copper, nickel and tellurium yields during leaching of anode slime

2010:119 CIV

M A S T E R ' S T H E S I S

Copper, Nickel and Tellurium YieldsDuring Leaching of Anode Slime

Jonas Bäckström

Luleå University of Technology

MSc Programmes in Engineering Chemical Engineering

Department of Chemical Engineering and GeosciencesDivision of Process Metallurgy

2010:119 CIV - ISSN: 1402-1617 - ISRN: LTU-EX--10/119--SE

2010-06-27

COPPER, NICKEL AND TELLURIUM YIELDS DURING LEACHING OF ANODE SLIME

Jonas BäckströmRCP

Rönnskärsverken

Supervisors:

Kjellåke Johansson Boliden Mineral AB, Rönnskärsverken, R1Ä

Åke Sandström Luleå University of Technology, Division of Extractive Metallurgy

AbstractThis project has been performed as a thesis work in the final year of an MSc Programme in Chemical Engineering with specialisation in Minerals and Metallurgical Engineering. The work has been done at New Boliden's smelter, Rönnskärsverken, near Skelleftehamn in Sweden.

The aim of the project was to get a better understanding of the copper, nickel and tellurium leaching process at the precious metals plant in general. But also to find out if there are improvement potential in the yield of tellurium in particular. Some process parameters have been studied in both laboratory and full-scale production units. Some of the adjacent processes have also been involved in the investigations.

Complementary to the practical work a thorough study of available literature in the field of interest has been done.

The results of the laboratory leaching tests gave some clues, but to be able to draw certain conclusions more work has to be done. Combined with the literature survey, full-scale trials and analysis of statistics it is however possible to say that: Silver and tellurium are leached simultaneously after most of the copper and nickel have been leached. The raw slime composition and mineralogy have an impact on the tellurium yield. The tellurium yield is often low, possibly due to re-precipitation or formation of tellurium bearing, refractory slime components.

It would be possible to leach a substantial amount of copper by supplying air and sulphuric acid in e.g. a storage tank before the slime is fed to the autoclave.

A washing experiment with leached slime shows that the observed occasional high copper content of the leached slime more likely is due to an insufficient washing rather than lack of dissolution during the leaching.

i Master Thesis in Process Metallurgy

Keywords:Leaching, tellurium, anode slime, sulphuric acid, oxygen, silver.

Acknowledgements

I would like to thank all who have been involved in this project. In particular I would like to mention my supervisor at Boliden, Kjellåke Johansson for all ideas and discussions. Thanks also to my supervisor at LTU, Åke Sandström for valuable input during report writing. Thank you Anne Blomgren and Lars Göran Bodén for help with the practical work in the lab. Thanks to the personnel at R1Ä and R1R for help during sampling and full-scale experiments. Many thanks to Malin Utterström for help with literature in Rönnskär's library. Thanks to Marie Holmberg and the personnel at RCA and RCP for analyses and discussions during the project.

Jonas Bäckström ii

Swedish summary, Sammanfattning

KOPPAR-, NICKEL- OCH TELLUR-UTBYTEN VID LAKNING AV ANODSLAM

Detta projekt har genomförts som ett examensarbete under det sista året av Civilingenjörsprogrammet i Kemiteknisk design med inriktning mot mineralteknik och processmetallurgi. Arbetet har utförts vid New Bolidens smältverk, Rönnskärsverken, i Skelleftehamn.

Syftet med projektet var att få en bättre förståelse av koppar-, nickel- och tellur-lakningsprocessen vid ädelmetallverket i allmänhet. Men också att utreda om det finns förbättringspotential för tellurutbytet i synnerhet. Ett antal processparametrar har studerats dels i laboratorieskala men även i den fullskaliga processen. Några av de närliggande processerna har också inkluderats i undersökningarna.

Utöver det praktiska arbetet har också en detaljerad genomgång av tillgänglig litteratur inom intresseområdet gjorts.

Resultaten av laboratoriearbetet gav några ledtrådar, men för att kunna dra säkra slutsatser måste ytterligare undersökningar göras. Kombinerat med litteraturstudien, fullskaleförsöken samt analysen av historiska processdata är det hursomhelst möjligt att säga att: Silver och tellur börjar lakas simultant efter det att större delen av slammets koppar och nickelinnehåll har lakats. Sammansättningen och mineralogin av anodslammet har en signifikant påverkan på lakutbytet av tellur. Tellurutbutet påverkas negativt, möjligen av återutfällning eller bildandet av tellurinnehållande refraktära föreningar.

Det är möjligt att laka en signifikant mängd av slammets kopparinnehåll genom att tillföra syra och luft (syrgas) i exempelvis en förlagstank redan innan slammet pumpas till autoklaven.

Ett tvätt-test av redan lakat slam påvisade att de ibland observerade höga halterna av kvarvarande koppar i det lakade och torkade slammet mer sannolikt beror på bristfällig tvättning av filterpressen snarare än utebliven upplösning under lakningen.

iii Master Thesis in Process Metallurgy

Table of contents

Introduction 1

Aim and purpose 1

Background 3

New BOLIDEN 3

Rönnskärsverken 3

Precious metals plant 5

Theoretical background/literature survey 9

The metals in focus 9

Slime overall composition and formation 11

Mineralogy and physical properties 13

Leaching processes 17

Possible additions 22

Other slime treatment processes 24

Experimental methods 27

Laboratory scale trials 27

The experiments in the laboratory autoclave 28

Pre-leaching experiments 29

Full-Scale trials 32

Autoclave samplings 32

Wash tests 33Jonas Bäckström iv

Data analysis of historical process data 34

Results 35

Laboratory scale trials 35

Autoclave leaching 35

Pre-leaching 37

Full-scale trials 37

Sampling number 1 37



Washing tests 48

Statistical analysis of historical data 49

Multivariate 49

One variable at the time 53

Discussion 60

Conclusions 68

Suggestions of improvements 68

Suggestions for future studies 69

References 70

v Master Thesis in Process Metallurgy

Appendix 1 Anode slime constituents

Appendix 2 Precious metals plant, flow sheet

Appendix 3 Risk analysis report

Appendix 4 Laboratory autoclave, Users guide

Jonas Bäckström vi

IntroductionThis project was performed during the spring of 2010 and was done as a thesis work in the end of the author's education at LTU, Luleå University of Technology. The MSc Programme in Chemical Engineering with specialisation in Minerals and Metallurgical Engineering specifies that the project should correspond to 20 weeks work or 30 hp. Within the project, experiments at both laboratory and full-scale are used together with a literature survey and statistical analyses of historical data to clarify some properties of the leaching process investigated.

Aim and purpose

This project is conducted because of the sometimes very poor yield of tellurium (Figure 1) in the precious metals plant's copper, nickel and tellurium leaching circuit. The aim of the investigations is to get a better understanding of the leaching process, especially how the Te-leaching is affected by different parameters. The tellurium yield was better but not total when the process was newly installed [1]. Another investigation of the leaching process giving a 75% yield has been done but the processes and thereby the slimes have been altered since then [2]. The best case scenario would be a complete leaching of tellurium, copper, and nickel without solubilising silver or selenium. A poor yield in the leaching contributes to a higher impurity load and other problems in the proceeding slime smelting and silver electrolysis.

Jonas Bäckström 1

Figure 1, Tellurium yields, monthly values since 2003

Jan 2003 Jan 2004 Jan 2005 Jan 2006 Jan 2007 Jan 2008 Jan 2009 Jan 2010

Month

0

10

20

30

40

50

60Tellu

rium

yie

ld [%

]

2 Master Thesis in Process Metallurgy

Background

New BOLIDENThe Boliden corporate group is a Swedish company focused on mining and smelting. The group is active in six countries and have producing facilities in Sweden, Norway, Finland and Ireland. The major products are zinc, copper, lead, gold, silver and sulphuric acid. New Boliden have 4467 employees (end of 2009) and the turnover is in the order of 30 billion SEK. The total production levels of some refined metals and the biggest by product, sulphuric acid, are compiled in Table 1 below. Boliden is the eighth largest zinc producer in the world and the third largest producer of copper in Europe.

Zinc [tonnes]Copper [tonnes]Lead [tonnes]Gold [kg]Silver [kg]Sulphuric acid [tonnes]

434 022302 35551 57415 028

539 5641 123 336

Table 1, Boliden production levels 2009

Within the corporate group there are four mining areas and five smelters. Besides this, several areas are prospected for mineralizations. The project described in this report is conducted at Rönnskärsverken, one of the Boliden smelters. Rönnskärsverken is located near Skelleftehamn in the northern part of Sweden.

RönnskärsverkenRönnskärsverken is a smelter producing mainly copper but also lead, zinc clinker, silver, gold and other by products. The raw materials for the smelter mainly consist of concentrates, both from the Boliden Aitik mine and externally bought concentrates. Also electronic scrap, metallic scrap, lead concentrates and residues from other industries are used as raw materials. The main processes can be seen in the flow sheet in Figure 2 below.

The copper concentrates that holds approximately 28% copper is pyrometallurgically treated to produce a so called blister copper (ca 99% Cu) which is cast to anodes and hydrometallurgically refined to Jonas Bäckström 3

>99.998% pure cathode copper (LME Grade A). The first step of the pyrometallurgical treatment consists of two possible ways; either the Outokumpu flash furnace or the fluidized bed roaster and electric furnace are used to smelt the concentrates and produce a matte. Matte from the smelting furnaces is, batch wise, fed to converters whereupon the sulphur content is reduced in a slag blowing and a copper blowing step. The copper is deoxidised by ammonia addition in an anode furnace whereupon the copper is cast into anodes weighing about 343 kg each. The hydrometallurgical refining is then done by electrolysis in a tankhouse. Each tank holds 44 anodes and 43 cathodes. The anodes are changed in an 18 days (16 days at maximum capacity) cycle during which a total of three sets of cathodes are produced. The capacity of the refining plant amount to 240 000 tonnes yearly.

The anode scrap is sent back to the converters as cooling scrap after adhering slime have been washed off. The impurities in the anodes are either soluble or insoluble in the electrolyte during the electrolysis. The insoluble impurities forms a slime which falls to the tank bottom when the copper surrounding the impurity (which is either an inclusion or in solid solution in the copper matrix) is electrolytically dissolved. Soluble or partially soluble constituents like nickel, arsenic, antimony and bismuth are enriched in the electrolyte and a bleed stream of electrolyte is therefore necessary. Silver is slightly soluble in the electrolyte especially the silver that originates from the copper matrix will solubilise. Because of this fact, the main impurity in the cathodes is silver. The slime is preferentially deposited on the cell bottom and flushed out during anode switches. The slime is collected in a storage tank and periodically pumped to the precious metals plant for further treatment. The processes by which the slime is treated at the precious metals plant are described more thoroughly below.

Lead concentrates are also refined at Rönnskär; this is done at the Kaldo plant where the concentrate is fed to a Kaldo furnace (also known as TBRC or TROF). The molten lead is further treated in a refinery where impurities are removed in the form of drosses. This plant is also used campaign wise to smelt electronic scrap. The black copper produced during these campaigns is transferred to the converters in the copper circuit. During 2010 and 2011 Boliden 4 Master Thesis in Process Metallurgy

invests in a new Kaldo furnace devoted solely to electronic scrap smelting. When this new facility is up and running Rönnskär will be by far the largest WEEE recycler in the world.

The zinc content of all materials processed will end up in the fuming plant where the zinc is reduced and vaporised in a furnace/boiler. When the zinc fume cools the metal is re-oxidised and a zinc oxide product is achieved. The oxide product is further processed into a zinc clinker product in a kiln. The zinc clinker is not refined to metallic zinc at Rönnskär but is transferred to Boliden Odda, a zinc smelter in Norway.

Figure 2, Main processes and material flows at Rönnskär

Precious metals plantThe precious metals plant at Rönnskär produces silver, gold, copper telluride, selenium and a platinum/palladium concentrate. Some by-products are recirculated either within the precious metals plant or to other parts of Rönnskär. The production is run in batch mode and

Jonas Bäckström 5

each cycle starts with the main raw material, anode slime, being pumped from a storage tank at the copper electrolysis plant. The slime consists of a large variety of elements combined in many different chemical substances (see also Appendix 1). A typical chemical analysis of Rönnskär's slime is provided in Table 2 below. The slime arrives together with some electrolyte (as transport media) to a receiving tank. The first process in the slime treating circuit is to leach the slime in a H2SO4/O2 environment. The process is known as the "Outokumpu process". The goal with this treatment is to solubilise copper, nickel and tellurium but keep as much as possible of the other valuable slime constituents in solid state. From the receiving tank slime pulp is pumped through a filter from which the liquid phase, tankhouse electrolyte, is returned to the electrolyte circulating system while the filter cake is returned to the slime tank. In this manner a proper liquid to solids ratio of the slime is built up. The slime is transferred to an autoclave whereupon suitable amounts of sulphuric acid (~ 67%) and water are added. One leaching batch corresponds to approximately 4.5 tonnes of dry slime. The autoclave is sealed and the batch is heated with direct steam. When a temperature just above 100°C is reached, the oxygen supply is started. The slime is leached for approximately 4.5 hours, excluding the heating time. The leaching is stopped when a set maximum pressure (6.8 bars) or temperature (170°C) is reached. In order to further leach the slime the autoclave is usually flashed down to about 3-5 bars followed by a repeated oxygen intake. This procedure is repeated once or twice for every batch.

