ES2574031B1 - RECOVERY OF HIGH PURITY SN BY ELECTROREFINO FROM SN ALLOYS CONTAINING PB - Google Patents

Abstract

Recovery of high purity Sn by electrorefin from Sn alloys containing Pb. # The present invention relates to a process of recovering high purity tin from tin alloys containing electrorefine lead. Therefore, the present invention falls within the technical sector of industrial waste recovery, specifically the use of electrical and electronic waste.

Description

The present invention relates to a process of recovery of high purity tin by electrorefin, from tin alloys containing lead.

Therefore, the present invention falls within the technical sector of industrial waste recovery, specifically the use of electrical and electronic waste.

STATE OF THE TECHNIQUE

After the dismantling of electrical, electronic circuits, devices, computers, etc., components containing tin are obtained, together with other metals. One way to handle them properly and reduce their volume is to melt them and obtain ingots. These ingots have little commercial value, since they are formed by tin and other metals (lead, silver and copper, mainly).

Industrially the separation and purification of tin (Sn) from alloys with a high content of 'lead (Pb) in the presence of other metallic elements such as copper and silver, is carried out through complex hydrometallurgical processes, which require a large number of stages, including stages of removal of polluting solvents [Studies on Metal (Cu and Sn) Extraction from the Discarded Printed circuit Board by Using Inorganic Acids as Solvents. Vijayaram R and Chandramohan K. J Chem Eng Process Technol, 4: 2, (2013) 1

Documents related to the separation and purification of Sn by electrorefin can be found in the literature.

An example is EP2548981 (A 1) which describes a method of obtaining ultra high purity Sn and ultra high purity Sn alloys with a very low alpha radiation content (sO.001 cph / cm2) comprising an electrorefin stage with which ultra high purity Sn materials suitable for the manufacture of semiconductor devices are obtained. Reducing the dose of alpha radiation of the materials

Ultra high purity Sn software errors of semiconductor devices are avoided.

Another example is CN101033557 (A) describes an electrorefin procedure for

5 recover Ag and Sn using a pure Sn cathode. The degree of purity of the Sn achieved in this process is not high, it also does not describe how to recover minor metals that are part of the starting sample, such as Ag and Cu, and how to separate Pb residues.

10 Therefore, it is necessary to develop a high purity tin recovery process that additionally gives value to other metals from alloys of different nature.

15 DESCRIPTION OF THE INVENTION

The present invention relates to a process of recovery of high purity tin from tin alloys containing lead by electrorefin.

In a first aspect, the present invention relates to the process for the recovery of tin (Sn) with a purity greater than 99% from an Sn alloy comprising Pb, characterized in that it comprises a step of recovering Sn by electrorefine which takes place in an electrochemical cell, where

The anode is formed by the starting alloy, the cathode by Sn with a purity greater than 99% the electrolyte is a solution of Sn in sulfuric acid comprising thiourea as an organic additive where the acid concentration is between 0.5 and 1.05M,

The concentration of thiourea is 0.13 M, and the concentration of Sn is 1.81.10-3 ± 0.05 _10-3 M; the electrolyte flow in the electrochemical cell being between 2.5 and 10 I / h

And where a current intensity is applied between the electrodes, between the anode and the cathode that form the electrochemical cell, between 0.6 and 1.2 amps_

Preferably, the electrolyte flow in the electrochemical cell between 4 and 7 1 / h.

By applying a direct current between the anode and the cathode, the tin is migrated from the alloy (anode) to the cathode, depositing on it. Using a

5 Sn cathode of high purity for the deposition of the Sn from the alloy further steps of separation of the pure Sn deposited from the cathode are avoided. The high purity Sn cathode with deposited high purity Sn is a direct product for use.

10 The use of Sn cathode instead of other metals, for example Ti, lowers the costs of the process, also allowing to work in H2S04 medium.

The present invention uses as a electrolyte a combination of a solution of H2S04 and thiourea as a crystalline growth inhibitor.

The use of H2 S04 electrolytes instead of Hel lowers the process and reduces the possible environmental problems resulting from the formation of chlorine vapors and the corrosion of the cathodes.

The operating conditions used during the electrorefin allow to treat Sn alloys comprising Pb, where Pb does not dissolve in the electrolyte. Pb is part of the anodic sludge.

Using a combination of a solution of H2S04 and thiourea is the most

25 effective in inhibiting the growth of dendritic or large crystals on the cathode surface in electrorefining processes and obtaining the highest separation performance of Sn.

In another preferred embodiment, the Sn alloy comprising Pb is an Sn-Pb alloy.

In another preferred embodiment, the Sn alloy comprising Pb is selected from Sn-Pb-Ag and Sn-Pb-Sb.

In a preferred embodiment, when the Sn alloy comprising Pb is an Sn-Pb-Ag alloy, the process described above comprises a step

additional to ") Ag recovery, after stage a) electrorefin, characterized by comprising the following stages:

a'1) Contact the anode sludge composed of impure Ag or a

5 Impure Ag-Pb alloy with concentrated nitric acid, a'2) Add distilled water to the solution obtained in step a), a'3) Filter a'4) Recovery of the Ag contained in the aqueous solution obtained in the

step a '3) by cementation with electrolytic Cu.

In another more preferred embodiment, the above process comprises an additional stage a) of recovery of the Cu contained in the aqueous solutions of the stage to O), characterized by comprising the following steps:

15 to "1) liquid-liquid extraction of Cu with an oxime, and" 2) of liquid-liquid re-extraction of Cu with sulfuric acid.

Preferably, the oxime of step a "1) is selected from the list comprising ketoxime, salicylalimexime or any combination thereof.

In another preferred embodiment, the method described above comprises a step of recovering Pb from the anodic sludge formed in the electrorefin step (a)

or of the anodic sludge after stage (a ") by conventional pyrometallurgical techniques.

Examples of conventional pyrometallurgical techniques for the recovery of Pb are the KALDO-TBRC process, the KIVCET process, QSL or the AUSMET / ISAMELT process.

The KALDO-TBRC process is specifically used to obtain lead from

30 of secondary materials and could be used for sludge treatment. In this process, which is discontinuous or flash, a TBRA converter (Top Blown Rotary With pouring) is used mixing oxidation and reduction stages through the use of reducing agents or coke.

Throughout the description and the claims the word "comprises" and its variants are not intended to exclude other technical characteristics, additives, components or steps. For those skilled in the art, other objects, advantages and features of the invention will be derived partly from the description and partly from the practice of the invention. The following examples and figures are provided by way of illustration, and are not intended to be limiting of the present invention.

BRIEF DESCRIPTION OF THE FIGURES

FIG. 1. Scheme of the facility used in laboratory level studies.

10 FIG. 2. Image of secondary electrons of the low Sb-Pb-Ag type alloy corresponding to another area of the sample.

FIG. 3. Image of secondary electrons of the Sn-Pb-Ag alloy with a high Pb content.

FIG. 4. Image of secondary electrons of the Sn-Pb type alloy.

FIG. 5. Image of secondary electrons from the tin deposit obtained at a current of 1.20 A and electrolyte flow: 5 Uh) (a and b).

20 FIG. 6. RX Diffraction diagram of the anodic sludge obtained after the electrorefin of an Sn-Pb-Ag type alloy of low Pb content.

FIG. 7. Image of secondary electrons of the anodic sludge (a and b).

FIG. 8. Image of secondary electrons from another area of the anode sludge sample (a and b).

FIG. 9. X-ray diffraction spectrum of the solid residue resulting from the thin layer leaching process.

FIG. 10. (a) Image of secondary electrons of the silver sample obtained by cementation on copper; (b) Image of high magnification (x6000) of secondary electrons of silver; (c) Image of high magnification (x6000) of electrons

35 backscattered silver.

