In this process (Figure 7.4), chalcopyrite concéntrate was treated with concentrated sulfuric acid at 200°C:
When leaching concentrates at ambient pressure and at high tem-perature but below the boiling point, then it is necessary to use a reflux condenser to avoid the escape of vapours. If the reacting mass solidifies after few minutes when the reaction started as for example in the case of chalcopyrite, then the baking process should be adopted and special equipment must be used.
Treadwell process for chalcopyrite
The action of concentrated H2S0^ on sulfide minerals received at-tention in the 1960s because of the fact that under certain conditions elemental sulfur can be formed and therefore pollution due to SOj that generally forms in smelters can be avoided. A process was developed on laboratory scale at Treadwell Corporation in Bronx, New York and was tested in Tucson, Arizona in 1970 at Anaconda Company. A 100 tonnes chalcopyrite concéntrate per day pilot plant was constructed for this purpose (Figure 7.3).
CuFeS, + 4H,S0^ -^ CuSO^ + FeSO, + 2 S 0 , + 2S + 4 H , 0
2 2 4 4 4 2 2
Make up HzSO^
H 2 0
-Sulfide concéntrate
Concentrated HsSO^
V Y t , so. , 7 " I
Baking [- -"^ Acid plant |
y V
Leaching
Filtralion [ — • - S u l f u r , gangue
Purification Spsnt
electtí)lyte I I Evapofatlon \^— --| Recovery |
Y
Metal
Figure 7.4 - Flowsheet for the treatment of chalcopyrite with concentrated sulfuric acid
Copper and iron in the mineral are converted to water-soluble sul-fates, while elemental sulfur is formed. The sulfur dioxide formed during the reaction must be converted to sulfuric acid, for recycle.
A number of side reactions may also take place if the conditions are not properly selected.
By using a two-stage process whereby the concéntrate was first agi-tated with a stoichiometric volume of 98% sulfuric acid. This step was quite short, taking only few minutes. In the second stage, the solidified mass was heated further until the reaction was complete and the product then leached with water to remove the soluble sul-fates from the gangue and elemental sulfur. A complicated system of bucket elevator and silica balls for heat transfer was used. The process was criticized by many metallurgists because SO2 was gen-erated and must be converted to acid for recycle and ferrous sulfate must be decomposed to genérate acid for recycle.
Figure 7.3 - Pilot plant at Tucson, Arizona
188 Pressure Hydrometallurgy
The situation was compounded further by the poHtical situation in Chile where the new socialist regime nationalized the copper indus-try and Anaconda lost its properties in Chuquicamata. The pilot plant was abandoned, and the process was dismissed as uneconomi-cal. At the same time, a new process was developed and became known later as the Arbiter process.
Incidentally, researchers at Kennecott Copper Corporation in Salt Lake City, Utah had a similar process under investigation but they used a bug mili for treating the sulfuric acid-chalcopyrite concén-trate and the process was abandoned before leaching a pilot scale.
Eievation
— __ f ü
y f p Plan Figure 7.5 - Bug mili
Arbiter process
The process (Figure 7.X) was developed in 1970 in a pilot plant at Anaconda Company in Tucson under the direction of Nathaniel Ar-biter, a former professor of mineral processing at Columbia School of Mines in New York City. A commercial 90 tonnes/day copper plant went in operation at Anaconda, Montana few months later.
