WO2017037625A1

WO2017037625A1 - A method and process of recovering metal values from waste monolithic ceramic capacitors

Info

Publication number: WO2017037625A1
Application number: PCT/IB2016/055189
Authority: WO
Inventors: Nitin Gupta; G. Prabaharan; Smruti Prakash BARIK; Bhuvnesh KUMAR; Jageshwar PRASAD
Original assignee: Attero Recycling Pvt. Ltd.
Priority date: 2015-09-02
Filing date: 2016-08-31
Publication date: 2017-03-09

Abstract

The present invention relates to the recycling process of electronic waste resources to recover metals of value. More particularly, it relates to eco-friendly yet efficient hydrometallurgical processes to recover metal values from waste monolithic ceramic capacitors. In the proposed method, the source material such as mother boards having MLCC is pulverized using a ball mill to dislodge various components such as monolithic ceramic capacitors (MLCC) to form a powdery mixture of components. The mixture of components is then sieved. The MLCC powder of particular size range is then subjected to hydrometallurgical process. The process then transfers the precious metals into solution by leaching process followed by selective precipitation, purification, cementation and final recovery of metals.

Description

"A METHOD AND PROCESS OF RECOVERING METAL VALUES FROM WASTE MONOLITHIC CERAMIC CAPACITORS"

FIELD OF THE INVENTION

The present invention relates to the recycling process of electronic waste resources to recover metals of value. More particularly, it relates to eco-friendly yet efficient hydrometallurgical processes to recover metal values from waste monolithic ceramic capacitors.

BACKGROUND OF THE INVENTION

Rapid technological growth and advancement in the electronics domain led to generation of new products and a resulting decrease in the life span of electronics. Per a study, more than 50 million tons of e-waste were discarded in 2009 and 72 million tons are expected to be disposed in 2014 (Ping Jiang et al.). E-waste contains precious and special metals, including gold, silver, palladium and platinum, titanium, nickel copper as well as potentially toxic substances such as lead, mercury, cadmium and beryllium. Hence, a large scope for recovering such precious metals that can be reused on their recovery.

Recycling of electronics allows for precious and special metals to be recovered, reduces the environmental impact associated with electronic manufacturing from raw materials, and ensures that hazardous and toxic substances are handled properly. Hydrometallurgical processing is emerging as a potential domestic solution for treating e-waste due to its nontoxic process.

The study of technologies for the recovery of raw materials from waste electronic components is therefore of great interest, in terms of environment protection and reuse of precious raw materials especially metals having a high commercial value. Hydrometallurgy technique is used for obtaining metals from their ores and divided into three general steps viz. leaching, solution concentration and purification, and metal recovery. Hydrometallurgical techniques, on the other hand, are generally performed at temperatures close to ambient temperature, with lower costs and limited atmospheric emissions.

Recycling of electronic waste can be divided into mechanical and physical separation, crushing, enrichment and purification of several stages. Usually the first electronic waste batteries, capacitors, resistors and other major components of the demolition to remove, and then broken into pieces its pre-inch size, and then fragments into the grinder pulverized powder pulverized available pyrometallurgical, hydrometallurgical, electrochemical method further enrichment, separation.

Pyrometallurgical include incineration, plasma furnace and blast furnace smelting, slag, sinter melting, sweating and high temperature gas-phase reactions. The basic principle is to use high-temperature metallurgical furnace heating stripped nonmetallic materials, precious metal is an alloy with other metals was out of state, re-refining or electrolytic process.

The main steps of hydrometallurgical process includes a series of pickling or alkali leaching to separate the solid material, and then subjecting the solution such as extraction solution, precipitation, cementation, ion exchange, distillation, filtration and separation processes to separate and enrichment of some important metals.

Monolithic ceramic capacitors in the e-waste have a good amount of precious metals like palladium as well as silver or nickel & tin and the like. There is a scope to recover and reuse such precious metals from the e-waste that is scrapped. Hence, an effective approach is needed which is economic and can effectively recover metals from the e-waste. OBJECT OF THE INVENTION

The main object of the present invention is to provide a method of recovery metal values from waste monolithic ceramic capacitors.

Yet another object of the present invention is to provide a method of recovering metals by hydrometallurgical process.

Yet another object of the present invention is to provide an improved method to separate out precious metal values using selective leaching process.

Yet another object of the present invention is to provide a convenient and eco-friendly approach for e-waste management. Yet another object of the present invention is to provide a method wherein the metals are selectively precipitated and separated from chloride medium using sodium hydroxide solution.