FIG. 11. Image of secondary electrons of the silver sample corresponding to the zone of homogeneous morphology.

FIG. 12. X-ray diffraction spectrum of cemented silver.

FIG. 13. X-ray diffraction spectrum of the wash water precipitate.

FIG. 14. Image of secondary electrons from the washwater precipitate (a and b)

10 FIG. 15. RX Diffraction diagram of the anodic sludge obtained after the electrorefin of an Sn-Pb-Ag type alloy with a high Pb content.

FIG. 16. Secondary electron image of the anodic sludge obtained in the electrorefin of an Sn-Pb-Ag type alloy with a high Pb content (a and b).

FIG. 17. Image of secondary electrons of the silver sample obtained by cementation on copper (a, b).

20 FIG. 18. X-ray diffraction spectrum of the anodic sludge obtained after the electrorefin of an Sn-Pb type alloy.

FIG. 19. Image of backscattered electrons of the anodic sludge (a and b).

25 EXAMPLES

The invention will now be illustrated by tests carried out by the inventors, which demonstrates the effectiveness of the process of the invention.

30 1.-Experimental systems used in the trials.

1. 1.-Description of the installation used in the electrorefine process

The starting alloys are subjected to an electrorefin process to obtain high purity Sn (> 99%) and anodic sludges comprising Pb, from which, metallic Pb can be obtained by conventional pyrometallurgical processes.

FIG. 1 shows a diagram of the installation used in laboratory-level studies, with an arrangement of two cathodes (4) and an anode (5). The rectifier (6) is composed of a Heinzinger power supply model LNG 32-6 and a continuous signal recorder GRAPHTEC model GL220 midi Logger. The electrolyte contained in the electrolyte tank (1) is pumped by a drive pump

(2) towards the electrochemical cell (3).

A cell of 5.3 liters capacity was used. The anode (that is, the different alloys studied), had dimensions of 14x11x2.5 cm, and a weight of 3,110 kg. Tin cathodes had dimensions of 15x11xO, 1 cm and a weight of 200 g. approximately.

1.2.-Characterization of the Sn deposits obtained:

The electrolytic Sn deposits obtained in the different tests were washed with distilled water and dried in an oven at 80 ° C. The wash waters were stored for later analysis and evaluation. The Sn was melted at 1200 ° C in an electric oven equipped with a graphite crucible. The melt was cast to obtain an ingot whose chemical composition was determined by optical emission spectrometry by spark in a Spectro SPECTROMAXx model spectrometer.

1.3-Characterization of anode sludge:

The different anodic deposits (anodic iodine) were removed from the corresponding anodes and by Pearl X-ray Fluorescence to determine their chemical composition. An X-ray fluorescence spectrometer, model S8 Tiger by Bruker, was used for wavelength dispersion. The spectrometer is equipped with tungsten tube, UF analyzer crystals and 4 kW generator. The pearl will be prepared by adding to the sample (1 g) 10 g of flux (34% U metaborate + 66% tetraborate U) and 2 drops of detachment (LiBr). Subsequently, this was melted in a flame pearler, at a temperature higher than 875 ° C, which is the melting temperature of the flux. Previously the sample had been calcined at 1 year to constant weight to eliminate volatile compounds.

The X-ray Fluorescence technique has been used to perform the analysis of this research because it is a versatile, fast analytical technique that manages to reach detection limits of up to 0.002% (20 ppm).

The mineralogical composition of the different anodic sludges was studied by X-ray Diffraction, using a Siemens model 05000 diffractometer, equipped with a Cu anode (Cu Ka radiation) and UF monochromator to eliminate K3 radiation from the samples that They contain iron. The generator voltage and current were 40 kV and 40 mA respectively. The measurement has been carried out continuously with a step of 0.03 ° and a time of 3 s for each step. The interpretation of the diffractograms was carried out with the Match program.

Finally, a microstructural analysis of representative samples of the different anodic sludges was performed by Scanning Electron Microscopy (FE8EM) in a HITACHI model 8-4800, using a voltage of 15 kV. The microscopy sample was embedded in bakelite (EPOXI resin) and polished with sandpaper 600, 1200 and 2000 (adding carnauba to these to protect the sample). Subsequently, the sample was polished with 3 and 1 IJm diamond paste and metallized with carbon in a JEOL JEE 48 model.

1.4.-Recovery of silver through leaching and cementation

To extract the Ag contained in the anodic sludge, a thin layer leaching process is carried out, bringing different amounts of mud and 5 mL of nitric acid (NH03) concentrated into contact for 24 hours at rest.

Then 200 mL of distilled H20 was added and the mixture was stirred for 2

h. The suspensions obtained after stirring were filtered, collecting a solid residue and an aqueous solution containing the leached Ag.

The solid obtained after washing and drying in an oven (100oC / 24h), was analyzed to determine its chemical composition by Pearl X-Ray Fluorescence. The mineralogical composition was studied by X-ray diffraction, using a Siemens D5000 diffractometer.

In this study, the solutions from the thin-layer leaching processes containing Ag were reacted with Cu electrol in continuous and gentle stirring at 20 ° C, studying the kinetics of the reaction taking samples at different times. At the end of the process, the samples were filtered to collect the Cementing Ag on the copper and after washing it, dried in an oven (10QoC / 24 h).

Cemented Ag is analyzed by ICP-OES Inductive Coupling Plasma Spectroscopy, on an ICP-OES optical emission spectrophotometer (Varian model 725-ES), to determine its purity. Previously a fusion of the sample with lithium metaborate at 110QoC is carried out, diluting subsequently with water and nitric acid (4 mL of HNO :) in 100 mL of ultrapure water). Silver microstructural analysis was also performed by Scanning Electron Microscopy (FESEM).

In the liquid samples taken at different reaction times, the concentration of Ag and Cu, by Atomic Absorption Spectroscopy, was analyzed in a flame spectrometer with a hydride generator (Varian SpectrM 220 FS (Fast Sequential) model).

1.5.-Recovery of copper contained in aqueous solutions from the process of cementation of silver on copper by liquid-liquid extraction of copper.

For the extraction of the Cu contained in the aqueous solutions in acidic medium from the cementation of Ag on electrolytic copper, a solvent extraction process (or liquid-liquid) is carried out.

20 mL of each solution with Cu, were contacted with 20 mL of an organic solution, under continuous stirring for 10 min at 20 ° C. The organic phases used in solvent extraction tests:

•

40% Cetoxima (LlX 84) in Exxsol 0100

•

40% Salicylalimexime (ACORGA M5640) in Kerosene

•

40% Salicylalimexime + Cetoxime (LlX 973N) in Exxsol Dl00

It can be seen that the aqueous phase, which was initially blue due to the copper in solution, loses this color after 10 min of stirring because the copper is in the organic phase. After 10 min the aqueous phase was collected and the concentration of Ag and Cu was analyzed, by Absorption Spectroscopy

Atomic in a flame spectrometer with hydride generator (model SpectrM 220 FS (Fast Sequentíal) from VARIAN).

Subsequently, the re-extraction of retained copper in the organic phase has been carried out,

5 by contacting 20 mL of each organic phase charged with copper, with 10 mL of 2M H2S04, at 20 ° C maintaining stirring for 10 min. After 10 min the aqueous phase was collected and the concentration of Ag and Cu was analyzed, by Atomic Absorption Spectroscopy in the same equipment that was used for the extraction tests. It is observed that the aqueous phase has a blue color due to copper re

10 extracted or recovered.

2.-Characterization of the alloys starting alloys that form the anode.