The leaching plant was composed of 10 intensely agitated vessels 14 m3 each and 5 counter-current decantation thickeners, with the first overflow filtered to form the pregnant solution. The process
Chapter 7 - Attempts to Avoid Autoclaves 189
is based on leaching chalcopyrite concéntrales with ammonia at 75-80°C in presence of oxygen to form copper ammine sulfate and i ron hydroxide (Figure 7.6):
2CuFeS2 + I2NH3 + SyzOj + 2 H p 2[Cu(NH3)J2* + 4 N H / + 4S0/- + ?ep^
NH,
CaO
Distillation Raffinate
Gypsum
Water-H,SO,
Chalcopyrite concéntrate
^ 2 ' ^
Leaching
"1~
Filtration V ir Solvent Extraction
Air
Solids
Washing M—Water
Organic
Washing
Stripping CuSO^ solution
Filtration . Dilute NH3 solution
Residue
Organic phase
Electrowinning
Copper
Figure 7.6 - Arbiter process
After filtering the solids, copper was extracted by LIX-65N, stripped by sulfuric acid, and electrolyzed to give metallic copper. Lime is then added to the raffinate, and the slurry boiled to distil off am-monia for recovery and precipítate gypsum for disposal:
(NHASO + Ca(OHL 2NH, + CaS0,.2H,0
3 4 2
TT
190 Pressure Hydrometallurgy
This ammonia recovery step as well as that from the residue wash-ings proved to be technically difficult and economically unsound and was the main reason for the shut down.
HydroCopper process
Chemists at Outotec, formerly Outokumpu Research Oy, in Pori, Finland have developed HydroCopper process for the treatment of copper sulfide concentrates avoiding the use of autoclaves (Figure 7.7). The process is based on leaching the concentrates in a strong NaCl solution containing Cu^* ion at pH 1.5-2.5 in agitated reac-tor at 85-95°C in presence of oxygen. Copper goes into solution as Cu* while iron is precipitated as hydroxide. After filtration and solution purification NaOH is added to precipítate CU2O, which is then slurried in water and reduced in autoclaves by hydrogen under pressure.
NaCl solution NaOH Electrolysis
H?
T
Copper
Figure 7.7 - Outokumpu, now Outotec, HydroCopper process
Chapter 7 - Attempts to Avoid Autoclaves 191
Leaching of sulfides in chioride media has been tried before in Clear Process developed by Duval Corporation in Arizona in the 1970s. The process, however, was not industrialized because the electrowinning of copper from chioride médium produced dendridic powder contaminated by silver that was difficult to process further.
It seems that this was the reason for Outokumpu chemists to avoid the electrowinning route and consider the production of Cu^O and its reduction.
The precipitation of copper by hydrogen under pressure from aque-ous chioride system is not effective. Another option is the thermal reduction of solid CuCI in a fluidized bed. Outokumpu researchers also found out that CUjO disproportionate in dilute H2SO4:
Cup + 2H* Cu + Cu"* + Hp
CuSO^ formed can be reduced by hydrogen in an autoclave by known methods. In this process, CaCI^ was a waste product for disposal.
In the Outokumpu process NaOH is used instead of Ca(0H)2 so that NaCl produced can be electrolyzed to recover hydrogen for reduc-tion, NaOH for precipitareduc-tion, and chlorine transformed into HCI. Iñ other words, an important section of the process is the regeneration of the reagents.
Galvanox process
In the Galvanox process leaching of copper concentrates is conducted at 80°C under atmospheric conditions using ferric sulfate as oxidant in presence of pyrite. The process takes advantage of the galvanic couple between pyrite and chalcopyrite which accelerates the rate of leaching (Figure 7.8). Elemental sulfur is produced. However, to achieve this accelerating effect, large amounts of pyrite must be added which involves material handling problems and increased size of reactors.
192
Pressure Hydrometallur^
Chalcopyrite Pyrjte
Chapter 7 - Attempts to Avoid Autoclaves 193
4Fe^-
4Fe=-Figure 7.8 - The galvanic couple between pyrite and chalcop3TÍte Galvanox process
Cuprex process
The Cuprex process was developed in 1988 as a joint venture between Técnicas Reunidas in Spain, ICI in United Kingdom, and Nerco Minarais. It uses ICI's selective DS 5443 extractant and Técnicas Reunidas' Metchlor electrowinning cell. Severa! two-to three-week runs on a variety of copper concentrates were completed at Técnicas Reunidas's facilities in Madrid, and the process was considered ready for commercialization. The advantage of this new extractant is that it permits the recovery of copper from chloride médium without the necessity to transfer from chloride to sulfate.