When the leaching is done the autoclave is flashed to atmospheric pressure and emptied into a cooling tank from which the slime is pumped to a filter press when a sufficiently low temperature, at about 80°C, is reached. The leachate is sent to a reactor where silver and selenium solubilised during the leaching procedure are reduced back into solid state. This is done by sulphur dioxide, SO2, addition while pumping the fluid through the filter press. The cementation is fine tuned by copper powder addition. The cemented silver is caught in the filter press and becomes a part of the proceeding batch of leached slime. The solution from the Ag/Se-cementation hopefully contains a large amount of the tellurium originally present in the slime. The tellurium content is also cemented with copper powder to form copper telluride, Cu2Te, which is sold as a product. The solution 6 Master Thesis in Process Metallurgy

which now is a sulphuric acid solution of copper and nickel is sent directly into the electrolyte circulating system of the tankhouse.

The leached slime that is recovered from the filter press is distributed on trays which are heated in drying ovens at 400°C for 12 hours in order to evaporate the remaining moisture. When the trays have cooled down, the slime is emptied together with slag formers like soda and additives into a mechanical mixer.

Raw anode slime analysis

Au[%]

Ag[%]

Cu[%]

Ni[%]

Pb[%]

Pd[%]

Pt[%]

Bi[%]

Sb[%]

As[%]

Se[%]

Te[%]

Average sample 2009

0.57822.9515.69 5.09 9.69 0.1030.0060.777 4.15 3.21 5.17 0.94

Table 2, Raw slime composition during 2009

The mixture, some recirculated materials and additives are charged to a small Kaldo furnace and smelted with oxygen and oil burners. A silver reduction step is then accomplished by coke addition where a precious metal smelt and a slag phase are produced. The slag from the first step, called A-slag, is withdrawn and recirculated to the copper or lead production line. The second step is a converting step where the metal smelt mainly consisting of silver is refined by charging sand as a slag former and then blasting the smelt with air through a lance. The so called B-slag is tapped, soda is charged after turning off the blast. The smelt is further heated with the burner and a second ladle of B-slag is tapped. The B-slag is recirculated to the smelting step. The smelt is transferred to an anode casting machine where silver anodes of 45 Kg are cast. The anodes are sent to the silver electrolysis department. For further info on the smelting process see [3].

A pure silver is produced by electrolysis in an AgNO3 electrolyte. The electrolysis is performed in cells similar to the so called Moebius cells, where silver crystals are formed on vertical cathode surfaces. The silver crystals are periodically scraped off and collected from the bottom of the cells, the silver is washed, melted and either cast into bars or granulated. The content of other precious metals is deposited as a slime in a filter cloth bag surrounding the silver anode.Jonas Bäckström 7

The silver electrolysis slime is first washed in several steps to remove residual nitrates. Then the filtered silver electrolysis slime is leached in a HCl/Cl2 environment to extract its content of gold and platinum metals. The leachate is treated to selectively precipitate the gold. This is done in two steps, the first step gives a pure gold sand ready for smelting and casting. The second step precipitate is not as pure as the first and therefore recirculated back to the leach tank and is leached with the proceeding batch of gold slime. The remaining solution holds the platinum metals and is treated in order to precipitate those giving a sellable palladium concentrate.

The selenium content of the Kaldo charge is virtually completely vaporised during smelting and the selenium is hence recovered in the gas cleaning system. The selenium is solubilised in the wet gas cleaning system whereupon the solution is taken care of and a pure selenium product is precipitated and filtered off.

Over the years the processes at the precious metals plant have been altered and modified [4]. The details of the plant's evolution will not be covered here, but the major modifications are described below.

In the mid eighties the existing gold refining electrolysis process was replaced by a wet chemical chlorination process. The improvement decreased the process time and gold inventory. At the same time a new gas cleaning system was introduced. Some years later the Outokumpu leaching process was introduced as a replacement of a soda roasting, selenium leach and copper leach process. At the same period of time the silver electrolysis was modernised and improved. The most recent large scale modification was done in 1997 when the Kaldo furnace was installed. This furnace replaced four process units, two roasting furnaces, a short rotary smelting furnace and the converting/refining furnace. The roasters are now instead used as drying ovens prior Kaldo charging. The furnace is built into an enclosure meaning that the furnace is not only giving rise to process improvements but also work environmental improvements and drastic emission reductions. A simplified process scheme of the precious metals plant at Rönnskär can be found in Appendix 2.


Theoretical background/literature surveyDuring the progress of this project a thorough investigation of the available literature in the field of interest was done. Literature describing different slime treatment methods, slime compositions, slime mineralogical characteristics and slime leaching mechanisms has been revised. Also some literature describing the basic chemistry and applications of copper, silver, tellurium and selenium have been reviewed. The knowledge found in the literature are summarised below.

The metals in focus

Copper, with atomic number 29 have a crustal abundance of about 60 ppm and is commonly mined as sulphidic or oxidic ores. The oxidic ores are predominantly treated hydrometallurgically while most plants treating sulphidic ores have a pyrometallurgical metal extraction followed by electrolytic refining. Copper have excellent electrical and heat conductive properties and is therefore often used in electrical wires and heat exchangers etc. Other applications are plumbing, roofing and coins.

Silver has atomic number 47, and a crustal abundance of about 0.08 ppm. Silver has excellent heat and electrical conductive properties, and is used in a wide range of applications such as jewellery, photography, electronics, anti bacterial materials and as investment metal. There are ores mined primarily for their gold and silver content but a substantial share of the world's silver and gold production is based on copper, zinc or lead refining by products.

Selenium and tellurium, atomic numbers 34 and 52 respectively, are among the rarest elements in the earth's crust. Their abundances are about 0.05 ppm for selenium and 0.01 ppm for tellurium respectively. Therefore they are in principal only produced as by products from copper electrorefining slimes.

Tellurium was discovered in 1782 by F.J. Mueller von Reichenstein in a Transylvanian gold ore. But the new mineral was not named until its discovery was confirmed several years later. The name tellurium

Jonas Bäckström 9

originates in the Latin word tellus meaning Earth. The element is in the same group as oxygen, sulphur, selenium and polonium in the periodic table. It resembles sulphur and selenium in many chemical characteristics but has more metallic properties. Tellurium is considered to be a metalloid in contrast to selenium, sulphur and oxygen which are considered as non-metals. The metalloid occurs in several minerals in nature often in association with Ag, Au, Bi, Pb, Sb, Cu and S. Some examples are Hessite (Ag2Te), Sylvanite (AuAgTe4), Richardite (Cu4Te3), Tellurobismuthite (Bi2Te3) and Tellurite (TeO2). There are no mineralization worth mining solely for its tellurium content but sulphidic copper and nickel ores are often tellurium bearing. One disadvantage of the minerals from tellurium extraction point of view is that usually 90% of the tellurium is lost already in the copper ore beneficiation, i.e. the sulphide flotation. The name selenium originates in the Greek word σεληνη, meaning moon. [5, 6, 7, 8, 9, 10]

In the processes considered in this project the dissolved species of tellurium most probably are of either Te(IV) or Te(VI), meaning tellurous acid H2TeO3 or orthotelluric acid, H6TeO6. The orthotelluric acid is sometimes denoted as either TeO3•3H2O or Te(OH)6, but H6TeO6 seems to be the most correct formula. The orthotelluric acid is soluble in acidic media, while the tellurous only is sparingly soluble. [5, 6, 7, 10, 11, 12]

The means of which selenium and tellurium can be separated are somewhat limited because of the substances many similarities. [6, 11, 13, 14, 15, 16].There are however some possibilities of which a couple are mentioned here. The elements can be pyrometallurgically separated by having the selenium and tellurium dioxides exposed to a temperature of more than 500°C. Selenium is under these conditions much more volatile and can thereby be removed to a large extent. The tellurium requires considerably higher temperatures to be gasified. To produce ultra high purity products more advanced methods are required.

Tellurium and selenium exists in four valence states (-II), (0), (IV) and (VI). This can be used to hydrometallurgically separate the elements, due to the differences in solubility at different pH. If both selenium 10 Master Thesis in Process Metallurgy

and tellurium are present as sodium salts of their negative divalent states they can be separated in an alkaline, reducing environment. These circumstances cause selenium to go into solution and tellurium to precipitate as particles of elemental tellurium, Te0. In their positive tetravalent states tellurium has a solubility minimum at pH 5.6 while selenium is soluble in the entire pH range. Tellurium in hexavalent state is only sparingly soluble in alkaline environments while selenium on the other hand, both tetravalent and hexavalent, are soluble in alkaline media. [6, 14]

Slime overall composition and formation

The anode slime composition does, obviously, vary from plant to plant depending on their processes and raw materials i.e. ore composition, secondary materials used etc. But the overall composition of the slime and the mineralogy might also change drastically within the same plant due to relatively small fluctuations in the raw materials and process disturbances. Most of the anode slimes from refineries around the world have compositions in the approximate ranges given in Table 3 below. Apart form the valuable elements listed in this table the slimes also contain substantial amounts of Ba, Si, O and S as well as traces of Cl, Sn, Fe and organic compounds. [17, 18, 19, 20, 21, 22, 23, 24, 25]

The main differences in compositions affecting the treatment methods chosen at different plants around the world are nickel and lead. These elements might in some cases require additional removal steps prior to slime treatment if the levels are considerably high.

Jonas Bäckström 11

Raw anode slime analysis

Typical slime composition range

Rönnskär 2009 sample

Au [%]Ag [%]Cu [%]Ni [%]Pb [%]Pd [%]Pt [%]Bi [%]Sb [%]As [%]Se [%]Te [%]

0–1 0.5785–20 22.95

10–30 15.690–20 5.093–20 9.690–0.2 0.1030–0.1 0.006

0.1–0.7 0.7770–8 4.15

0.5–5 3.212–15 5.170.3–3 0.94

Table 3, Typical overall composition of anode slime and Rönnskär 2009 average sample

The slime is continuously formed during the electrolytic refining of copper. Insoluble anode constituents falls to the bottom of the tank as well as substantial amounts of baryte, BaSO4 which is used as a lubricating mould wash during anode casting. Some elements like silver, nickel, arsenic, bismuth and antimony are partially soluble in the electrolyte and are distributed between the slime and the electrolyte. Some of the anode impurities are chemically altered during contact with the electrolyte whereas some are rather unchanged even morphologically. Other slime constituents are directly precipitated from the electrolyte solution.

Silver in the anodes can mainly be found either in metastable solid solution in the copper matrix or in impurity inclusions at the grain boundaries. Much of the silver in solid solution will be dissolved during electrolysis. This silver is to a large extent cemented as a fine grained slime due to the presence of elemental copper particles or copper(I)-ions originating from dissolution of copper oxide. These processes can be described by the following reactions [26, 27, 28],

(1) Ag+ + Cu0 ⇒ Ag0 + Cu+ (2) Cu2O + 2H+ ⇒ 2Cu+ + H2O ,


(3) Ag+ + Cu+ ⇒ Ag0 + Cu2+ .

But the electrolyte's silver content can also be lowered by substitution in copper/silver selenides in the slime layer,

(4) AgxCuySe + zAg+ ⇒ Agx+zCuy-zSe + zCu+ .

The As, Sb and Bi content of the anode can in many instances form a so called floating slime at the electrolyte to atmosphere interface. This slime will also sink to the bottom of the tanks and become a part of the slime layer.

A large amount of the copper content of the slime does not originate from refractory copper compounds but rather anode fragments, shortcuts (dendrites) and crystalline copper sulphate. The copper sulphate content may be high if the electrolyte temperature of some reason has dropped significantly and thereby sulphates have been precipitated due to decreased solubility. Copper sulphate is also present as a co-precipitate in agglomerates with other slime particles.

Mineralogy and physical properties

Several articles by among others Chen and Dutrizac describes the mineralogical composition of copper anode impurities and anode slimes. The mixture of components is often complex and highly variable. A compilation of most of the possible slime constituents found in the literature can be seen in Appendix 1. Some major constituents are though more or less abundant in all similar slimes. The anode slime is, apart from the electrolyte components, mostly a mixture of metallic silver and copper, nickel oxide, copper sulphate, lead sulphate, baryte, floating slime, a complex oxidate phase and various copper/silver-selenides/tellurides. The so called oxidate phase is described by Chen and Dutrizac [26, 27, 29, 30, 31] as a "Cu-Ag-Pb-Au-sulphate-arsenate-antimonate-selenite-tellurite-material". The material seems to be an agglomerate of very small particles of highly variable composition. [26, 27, 28, 29, 30, 31, 32, 33, 34, 35, 36, 37, 38, 39]


The slime has often very small particle size ranges with the smallest particles well below one µm. This size corresponds well to the size range of the impurity inclusions at the grain boundaries in the copper anodes. Some slime constituents are though larger, for example baryte particles have been seen to agglomerate quite extensively. The metallic copper is to a large extent made up of anode fragments and dendrites which can be up to sizes of several centimetres but at the Rönnskär electrolysis plant an 8 mm sieve disc catches the largest pieces before the slime reaches the storage tank. The particle sizes of slimes have been investigated by several researchers [40, 41, 42] and the results corresponds well to the statements made above. Most of the slime particles are in the size range up to about 20 or 30 µm. Only a very small fraction is above 150 µm and this fraction consists mostly of baryte and copper.