The chemical composition of the starting alloys studied that constitute the

The anode and the chemical composition of high purity tin used as a cathode in the electrorefin process was determined by optical emission spectrometry by spark in a Spectro SPECTROMAXx model spectrometer. The alloys studied can be classified into three representative types, Sn-Pb, Sn-Pb-Ag and SnPb-Sb. Table 1 summarizes the chemical composition, expressed in% weight, of the

20 anode and cathode.

Table 1 Chemical composition of anode and cathode (% by weight) The morphology of the sub-surface of the different types of alloys were studied by scanning electron microscopy. FIG. 2. shows the surface of the low-Pb-type Sn-Pb-Ag alloy where Sn appears as a base element and Ag Y

Element

Anode (%) Cathode (%)

Alloy No. 4 Type Sn-Pb-Ag Low cont. Pb

Alloy No. 2 Type Sn-Pb-Ag With !. Half Pb Alloy No. 7 Type Sn-Pb Alloy Type Sn-Pb-Sb Low cont. Pb

Sn

95.91 89.87 64.38 91, 90 99.83

Pb

0.32 7.38 35.49 0.1 <0.01

Cu

0.79 0.47 0.133 0.89 0.016

Ag

2.93 2.26 <0.05 <0.005 0.055

Neither

0.036 0.021 <0.001 <0.005 -

Sb

<0.05 <0.05 <0.05 7.10 0.038

Bi

0.011 <0.01 <0.01 0.003 <0.05

Faith

<0.01 <0.01 <0.01 0.03 0.033

CD

<0.005 <0.005 <0.005 <0.005 <0.005

In

<0.005 <0.005 <0.005 <0.005 0.007

Ge

<0.005 <0.005 <0.005 <0.005 -

Zn

<0.002 <0.002 <0.002 <0.005 <0.005

5 Cu as major elements in the chemical composition of the alloy (See Table 1). Cu appears mainly in circular deposits of a darker color than the rest of the surface. The silver is distributed mainly forming a kind of chains with a rough appearance and lighter color, formed in turn by particles of small size.

10 FIG. 3. shows the surface of the Sn-Pb-Ag type alloy with a high Pb content where two zones of different hue are distinguished, corresponding to the two major elements of the alloy (Sn and Pb). The Sn appears as a base element being distributed throughout the entire sample and the Pb is mainly

15 in light colored areas. The Ag contained in the alloy is distributed mainly forming a kind of chains.

In FIG. 4 a homogeneous morphology of the Sn-Pb type alloy is observed in which two zones of different hue are distinguished, corresponding to the two

20 major elements of the alloy (Sn and Pb). The Sn appears as a base element being distributed throughout the entire sample and the Pb is distributed in light colored areas.

3.-Electrolyte

The base electrolyte was prepared from a solution of Sn in H2 S04. For this, tin shot was dissolved in 18 M sulfuric acid (1 L of H2S04 per 250 g of tin shot). The mixture was stirred for 72 hours and subsequently the precipitate formed (SnS04) was separated. An aliquot of the remaining solution was diluted with

distilled water, obtaining an electrolyte with a concentration of 0.85 M H2S04 and a concentration of 215 ± 6 mg / L Sn (1.81 -10.3 ± 0.05 -10.3 M). This value was obtained by optical emission spectrometry with ICP excitation source (ICP-OES), using a Perkin Elmer spectrometer, model OPTIMA 3300 DV. To the electrolyte solution 5, 1 gIL of organic inhibitor / additive agent was added. This solution constitutes the initial electrolyte arranged in the tank. The tank was filled with 3,840 liters of electrolyte and the flow regulator tank, with 5 liters. The cathode area in

Contact with the electrolyte was, in all the tests carried out, 110 cm.

10 3. 1.-Choice of organic additive

The choice of the organic electrolyte additive was carried out by adding 1 gIl of organic additive to the electrolyte and evaluating the tin yield obtained by electrorefin when starting from an Sn-Pb type alloy. Table 2 compiles the

15 results obtained for the different organic additives tested and the yield obtained with each of them

Table 2. Performance in Sn (gth)

Organic Additive (1g / l)

Yield in Sn (g / h)

Thiourea

2.04 (Final -96 h)

Horse

1.96 (Final -72 h)

Polyacrylamide

1.27 (Final -54 h)

Beta-naphthol

0.80 (Final 72 h)

20 The best performance was obtained for thiourea.

4. TYPE Sn-Pb-Ag OE LOW TYPE CONTENT IN Pb.

25 4.1.-Influence of the current intensity and electrolytic flow 4.1.1.-Electrorefin tests in discontinuous

The influence of the current intensity applied to the electrodes was evaluated under the 30 experimental conditions that are compiled in Table 3 using an Sn-Pb-Ag type alloy of low Pb content.

Table 3. Experimental conditions used to determine the influence of the current intensity in the final deposit of Sn.

Test Current intensity (A) ODS 0.6 SA1 1.2 SA2 0.9 SA3 0.75

Experimental conditions Separation Flow Density cathode-anode (A / m ') current (Uh) (cm) 54.5 10 4 109.1 10 4 81.8 10 4 68.2 10 4 Time (h) 24 24 24 24

For an intensity of 0.60 A, a deposit with a small amount of sample and not homogeneous is observed. Increasing the intensity shows an increase in the amount of Sn deposited. From an intensity of 0.90 A, more homogeneous, uniform and thicker deposits are observed. The increase in current intensity produces greater formation of dendritic crystals, mainly in the three tests carried out at lower intensity, with large dendrites clearly visible in the lower part of the cathode in the test tank carried out at 0.90 A. For an intensity of 1.2 A the deposit obtained is homogeneous and uniform with preferential areas of crystalline growth in the form of

15 acicular crystals, but without the presence of large dendrites as in the case of the deposit generated in the test conducted at 0.90 A.

The tests were performed discontinuously. After each test, the deposit produced in the cathodes was removed, washed with distilled water and dried in an oven to

20 80 ° C. Subsequently, they melted and strained and the ingot purity was determined. The anode was also cleaned by removing the anodic deposit generated in the assay and analyzed as described above.

Table 4 shows the chemical composition of the Sn deposits obtained in function of the current intensity, expressed in A, for a constant flow of electrolyte of 10 Uh and a separation between anode-cathode of 4 cm

Table 4: Chemical composition, expressed in% by weight, of tin deposits

depending on the current intensity and amount of Sn deposited in the cathode

(electrolyte flow: 10 Uh, anode-cathode separation: 4 cm)

Test

Intensity Tin Silver Copper Lead Iron Sulfur Deposit

from

(%) (%) (%) (%) (%) (%)

from

stream

tin

(TO)

(g / h)

SAO

0.60 - - - - - - 0.44

SA1

1.20 99.95 <0.01 <0.005 0.048 <0.01 0.01 2.15

SA2

0.90 99.93 <0.01 <0.005 0.040 <0.01 0.01 1.69

SA3

0.75 99.90 <0.01 <0.005 0.028 <0.01 0.01 1.29

* The chemical composition of the tin deposit obtained in the SAO test was not analyzed due to the small amount of deposit obtained, which proved insufficient to perform the analysis by optical emission spectrometry by spark.

10 It is observed that the element that is mobilized in greater proportion and that constitutes the most important impurity is Pb. Keep in mind that the electrochemical potential of this element (-0.13 V) is similar to that of Sn (-0.14 V). The increase in current density (Alm2) produces an increase in the amount of Sn

15 deposited. For an intensity of 0.60 A (54.5 Alm2), the amount of Sn deposited equals 0.44 g / h, which increases to 1.69 g / h for 0.90 A (81, 8 Nm ' ) and 2.15 g / h for a current intensity of 1.2 A (109.1 Alm2).

The results in Table 4 indicate a higher purity of the deposited Sn when increasing

20 the current intensity, although this increase has associated defects in deposition due to the formation of dendritic crystals, especially in the three tests carried out at a lower current intensity.