In this process, copper concéntrate is leached in two stages with ferric chloride solution at 95°C and atmospheric pressure to produce a solution of cupric chloride, ferrous chloride, and elemental sulfur:
CuFeS^ + 4Fe3* -> Cu^^ + SFe^* + 2S
The leach residuo consists of gangue, pyrite and sulfur. The pregnant solution, is sent to the solvent extraction circuit where it is contacted with a kerosene solution of DS 5443. Using three extraction stages, copper in the aqueous phase is reduced to less than 0.5 g/L.
The loaded organic is scrubhed with spent anolyte from the elec-trowinning cell and then stripped by contacting with water at 65°C.
Three strip stages produce an aqueous copper chloride solution grading over 90 g/L Cu. The chloride ion content is then increased
by adding sodium chloride, which enhances conductivity and mini-mizes the possibility of the precipitation of cuprous chloride, and the solution is sent to the Metchlor electrolysis cell. The cell has two compartments separated by a catión exchange membrane. It uses a titanium cathode and an inert anode. Copper granules are deposited in the cathode compartment, and the electronic balance in the catholyte is maintained by transfer of sodium ions from the anolyte through the membrane.
Spent catholyte leaving the cell is a sodium chloride solution con-taining about 10 g/L Cu roughly divided between the cupric and cuprous oxidation states. This proceeds to the reforming stage where it is treated with chlorine gas from the anode compartment of the electrowinning cell to oxidize the cuprous ions to cupric. Cupric chloride is removed from the reformed catholyte in the depletion section by contacting it with copper-free organic from the stripping section of the solvent extraction. A two-stage depletion at a high 8:1 organic/aqueous ratio reduces copper in the spent catholyte to less than 0.1 g/L. The organic, containing relatively little copper, is recycled to the extraction unit and the aqueous raffinate becomes anolyte in the electrowinning cell.
Any silver in the original concéntrate reports in the raffinate from the solvent extraction stage and can be recovered by zinc dust pre-cipitation. Excess iron from the leaching of chalcopyrite is removed as goethite in a subsequent pressure oxidation step which simultane-ously regenerates part of the leachant. The remaining ferric chloride leachant is regenerated by chlorine from the electrowinning cell.
Zinc concentrates
In spite of the success of the aqueous oxidation process of zinc sul-fide in three operating plants, there are still attempts to avoid using autoclaves. For example, a plant under construction in San Luis Potosi in México will use a Finnish technology that uses four large
194 Pressure Hydrometallurgy Chapter 7 - Attempts to AvoidAutoclaves 195
reactors operating at 90°C instead of one autoclave (Figure 7.9).
Instead of two hours residence time in an autoclave the múltiple re-actors will be operating for 12 hours under continuous air injection.
The reaction taking place:
ZnS + 'ÁO^ + 2H^ ^ Zn2- + S + H p
It is believed that such operation cannot compete with pressure leaching.
OFF ON
. Ah
Motor Gas inlet
Gas hold-up
Measurement iocation
Motor I Gas inlet
Figure 7.9 -A reactor designed for leaching zinc sulfide concéntrate at 90°C.
Leñ: before introducing oxygen, right: after introducing oxygen
Albion process for gold
To avoid using autoclaves researchers proposed the Albion process for liberating gold from pyrite. The process is named after the sub-urb where it was developed. It involves fine grinding of ore and using oxygen in leaching at atmospheric pressure in conventional agitated tanks at 90°C. Oxygen is introduced by a special supersonic gas
>
jet. Special fine grinding equipment is used (Figure 7.10). As an approximate guide, it is expected 2-2.5% oxidation of the sulfide matrix per hour. To achieve 80% oxidation requires 34-40 hours for a typical refractory gold sulfide concéntrate at a grind size of 80%
passing 11 microns. This compares with 2-3 hours in an autoclave for a material ground to 80%) passing 44 microns.
Figure 7.10 -Albion fine grinding mili
The process was originally designed for zinc sulfide and extended to refractory gold ores. It is claimed that the capital cost of an Albion plant can be lower than a comparable pressure or bacterial leach, due to the simplicity of the process flowsheet. It should be noted however that the solubility of oxygen at 90°C at atmospheric pressure is low and no data are known regarding the filtration of the fine residue.