The particle shapes varies a lot between different constituents; some are spheroidal or ring shaped while others are more irregular or needle shaped. Some of the particles are euhedral and some appear to have a less well defined morphology and are more amorphous. The selenides in the cast anodes are often spherical or ring shaped. The selenides often retain their shape during the refining and the round shapes can often be found in the anode slimes. The selenide particles are often not homogeneous but have cores of Cu-Bi-As-oxide, lead oxides or lead sulphates. Naturally the baryte agglomerate and the anode fragments are highly irregular in shape.

The selenium present in the anodes is in the form of copper selenides i.e. Cu2Se. Although the selenides retain their morphology during electrolysis the composition changes. The copper in the selenides are progressively substituted by silver ions present in the electrolyte. The majority of these reactions occur within a few hours of contact with the electrolyte. This results in a cell bottom slime containing mostly AgCuSe and Ag2Se. The overall process described by reaction (4) really occurs in several steps which are completed to different extents. The step wise reactions giving this compositional change are:


(5) Cu2(Se,Te) + xAg+ ⇒ Cu2-xAgx(Se,Te) + xCu+ ,(6) Cu2-xAgx(Se,Te) + (1-x)Ag+ ⇒ AgCu(Se,Te) + (1-x)Cu+ ,(7) AgCu(Se,Te) + (1-x)Ag+ ⇒ Ag2-xCux(Se,Te) + (1-x)Cu+ ,(8) Ag2-xCux(Se,Te) + xAg+ ⇒ Ag2(Se,Te) + xCu+ .

This successive change is also evident from differences in slime analysis of cell bottom slime versus slime collected from the corroding anode. According to Scott [37] the selenide also undergo a minor dissolution on the tank bottom, the dissolved selenium is supposedly immediately precipitated in the heterogeneous oxidate phase.

The majority of the anode's silver content is in solid solution in the copper matrix but metallic silver has also been found as thin rims around the spheroidal selenide inclusions [43]. The silver that are in solid solution in the copper matrix of the anode must undergo a dissolution during electrolysis and will thereafter react with other slime or electrolyte components such as the selenides, elemental copper, copper or silver ions. The metallic silver found in some slime samples is often present as tiny particles in larger clusters or agglomerates. According to an investigation by Hiskey and Campin [28] the elemental silver content in the anode slimes corresponds to 13-15% of the total silver content. The remainder of the silver content is mainly in the form of selenides, whose presence has been confirmed by XRD analysis in the same research. The tiny silver particles are produced according to the reactions (1) and (3) mentioned above. The elemental silver particles might also undergo some dissolution in the slimes layer and thereby contribute to the silver enrichment of the selenides. Although elemental silver is abundant in the slimes the selenides are the main silver carrier.

The main part of the tellurium found in raw anode slimes is in solid solution in the copper-silver-selenides. In addition the oxidate phase is a minor tellurium carrier. Traces of tellurium are found in several other slime components but they are overall insignificant as tellurium carriers. In the anode itself the selenide/telluride components are the single most important tellurium bearing compound although some tellurium seems to be in a metastable solid solution in the copper


matrix. Despite the selenides shift in silver content during refining the morphology and degree of tellurium substitution remains essentially the same. At least if the tellurium content of the anode is moderate compared to its selenium content. Some dissolution of tellurium and an increase in the Se/Te ratio might occur but only to a lesser extent. In other cases, when the tellurium content is high, the tellurium forms a porous mass of silver telluride. [26, 29, 44]

The lead content of the anode is mainly found in oxidic inclusions although some lead is present in solid solution in the copper matrix. Lead bearing phases in the raw slime is mainly lead sulphate and similar compounds. The lead sulphate phase is commonly found either as euhedral crystal clusters or as a core in selenide particles. The lead sulphate cores in the selenides originates from oxidic lead bearing phases present in the selenide's cores in the anode inclusions. In some selenide particles where the selenide completely covers the core of Cu-Pb-oxide, Cu-Pb-As-oxide or Cu-Pb-As-Sb-Bi-oxide the particle may pass into the slime with unaffected core. Minor amounts of lead can also be found in oxides and arsenates precipitated during electrorefining.[26, 43, 45]

Anodes often contain significant amounts of arsenic, antimony and bismuth. These elements are generally found in different complex oxides with copper and lead. Antimony will form Cu-Ni-Sb-oxide (kupferglimmer) inclusions if the anode also has a high level of nickel.

The nickel present in the anode slime is found in different configurations depending on the nickel, antimony and oxygen content of the anodes. The main nickel carrying components are the copper matrix, NiO and kupferglimmer. Anodes containing high levels of both antimony (>200 ppm) and nickel (>0.3%) have significant amounts of kupferglimmer present as impurity inclusions. In this case also the cooling rate in the anode casting might have a significant effect on the amount of formed kupferglimmer. Also NiO is present and its formation is logically favoured by high nickel and oxygen levels. When the nickel content of the anodes exceeds 0.25% the oxide is formed, while a lower content permits most nickel to be in solid solution in the copper matrix. The nickel compounds formed in the anodes during solidification are as said strongly dependent on both 16 Master Thesis in Process Metallurgy

nickel and dissolved oxygen content but the amount of iron and antimony also affects the formation of some inclusion substances. During electrolysis the compounds in which nickel occurs do have an impact on the distribution of nickel between the slime and the electrolyte. Nickel oxide, NiFe2O4, Cu-Fe-Ni-oxide, Cu-Ni-oxide, Cu-Sn-Fe-Ni-Zn-oxide, Cu-(Sn,Sb)-(Ni,Fe)-oxide and Cu-Ni-silicates have been found to be refractory during electrolysis. Nickel in solid solution dissolves completely but re-precipitates partly. The kupferglimmer phase is partly refractory. [43, 46, 47, 48]

The gold content of the anode and its slimes are mostly not high enough to make it possible to draw any certain conclusions from electron-microscopic and X-ray investigations. Some attempts to study the gold occurrence in anodes, raw slimes as well as leached slimes by Mössbauer spectroscopy have been done by Sawicki et al. [49]. They found that gold in anodes are in a dilute copper alloy. In the raw slime the gold content was distributed into two phases. One was fischesserite, Ag3AuSe2, or possibly a similar compound. The other phase was a gold-silver, gold-copper or possibly a ternary gold-silver-copper alloy. The leached slimes indicated similar compounds but in the slime leached under the most severe conditions all the gold seems to have been converted into the selenide phase. This suggests that gold is dissolved and reacted with selenide particles during high temperature H2SO4 leaching. [49, 50]

Leaching processes

Some patents that describe varieties of the used leaching method exist. Claessens and White [51, 52] describe a method of leaching at 10 bars pressure and 100-200°C. In the patent documents they claim to be able to leach virtually all tellurium present. This process configuration is similar to the one applied at Rönnskär the main difference was that a more dilute acid was used. This process variety has previously been tested at Rönnskär without any significant difference. The most probable reason for the differences is that the slime composition and mineralogy differed between Rönnskär and Noranda where the process was developed. A similar method is patented by Morrison et al. [53].


Also Baboudjian et al. [54] describes a leaching method, in this case it is a two step process in which nickel is solubilised anaerobically at 160°C and 5 bars in a first step. The second step aims at aerobically solubilise tellurium and copper at a lower temperature, 120°C, and an oxygen pressure of 3 bars. They managed to solubilise 66% of the tellurium without any silver or selenium being leached. No solid liquid separation is performed between the two steps.

The leaching procedure itself alters both the overall chemical analysis of the slime and the composition and occurrence of several compounds in the slime. Prior to leaching the most abundant slime constituents are: metallic silver and copper, nickel oxide, copper sulphate, lead sulphate, baryte, floating slime, the complex oxidate phase and various copper/silver-selenides/tellurides. In the leaching procedure sulphuric acid and oxygen is added which causes oxidation and dissolution of some elements. The reactions and processes occurring in the H2SO4/O2 leaching process are studied and described in some articles [23, 25, 30, 33, 35, 36, 55, 56, 57, 58, 59]. The main objective, as previously stated, is to remove as much as possible of copper, nickel and tellurium while avoiding extensive dissolution of silver and selenium. The oxidation potentials for the main elements of interest are in decreasing order Au >Ag >Se >Te >Cu. If only the oxidation potential is considered it would be possible to dissolve all copper (and nickel) without affecting silver. But due to reaction kinetics, possible formation of refractory compounds in the electrolysis and re-precipitation during leaching and cooling this is not the case in real processes.

The copper content is readily dissolved and would be so even under milder conditions, since merely the tankhouse electrolyte have been used to leach copper from slimes. The nickel requires a higher temperature in order to go into solution, whereas it does not require oxygen since only acid is required to dissolve the nickel oxide.

Copper present in the slime is more or less completely leached and the reaction mechanisms for some major copper bearing substances are:


(9) Cu + 0.5O2 + H2SO4 ⇒ CuSO4 + H2O ,(10) Cu2O + 0.5O2 + 2H2SO4 ⇒ 2CuSO4 + 2H2O ,(11) Cu2Se + O2 + 2H2SO4 ⇒ 2CuSO4 + Se0 + 2H2O ,(12) 2CuAgSe + O2 + 2H2SO4 ⇒ 2CuSO4 + Ag2Se + Se0 + 2H2O .

The reactions listed above are all contributing to the heating of the batch since they are all exothermic under the given conditions. The end products might be inferred by the formulas but other forms are also possible by further reaction or varieties of the mechanisms. The nickel present as nickel oxide is dissolved in the acid at elevated temperatures,

(13) NiO + H2SO4 ⇒ NiSO4 + H2O .

Also nickel present as kupferglimmer is slowly attacked during autoclave leaching.

Arsenic might form arsenic acid through oxidative dissolution. Antimony is oxidised to its pentavalent state and form a sparingly soluble oxide specie. Bismuth on the other hand is not as easily oxidised and is therefore not extensively dissolved. The bismuth often occurs as arsenate, BiAsO4. But if arsenic is absent a larger portion of bismuth might go into solution under high pressure and temperature. Some dissolved bismuth may during cooling re-precipitate as sulphate or oxysulphate. [36]

The tellurium present as copper-silver selenide-tellurides can dissolve according to the overall formulas:

(14) Cu2Te + 2H2SO4 + 2O2 ⇒ 2CuSO4 + H2TeO3 + H2O ,(15) Cu2Te + 2H2SO4 + 5/2O2 +H2O ⇒ 2CuSO4 + H6TeO6 .

Most probably a mixture of (hexavalent) orthotelluric acid and (tetravalent) telluric acid will form. Some literature refers to the hexavalent tellurium as the metatelluric acid H2TeO4 which other sources claim to be unknown other than in the form of a polymer sometimes denoted as (H2TeO4)n (where n≈10). The other form of Jonas Bäckström 19

hexavalent telluric acid is the orthotelluric acid, H6TeO6 which seems to be more commonly described in literature although the formula is sometimes written differently as Te(OH)6 or TeO3•3H2O. The form of the tellurium compound will probably affect the yield of tellurium as well as the further processing of the solution. The hexavalent form would be preferable since it has a higher solubility in the acid media.[5, 6, 7, 9, 10]

Some articles suggest a couple of explanations for the sometimes poor tellurium yields during autoclave leaching. Chen and Dutrizac [29, 30, 32, 34] have observed indications of re-precipitation of tellurium during leaching. According to their results most of the tellurium might have been leached at some point during the treatment but later re-precipitated as other compounds. The dissolved tellurium reacts with Ag, Au and Se and forms a tellurium bearing silver-gold-selenium phase with a suggested formula of Ag3Au(Se,Te)2. Some of the dissolved tellurium also reacts with Sb, Bi and As under formation of various oxide compounds. Chen and Durtrizac also observed that formation of more silver or gold rich copper/silver-selenide/telluride compounds retards the dissolution of tellurium. The re-precipitation of tellurium has also been observed by others [23, 33]. Jennings et al. [23] have also observed this kind of re-precipitation, in their experiments the tellurium yield markedly decreased at temperatures above 135-160°C. Silver ions are present in the solution after leaching and might also contribute to the precipitation of tellurium. In fact silver ions have been suggested [60, 61] as titration reagent for tellurite and tellurate anions. The ions precipitate as silver salts. Although the experiments were done in alkaline media and the solubility are probably pH dependent.

Hoffmann [36] suggests another condition contributing to a limitation of the tellurium content of the slime. Some tellurium occurs in solid solution in metallic silver particles which (preferably?) do not dissolve completely during slime leaching. Hence this part of the tellurium content is locked into the silver particles.

The tellurium that re-precipitates in the oxidate phase formed during leaching is possibly precipitated as a Ag-Pb selenite-antimonate-tellurate. The original selenide phases are according to some


investigations to some extent depleted of its tellurium content during leaching. [30]

Selenium is preferentially present as selenides which are in principal not dissolved during the leaching process. However, if the leach is not stopped in time some selenium might end up in solution. This selenium is present in its tetravalent state, more severe conditions are required in order to further oxidise it to hexavalent selenium. The selenides are whatsoever not unaffected by the leaching. The selenides are close to completely depleted for their copper content and the degree of tellurium substitution might change. The selenides are although morphologically stable and despite the intense treatment some particles manage to retain their spheroidal shape gained during anode casting. [30, 36] The principle of dissolution is

(16) Cu2Se + 2H2SO4 + 2O2 ⇒ 2CuSO4 + H2SeO3 + H2O .

The amount of selenium in solution may originate from a reaction mechanism similar to reaction (11) and reaction (12). But in this case selenite ions are formed instead of the elemental selenium

(17) 2AgCuSe + 2H+ +2O2 ⇒ Ag2Se +2Cu2+ + SeO32- + H2O .