Yet another object of the present invention is to provide a method for recovering precious metals including Barium, Titanium, Tin, Copper, Nickel, Gold, Palladium and Silver and other metals present in e-wastes.

Yet another object of the present invention it to provide a metal recovery process ends with minimum solid waste and with no liquid discharge.

Yet another object of the present invention is to provide an efficient metal recovery method with high extraction rate of metal values.

SUMMARY OF THE INVENTION

The present invention relates to an improved method of recovering metal values from waste monolithic ceramic capacitors. The invention extracts metals in waste monolithic ceramic capacitors (MLCC) rendered as e-waste. The method is specific to waste MLCC especially from mother boards of computers, laptops and in other MLCC based electronic equipments.

In the most preferred embodiment of the invention, an improved hydrometallurgical method is provided to segregate valuable metal from all kind of waste Monolithic ceramic capacitors (MLCC). In this method, the source material such as mother boards having MLCC is pulverized using a ball mill to dislodge various components such as monolithic ceramic capacitors (MLCC) to form a powdery mixture of components. The mixture of components is then sieved using a sieve having pore size of 300 microns. The MLCC powder of below 300 microns is then subjected to hydrometallurgical process. The process then transfers the precious metals into solution by leaching process followed by selective precipitation, purification, cementation and final recovery of metals.

BRIEF DESCRIPTION OF THE DRAWINGS

Figure 1 elucidates the process flow-chart for recovering of total metal values from waste MLCCs.

DETAILED DESCRIPTION OF THE INVENTION

Reference will now be made in detail to the present embodiment of the present invention, by way of accompanying drawings and claims, examples of which are illustrated in the description below. The embodiments are described below in order to explain the present invention.

In an embodiment of the invention, an improved hydrometallurgical method is provided to separate out metal values from all kind of waste Monolithic ceramic capacitors (MLCC). In this method, the source material is pulverized using a size reduction apparatus preferably a ball mill to liberate the scrap monolithic ceramic capacitors (MLCC) in the form of a mechanical mixture of particles. The mixture of particles is then sieved using a sieve having pore size of 300 microns. The MLCC powder of below 300 microns is then subjected to hydrometallurgical process. The process then transfers the precious metals into solution by leaching process followed by selective precipitation, purification, cementation and final recovery of metals.

In another embodiment, as described in Fig. 1, the invention provides a process of metal recovery and comprises the following steps of: a. Pulverization of Waste monolithic ceramic capacitor (MLCC) using a ball mill. b. Sieving of pulverized MLCC obtained in step a. with mesh of pore size 300 microns.

c. Primary leaching of MLCC powder of below 300 microns size obtained in step b. with HCI/H₂0₂ for metal values other than precious metals.

d. Leaching of gold and palladium from the above residue of the step c. with aquaregia.

e. Leaching of silver from the above residue of step d. after drying and roasting with Nitric acid.

f. Selective Precipitation of each metal with pH variation using sodium hydroxide solution (except barium which is precipitated with Sulphuric acid as barium sulphate).

g. Precipitation of Gold from the leach liquor of step d. with Urea and sodium metabisulphite.

h. Cementation of Pd from the above gold free liquor of step g. with zinc dust. i. Purification of Pd by leaching with aqua regia followed by precipitation with sodium formate.

j. Precipitation of silver from the leach liquor of step e. with sodium chloride, k. Recovery of Cu from Pulverized MLCC of above 300 microns size of step b., by Electro refining method. In preferred embodiment of the present invention, the waste MLCC scrap is initially crushed by any crusher, preferably a ball mill. Sieving is then carried out using a sieve of pore size 300 μ. The MLCC powder having size less than 300 μ is then subjected to primary leaching process using hydrogen chloride and hydrogen peroxide solution. The purpose of such a primary leaching is to extract non precious metals of value. The precious metals remain in form of residue.

Selective leaching for the metals like gold, palladium, silver is then carried out from the residue obtained from primary leaching process. Gold and palladium are leached out using Aqua Regia. Leaching of silver is then carried out from residue of step d. after drying and roasting with Nitric acid. Once leaching process is done, each metal is selectively precipitated thereafter. The selective precipitation is achieved under varying pH environment. The selection of precipitant(s) for selective precipitation process depends upon the metal to be precipitated. For example, barium is precipitated with sulphuric acid and barium sulphate; gold is precipitated with urea and sodium metabisulphite; silver is precipitated with sodium chloride. Palladium is cemented from the above gold free liquor with zinc dust.