The influence of electrolyte flow was evaluated under the experimental conditions that

25 are compiled in Table 5 using a low Pb content Sn-Pb-Ag alloy.

Table 5. Experimental conditions used to determine the influence of electrolyte flow in the final Sn deposit.

Test Current intensity (A) SA1 1.2 SA4 1.2 SA5 1.2

Experimental conditions Density of Separation Flow current cathode-anode electrode (Alm ') (Uh) (cm) 109.1 10 4 109.1 5 4 109.1 2.5 4 Time (h) 24 24 24

5 Table 6 shows the chemical composition of the Sn deposits as a function of the electrolytic flow, for a current intensity of 1.20 A and an anodecathode separation of 4 cm.

Table 6. Chemical composition, expressed in% by weight, of Sn deposits in

10 function of the flow of the eleetrolite and quantity of tin deposited in the cathode (current intensity: 1.20 A, anode-cathode separation: 4 cm).

Test

Electrolyte flow (Llh) Sn (%) Ag (%) Cu (%) Pb (%) Faith (%) S (%) Sn deposit (gth)

SA1

10 99.95 <0.01 <0.005 0.048 <0.01 0.01 2.15

SA4

5 99.92 <0.01 <O 005 0.055 <0.01 0.01 2.33

SA5

2.5 99.93 <0.01 <0.005 0.044 <0.01 0.02 2.39

It is observed that the element that constitutes the most important impurity is again Pb. The increase in electrolyte flow produces a small decrease in the amount of tin deposited.

For a flow of 5 L / h, a uniform and homogeneous deposit with small acicular crystals but without dendrites is observed. If we compare this deposit with that obtained 20 with a flow of 10 Uh. it is appreciated that it is less agglomerated and the tin obtained is of higher quality since the acicular crystals are smaller and there are no dendritic formations. For a flow 2.5 Uh the deposit obtained is homogeneous but less uniform than in the case of flow of 10 and 5 L / h. With the flow of 2.5 Llh you

they can appreciate more quantity of acicular crystals, especially in the lower part of the cathodes, which form certain dendrites in the lower part of these.

Figure 5a and 5b shows the morphology of the Sn deposit obtained by

5 Scanning Electron Microscopy (FESEM) in a HITACHI model 8-4800, using a voltage of 15 kV, for the sample of Sn obtained in the SA4 test. The tin sample has homogeneous laminar morphology. In the EDS spectrum, no signals corresponding to elements other than tin are observed.

10 Studying these results, taking into account not only the composition of the tin deposits but also their morphology, it follows that the optimal operating conditions in the tests carried out on a laboratory scale are obtained for a current intensity of 1.20 A, an anode-cathode separation of 4 cm and

15 an electrolytic flow of 5 L / h.

4.1.2.-Continuous electrorefin tests

Different tests were carried out continuously using an Sn-Pb-Ag alloy of

20 low Pb content until the anode studied is exhausted. In total, 845 h of electrodeposition test was required until the anode used was exhausted.

Table 7 shows the list of tests carried out.

25 Table 7. Description of continuous tests. Operating conditions

Test SA6 SA7 SA8 SA9 SA10

Current intensity (A) 1.20 1.20 1.20 1.20 1.20 Test characteristics Separation Flow Density current electrolyte cathode-anode (Alm ') (Uh) (cm) 109.1 5 4 109.1 5 4 109.1 5 4 109.1 5 4 109.1 5 4 Time (h) 149 96 72 96 72

SA1 1 SA12 SA13

1.20 1.20 1.20 109.1 109.1 109.1 5 5 5 4 4 4 72 72 72

The operating conditions were: current intensity 1.20 A, electrolyte flow 5 Llh and anode-cathode separation 4 cm, with duration of the test the number of hours indicated in Table 7, discharging the cathodes of the Sn tank every 24 5 h.

Throughout the test period, the current intensity remained constant at 1.20 A. The voltage tends to decrease as the reaction time increases, especially at the beginning of the test. During the first hour of rehearsal the

10 voltage tends to decrease and then stabilizes until the end of the test. This trend was also maintained in the trial that lasted 6 days (SA6) (discharging cathodes every 24 h).

Table 8 shows the chemical composition of the deposits obtained and the amount

15 of Sn deposited, compared to the same results of the equivalent discontinuous test (SA4). It is noted that the operation of the installation in continuous mode provides greater purity of the Sn, as well as a similar amount of tin deposit per test hour.

20 Table 8. Chemical composition of the Sn deposits obtained continuously and discontinuously · depending on the current intensity and amount of Sn deposited at the cathode (current intensity: 1.20 A, anode-cathode separation: 4 cm )

Test

Electrolyte flow (Llh) Sn (%) Ag (%) Cu (%) Pb (%) Faith (%) S (%) Tin deposit (gth)

SA6

5 99.96 <0.01 <0.005 0.016 <0.01 0.01 2.33

SA7

5 99.96 <0.01 <0.005 0.013 <0.01 0.01 2.26

SA8

5 99.94 <0.01 <0.005 0.015 <0.01 0.02 2.32

SA9

5 99.95 <0.01 <0.005 0.016 <0.01 0.03 2.20

SA10

5 99.95 <0.01 <0.005 0.015 <0.01 0.02 2.36

SA11 SA12 SA13 SA4 '

5 5 5 5 99.96 99.95 99.95 99.92 <0.01 <0.01 <0.01 <0.01 <0.005 0.008 <0.005 <0.005 0.014 0.015 0.017 0.055 <0.01 <0.01 <0.01 <0.01 0.02 0.01 0.02 0.01 2.31 2.47 2.27 2.32

A study of the cathodic deposits was carried out as a function of time, demonstrating that the greatest amount of the electrodeposit is produced in the first 6 hours of each 24-hour process, with little change in appearance after these

5 hours, obtaining a homogeneous deposit. In the deposits there are hardly any dendrites and the tin crystals are smaller than 1 cm.

All deposits obtained after 24 hours until the corresponding hours of each trial are completed are very similar; The results show that, in 10 operating conditions used in continuous tests, it is possible to obtain a good tin deposit with very low levels of impurities.

Applying the Faraday Law, in the case of the test in which a greater quantity of tin has been obtained = 177.56 g (SA 12), the theoretical weight of deposited tin should have been 191, 32 g, which indicates that, under the conditions tested, 92.81% of the theoretical weight of Sn has been obtained.

A summary of the process balance is presented in Table 9

20 Table 9. Summary of the balance of the tin recovery process contained in the low Pb Sn-Pb-Ag alloy.

Sn in initial alloy (g)

Sn recovered by electrorefin (g) Sn (%) Sn recovered (g) Sn recovered from the initial alloy (%)

2284.58

1871, 30 99,944 1870.23 81.9

"Total mass of electro-refined Sn after 864 h of process until the anode runs out Under the experimental conditions used, the recovery of the Sn contained in the initial alloy has been 81.9% (785.2 kg Sn / t, initial alloy) .19

During the entire electrorefin process, the same electrolyte was used, maintaining the pH value between 0 and 1. The amount of final Sn in the electrolyte was determined by Atomic Absorption Spectroscopy. The analysis of the results, comparing the values of the chemical composition before and after the process, indicates that a part

5 of the Sn contained in the alloy passes to the electrolyte. In this case, it has been determined that the amount of Sn that passes into the electrolyte is 242.37 g., Which implies a loss of 10.61% of the Sn contained in the initial alloy.

4.2.-Characterization of the anode / ode

10 Iodine is a powdery solid, blackish in color, deposited on the surface of the anode and must be constituted by the impurities that accompany the Sn. The wear of the anode was approximately 21.3 IJm thickness / h and the weight of sludge obtained in the electrorefin process until the anode studied was exhausted was 193.03

15 g This weight is adjusted to the ratio 0.10 kg sludge / kg Sn deposited in the cathode.