Silver in the slimes is generally present in elemental form or as selenides, an extensive dissolution of silver selenides indicates that the leaching has gone too far. The metallic silver dissolves to some extent and might precipitate as silver sulphate and become a part of the post leach oxidate phase. The solubility of silver sulphate is strongly dependent on temperature (see Figure 3). Hence, the final silver concentration in the leachate is also dependent upon the filtration temperature and the residual acid concentration rather than only the extent of dissolution.


Figure 3, Solubility of silver sulphate as function of temperature and sulphuric acid concentration, diagram made with data from Linke, W.F. [62]

The solid residue after the pressure leaching process is generally a mixture of silver sulphate, elemental silver, lead sulphate, baryte, silver selenide, elemental selenium, a new precipitated oxidate phase and different complex oxides. [30, 34, 38]

Selenium and tellurium present in the solution after the leach is commonly precipitated by e.g. copper powder additions. Selenium precipitation may also be done by SO2 gas, this method has also been tested on tellurium precipitation [63]. Some varieties of the process exist but the overall theme is similar. [64]

Possible additionsIn several articles and patents, processes involving additions of other oxidising agents or reagents used in leaching are described. In an article by Schloen and Elkin [17] dated as far back as 1950 several possible additions and leaching operations are mentioned. As an example, a method of adding ferric sulphate to the slurry in the end

0 50 100 150 200 250

Temperature [°C]

0

50

100

150

200

Solu

bili

ty o

f A

g!S

O" [g

/l]

1,0 M H!SO"

0,5 M H!SO"

0,1 M H!SO"

Water


of an atmospheric leaching is mentioned. This addition aids the leach by reducing Fe(III) to Fe(II) while the copper present in the slime is oxidised. Another method discussed is leaching in ammonium hydroxide and ammonium carbonate, but no advantage over sulphuric acid is found. Leaching in cyanides or using MnO2 as an oxidant is found to be expensive and also introducing a contaminant into the electrolyte. Another article by Hait et al. [65, 66] also describes MnO2 as a possible leaching reagent. They found that an addition of MnO2 vas beneficial for the leaching of copper, selenium and tellurium. A combination of MnO2 and NaCl was also investigated giving even higher recoveries but also extensive leaching of gold. They have also by means of SEM and XRD techniques done a characterisation of the slime mineralogy before and after leaching. Chen and Dutrizac [30] have also studied slime leached with NaCl additions and the major difference in the leach residue was found to be that silver mainly occurs as silver chloride instead of sulphates.

There are other factors to consider when dealing with these kinds of additions, there is an imminent risk of chlorine gas evolution which is not only highly toxic but also corrosive on the process equipment. The chlorine gas is probably consumed in leaching reactions but the risk of chlorine gas leakage must be considered. Another disadvantage is the possible co extraction of gold and other PGM:s which in this stage is disastrous for the yield of precious metals. The addition of a substance as a leaching aid might also give difficulties and require removal steps in the further treatment of the leach products.

The usage of ozone in the leaching process has been mentioned by Dönmez et al. [25]. This would probably contribute to a more aggressive leaching but would also be more expensive and demanding considering process equipment. Using hydrogen peroxide instead of oxygen gas would logically give a better contact between the oxidant and the slime particles. No slime treatment process using hydrogen peroxide has been found and hopefully the stirring in the autoclave is sufficient to give optimal oxygen mass transfer into the solution.


Another possible addition is described by Stafiej et al. and Decker et al. in [67 and 68]. The method suggests that the leaching of tellurium could be improved while a minimum of silver is leached. The reagent used is selenium (or sulphur), the idea is that addition of selenium would help to stabilise silver in the form of silver selenide as well as favouring a selective extraction of the tellurium. The aim is to maintain the silver/selenium ratio at ≲ 2 to give the desired conditions. The leaching is otherwise conducted in a similar way as normal. The silver to selenium ratio in the slime might affect the silver content of the selenides in the slime [69] and thereby also their leachability.

Other slime treatment processes

Around the world, slimes are and have been treated in other processes than the scheme applied at Rönnskär. Common process solutions are different combinations of hydrometallurgical and/or pyrometallurgical operations. Many researchers have done comparisons and descriptions of the different methods.[4, 14, 19, 20, 21, 18, 55] Below follows a brief explanation of the main unit processes in the different process alternatives.

Direct smelting of the raw anode slime was one of the first methods applied. This method is however not used today since the copper content of the slime has to be really low if problems are to be avoided in the smelting and proceeding silver refining processes.

Pyrometallurgical pre-treatment before hydrometallurgical processes can be done in a number of ways. Sulphatising roast is one common process, which means that the dewatered slime is heated during contact with concentrated H2SO4 in presence of air (as oxidant). This causes sulphatising of the base metals and oxidising of the selenium and tellurium as well as volatilisation of the selenium content. The sulphuric acid is regenerated and elemental selenium is recovered in a wet gas cleaning system. [59] Oxidising roast on the other hand transforms copper, nickel and tellurium into acid soluble oxide compounds whereas the selenium once again is volatilised and recovered from the gas phase. The oxidising roast can also be conducted as a separate deselenizing treatment after an initial sulphuric acid leach [55, 58]. The process can be performed in e.g. a 24 Master Thesis in Process Metallurgy

fluidized bed but in other cases the slime can be extruded, pelletized or pressed into briquettes to simplify the oxidising treatment in other furnace types. These two roasting procedures are commonly followed by some hydrometallurgical process.

Soda ash roasting is a process where the slime is heated and reacted with sodium carbonate in order to transform selenium and tellurium into their hexavalent states. The roast is performed after a decopperizing acid leach. The slime is dewatered and mixed with carbonate to a paste which is agglomerated as pellets or extruded before the low temperature roasting at 530-650°C. The selenium can thereafter be dissolved in an alkaline water leach. This procedure offers a method of separating selenium and tellurium since the hexavalent tellurium is insoluble in strong alkaline solutions.

Another process mentioned in the literature [55, 70] is oxidising roast with potassium nitrate which should convert the tellurium, selenium, antimony and arsenic into potassium salts. The reaction is conducted continuously in a kiln at 400-450°C. The slime should then be leached in sulphuric acid or sodium hydroxide (if the Te content is low) to extract these elements.

Hydrometallurgical pre-treatment followed by pyrometallurgical processes in the sense of sulphuric acid leaching followed by smelting is the worldwide dominating configuration of slime treatment processes. The so called Outokumpu process (the process applied at Rönnskär) is one of several varieties of this theme. Other varieties are used at Norddeutsche Affinerie, Canadian Copper Refiners, Hitachi Smelter and Refinery as well as Indian Copper Complex. The process variety at Indian Copper Complex is divided into two leaching steps, one dilute acid leach and one concentrated acid leach. The first step removes copper whereas the second step solubilises most of the nickel. The leaching at Hitachi Smelter is a ferric leaching where copper is solubilised with aid of added ferric ions. Then the copper is won by electrolysis and the ferric ions simultaneously regenerated. An obvious disadvantage with this setup is that the leach solution cannot be sent back to the copper electrolysis plant but must be treated in a separate circuit. This is due to the ferric and ferrous ions ability to significantly decrease the current yield during copper electrolysis. [22, 55]Jonas Bäckström 25

A pyro-hydro-pyrometallurgical process is described [19, 55] as an efficient way of separating the slime constituents in high grade fractions at high recovery. The suggested process involves an ammoniacal leaching to extract copper, selenium and some other by-products. The leachate is distilled and the solution returned whereas the residue is once again leached, this time in hydrochloric acid. This leachate is reduced by sulphur dioxide to extract selenium and the filtrate can thereafter be treated to recover copper and other by-products. The leach residue from the ammoniacal leaching is soda ash roasted and deselenized by water leaching. The selenium containing solution is added to the selenium reduction treatment. The deselenized slime is then once again leached in hydrochloric acid to extract the base metals i.e. lead. The remaining residue is now a noble metals concentrate. In the initial ammonia leach >96% Cu, >65% Se and <1% Ag have been found to be extracted. This process was developed for high lead but low nickel slimes.

Apart from the processes mentioned above some different approaches to purely hydrometallurgical processes exists. One of them, the chlorination process is performed by treating the slime with a chlorine bearing reagent and oxidants which converts most of the metals to its chlorides. Both dry and wet chlorination processes exist but merely the wet process have been widely accepted as an industrial application. Some constituents go into solution while some remain in the solid residue. The way of which the elements are further separated differs between different plants but several steps in quite complex schemes are required. [20, 55] Other purely hydrometallurgical processes exist. In these processes quite complex schemes are required to separate all valuable slime components.

Some processes where the tellurium content is left in the slime during smelting comprise instead a tellurium leaching step of the slag produced. The leaching is generally performed at high pH and an intermediate product of sodium tellurate or tellurium dioxide is produced. [12, 71] The slag leaching processes mentioned above seems to have quite high tellurium yields in general.


Experimental methodsLaboratory scale trials

The laboratory trials performed within this project was mainly done in a laboratory scale autoclave. The autoclave is equipped with a heating jacket, gas inlet, thermocouple, pressure gauge, flash vent and a safety rupture disc. The work with the autoclave was preceded by a risk analysis as well as a thorough study of instructions and the available documentation [72, 73]. This study resulted in a written users guide with a description of the autoclave, its functions and some tips for future users. See also Appendix 3 and 4 (written in Swedish). Much work have also been done to restore the autoclave itself, driving belts have been replaced and other parts have been looked over etc.

The oxygen gas purged into the autoclave was supplied from a gas tube and the gas usage was approximated by weighing the tube. The laboratory autoclave setup is illustrated in Figure 4. After an initial test run three autoclave leaching experiments was done in an attempt to replicate the full-scale production unit. These experiments would also give an idea of the repeatability and the possible difficulties with the practical work with the autoclave.

Figure 4, Laboratory autoclave setup [73]Jonas Bäckström 27

The slimes used in the experiments were sampled directly from the copper electrolysis tanks. This procedure gives a better correlation of the slime back to the analysis of the anodes. i.e. mixing of slimes with slightly different compositions is avoided. Directly after sampling the slime was slightly diluted in order to avoid copper sulphate crystallisation during cooling. The bulk sample was thereafter divided into smaller samples by alternately pouring the slime into two different containers. The procedure was repeated until small samples corresponding to approximately 100 g of dry slime were obtained.

When an experiment was finished the leachate and the residue was sent to Rönnskär's laboratory for analysis. The samples was either analysed chemically or by XRF spectrometry, the latter only giving approximate analyses. The slime used in the autoclave trials was analysed in two different reference samples.

The experiments in the laboratory autoclaveAutoclave leaching, run: A1The general purpose of this experiment was to see if it was possible to replicate the full-scale leaching process in the laboratory unit as well as testing methodology and the autoclave functions.

Autoclave leaching, run: A2, A3 and A4An attempt to run three reference points under the same conditions was done. In each of these autoclave leaching tests a slime sample corresponding to about 300 g dry mass was used. The slime was mixed with 350 ml (1000 g/l) sulphuric acid and heated to approximately 115°C before oxygen addition was started. The temperature was successively risen to about 170°C and the stirrer was set at 300 rpm during the whole process. The batch had a volume of 1.3 litre including acid, electrolyte and slime. When the autoclave was sealed the heater and stirrer was started and after 45 minutes the desired start temperature was achieved. The oxygen regulator was set at 8 bars and the inlet valve was slowly opened. Oxygen was supplied for approximately 2 to 2.5 hours, and then the autoclave was flashed and opened. The sample was immediately diluted and cooled with 500 ml of water to enable safe handling. The slime was filtered in a Büchner funnel and the solid residue was dried at <100°C, all samples were weighed and the volume of the diluted


leachate was measured. The leachate was analysed for Ag, Se and Te while the solid residue was analysed for Ag, Cu, Ni, Se and Te.The oxygen consumption was estimated by weighing the gas tube. The purpose of these three trials was to determine the repeatability of the experiments and to find difficulties with the method.

Autoclave leaching, run: A5An experiment with milder leaching conditions but a longer run time (~ 4 hours) and a lower amount of solids was tested. This time 200 g slime was added to 400 ml acid and an equal amount of electrolyte as in the previous tests. The gas regulator was set to 6.5 bars and the maximum temperature was maintained at 130°C. The reaction products were treated as in the previous tests.

Autoclave leaching, run: A6Next trial was performed under even milder conditions regarding the operating temperature. The slurry was prepared as in the first experiments and the gas pressure was set to 8 bars. The temperature was set at 80°C and the time was at about 2.5 hours. Also in this case the products was collected and analysed in the same way.

Autoclave leaching, run: A7In the last autoclave run an addition of selenium was made to give an approximate Ag/Se ratio of 2. The aim was to test the possibilities of stabilising silver and extract more tellurium according the theory described earlier in the section covering leaching additions. All other parameters were similar to the three reference points. The leachate and the residue were treated as previously.

Pre-leaching experimentsIn addition to the autoclave leaching tests a setup was made to investigate the possibilities of a pre-leach of the slime before autoclave filling. The possibilities in the full-scale production unit is to aerate (or possibly oxygenate) the slime during storage in the slime tank. Other plausible ideas are dilution and acid additions. To be able to test the effects of such process alterations a laboratory setup was done. The setup consisted of a reaction vessel on hotplate with a magnetic stirrer. The vessel was equipped with a contact thermometer and a gas sparger close to the bottom. The reactor was Jonas Bäckström 29

sealed apart from a condenser which was cooled in order to condense and recirculate any evaporated fluid. The gas sparger was connected to a pressurised air supply via a flow indicator. Experiments were performed at different temperatures and acid content

Figure 5, Laboratory pre-leach setup

Pre-leach test 1The first test was an orienting trial to test the setup. A dilute slime sample was used and air was sparged through the slurry during heating and stirring. The leachate and the residue were sampled.

