74 By spectrophotometry: K=0.63 ----------------------------------------------------------------------------- Mo(V) kin NaClO4 25°C 2.00M U T 1975STa (14859) 19 K(Mo2O4+L=Mo2O4L)=2.38 Medium: LiClO4 ----------------------------------------------------------------------------- Mo(V) sp non-aq ? 100% U K1=2.88 1970BRb (14860) 20 Medium: (EtO)2PSSEt + EtOH(4:1) ----------------------------------------------------------------------------- Mo(V) nmr NaClO4 23°C 2.0M U M 1968MDf (14861) 21 K(MoOL4+A=MoOL3A+L)=-1.64 K(MoOL4+2A=MoOL2A2+2L)=-3.24 K(MoOL4+3A=MoOLA3+3L)=-6.19 Medium: HClO4. A=(NH2)2CS ----------------------------------------------------------------------------- Mo(V) sp non-aq ? 100% U I K1=5.0 B2=9.40 1965ULa (14862) 22 K3=4.0 K4=3.4 Medium: Me2CO, Mo as MoCl5. In MeOH: K1=3.85 ----------------------------------------------------------------------------- Mo(V) sp oth/un ? 3.25M U I 1959NAb (14863) 23 K6?=1.35 Medium: H2SO4. In 3.1 M (NH4)2SO4 K3*K4*K5?=2.25 ----------------------------------------------------------------------------- Mo(V) sp mixed ? 60% U K1=3.2 B2=6.2 1958PEb (14864) 24 K3=ca.2 K4=-1.6 Medium: 60% w/w acetone/H2O ----------------------------------------------------------------------------- Mo(V) sp mixed 20°C 60% U K1=3.2 B2=6.2 1958PEb (14865) 25 K3=1.85 Meedium: 60% w/w acetone/H2O, 1 M HCl. Also by electrical migration ----------------------------------------------------------------------------- Mo(VI) nmr oth/un ? var U M 1969MDb (14866) 26 K(MoOL4+A=MoOL3A+L)=-1.5 K(MoOL4+2A=MoOL2A2+2L)=-3.1 K(MoOL4+3A=MoOLA3+3L)-5.1 K(MoOL4+4A=MoOA4+4L)=-7.6 A=Br-. Other ternary complexes also reported. Method: esr ----------------------------------------------------------------------------- REFERENCES 1997NCa M Sokolov,N Coichev,H Moya,A Sykes et al; J.Chem.Soc.,Dalton Trans.,1863 (1997) 1994RLa A Roodt,J Leipoldt,L Helm et al; Inorg.Chem.,33,140 (1994) 1993HLa M Hong,Y Li,J Lu,M Nasreldin et al; J.Chem.Soc.,Dalton Trans.,2613 (1993) 1993LMb G Lamprecht,M Martinez,M Nasreldin; J.Chem.Soc.,Dalton Trans.,747 (1993)
90
1993VSa J Varey,A Sykes; J.Chem.Soc.,Dalton Trans.,3293 (1993) 1986CZa Chen Lianshan,Zhao G L,He,Z L,Zhao H G; Acta Chimica Sinica,520 (1986) 1976CSa G Cayley,A Sykes; Inorg.Chem.,15,2882 (1976) 1976OSa J Ojo,Y Sasaki,R Taylor et al; Inorg.Chem.,15,1006 (1976) 1975STa Y Sasaki,R Taylor,A Sykes; J.Chem.Soc.,Dalton Trans.396 (1975) 1974FIb F Ferranti,A Indelli; J.Solution Chem.,3,619 (1974) 1974SSd Y Sasaki,A Sykes; J.Less Common Metals,36,125 (1974) 1973BKa M Beg,Kabir-ud-Din et al; Australian J.Chem.,26,671 (1973) 1973MHa G McKnight,G Haight; Inorg.Chem.,12,1934 (1973) 1972KTa K Kustin,D Toppen; Inorg.Chem.,11,2851 (1972) 1971JSb D Joshi,K Sharma; Z.Phys.Chem.,246,281 (1971) 1970BRb A Busev,T Rodionova; Anal.Lett.,3,325 (1970) 1969KBc Kabir-ud-Din,M Beg; J.Indian Chem.Soc.,46,503 (1969) 1969MDb I Marov,Y Dubrov,A Ermakov,G Martynova; Zh.Neorg.Khim.,14,438(E:224) (1969) 1968DBb Kabir-ud-Din,M Beg; J.Indian Chem.Soc.,45,455 (1968) 1968MDf I Marov,Y Dubrov,A Ermakov,G Martynova; Zh.Neorg.Khim.,13,3247 (1968) 1968PNb J van de Poel,H Neumann; Inorg.Chem.,7,2086 (1968) 1965ULa N Ulko; Ukr.Khim.Zh.,31,887 (1965) 1959NAb B Nabivanets; Zh.Neorg.Khim.,4,1797 (1959) 1958PEb D Perrin; J.Am.Chem.Soc.,80,3540 (1958) 1958SEa A Sergeeva; Nauk Zapiski L'vov Inst.,50,22 (1958) EXPLANATORY NOTES DATA Flags are :- T Data at other TEMPERATURES I Data with various BACKGROUNDS H Data for THERMOCHEMICAL quantities M Data for TERNARY Complexes EVALUATION Flags are :- T or IUP=T signifies EVALUATION RATING = Tentative by IUPAC ----------------------------------------------------------------------------- END Experiments recorded for Boliden Mineral AB, Boliden, Sweden from SC-Database on Monday, 20 July, 7-02 at 08:51:38