The composition of the anodic sludge (Table 10) shows that the Ag contained in the initial alloy has been concentrated in the sludge after the Sn extraction process, reaching a value of 34.2% by weight. In smaller percentage they appear

20 also Cu (9%) and Pb (0.9%), with 13.8% of 0, indicating the presence of oxides in its mineralogical composition. FIG 6 shows the RX Diffraction diagram of the Iodine.

Table 10. Chemical composition of the anodic sludge, expressed in% by weight (only 25 major elements,% weight 2 ::: 1.00).

Element

Anodic sludge (%)

Sn

41, 1

Ag

34, 16

Cu

9.1

Pb

0.91

OR

13.17

It is observed that the sludge is formed by the following crystalline phases: thermetallic compounds (Ag, Sn) (record: 00-071-0530) and (Cu, "Sn,) (record: 00-047-1575) and

casiterita (8n02) (record: 00-041-1445). The results of the chemical composition of the mud are consistent with the result obtained in the mineralogical study of the sample.

5 The microstructural analysis of the anodic sludge sample by Scanning Electron Microscopy (FESEM) is shown in Figures 7 and 8. A heterogeneous aspect surface formed by particles of different size and appearance is observed. Figures 7a and 8a indicate some areas where the majority elements contained in the mud (Sn, Ag and Cu) are found, as the studies of

10 EDS. These elements are distributed more or less uniformly throughout the sample.

4.4.-Recovery of Ag by thin layer leaching and cementation

15 Table 11 shows the amounts of mud and nitric acid used in the leaching process.

Table 11 Amounts of anodic sludge from the Sn-Pb-Ag alloy with low Pb content and concentrated nitric acid (NH03) used in various 20 thin-layer leaching tests.

Test

Anodic mud NH03 cc

(reference)

(g) (mL)

M-A

5 5

M-B

10 5

A4-C

2.5 5

M-O

2.5 + 2.5 " 2.5 + 2.5 '

I

5 5

"The process is carried out in two stages of 24 h each.

25 Table 12 shows the cementation tests carried out and the operating conditions used.

Table 12. Cementation tests performed with each solution obtained in the leaching process.

Test

Weather Dissolution volume Copper mass used

(reference)

(min) (mL) (g)

A4-A

0, 10, 30, 60, 120 95 1.76

A4-B

0, 10, 30, 60, 120 97 6.04

A4-C

0, 10, 30, 60, 120 97 4.40

A4-O

0, 15, 30, 60, 120 100 1.53

A4-E

0, 10, 30, 60, 120 96 8.62

At the end of the process the samples were filtered to collect the silver that cements on the Cu and after washing it, dried in an oven (10QoC / 24 h).

Table 13 shows the chemical composition of the residue obtained in the thin layer leaching process.

Table 13. Chemical composition of the residue obtained in the leaching process in thin layer (only major elements,% weight ~ 1.00).

Element

Waste of lixi viación Leaching residue

(LA4-O) (%)

(LA4-0 f (g)

Sn

57.58 78.02

Cu

12.58 17.05

Ag

1.34 1.82

Pb

1.00 1.36

OR

20.48 27.75

P.xC

0.561 -

P.xC. = Losses due to calcination ·· leaching residue mass (LA4-D) = 135.51 9

It is observed that the percentage of Ag that has remained in the solid residue is 1.34% (1.82 g), which represents approximately 2.77% of the Ag in the starting anode sludge; the rest of the Ag contained in the sludge being dissolved, so that 96.60% of the Ag that had the initial anodic sludge has leached out

to dissolution. To extract the silver contained in these solutions, a process of cementing Ag on electrolytic Cu was carried out.

FIG. 9. shows that the crystalline cassiterite phases exist in the residue studied

5 (SnO,) (record: 00-041-1445) and an intermetallic compound (Cu, Sn,) (record: 00-0451488), formed by the two major elements present in the residue. The results of the chemical composition of the solid leaching residue are consistent with the result obtained in the mineralogical study of the residue.

10 Table 14 shows the contents of Ag and Cu in the samples obtained in the cementation process at different reaction times.

Table 14. Concentration of silver and copper in solution in the samples at different reaction times and A9ac ratio (initiates IYCU (tC (final) at the end of the cementation process 15 (t = 120min).

Sample (reference)

Time {min) P1ata {gIL) Cemented Silver (%) Copper {SIL) Agac (nC.! ICuac (! Ni)

AA · A

OR 5.24 ± 0, 05 100.00 O, 20 ± 0, Ol

10

4, 31st, O ~ 82, ~ O, 63 ± O, Ol

30

2.72 ± O, O ~ ~ 1, 91 1.30 ± 0.01

60

1.28 ± O, 02 24.43 1.86 ± 0, O2

120

28.9.0.05 mg / L 0.55 2.52 ± 0, O2 2.08

AA · B

OR 8.33 ± O, 05 100.00 O, 15 ± O, Ol

10

7, OO ± O, 05 84.03 O, 95 ± O, Ol

30

3.85 + 0.05 40.22 1.92 + 0.01 m

AA-C

60 120 O 2.43 ± O, 05 40.2 ± O, 05 mg / L 2.86 ± O, 05 29.17 0.48 100.00 3.23 ± 0, O2 4.44 ± 0, O2 98.4 ± 0.5 mgIL 1.90 [/) N '"'" ..

N ...

10 30 1.43 ± O, 02 0.17.0.02 ~, OO ~, 94 O, 69 ± O, Ol 1.18 ± O, O1 or eN

60

34.2 ± 0.5 mgll. 1.20 l, 18 ± O, Ol P

120

3.1.0.1ng / L 0.11 l, 32 ± O, Ol 2.17

A4. [)

OR 8.54 ± O, 10 100.00 O, 22 ± O, Ol

fifteen

5.64 ± 10, 66.04 O, 73 ± O, Ol

30

4.26 ± 10 49.90 1.07.0.02

60

2.44 ± O, 05 28.57 l, 58 ± O, O2

120

40, O ± O, 05 mglL 0.47 4.20 ± 0, O2 2.03

AAE

OR 5.9O ± O, 05 100.00 O, 16 ± O, OO

10

2, ~, O ~ 48.31 l, ~ 7th, 02

30

O, 12 ± O, OO 2.03 2.51.0.02

60

5.84 ± O, Ol mg / L 0.10 2.57.0.02

120

0.67.0.01 mgIL 0.01 2.67.0.02 2.21

The thinner leaching process corresponding to the A4-D sample was chosen as the most efficient process to leach the silver contained in the Iodine, because assuming that the leaching has been carried out at 100%, this process is the one that has leaching virtually all of the silver (94.54%) contained in the initial alloy (or

5 anode).

In all the samples studied, it is observed that increasing the time of the cementation process increases the amount of cemented silver on the electrolytic Cu, because the concentration of silver in solution is decreasing. After

10 120 min of process, the concentration of silver in solution (Cemented cement) is very small (between 0.55% and 0.01% of the initial concentration). At the same time as Ag cements, the concentration of Cu in solution increases, being in all samples, after 120 min, the A9ac ratio (initial ¡fCUac (final approximately 2.

15 Table 15 shows the influence of the amount of Cu used to perform the cementation process in relation to the Ag existing in the initial solution.

Table 15. Percentage of Ag cemented at 60 and 120 minutes according to the Ag / Cu molar ratio used in the cementation process.