Pre-leach test 2In this test about 100 g slime was used. To this 250 ml (400 g/l) sulphuric acid was added. The total volume of the sample slurry including acid, slime and electrolyte was adjusted to about 800 ml by water addition. The slurry was heated to 55°C during stirring, then air 30 Master Thesis in Process Metallurgy

was sparged during 90 minutes. Samples of the slurry were extracted several times during the leaching procedure.

Pre-leach test 3This experiment was performed in a manner similar to the second test, the same amount of slime and acid was used. But in this case a temperature of 85°C was used and the batch leached for totally 180 minutes.


Full-Scale trials

Autoclave samplingsFull-Scale trials/samplings have been conducted at the precious metals plant. The production autoclave was sampled through a dive tube from which a sample was led into a bucket with a predetermined amount of water. This procedure ensures that the sample is instantly cooled and diluted so that precipitation is avoided.

The first trial was done in order to examine how the different components were leached relative to pressure, temperature and time. In this first full-scale trial, the previously used maximum pressure of 8.5 bars was applied. At the time where the autoclave normally would have been flashed the first time a hold time of one hour was introduced. After the one hour holdup the autoclave was flashed down to about four bars. This hold time was applied in order to investigate whether it is the renewed oxygen intake and sparging or the increased process time that is the main factor during the leaching process. The batch was sampled with a total of 11 samples, taken before, during and after the leach. One of the samples was taken as a replicate sample to ensure that the samples were taken in a representative way. The sample points are described in Table 4.

No Sampled:

1

234567891011

11.5 hours before leach (heating) startafter heating but before oxygen startat ~3.5 bars pressureat ~6.5 bars pressureat maximum, 8.5 bars pressure(replicate of no 5)after first flash at ~4.5 barsat ~8.5 bars just before last flashafter flash before the batch was released to the cooling tankin the cooling tank after coolingfrom the filter press at a later point than sample 10

Table 4, Samples taken during full-scale leaching


The results from the first trial led to the need of a second sampling where samples were extracted more frequent during the last part of the leach. In this trial the autoclave was run to the normal maximum pressure of 6.8 bars. Three samples were taken during the last part of the leach. In addition to these the production samples taken after filtration and after Ag/Se precipitation are also included in the evaluation of the results.

A third test was done in order to investigate the importance of the cooling time after the leach. According to Wesstrom [33] a prolonged holdup at about 100°C might be beneficial for the yield of tellurium. This test was performed after an ordinary leach. The batch was sampled just before the last flash, then the autoclave was flashed but the batch was left in the autoclave to cool. No external cooling was applied. One hour after the first sample a second sample was taken and a third sample was taken after another hour. After this sample the batch was released to the cooling tank and further treated as usual. The ordinary sample taken after filtration is also included in the evaluation.

Wash testsA possible cause of the sometimes relatively high copper content of the leached, washed and dried anode slimes was investigated. Either insufficient leaching or washing and filtering of the slime are the only probable causes. Therefore a sample of the washed and filtered slime was taken before it entered the drying ovens. The sample had a consistency of clay and was therefore crushed (divided) to a lump size smaller than 6 mm before it was soaked in water. In the first round two smaller samples was extracted from the bulk sample and treated by stirring in hot and cold water respectively. The water was filtered off and analysed for copper. Since the amount of water, slime and the moisture content of the slime was known the amount of copper washed out of the slime could be calculated. The test was repeated with stirring in cold water for another slime sample in connection with one of the full-scale sampling campaigns.


Data analysis of historical process data

Historical data from 2003 up until 2009 have been studied using multivariate statistical software. The values used are unfortunately not batch-wise data but monthly average samples. The reason to this is simply that this is the only data available for the raw slimes as well as for the leached and dried slime. The statistical software used is calledSIMCA-P+11® and was developed by Umetrics during the eighties and nineties. The software use PCA, principal component analysis, to "plot" the datapoints of a large dataset in an n-dimensional space where n is equal to the number of variables. The method is similar to regression analysis but with multiple variables investigated simultaneously. The main directions of the dataset in this n-dimensional space are called principal components. The other analysis method used in the software is the PLS where the dataset is modelled in the same way but this time divided into an x- and a y-space. The relation between the x- and y-spaces can thereafter be investigated. This type of software is a powerful tool which enables investigation of the correlations between many variables and development of predictive models for complex systems. [74]

Correlations between different slime components and properties have also been investigated with simpler spreadsheet software as for example Excel®. With this kind of software different parameters can be plotted against each other and the resulting graph can be fitted with trendlines (regression lines). This is a good way of getting a feeling of the dataset and see which variables that might be of larger interest.


ResultsLaboratory scale trials

Autoclave leachingThe chemical analyses of the leaching residue as well as XRF-analyses of the leachate from the different laboratory scale leaching experiments are compiled in Tables 5-7 below. The leach yield of some selected elements are also calculated.

Sample

Au [%]

Ag [%]

Cu[%]

Fe [%]

Ni [%]

Se [%]

Pd [g/t]

Pt [g/t]

Te [%]

Sb [%]

Ba [%]

Cl [%]

A1A2A3A4A5A6A7

Ref. 1Ref. 2

0.59 15.6 1.1 0.17 1.9 4.35 1620 103 0.8 5.5 4.1 <0.1 *31.7 1.57 0.41 5.2 0.65 **39.9 4.48 0.46 4.75 0.93 **31.1 1.72 0.44 5.3 0.65 **30.7 0.82 0.23 5.8 0.82 **31.5 1.56 2.72 5.2 0.92 **40.2 0.71 0.18 0.18 0.66 **/***

0.48 29.3 15.5 0.11 5.1 3.94 570 39 0.740.46 27.9 15.7 5.1 3.66 0.69 *** Another slime sample used, the separate reference not included here** Ag analysis also includes small amounts of lead, bismuth and other precious metals*** 0.64g Ag2Se precipitate not included

Table 5, Analysis results from laboratory leach tests, solid residues and reference samples

Sample Ag [g/l] Se [g/l] Te [g/l]

A1

A2A3A4A5A6A7

1.8 0.293 0.07621.13 0 0.55

0 0 022.56 0 0.53

* 20.44 0.1 0.1725.14 0 0

5.8 11.62 1.06* Value recalculated because of post filtration precipitate

Table 6, Leachate XRF analysis from laboratory scale tests


Different analyses of the leach residues and leachates were used together with measured weights and volumes to calculate the yields of the most important elements. The yields in Table 7 are based on the leach residue analyses.

Sample Cu [%] Ni [%] Te [%] Ag [%] Se [%]A2A3A4A5A6A7

93.7 95.0 45.8 33.2 18.580.4 93.9 14.7 7.4 18.193.7 95.1 50.4 40.0 24.097.4 97.8 44.8 47.8 26.793.8 66.9 22.8 33.1 18.197.0 97.7 32.1 11.4 -

Table 7, Calculated leach yields for selected elements

The selenium yields in the last run is not declared because of that the result gave a negative value indicating that the selenium balance is not correct enough to calculate a yield. The amount of selenium in the leachate is significantly higher in this run, not surprising considering the relatively large amount of selenium added. In the last leaching experiment an interesting observation was made. Some nugget shaped particles a few millimetres (2-3) in diameter had formed during leaching, see Figure 6. An XRF analysis of the material gave the stoichiometric composition of Ag2Se. The particles were smooth on their outer surface and hard but quite brittle. A total of 0.64 g of this precipitate was picked out of the slime sample after leaching. Another observation made during this leaching was that the slime sample afterwards was very difficult to filtrate.

Figure 6 A, Silver selenide "nuggets" formed in autoclave test A7 and Figure 6 B, close-up


A B

Pre-leachingPre-leach test 1This trial indicated only a very limited leaching. The copper, nickel and tellurium concentration of the leachate and residue was not significantly changed compared to the reference samples.

Pre-leach test 2 and 3The resulting leach residue analysis of the second and third pre-leached sampled as well as the reference values are compiled in Table 8 below. The values show a minor copper leaching in the moderately heated sample while the tellurium content are unchanged and nickel slightly decreased. The test performed at higher temperature i.e. "Pre-leach 3" shows an extensive decrease in copper, almost 2/3 is leached, while nickel and tellurium content is slightly increased due to the weight loss.

Sample Cu [%] Ni [%] Te [%]

Pre-leach 2

Pre-leach 3Reference

13.3 5.4 0.746.2 6.2 0.86

15.5 5.1 0.74Table 8, Leach residue analyses

Even if the slime has been leached for a relatively long time, and the final concentrations are accounted for in Table 8, the leaching process was rapid. This can be seen by the colour of small slime samples extracted during the leaching. The leachate is enriched in copper sulphate, by estimate, within the first 20 minutes of treatment.

Full-scale trials

Sampling number 1The process parameters during the first full-scale experiment are shown in Figure 7 and 8 below. Figure 7 shows the pressure during the leach, in this diagram the different steps in the leaching procedure can be identified. The first pressure increase up to about 1 bar is entirely due to heating, no oxygen has been added yet. Then the pressure increases quite steeply up to 8.5 bars due to oxygen sparging. The following hour the pressure decreases about 0.5 bars because of oxygen consumption. Then the autoclave was flashed and the oxygen intake renewed. The second peak has a less steep Jonas Bäckström 37

rising because of a lower oxygen flow during this part of the leaching. Finally the autoclave is flashed to atmospheric pressure. The pressure decrease was rapid in the beginning when the oxygen overpressure is released then the steam pressure slightly limits the rate of the pressure drop. Figure 8 shows the total amount of oxygen used and the temperature. In this particular case the batch was heated with steam to a temperature just above 100°C before the oxygen addition started. The further temperature increase is due to the reaction heat and is generated by the slime in the autoclave.

Figure 7, Pressure in the first full-scale autoclave run

16:00 20:00 00:00 04:00 08:00 12:00 16:00

Time

0

2

4

6

8

Pre

ssure

[bar]

Pressure


Figure 8, Temperature and oxygen consumption in the first full-scale autoclave run

Figure 8 above shows that the temperature was still rising when the vessel was flashed to atmospheric pressure. The different inclinations in the oxygen curve are due to the altered setting in gas flow.

16:00 20:00 00:00 04:00 08:00 12:00 16:00

Time

0

50

100

150

200

250

300

Tem

pera

ture

[°C

], O

xygen c

onsum

ed [kg]

Temperature

Oxygen consumed


Figure 9, Ag and Cu analyses in the first full-scale autoclave run

The diagram above (Figure 9) shows the silver and copper concentration as a function of time. The copper is mainly leached during the first pressure peak while silver is not affected at all until the second pressure increase. The amount of solubilised silver decreases substantially during cooling. The tellurium is also leached in last part of the leaching while the nickel is solubilised before the leach start when the acid concentration is the highest and during the second part of the first oxygen sparging.

16:00 20:00 00:00 04:00 08:00 12:00 16:00

Time

0

20

40

60

80C

oncentr

ation [g/l]

Cu

Ag


Figure 10, Te and Ni analyses in the first full-scale autoclave run

In the diagrams of element concentration from all three full-scale samplings the tellurium and silver values are based on XRF analyses while copper and nickel are analysed by means of AAS or titration. In addition to the analyses illustrated in the figures above, sample number 1 and number 10 were analysed for some additional elements to see what else might have been leached. The results are listed in Table 9 below.

Sample Sb [mg/l] As [mg/l] Se [mg/l] Cl [mg/l] Te [mg/l]

1

1014.3 0.614 - 77 -<15 9.5 286 <7 673

Table 9, Leachate chemical analysis

As seen in the table above the amount of arsenic have increased while the chloride content has decreased. The antimony analysis of sample 10 is only given as <15 because the dilution of the sample made the concentration less than the detectable limit. Both selenium

16:00 20:00 00:00 04:00 08:00 12:00 16:00

Time

0

2

4

6

8

10C

oncentr

ation [g/l]

Ni

Te


and tellurium was leached but the values of the chemical analysis do not exactly match the XRF analysis values.

Sampling number 2The process data as well as the analyses of the samples are shown in Figures 11-13. The maximum pressure was 6.5 bars and the autoclave was flashed three times. The batch was relatively small and therefore the oxygen consumption was lower than in the first full-scale test. The maximum temperature of about 150°C was reached after approximately three hours of oxygen sparging.

Figure 11, Pressure in the second full-scale autoclave run

08:00 10:00 12:00 14:00

Time

0

1

2

3

4

5

6

7

Pre

ssure

[bar]

Pressure


Figure 12, Temperature and oxygen consumption in the second full-scale autoclave run

08:00 10:00 12:00 14:00

Time

0

50

100

150

200

Tem

pera

ture

[C

], O

xygen c

onsum

ed [kg]

Temperature

Oxygen consumed


Figure 13, Leached elements in the second full-scale sampling.

Sampling number 3The process data as well as the analyses of the samples are shown in Figures 14-16. The maximum pressure was 6.5 bars and the autoclave was flashed three times. The pressure peaks visible after the last flashing are due to the sampling when the autoclave was pressurised with air to be able to extract a sample. The maximum temperature of about 150°C was reached after just over four hours of oxygen sparging.

11:00 11:30 12:00 12:30 13:00 13:30 14:00

Time

0

10

20

30

40

50

60C

oncentr

ation A

g, C

u, N

i [g

/l]

0

0.1

0.2

0.3

0.4

0.5

0.6

Concentr

ation T

e [g/l]

Cu

Te

Ag

Ni


Figure 14, Pressure in the third full-scale autoclave run

08:00 10:00 12:00 14:00 16:00

Time

0

1

2

3

4

5

6

7P

ressure

[bar]

Pressure


Figure 15, Temperature and oxygen consumption in the third full-scale autoclave run

In Figure 15 it can be seen that the attempt to post digest the slime at about 100°C was successful in terms of the process conditions. The concentration of silver in the solution is constantly decreasing during the cooling while the tellurium analysis is variable.

08:00 10:00 12:00 14:00 16:00

Time

0

50

100

150

200

250

Tem

pera

ture

[C

], O

xygen c

onsum

ed [kg]

Temperature

Oxygen consumed


Figure 16, Leached elements in the third full-scale sampling.

12:12 13:01 14:02 15:03 16:04 17:05 18:06

Time

0

5

10

15

20

25

30C

oncentr

ation A

g [g/l]

0

0.5

1

1.5

2

2.5

3

Concentr

ation T

e [g/l]

Te

Ag


Washing tests

The results of the slime washing tests are accounted for in Table 10 below. The table shows the calculated amount of copper washed out from the two slime samples. The available amount of copper inferred by an XRF analysis of the dried slime gave a copper content of 1.3% for the first sample and 2.3% for the second, respectively. The stirred hot water wash gave the highest amount of extracted copper. However, all tests were successful in terms of extracting a significant amount of copper from the filtered slime.

Moisture content

calculationSample 1

Warm and stirred

Sample 1

Cold and stirred

Sample 1

Moisture content

calculationSample 2

Cold and stirred

Sample 2

Amount added water [ml]Weight of slime [g]Weight of dried slime [g]Calculated % solidsCopper analysis of wash solution [g/l]Extracted copper [g]Extracted copper[%-units]

- 2000 2000 - 2000

430.49 200.00 209.04 250.30 329.67

340.43 158.16 165.31 202.97 267.33

79.08 - - 81.09 -

- 1.4 1.05 - 2.00

- 2.80 2.10 - 4.00

- 1.77 1.27 - 1.50

Table 10, Results from the two separate slime wash tests


Statistical analysis of historical data

MultivariateThe multivariate data analysis was performed using chemical analyses of monthly samples from the precious metals plant. Analysis of raw slime, dried slime, copper telluride and the Kaldo slag as well as the total amounts of the different materials was included in the initial dataset. A calculated tellurium yield was used as response variable. No obvious and strong correlations were immediately found but after some work with the dataset a model with a degree of explanation, R2Y,at 60% was achieved. The predictive power, Q2Y is in the order of 0.56. The significant variables left in the model after removing the insignificant are: the antimony, selenium and tellurium content of the raw slime, the gold, silver, lead, antimony and platinum content of the dried and leached slime and the silver content of the copper telluride (see Figure 17-18). According to this model the silver and antimony have negative effect on the tellurium yield, whereas the other variables have a positive effect. The model validation plot (Figure 19) looks good, the model can be said to be valid and most scores are within 2 standard deviations (Figure 20). The distance to the model in the X-plane is within the critical level for most observations (Figure 21).


Figure 17, VIP-plot shoving the most important variables. From the left: silver in the copper telluride, selenium in raw slime, silver in leach residue, gold in leach residue, lead in leach residue, antimony in raw slime, platinum in leach residue, tellurium in raw slime and antimony in leach residue.

Figure 18, Coefficients plot illustrating the model and the effect of the different variables. From the left: antimony, selenium and tellurium in raw slime, gold, silver, lead, antimony and platinum in leach residue and silver in the copper telluride.


Figure 19, The points in the left of the validation plot are all well below the reference points to the right

Figure 20, Scores plot, all but two points are within +/- 2 SD


Figure 21, DModX-plot, all observations but three are within the critical level of the distance to the model plane

Figure 22, Observed vs. predicted yield of tellurium


The observed versus predicted plot in Figure 22 are somewhat scattered but the points still follows a clear trend. The multivariate model cannot be said to be a very good explanation of the tellurium yield but it still gives some clues.

One variable at the timeOne logical assumption could be that the tellurium yield is dependent on the absolute tellurium content of the incoming material. If the yield is plotted versus the tellurium content a diagram as shown in Figure 23 is achieved. There might be a trend but there is not possible to draw any certain conclusions concerning the correlation. Looking at the area in the plot where the tellurium content are in the range of 1.8 to 2.1 it can easily be seen that the variation in the yield is very large. If the values in this range are extracted and treated separately it might be possible to see which other parameters that affect the yield when the tellurium content is in the same range (~ constant).

Figure 23, Fraction of leached tellurium as a function of the raw anode slime tellurium analysis

0.5 1 1.5 2 2.5 3

Raw anode slime tellurium content [%]

0

0.1

0.2

0.3

0.4

0.5

0.6

Fra

ction tellu

rium

leached


The following graphs (Figures 24-29) are made from the data points where the tellurium content is between 1.8 and 2.1. As seen from the Figures some possible correlations were found. Some of the indicated parameters coincide with the SIMCA model described earlier. If the correlations are true, as opposed to random, a high bismuth content of the raw slime would be favourable for the tellurium yield. Also a high level of barium, lead and gold in the leached and dried slime correlates with a good tellurium yield. A high amount of nickel and antimony seems to be disadvantageous for tellurium leaching. The regression lines in the plots below are no perfect correlations but the degree of explanation, R2, are in the order of at least 15% for each plot.

Figure 24, Fraction of leached tellurium plotted as a function of raw anode slime bismuth content (at Te ≈ 2%)

Bismuth content of the raw slime indicates a positive correlation to the tellurium yield of the pressure leach.

0.4 0.6 0.8 1 1.2 1.4

Raw anode slime bismuth content [%]

0

0.1

0.2

0.3

0.4

0.5

0.6

Fra

ction tellu

rium

leached


Figure 25, Fraction of leached tellurium plotted as a function of dry anode slime barium content (at Te ≈ 2%)

The barium content of the dried slime seems to be positively correlated to the tellurium yield although the yield at lower barium content are quite scattered.

4 5 6 7 8

Dried anode slime barium content [%]

0.1

0.2

0.3

0.4

0.5

Fra

ction tellu

rium

leached


Figure 26, Fraction of leached tellurium plotted as a function of dry anode slime lead content (at Te ≈ 2%)

The correlation to the post leach lead content also seems to be positive, one of the points deviates strongly from the general trend.

10 12 14 16 18

Dried anode slime lead content [%]

0.1

0.2

0.3

0.4

0.5

Fra

ction tellu

rium

leached


Figure 27, Fraction of leached tellurium plotted as a function of dry anode slime gold content (at Te ≈ 2%)

The positive correlation towards gold content looks a bit uncertain and the points are quite scattered but the R2 is as high as 21.6%. Generally this is not considered a very good correlation but taking into account the fact that it is one of many factors and few points it is surprisingly good.

0.4 0.6 0.8 1 1.2 1.4

Dried anode slime gold content [%]

0

0.1

0.2

0.3

0.4

0.5

0.6F

raction tellu

rium

leached


Figure 28, Fraction of leached tellurium plotted as a function of raw anode slime antimony content (at Te ≈ 2%)

A negative correlation between antimony content of the raw slime and the tellurium content are inferred by the plot. Although, the points are very scattered.

2.5 3 3.5 4 4.5 5 5.5

Raw anode slime antimony content [%]

0

0.1

0.2

0.3

0.4

0.5

0.6F

raction tellu

rium

leached


Figure 29, Fraction of leached tellurium plotted as a function of raw anode slime nickel content (at Te ≈ 2%)

A high nickel content of the raw anode slime has a negative effect on the tellurium yield while the points are spread in the entire range at lower nickel content. The points with the highest nickel content are strongly influencing the regression in this plot. Thereby it is of great importance that these points are correct if the results are to be trusted.

0 2 4 6 8 10 12

Raw anode slime nickel content [%]

0

0.1

0.2

0.3

0.4

0.5

0.6F

raction tellu

rium

leached


DiscussionAs is evident from the introduction the Rönnskär plant is quite complex. Virtually all residue materials are recycled within the plant. Therefore every change to the existing process must be thoroughly thought through because of the possible effects it might have on other processes. First of all it should be said that the experiments performed within this project cannot be assumed valid for all possible batches of slime due to the large variations. The results might anyhow give valuable input and knowledge of which parameters that are of greater importance. A percentage unit more or less in tellurium yield might seem unimportant at a first look. But if one consider the whole picture it becomes more obvious why it is of importance. A lower yield means a poorer product quality of the copper telluride and a higher load of tellurium in the smelting process. This in turn contributes to larger amounts of slag produced and thereby larger losses of precious metals. The increased slag volume gives a circulating load in the smelting cycle and decreases the furnace capacity. The amount of tellurium ending up in the selenium product will of course also increase and hence lowering its quality. A higher impurity (copper, nickel or tellurium) content in the anode silver makes additional electrolyte withdrawal necessary and thereby additional electrolyte treatment costs as a consequence. It is therefore of importance to the subsequent processes to get an as good as possible result in the leaching treatment.

A considerable part of this project has been the literature survey. Many articles dealing with anode slimes have been found. The mineralogy of slimes has been extensively examined by some, not least Chen and Dutrizac. Conclusions and assumptions of the slimes at Rönnskär can be drawn from their results. The slime really consists of a mixture of several different slimes such as refractory compounds, precipitates, baryte and anode fragments. The selenium and tellurium content are probably mostly bound in the common copper-silver selenides-tellurides. The nickel content of Rönnskär's anodes are sufficiently high to assume that both nickel oxide and kupferglimmer as well as nickel in solid solution in the copper matrix are present. The slime composition at Rönnskär does not differ significantly from the "normal case" and therefore similar mineralogy


would be expected. The baryte and copper content of the slime are probably varying dependent on whether or not the slime storage is completely emptied when slime is pumped from the electrolysis plant. The baryte and copper content are most likely enriched in the bottommost layer of slime due to their larger particle size and thereby rapid sedimentation.

The mineralogy and composition in detail may have severe effect on the slime processing and are dependent on several factors such as: the ore feed composition, the smelting and converting processes, the anode furnace deoxidisation, the anode casting temperature and cooling rate, the electrolysis temperature and current density as well as the duration time in the refining cells. An attempt to once and for all optimise the process parameters in full-scale leaching is therefore doomed to fail. Better then is to try to find clues to how and which parameters that affect the result. This enables a better understanding of the overall process and might in the long run lead to a better process.

The processes occurring in the leaching are oxidation and dissolution of some of the slime constituents. Selective leaching of elements out of some compounds does also occur. Some re-precipitation of dissolved substances might also occur either in the autoclave during leaching or during cooling afterwards. The tellurium content of the slime is mostly present in the selenides and in the so called oxidate phase. The recovery of tellurium might therefore be dependent on the composition and properties of these phases. The selenide structures are primarily formed during anode casting. The selenides described in the literature are quite similar in different slimes even if the overall analysis differs. Although, the amount of tellurium in the selenides must be dependent on the total Se/Te ratio of the slime. The selenides undergo alteration in composition due to substitution during electrolyte contact. The copper in the selenide structure is replaced by silver and hence the selenide particles are enriched in silver. This fact might seem advantageous from a slime leaching perspective; the less copper in the slime to be leached. But that is not certainly the case since the silver rich selenides have been found to be more difficult to leach, and act as a retardant for the desired selective tellurium dissolution. [34] Much of this substitution will have taken place within a few hours slime electrolyte contact but further Jonas Bäckström 61

substitution might occur during a longer period of time. This implies that ageing of slime might be disadvantageous for the leaching process even if the initial substitutions are immediate. The other major tellurium carrier, the oxidate phase, is an intimate mixture of a Cu-Ag-Pb-Au-sulphate-arsenate-antimonate-selenite-tellurite-material and its composition would logically be more dependent on the overall chemical analysis of the slime and the electrolysis parameters. Then it is very likely that the leachability of tellurium from this phase will depend upon several parameters. Some of the parameters that affect the oxidate phase might be the parameters inferred in the investigation of historical process data. One of the parameters that have an indicated effect on the tellurium yield is the amount of tellurium in the raw slime. This is a correlation that feels quite logical and would be the result of some refractory tellurium compound present to a certain extent in the slime also at lower total amounts of tellurium. If a refractory tellurium bearing compound is formed and holds an equal amount of tellurium despite a lower total tellurium content the implication would be that the tellurium content of the slime cannot be reduced below a certain level. Thereby the yield would be lowered if the total amount is. Also if the total amount is lower but the leachate volume is the same the concentration in the leachate would be lower. If then the residual concentration of tellurium in the electrolyte sent back to the tankhouse also is the same this will mean that a larger portion of the tellurium is not precipitated in the copper telluride production. Hence lowering the total yield. The lower concentration of tellurium in the leachate might also be unfavourable for the telluride precipitation with difficulties in accurate copper powder dosages.