Sample

Ag / Cu Cemented Silver (%) Cemented Silver (%)

t = 60 min

t = 120min

A4-A

0.165 75.6 99.4

A4-B

0.079 70.8 99.5

A4-C

0.037 98.8 99.9

A4-0

0,328 71.4 99.5

A4-E

0.038 99.9 100

It is observed that at 120 min of reaction the extraction of Ag (as cemented Ag) is greater than 99.45% in all tests. The percentage of Ag extracted

or recovered from the anodic sludge has been 99.5% considering the thin layer leaching process that was chosen as more efficient (A4-D).

It is also shown that the cementation process produces a higher percentage of cemented Ag for t = 120 min the lower the Ag / Cu molar ratio. Therefore, for the same concentration of Ag in the initial solution the cementation will be more

favored for larger quantities of Cu. The mass of Cu used also influences the speed of cementation of Ag. For a cementation time of 60 min, it is observed that the most effective processes occur in samples A4-C and A4-E, which are the lowest Ag / Cu molar ratio used. In both cases, the

5 Ag recoveries are greater than 98%.

Table 16 shows the process balance.

Table 16. Summary of the recovery process balance of the silver contained in the initial alloy from which the anodic side comes.

Initial Alloy Silver (9)

69.79

Silver in Anodic Jodo (9)

65.94

Silver leaching residue (9)

1.82

Silver in leaching solution (9)

64, 12

Cemented Silver (%)

99.53

Silver extracted (9)

63.82

Silver extracted from the mud (%)

96.78

Silver recovered from initial alloy (%)

91.45

The designed process allows 97% of the silver contained in the anodic sludge to be recovered, which represents a recovery of 91% of the Ag existing in the starting alloy 15, that is, 26.7 kg.Ag/l. alloy

The chemical analysis of Ag, performed by Inductive Coupling Plasma Spectroscopy (ICP), indicates that the impurity content (Sn and Cu, mainly) is below 0.4%, thus presenting a high degree of purity (99.6%).

20 The study conducted by scanning electron microscopy shows images of the morphological appearance of the cemented Ag sample (FIG. 10). In FIG. 10. A homogeneous aspect surface is observed, formed by particles of different sizes (Fig. 10a). The image of backscattered electrons (10c) shows a

25 homogeneous composition of the sample.

Figure 11 shows an image of secondary electrons of the silver sample corresponding to the homogeneous morphology zone. In the chemical composition only the presence of Ag is observed.

5 The X-ray diffraction spectrum of cemented silver is shown in Figure 12. The only crystalline phase that appears is silver (Ag) (tab: 03-065-2871).

4.5.-Recovery of liquid-liquid extraction of Cu contained in aqueous solutions from the process of cementing Ag on Cu

10 Tables 17 and 18 show the Cu and Ag contents in the samples obtained after the extraction-re-extraction process under different operating conditions.

15 It is observed that the percentage of Cu extracted in the three cases studied, is higher than 92%, being in sample A4-B where the highest percentage of Cu re-extracted (99.24%) is obtained. Therefore, the most effective extraction agent is that formed by the 40% LlX 84 mixture in Exxsol D100. In this case, re-extraction with H2S04 2M solution allows 99% of the copper contained in the phase to be recovered

20 organic

Table 17. Concentration of Cu in solution after extraction extraction process.

Sample from which copper dissolution comes

CUolution after extraction (mg / L) CUextracted (%) CUsolution after re-extraction (giL) CU, ee>: t, acid (%)

A4-B (Ag: 40.2 ± 0.05 mg / L, Cu: 4.44 ± 0.02 giL) A4-e (Ag: 3.1 ± 0.1 mglL, Cu: 1.32 ± 0 , 01 giL)

344.0 ± 0.5 69.8 ± 0.5 92.25 94.71 8.13 ± 0.02 2.20 ± 0.01 99.24 88.00

A4-E (Ag: 0.67 ± 0.01 mg / L, eu: 2.67 ± 0.02 giL)

66.6 ± 0.5 97.51 3.28 ± 0.02 63.00

In the case of silver (Table 18) the extraction is greater than 71% when the extraction agent formed by 40% LlX 84 is also used in Exxsol 0100. The reextraction with H2S04 2M is higher for sample A4-E, obtaining 23% of re-extracted silver.

Table 18. Concentration of Ag in solution after extraction extraction process.

Sample from which copper dissolution comes

AQsolllción after extraction (mg / L) Agexlraide (%) Ag $ Oluation after re-extraction (mg / L) Ag, ee>: t, aida (%)

A4-S (Ag: 40.2 ± 0.05 mg / L, eu: 4.44 ± 0.02 giL) A4-e (Ag: 3.1 ± O, 1 mg / L, eu: 1.32 ± 0.01 giL) A4-E Ag: 0.67 ± 0.01 mg / L, (eu: 2.67 ± 0.02 giL)

11.3 ± 0.2 0.82 ± 0.01 0.17 ± 0.01 71, 89 73.54 74.63 0.91 ± 0.01 0.24 ± 0.01 0.23 ± 0.01 1.57 5.26 23.00

Because both Cu and Ag are contained in the solution from the cementation, the most efficient organic phase for extracting the Cu and Ag is 40% LlX 84 in Exxsol 01 00 which allows 92% of the Cu and 72% to be extracted of Ag. Re-extraction with H2S04 2M solution allows recovering the entire Cu (99.2%) while the

15 Ag remains mostly in the solution.

The designed process, allows to recover the u content in the resulting solutions

of the cementation process, obtaining a reusable organic phase and a phase

aqueous with Ag contents less than 1 mg / L.

4.6.-Study of the washing waters of the electrorefine process

The removal of the anodic and cathodic deposits obtained in the electrorefin process is carried out by means of a mechanical process (scraping). The deposits are subsequently washed to remove the remains of electrolyte that may have been deposited therein.

Washing waters are, therefore, solutions that may contain Ag and is

It is necessary to quantify its production and study its composition, with a view to

completion of the process on a larger scale.

10 The wash waters generated in the study of the alloy (until exhaustion), were stored in a tank. As a consequence of the pH variations, precipitation of a part of the Sn from the electrolyte occurs.

For the study of washing waters, a representative part of them, is

15 filtered on a pressure filter, obtaining a solid (precipitate) and an aqueous solution, which were analyzed separately. The precipitate was analyzed by X-ray Fluorescence and X-ray Diffraction, to determine its mineralogical composition. The solution was analyzed by Inductive Coupling Plasma Spectroscopy (ICP)

20 The total volume of washing water generated in the electrorefin treatment of the Sn-Sb-Ag alloy of low Pb content was 54.31 L, its pH being 1.2. After filtration, the precipitate weight was 66.80 g. The chemical analysis of the aqueous solution and the precipitate is shown in Table 19.

25 Table 19. Chemical analysis of the solid and liquid fractions obtained in the filtration of the Sn-Sb-Ag alloy wash waters of low Pb content.

Element

Concentration

Solid (% weight)

Liquid (mg / L)

Ag

0.00 0.00

Cu

0.03 0.27

Pb

0.00 1.71

Sn

67.12 0.08

It is observed that the Sn contained in the wash waters is concentrated in the precipitate, which is formed only by 8n02, as it is deduced from the study by RX Diffraction (FIG. 13), where a certain amorphous character is also observed of the oxide, no doubt due to the precipitation conditions. Study 5 by microscopy (FIG. 14), shows particles smaller than 1 I-Im and with spherical morphology, without the presence of impurities detected.

Taking into account the weight of the precipitate and its Sn content, it follows that the 54.3 L of wash waters contain 44.84 9 Sn. This means a loss of the Sn 10 existing in the electrolyte, which does not represent more than 2% of the total.

Therefore, it is possible to obtain Sn02 of high purity microcrystalline from a washing step, which would be a secondary product with commercial value.

5. TYPE Sn-Pb-Aq HIGH TYPE CONTENT IN Pb.

5.1.-Results of the electrorefin.