A possible explanation to the large variation in the leaching of tellurium is that the tellurium content could either be oxidised to Te(IV) or Te(VI). The different ions have different solubility in the acid media and might also have different tendencies to react with other substances in the slime. If the hexavalent tellurium is preferable as some of the literature have inferred it is possible that the prolonged digestion described is a somewhat slower oxidation reaction of the tetravalent tellurium into hexavalent. This statement is merely a theory and has not been confirmed by any studies.


The first full-scale trial shows that copper is leached almost immediately and will continue to leach to some extent during the latter part of the leaching process. The nickel seems to have been leached to large extent during the night while the batch was prepared and stored in the autoclave. Since no sample was taken at 4 a.m. it cannot be said if this leaching occurred during the prolonged storage in (quite strongly) acidic environment or during heating of the batch in the morning. The most probable explanation considering other results is that the nickel is leached largely because of the long time in the acid since the heating is not performed to more than around 100°C. This implies that the high temperature required for nickel leaching to some extent might be replaced by a longer reaction time and/or higher acid concentration. The silver content is leached after the first autoclave flashing and significantly re-precipitated during cooling after the second flashing. The precipitation is probably due to the temperature decrease affecting the solubility of silver sulphate. No leaching of tellurium occurs before the first flashing and the content is increasing in the last samples. The resolution is not large enough to get a view of the tellurium leaching. Therefore a second batch is sampled with a narrower interval at the end of the process.

In this second batch the tellurium content increased from the second flash to the third but the concentration in the filtrate was significantly lower in the beginning. Then, later on, the concentration of the filtrate rose again. This combined with some comments in revised literature gave rise to a theory, maybe the tellurium is precipitated during leaching or in the cooling/depressurising treatment and further re-leached in the cooling tank during filtration. To test this, a third full-scale experiment was done.

Another possible explanation of the low yield is, as mentioned in the literature survey section, that tellurium is re-precipitated simultaneously as it is leached. The third full-scale trial was aiming at finding out if a prolonged digestion at temperatures ≳ 100°C could dissolve a possible precipitated phase. But the test did not give a clear answer to this question. The first sample taken after the filtration of the slime indicated a drastic decrease in tellurium concentration. The second sample taken one hour after the flashing was started indicated a higher tellurium than the first sample. The next two samples were lower in tellurium concentration whereas the Jonas Bäckström 63

final sample taken after Ag/Se cementation had a considerable higher tellurium concentration

In the third full-scale sampling a confirmation of a previous observation was made; it seems like the sample taken from every batch in the beginning of the filtration of the leached slime generally holds a lower amount of tellurium than samples taken both later and earlier. This has sometimes been observed in samples taken during production. The only possible conclusion is that tellurium is "consumed" in the beginning of the filtration when the slurry is pumped from the cooling tank via the filter to the Ag/Se precipitation reactor. Since the cemented silver from the previous batch already is present in the filter press it is plausible that a surplus of copper added in the silver cementation will deplete the first solution run through the filter of some of its tellurium content. Observations of copper powder still present in the cemented silver have been done. This mechanism will obviously lower the yield of tellurium to some extent, how much is difficult to say without further investigations. This theory is further supported by statistics indicating some degree of negative covariation between the tellurium and silver content of the leachate. The SIMCA-model also has silver content of the slime as a significant variable. This indicates that the dosage of both SO2 and copper powder in the silver cementation process are important but also difficult if the analysis of the first leachate is doubtful. According to this theory an overdosage will lead to copper powder in the filter which in turn leads to lower concentrations in the sample taken during the next batch. This leads to an initial underdosage of SO2 since the sample is used as a guide. This means that more copper powder is used to adjust and precipitate the last silver and thereby also a greater risk of overdosage.

Some sources of error must be mentioned in connection with the full scale samplings. The sample was extracted from the pressurised autoclave through a dive tube, the tube was led into a bucket filled with a known amount of water for instant cooling and dilution. The final volume was measured by markings in the bucket and a part of the sample was filtered off. This procedure gives a slightly rough calculation of the sample dilution and may cause marginal errors in the results. The tube was rinsed with slime before each sampling to avoid contamination by older slime. Samples taken after the 64 Master Thesis in Process Metallurgy

autoclave was emptied were mostly taken at the inlet to the Ag/Se precipitation reactor. If the sample taken in the beginning of the filtration is omitted from the diagram in Figure 16 it can be seen that a rising trend in tellurium concentration might be true. The results from the sampling can therefore not conclusively confirm nor disprove the theory of a re-leaching of precipitated tellurium.

Due to the very complex chemistry of tellurium and the diverse mixture of elements in the slime and leachate there are plenty of candidates for substances that tellurium might react with and precipitate. For instance lead, silver, antimony, bismuth and selenium are all possible candidates.

The statistic analysis gave indications of antimony in the slime as a retardant of tellurium dissolution. One possible cause, is an altered composition of the oxidate phase which is the second most important tellurium carrier but also a significant antimony carrier. Another explanation is some kind of precipitation occurring in the leach. The second substance indicating negative influence is silver, some possible reasons are mentioned above and another is the composition of the selenides which might be altered by the higher silver content. The SIMCA model had a degree of explanation at almost 60% that is a good value considering the fact that monthly averages was used which means that much of the variation may be hidden in the averages. However, from a model point of view it is not a high value. The statistical analysis may in the future be done in a better way if the raw slime also is sampled on a batch wise basis and analyses are made carefully. Platinum and gold are quite surprising to find in the model describing the tellurium yield. But tellurium can, according to Lister and McLeold [75], form complexes with palladium and platinum. Gold can also substitute for silver in selenide compounds and affect their leachability.

Of all the conducted experiments it is the washing test that has given the most straightforward results. The copper content in the dried slime is almost entirely due to a lack of slime washing, not a consequence of a lack of leaching. The sometimes deficient washing of the slime will of course have the same effect on the nickel and tellurium content but that are of lesser importance since their concentrations in the leachate are much lower. The copper content of Jonas Bäckström 65

the slime will never be zero but with a controlled washing the residual content might be very small. The copper washed out is likely in the form of copper sulphate. Other sources of copper like refractory precipitates or copper powder residues in the cement silver can obviously not be washed away by water. Some residual sulphate will probably also be left in the slime even after washing in copious amounts of water. Although a copper content of 2% can be significantly lowered by thorough washing.

The laboratory autoclave leaching tests does confirm some things. The nickel content is leachable at high temperatures even if the oxygen supply is low. Copper is leachable also during lower temperature but the oxygen supply is crucial. The tellurium yield seems to be disfavoured by both lower temperatures and lack of oxygen. The silver content of the samples are likely at least partly a function of filtration temperature. Although, the oxygen supply is a necessity for the silver dissolution. The silver leaching also seems to be favoured over the tellurium leaching in more dilute batches. The silver content of the leachate was drastically reduced by the selenium addition, probably due to precipitation of Ag2Se (which was observed in the residue). The laboratory autoclave is not ideal for this kind of experiments and the full-scale process is difficult to exactly replicate. Some difficulties during leaching also give arise to sources of error in excess of the scale differences. The oxygen inlet was equipped with only a small opening to give smaller bubbles but the consequence was that the tube sometimes clogged giving problems with the oxygen supply. Another difference is that the full-scale autoclave manages to self regulate its temperature after the initial heating while the laboratory unit cannot. The temperature must be risen manually and will therefore not follow a similar curve as in the production unit. Also, if the temperature curve is steeper this means that less oxygen will enter the autoclave due to the added vapour pressure.

The methodology for the pre-leaching studies worked quite well besides the added uncertainty due to minor water evaporation. The tests show that it probably is possible to pre-leach much of the slime's copper content. The following question is then if that is a good thing. A lesser amount of solid copper entering the autoclave means that a lesser amount has to be leached and thereby a possibility of saving autoclave time. But if the copper content of selenide particles 66 Master Thesis in Process Metallurgy

are leached out there might be a negative effect on the tellurium leaching.

The stirring and positioning of the oxygen lance in the production autoclave are two other parameters that would affect the overall leaching result. It is of great importance to get an as good as possible contact between oxygen and pulp in order to maximise oxygen dissolution and utilise the oxygen to a maximum. The results of the first full-scale trial enhance this statement. The results show that the leaching continues if the autoclave is left at hold, instead of flashing it. But the reaction kinetics is probably somewhat slower when the oxygen is not sparged through the pulp.

During the course of this investigation the tellurium content of the slime has been quite stable at a low level. Some samples of higher tellurium content would have been interesting to investigate. The effects of some parameters would probably be easier to see if the total content is higher.

In the literature several other leach reagent additions have been mentioned but most of these were not further investigated in the practical part of this project. This because of the fact that these reagents would risk causing severe problems in other parts of the processes or because of the additional hazards some of the chemicals would cause.

Another source of error concerning all tests which that was discovered after the completion of the project was that the XRF analysis of silver, selenium and tellurium concentrations of the solutions are not calibrated for samples containing less than 0.5 g/l Te. Most samples analysed in this project as well as samples from the production are around or below this value. The analyses however have a good repeatability even if the absolute values might be biased by the absent calibration.


Conclusions

• The leaching of tellurium is possibly affected by the slime's content of antimony, silver and lead.

• Much of the residual copper content after leaching is possible to wash out.

• The experimental methods used are applicable and can give valuable results but the trials in the laboratory autoclave are merely indicative due to scale differences.

• Some tellurium is probably re-precipitated in the post-leach filter press.

• Copper and nickel are the first elements to be leached. Tellurium and silver are leached simultaneously after the main part of copper and nickel have been solubilised.

• A significant amount of copper could be leached in the slime bulk by supplying air. However, the consequences of the pre-leaching for other processes are not clarified.

• The analysis of process data indicates that there are some correlations of the slime analysis to the tellurium yield. But to be able to understand the process better and clarify the correlations a thorough sampling on batch level is required. The monthly averages are not detailed or reliable enough for this purpose.

Suggestions of improvements

• The XRF analysis method needs to be calibrated, also for lower concentrations of tellurium.

• The routines for sampling and filter press washing needs to be looked over.


Suggestions for future studies

The most interesting topics for further studies are:

The possibilities of decreasing the copper, nickel and tellurium content of leached slime through modifications in the filter press washing procedures.

The possibilities of post leach digestion of slime.

The influence of other slime constituents. A batch wise sampling of raw slime is a requirement for better analyses.


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Appendix 1Anode slime constituents

The anode slime fed to the precious metals plant really consists of a large number of different chemical precipitates. The precipitates have different compositions and properties and are formed in different ways. A more correct denotation would therefore be "anode slimes" rather than anode slime. some of the slime constituents are precipitated from the electrolyte while some are inclusions from the anode that have remained unchanged during electrolysis. Some substances also undergo extensive reactions and compositional change due to prolonged contact with the electrolyte. According to the literature the following substances have been found in anode slimes originating from copper electrolysis. The numbers in the list below refers to the reference list in the main report.

Elements:Au [29, 76]Ag [21, 24, 26, 27, 29, 39, 76]Cu [26, 27, 39, 55, 76]Te [24, 76]Se [24, 76]S (with Se and Te impurities) [24]Ag, Au "alloy" [29]Ag, Au, Cu "alloy" [42]

Compounds:NiO [21, 26, 27, 55, 66, 76](Ni,Cu)O [26, 55, 66]CuSO4 [24, 26, 27, 29, 39, 42, 55, 66]NiSO4 [24]PbSO4 [21, 24, 26, 29, 39, 42, 55, 76]BaSO4 [21, 26, 27, 29, 39, 66]CaSO4 [24, 42]Ag2Te [27, 29, 55, 76]Ag2Se [26, 29, 39, 55, 76]

1

AgCu(Se,Te) [29](Ag,Cu,Ni)2(Te,Se) [27]Cu2Te or Cu2Se or Cu2(Se,Te) [26, 27, 39, 42, 55, 76](Ag,Au,Cu)2(Se,Te) or AgxCuySe or AgxCuyTe [24, 26, 27, 29, 39, 42](Cu,Ag)2OSeO2 or Ag2SeO3 or CuSeO3 [24, 26, 27, 29, 55]CuSe [66]CuAgSe [39, 55, 76](Ag,Au)Te2 [76]Ag3Sb [24]AgCl [24, 26, 76]CuCl or Cu2Cl2 [26, 29, 39, 76]~PbO(SbxAs1-x)2O5 [24]Al2O3 [24, 42]SiO2 [24, 26, 42, 76]Cu-silicate [77]Cu-Ni-silicate [77]Ca-Cu-Ni-silicate [77]K-Al-silicate [77]Al-silicate [77]Ba-Al-silicate [77]Ba-Al-SO4-silicate [77]Na-Al-silicate [77]Ca-Al-silicate [77]Fe-Mg-Ca-Al-silicate [77]Ca-K-Al-Na-silicate [77]Fe2O3 [24, 42](Sb2O3)x•(As2O5)1-x (so called floating slimes) [24, 42]Sb2O3 [76]Bi2O3 [76]AuTe [24]CuAs [24]Cu2O [26, 27, 29, 39, 55]6PbO•2As2O5•3CuO [37]Pb5(AsO4)3(OH,Cl) [39]Cu2O•Sb2O5 or 5Cu2O•Sb2O5 [2, 42, 55]Cu-Ni-Sb-oxide [21, 55] (= kupferglimmer, Cu3Ni2-xSbO6-x where x = 0,1 - 0,3)Cu2S [42, 76]CuFeS2 [24]2

ZnS [24]MgO [24]NiFe2O4 [43](Ni,Fe)Cr2O4 [24]SnO2 [21, 26, 27]SbAsO4 [26, 39, 55, 76]SbAsO3 [39]BiAsO4 [26, 76]As2O3 [76]Cu-AsO4 [26]Sn-arsenate [26]various Cu-Pb-As-Bi-Sb-Sn-oxides like: [26, 27, 29, 39, 76] Pb2Sb2O6 Sb-As-oxide Cu-Pb-As-Bi-Sb-oxide Pb-As-Bi-Sb-oxide Sb-As-Bi-oxide Cu-Pb-As-oxide Cu-Pb-Bi-oxideRare Zn-Fe-Ni-Sb-oxide [29]"Oxidate phase" [26, 29, 27, 31](= varieties of Cu-Ag-Pb-Au-sulphate-arsenate-antimonate-selenite-tellurite-material)Often associated with other precipitates e.g. CuSO4.