The tests carried out with the alloy studied were carried out continuously for 864 h, until the tested anode was exhausted. Table 20 shows the test operating conditions.

Table 20. Operating conditions used in the electrorefin of an alloy -Sn25 Pb-Ag with high Pb content.

Test

Trial characteristics

Current Intensity (A)

Current Density (Alm ') Electrolyte flow (Uh) Cathode-anode separation (cm) Time (h)

SAN2

1.20 109.1 5 4 864

Table 21 shows the chemical composition of the tin deposit obtained and the amount of tin deposited.

Table 21 Chemical composition of the tin deposit (expressed in% by weight) obtained continuously according to the flow of the eleetrolyl and amount of tin deposited in the cathode (current intensity: 1.20 A, anode-cathode separation: 4 cm).

Test

Electrolyte flow (Llh) Tin (%) Silver (%) Copper (%) Lead (%) Match (%) Sulfur (%) Tin deposit (gth)

SAN2

5 99,930 0.007 0.007 0.033 0.005 0.020 2.89

Sn electrodeposites have a homogeneous appearance, sometimes with

some acicular crystals located in the lower part of the cathode; they are not detected

dendritic formations. All deposits obtained during the 864h of

10 operation are very similar. These results show that, under the operating conditions used in the continuous tests, it is possible to obtain a good tin deposit with a purity of 99.93% (Table 21).

A summary of the electrorefine process balance, for an Sn-Pb-Ag 15 alloy with a high Pb content, is shown in Table 22.

Table 22. Summary of the balance of the tin recovery process contained in a high Pb Sn-Pb-Ag alloy.

Tin initial alloy (g) 2246.75

in Tin obtained Tin Tin by elecrorefin (%, weight) recovered (g) (g) '2056.99 9 99.930 2055.55 Tin recovered from initial alloy (%) 91, 50

twenty

"Total mass of electro-refined Sn after 864 h of process until the anode runs out

25

The initial Sn recovery. represents the 91.5% of the tin contained in the alloy

The same electrolyte was used throughout the electrorefin process, maintaining

the value of its pH between O and 1. The analysis of the final electrolyte, after 864 h of 31

operation, indicates a slight increase (3 gi L) of the concentration of Sn with respect to the initial Sn content.

5.2.-Characterization of the anodic Jodo

The wear rate of the anode was approximately 21.4 I-Im thickness / h and the weight of sludge obtained in the electrorefin process until the anode studied was exhausted was 398.47 g. The mass ratio is 0.19 kg sludge / kg tin deposited in the cathode.

The composition of the anodic sludge (LAN-2) is shown in Table 23.

Table 23. Chemical composition of the anodic sludge, expressed in% by weight (only major elements,% weight; :: 1.00).

Element

Anodic sludge (LAN-2)

('the)

Pb

49.85

Sn

15.06

Ag

13.50

Cu

3.78

It is observed that the anodic sludge is mainly formed by Pb, Sn and Ag Y in a smaller proportion, by Cu.

20 The mineralogical composition, determined by RX Diffraction (FIG. 15), indicates that it consists of intermetallic compounds of Sn-Ag-Cu, (Ag3Sn) (data sheet: 00-071-0530) and (Cu6.26SnS) (data sheet : 00-047-1575). Lead is concentrated as sulfate (anglesite (PbSO,) (data sheet: 00-036-1461)), and there is also Slavic oxide ((SnO,) (information: 00-033-1374).

25 The study by FESEM (FIG. 16) shows a heterogeneous aspect surface, formed by particles of different size and appearance. In FIG 16a, some areas where the majority elements contained in the mud (Pb, Sn, Ag and Cu) are found, confirmed by EDS analysis. There is a clear

difference in grain size and appearance in the area where there is lead, this being a smaller area of grain compared to areas where there is no lead.

5.3. Extraction of the silver contained in the anodic Jodo 5

Table 24 shows the chemical composition of the residue obtained in the thin layer leaching process. It is observed that the percentage of undissolved Ag is 0.10%. The Pb, together with the Sn, are retained in the leaching residue, and can subsequently be recovered by pyrometallurgical processes.

10 Table 24. Chemical composition of the residue obtained in the thin layer leaching process.

Element

Leaching residue (LA2-D) (%) Leaching residue (LA2-D) (g)

Sn Pb Cu Ag

23.08 17.77 0.26 0.10 58.86 45.32 0.66 0.26

15 • maS8, leaching residue (LA2.0) = 255.02 9

The solution from the thin-layer leaching process containing Ag was contacted with electrolytic Cu under continuous and gentle stirring at 20 ° C. Reaction kinetics is performed by taking samples at different times. Table 25 shows

20 can observe the operating conditions of the test carried out.

Table 25. Cementation process conditions.

Sample

A2-A

Time (min)

0, 10, 30, 60, 120

Solution Volume (mL)

98

Copper mass (9)

1.64

25 Table 25 shows, depending on the reaction time, the contents of Ag and Cu in the cementation solution. After 120 min of process, the

Ag concentration in solution (cementless Ag) is negligible (0.01% of the initial concentration).

Table 26. Concentration of Ag and Cu in solution at different reaction times.

Sample (reference) A2-A

Time (min) O 10 30 60 120 Ag (gi L) 3.78 ± 0.02 1.7HO, 02 0.644 ± 0.002 0.174 ± 0.00 1 0.395 ± 0.002 mg / L Cemented Ag (%) 100.00 45.24 17.04 4.60 0.01 Cu (giL) 33, HO, 50 mg / L 0.29 ± 0.02 0.85 ± 0.01 1.02 ± 0.01 1.39 ± 0.01

The speed of cementation is, in this case, very fast. After 60 minutes of reaction, the cemented Ag represents 95.40% of the initial Ag existing in the solution. After 120 minutes, the cementation of the Ag is total (Table 27).

Table 27. Percentage of cemented silver at 60 and 120 minutes.

Sample

Ag / Cu Cemented Ag (%) Cemented Ag (%)

t = 60 min

t = 120min

A4-A

0,130 95.40 99.99

Finally, Table 28 shows a summary of the balance of the 15 cementation process.

Table 28. Summary of the balance of the Ag recovery process contained in the initial alloy from which the anodic sludge is derived.

Silver in initial alloy (g)

56.5

Silver in anodic sludge (g)

53.79

Silver leaching residue (g)

0.26

Silver in leaching solution (g)

53.54

Cemented Silver (%)

99.99

Silver extracted (g)

53.53

Silver extracted from the mud (%)

99.52

Silver recovered from initial alloy (%)

94.75

Regarding the anodic sludge, the recovery of Ag is 99.5% and with respect to the initial alloy, 94.7%, which represents an Ag production of 21, 30kg / t. Alloy.

The chemical composition of Ag recovered by cementation, obtained by Inductive Coupling Plasma Spectroscopy (ICP), indicates an impurity content <1%.

10 Morphologically, Ag shows a particulate and homogeneous appearance (F1G. 17). The EDS spectrum does not detect significant impurities.

5.4. Liquid-liquid extraction of copper contained in aqueous solutions from the process of cementation of silver on copper

The extraction agent used was the 20% mixture of Oxima (ACORGA PT5050) in Exxsol D-100.

Table 29 shows the contents of Cu and Ag in the solutions obtained after the extraction-re-extraction process.

Table 29. Concentration of Cu and Ag in solution after the extraction and extraction process.

Sample from which the A2-A copper solution comes

(Ag: 0.395 ± 0.002 mg / L, Cu: 1.39 ± 0.01 giL)

CUsolution after extraction (giL)

0.137 ± 0.002

CUextraióo (%)

95.07

CUsolution after re-extraction

2.50 ± 0.01

(giL)

C UreeKlraido (%)

99.76

Agsolution after extraction (mg / L)

0.78 ± 0.01

Agextraid (%)

1.27

It is observed that the percentage of Cu extracted in the organic phase is 95.07% and its re-extraction with 2M H2S04 solution is greater than 99%.

5 The content of Ag in the solution after extraction is 1.3. In this way, it is possible to extract the Cu contained in the starting aqueous solution in a high percentage (95.07%), cleaning the organic phase from Cu and allowing its reuse. The aqueous phase has a residual Ag content of 0.39 mg / L.

10 6. TYPE Sn-Pb.

6.1.-lfluence of the concentration of H2 S04 / ibre on the electro / ita.

Various electrolytes with different concentrations of H2S04 were prepared. All the

15 electrolytes had pH <1. To all electrolytes were added 19 / L of Tiourea {thiourea concentration = 0.13 M) and electrolysis of an Sn-Pb type alloy was carried out. under the following conditions (1: 1.20 A; electrolyte flow: 5 I / h; anode / cathode separation: 4 cm).

In each electrolyte the concentration of free acid was determined, determining the Total Acid Number (TAN), which expresses the mi of NaOH of concentration 0. 1 N necessary to neutralize 10 ml of electrolyte solution.

So the total acid number is expressed in me. Taking into account the definition, it can be established that:

VNaOH N NaOH = VEleClrOlito. N Electrolyte

NE1e <: lffililO = (VNaoH. NNaoH) NE1e <: troll

30 According to the definition of TAN: VNaoH = TAN (mi), NNam.¡ = 0.1N, VEte <: lrotito = 10 mi;

NE "" "" ,, (N) = (SO mi. 0.1 N) / 10 mi = SO · 0.01

For sulfuric acid NE ~ ctfomo (N) = M Elactrolyte (M) "2; M E1e- = trol ito (M) = NE1e <: trolilJ2

At the end of each electrolysis test, the deposited tin yield was determined, obtaining the values set out below in Table 30:

10 Table 30: yield of Sn deposited as a function of the total acid number and / or concentration of the acid (M).

SO (mi)

[acid] (M) Yield Sn deposited (% weight)

100 150 196 206 214 250

0.5 0.75 0.98 1.03 1.07 1.25 60.7 83.7 93 75.4 57.2 35.2

To obtain a high yield in Sn, the electrolyte must contain a certain concentration of free sulfuric acid, expressed by the measurement of TAN and its corresponding molar concentrations.

6.2.-Results of the electrorefin.

20 The tests carried out were carried out continuously for 264 h, until a loss of the anode mass of 1020 g was achieved. (25% of the initial mass). The operating conditions are summarized below:

• Current intensity (A): 1.20

• Current density (Alm2): 109.1 25 • Electrolyte flow (Llh): 5

•

Cathode-anode separation (cm): 4

•

Time (h): 264

The speed of the electrorefin process is, in this case, 3.9 g / h (Table 31), obtaining a metallic tin of 99.93%.

Table 31 Chemical composition of the tin deposit (expressed in% by weight) obtained continuously in the electrorefin of an alloy type Sn-Pb.

Test

Electrolyte flow (Llh) Sn (%) Cu (%) Pb (%) P (%) S (%) Sn deposit (g / h)

SAN7

5 99,930 0.002 0.029 0.007 0, Q30 3.86

The cathodic deposits are formed by large sheets of metallic Sn, homogeneous, without dendrite formation.

10 The balance of the process (Table 32) indicates that the recovery of Sn represents 83.7% of the tin contained in the initial alloy.

Table 32. Summary of the balance of the recovery process of the Sn contained in the initial alloy.

Initial alloy Sn (g) Electro-refined Sn (g) Sn (%) Recovered Sn (g) Recovered initial alloy Sn (%)

656.65 550.21 9 99.93 549.82 83.73

· Total mass of electro-refined Sn after 264 h of process until a loss of the anode mass of 1020 g is achieved

20 During the entire electrorefin process, the same electrolyte was used, maintaining the pH value between 0 and 1. The electrolyte analysis, after 264 hours of operation, indicated an increase in the concentration of Sn equivalent to 7.6 gil.

6.3. Anodic mud characteristics

The sludge obtained after the electrorefining process is a solid with a dense, compact appearance, very attached to the anode surface and grayish in color. Production 5 of the sludge was 0.84 kg sludge / kg tin deposited in the cathode.

The chemical composition (Table 33) shows a clear enrichment in Pb, which reaches 77% by weight. This iodine can be treated by pyrometallurgical techniques for the recovery of lead.

10 Table 33. Chemical composition of the anodic sludge (only major elements,% weight 2 ::: 0.48) obtained in the electrorefin of an alloy of the Sn-Pb type.

Element

Anodic sludge (LAN-4) (%)

Pb

76.60

Sn

13.00

Cu

0.48

15 From a mineralogical point of view, the sludge is formed by lead sulfate (anglesite (PbSO.) (File: 00-036-1461)), tin oxide (SnO,) (file: 00-033-1374), and to a lesser extent, tin (Sn) (record: 00-004-0673) and tin sulfate (00-0170935) (FIG. 18).

20 Molfologically, mud is a heterogeneous material. The concretion zones of the Sn have two different molfologies. A globular mass, probably of metallic Sn and particles of size between 10-20 IJm, probably of Sn02 (FIG. 19). Pb is concentrated in the form of small, dark particles of hexagonal molfology, included in the surface of the Sn02 grains.

Claims (6)

1. Procedure for the recovery of tin (Sn) with a purity greater than 99% from an Sn alloy comprising Pb, characterized in that it comprises a step of recovering Sn by electrorefining which is carried out in an electrochemical cell where the anode is formed by the starting alloy, the cathode by Sn with a purity greater than 99% The electrolyte is a solution of Sn in sulfuric acid that comprises thiourea 10 as an organic additive where the acid concentration is between 0.5 and 1.05 M, the concentration of thiourea is 0.13 M, And the concentration of Sn is 1.81 · 10 · 3 ± 0.05 -10.3 M; the electrolyte flow in the electrochemical cell being between 2.5 and 10 I / h 15 And where a current intensity is applied between the electrodes, between the anode and the cathode that form the electrochemical cell, between 0.6 and 1.2 amps. 2. A method according to any one of claims 1, wherein the Sn 20 alloy comprising Pb is an Sn-Pb alloy. 3. The method according to any one of claims 1 or 2, wherein the Sn alloy comprising Pb is selected from Sn-Pb-Ag and Sn-Pb-Sb. Method according to the preceding claim, wherein the Sn alloy comprising Pb is an Sn-Pb-Ag alloy comprising an additional stage a ') of Ag recovery, after electrorefin stage a), characterized by understand the following stages: 30 to "1) Contact the anode sludge composed of impure Ag or a Impure Ag-Pb alloy with concentrated nitric acid, a "2) Add distilled water to the solution obtained in step a), to "3) Filter a "4) Recovery of the Ag contained in the aqueous solution obtained in step a'3) by cementation with electrolytic Cu. 5. The method according to the preceding claim, which comprises an additional stage a ") of recovery of the Cu contained in the aqueous solutions of stage a O), characterized by comprising the following steps: 5 to "1) liquid-liquid extraction of Cu with an oxime, and a "2) of liquid-liquid re-extraction of Cu with sulfuric acid. 6. Method according to the preceding claim, wherein the oxime of step a "1) is Select from the list that comprises ketoxime, salicylalloxime or any of its 10 combinations. Method according to any one of claims 1 to 6, which comprises a Pb recovery stage of the anodic sludge formed in the electrorefin stage (a) or of the anodic sludge after step (a ') by conventional pyrometallurgical techniques,