Electrolyte constituents:H2SO4H2OGlue, and its decomposition products.Thiourea, and its decomposition products.NaCl

The most abundant of these compounds are in most cases probably: metallic silver and copper, nickel oxide, copper sulphate, lead sulphate, barite, floating slime, the complex oxidate phase and various copper/silver-selenides/tellurides.

3

After the Cu-Ni-Te leaching the slime of course have a changed composition with silver sulphate, elemental silver, lead sulphate, barite, silver selenide, elemental selenium, a new precipitated oxidate phase and different complex oxides as the most abundant constituents. [30, 34, 38]

4

Appendix 2Rönnskär precious metals plant flow sheet.

1

Appendix 3

Rapport riskanalys Sida 1(3)

Distribution:

Objekt/aktivitet/Plats Autoklav/Slamlakningsförsök i laboratorieautoklav/C-lab. vån 3

Datum 2010-02-08

Bakgrund (Varför görs riskanalysen? Kort beskrivning av arbetet och utrustning, Farliga arbeten, ämnen etc.)

Ett antal lakförsök kommer att genomföras i Laboratorieautoklav, Parr no 4551. Försöken genomförs inom ett examensarbete. Anodslam från kopparelektrolysverket lakas i svavelsyramiljö med varierande tryck, temperatur, syrahalt, tillsatser etc.

Riskanalysmetod Arbetssäkerhetsanalys ASA

Deltagare vid analysen Kjell-Åke Johansson, R1Ä. Jonas Bäckström, RCP. Anne Blomgren, RCP. Cecilia Lindmark, RMA.

Referenser (foton, ritningar, rapporter, platsbesök etc.)

Översättning: ”Instruktioner och säkerhetsföreskrifter för laboratorieautoklav PARR 4551” 1978-09-07

Resultatsammanfattning: (Kort beskrivning av de farligaste riskerna som måste åtgärdas. Övriga synpunkter av intresse) De flesta potentiellt farliga situationer som kan uppstå är resultat av oaktsamhet och/eller okunskap. De kan i de flesta fall undvikas med hjälp av sunt förnuft, försiktighet, rätt förberedelser och användande av korrekt skyddsutrustning.

Ansvarig för uppföljning Uppföljning datum Jonas Bäckström

Vid varje användningstillfälle

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3

Appendix 4

Användarinstruktion Laboratorieautoklav 4551Jonas Bäckström, RCP

2010-04-23

1

1. Autoklavens utseende och funktioner

Innan användning skall autoklaven vara besiktad, provtryckt och godkänd, kärlet är ursprungligen provtryckt och godkänd för 70 bars tryck vid 300 °C.Autoklavkärlet står i ett golvstativ, i detta stativ finns det en värmemantel samt en drivmotor till autoklavens omrörare. Bägge dessa styrs från en kontrollpanel på stativet, önskad temperatur ställs in och värmaren regleras så att temperaturen hålls vid denna nivå (endast värmaren styrs, om kylning är nödvändig måste denna regleras manuellt). På kontrollpanelen kan kärlets temperatur avläsas inom ett +/- 50 °C intervall från det inställda värdet. Omrörarens hastighet kan ställa in med hjälp av ett vred och dess hastighet kan avläsas på en mätare. Kom ihåg att kontrollera att drivremmen är hel innan försöket påbörjas. I botten av kärlet finns en dräneringsventil, givetvis måste denna vara stängd för att kärlet skall kunna trycksättas. Om en mycket aggressiv miljö förekommer under laborationerna kan kärlet kompletteras med en glasliner. Glaslinern sätts försiktigt på plats innan materialet fylls i autoklaven.

I autoklavens lock finns det ett antal anslutningar:I. OmröraraxelII. ManometerIII. FlashventilIV. Gastillförsel/provuttagsventilV. Ficka för termoelementVI. Nippel för sprängblecksmonteringVII. genomföringar för att möjliggöra montering av en

kylslinga(Se även Figur 1 nedan)

2

Figur 1, Laboratorieautoklavens utseende när alla delar monterats

På omröraraxelns lager finns det tre nipplar, till de två större skall kylvatten anslutas för att undvika överhettning av det täta lagret. Den mindre nippeln är till för att kunna upptäcka läckage. Från denna nippel ansluts en slang som leds till en bägare med några centimeter vatten, därigenom kan läckage lätt upptäckas som bubblor. På omröraraxeln monteras en lämplig omrörare. Manometern som sitter monterad på autoklavens lock möjliggör avläsning av autoklavens inre tryck upp till ca 16 bar. Flashventilen används för att sänka trycket i kärlet antingen under drift eller vid avslut. Till denna ventil bör en slang som leds till avlopp i dragskåp anslutas, detta för att undvika att giftiga och/eller heta gaser och vätskor sprids när ventilen öppnas. Gastillförseln sker genom en ventil, på undersidan av autoklavens lock sitter ett dykrör kopplat till 3

denna ventil. Detta medför att gastillförseln sker under vätskeytan i reaktorn. I samma anslutning som gasventilen sitter en annan ventil som möjliggör provtagning under drift (vid provtagning bör en slang vara monterad på denna ventil). På undersidan av locket sitter också en ihålig stav i vilken värmestyrningens termoelement kan föras in från utsidan av reaktorn. I locket finns också en hållare för ett sprängbleck (se även Figur 2), i denna skall ett sprängbleck av korrekt storlek och typ vara monterad, detta är en viktig säkerhetsdetalj då blecket skall brisera om trycket blir alltför högt i reaktorn. Till denna hållare kan också en slang anslutas för att undvika skaderisk (personer och material) om olyckan skulle vara framme och blecket sprängas. Sprängblecket är kalibrerat att brista vid ett tryck av 1000 psi (= 69 bar). Till varje sprängbleck finns det en skylt som anger dess bristningstryck och andra data. Denna skylt skall sitta på ett synligt ställe på autoklaven; när sprängblecket byts skall givetvis den nya skylten ersätta den gamla .I locket finns det också två genomföringar för att kunna montera en kylslinga (denna kan dock inte användas samtidigt som glaslinern på grund av utrymmesbrist). Kontrollera även packningen mellan lock och cylinder. Alla genomföringar som eventuellt kommer att användas för att evakuera kärlet (avsiktligt eller oavsiktligt) bör ha kran eller rör vänt bort från alla utrymmen där personer vistas samt vara försedda med en slang för att leda bort gas och vätska.

4

Figur 2, Hållare för sprängbleck

2. Drifttagande av autoklaven

När alla delar till autoklaven är monterade är den redo att tas i bruk. Om glaslinern skall användas sätts denna först försiktigt ner i reaktorkärlet. Sätt på locket och kontrollera att alla delar passar och får plats, omröraren får inte ta i dykröret eller termoelementfickan. Kontrollera även så att omröraren och dess axel inte skrapar emot botten av glaslinern. Öppna locket och häll varsamt ned lagom mängd material. En ”lagom mängd material” kan ju variera beroende på vad som skall behandlas, temperatur, tryck, förväntade reaktioner etc. använd sunt förnuft och tänk efter före. Fyll aldrig autoklaven mer än till ca ⅔ - ¾, detta eftersom det måste finnas en gasfas ovanför vätskeytan som kan komprimeras när trycket ökar. Överfyllning ger upphov till en risk att trycket skenar okontrollerat.

När autoklaven fyllts och locket satts dit lyfts de bägge ringhalvorna på plats, när det kontrollerats att inget är i vägen för bultarna kan dessa dras fast. Dra åt bultarna växelvis, och vänta några minuter så att packningen hinner sätta sig (särskilt viktigt om packningen är ny) dra sedan efter bultarna växelvis ännu en gång. Innan något startas bör alla slanganslutningar 5

kontrolleras, sätt sedan igång kylvattnet till lagret. Se till så att termoelementet sitter på plats och slå sedan igång värmaren och ställ in önskad temperatur. Vrid drivarmen för omröraren i rätt läge och fäst omröraraxeln genom att trycka ner bussningen över axelns topp. Slå på omröraren, om reaktionsblandningen är mycket viskös kan det vara fördelaktigt att för hand hjälpa omröraren igång.

Om prover skall tas under drift, så görs detta med hjälp av gastillförselgenomföringen. Med hjälp av autoklavens inre tryck kan ett prov tryckas bakvägen genom dykröret till provuttaget.För att ta ut ett prov måste först gaskranen på autoklaven stängas, sedan leds slangen från provuttaget ner i provkärlet. Därefter kan provkranen försiktigt öppnas tills önskad provmängd erhållits. Stäng kranen och öppna sedan åter gaskranen för att tömma ur dykröret och återställa trycket. Tänk på att provet kan vara mycket varmt och att trycket ur slangen kan bli mycket högt om kranen öppnas för mycket. Använd därför lämplig skyddsutrustning.

3. Avslut av autoklavförsök

Innan autoklaven kan öppnas för att ta ut reaktionsprodukterna måste såklart trycket sänkas. Detta görs genom att öppna flashventilen och låta gas strömma ut ur autoklaven. Men innan denna ventil kan öppnas bör värmaren slås av och gastillförseln stängas. Gasen bör stängas på autoklavens gaskran först för att undvika att autoklavens innehåll trycks bakvägen upp i gasslangen. OBS! Detta innebär att gasslangen är trycksatt, den måste följaktligen tömmas på gas genom att åter öppna gaskranen sedan trycket i autoklaven sänkts. Om så inte sker föreligger en skaderisk när gasslangen sedan kopplas lös. För att sänka autoklavens tryck 6

öppnas försiktigt flashventilen. Gasen strömmar ut och det minskande trycket kan övervakas på manometern. Stoppa och fäll undan omröraren, se till så att inget resterande tryck finns i autoklaven innan bultarna lossas. När sedan de bägge ringhalvorna lossats och lyfts bort är det dags att lyfta på locket. Detta moment bör utföras med stor försiktighet eftersom het ånga kan komma att frigöras då locket lyfts (använd personlig skyddsutrustning). Om locket sugit fast i packningen kan ett verktyg användas för att försiktigt bända lös locket.

Om provet fortfarande är mycket varmt bör det få stå och svalna ytterligare ett tag innan glaslinern varsamt lyfts upp (en annan möjlighet är att kyla provet genom att späda med kallt vatten). Provet kan sedan efterbehandlas på önskat sätt. Innan autoklaven lämnas måste den rengöras för att undvika att rester av korrosiva material finns kvar. Skölj av undersidan av locket, omröraren, dykröret och termoelementfickan. Det kan vara nödvändigt att skölja igenom dykröret med vatten, särkilt om prov tagits ut under försökets gång. Skölj insidan av autoklaven även om en glasliner har använts. Detta kan enkelt göras genom att fylla kärlet med vatten och sedan öppna dräneringsventilen i botten.

4. Allmänna råd och kommentarer

SäkerhetGällande användning av autoklaven skall sunt förnuft användas och kom ihåg att tänka efter före. Röraranslutningar på autoklavens lock där heta gaser och/eller vätskor kan komma ut såsom sprängblecket provtagnings- och flashnings-ventilerna bör för säkerhets skull vändas bort från alla personer som kan tänkas vistas i närheten.

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För att underlätta öppnande av autoklaven kan de slangar som finns anslutna monteras lös. Men i detta läge är det viktigt att komma ihåg att tömma eventuellt trycksatta gasslangar innan de monteras lös. Detta är viktigt att komma ihåg då gastillförseln strypts på kranen vid autoklaven innan kranen på gasflaskan stängts. Ibland kan det vara nödvändigt att göra så för att undvika att material trycks baklänges in i gasslangen då kranen på flaskan stängs.

Glaslinern och locketEftersom alla delar sitter ganska tajt inne i autoklaven kan det ibland vara fördelaktigt att fixera glaslinern med ett dämpande (värmetåligt) material för att undvika att den slås sönder av skakningar som uppkommer då omröraren körs. Någon typ av glasfibermaterial kan fungera för detta syfte. Hantera locket varsamt för att undvika att böja omröraraxeln eller skada manometern eller någon av de anslutna kranarna. Det är en god ide att se till att ha någonstans att lägga (de varma) metalldelarna, ringhalvorna och locket, innan autoklaven öppnas.

RengöringKranar och dylikt kan rengöras relativt enkelt genom att koppla en vattenslang till anslutningen och helt enkelt spola igenom alla rör och kranar med mycket vatten. Tänk på att skölja av alla skrymslen på insidan av autoklavens lock. Själva kärlet (tänk på att det kan vara varmt långt efter att värmaren stängts) kan fyllas med vatten borstas ur och tömmas genom bottenventilen. Lossa gärna omröraren från axeln för att kunna rengöra däremellan.

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copper, nickel and tellurium yields during leaching of anode slime

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