In another embodiment of the invention, leaching process is provided primarily to leach out precious metals from other non-precious ones and then series of selective leaching steps take place depending on the type of metal to be leached. The method provides an economical and eco-friendly approach for e-waste management in the form of metal recovery from waste MLCC.

The method also provides selective precipitation approach to precipitate out precious metals including Barium, Titanium, Tin, Copper, Nickel, Gold, Palladium and Silver and other metals abundant in waste MLCC.

In an embodiment of the invention, the method provides an efficient method leaving minimum solid and no liquid wastes after completion of the recovery process. The average recovery of metallic fraction containing precious metals into basic concentrate using this method is above 90 % and which is effective.

In principal, the present invention provides for a process for recovering metals from waste monolithic ceramic capacitors comprising the steps of: a) pulverizing of the waste monolithic ceramic capacitors to obtain particles of preferred size;

b) sieving of the pulverized capacitors;

c) preparing an aqueous slurry of sieved powder obtained in step b);

d) leaching of the aqueous slurry of step c) with hydrochloric acid and hydrogen peroxide by agitating it for 4-6 hours at a preferred range of temperature to obtain a leached slurry;

e) filtering the leached slurry obtained in step d) through a filter press to obtain a leached liquor containing metals including alkaline earth metals and transition metals and a residue containing precious metals including gold, silver, and palladium;

f) treating the leached liquor of step e) with sulphuric acid for selective precipitation of Barium as barium sulphate;

g) treating the barium free liquor of step f) with sodium hydroxide solution for selective precipitation of transition metals at their respective pH ranges;

h) agitating the residue of step e) with aqua regia at a preferred temperature range for 3-4 hours to obtain a silver rich precipitate and a supernatant containing gold and palladium;

i) treating the supernatant of step h) with urea and sodium metabisulphite and filtering the solution to obtain a palladium containing liquor and gold precipitate; j) cementing the palladium from the liquor of step i) with zinc dust and treating the cemented palladium with aqua regia and sodium formate at a preferred range of temperature to obtain palladium powder; k) melting the palladium powder of step j) at a temperature ranging from 1200- 1800 °C to obtain pure palladium.

The invention will now be illustrated by the following non-limiting examples. Example 1: In a 10 kg batch of waste monolithic ceramic capacitors, the process was performed. Initially, the sample was pulverized in a ball mill containing eight iron balls weighing each 18 kg for 6 hours at a speed of 40 rpm and finally sieved to obtain a particle size of less than 300 μιτι (8.07 Kg). The chemical analysis of the sieved material is shown in table 1.

Table 1 Chemical analysis of the sieved MLCC

About 2 Kg sample of the sieved material was taken for the study in different batches. Slurry is made by dissolving the sample in water (about 10 liters). The slurry is then used for primary leaching with hydrochloric acid (about 5.4 liters) and 50% w/v hydrogen peroxide (about 132 ml). The mixture is agitated for five hours at a temperature ranging from 70 to

80 °C. After predetermined time, the slurry was filtered to obtain leach liquor, L_A (about 15.9 liters) and residue, R_A (about 0.54 Kg). The analysis of the leach liquor, L_A is represented in table 2. The leach liquor, L_A was taken for selective precipitation of metals. Whereas, the residue, R_A was taken for recovering precious metals like gold, silver, and palladium.

Table 2. Chemical analysis of the leach liquor, L_A

Elements Cu Ni Zn Sn Fe Ba Al Mn Cr Ti g/l 3.24 12.75 0.41 4.40 4.63 21.00 0.19 0.19 0.38 32.75 The leach liquor, L_A was taken for selective precipitation of metals including alkaline earth metals and transition metals present in the liquor. To recover alkaline earth metals, leach liquor, L_A is treated with sulphuric acid at room temperature (25±3) °C. The treatment of L_A with sulphuric acid (1.6 time stoichiometric amount of Ba in the liquor) results into precipitation of barium as barium sulphate. More than 99% precipitation efficiency was observed and about 0.569 Kg of barium sulphate was collected with 97.8% purity.

In the subsequent step, barium free liquor was taken for recovery of remaining metals that includes transition metals along with other metals. The barium free liquor is treated with sodium hydroxide at a temperature of 90 °C and a pH of 0.5 for atleast 3 hours. The slurry was then filtered to separate residue containing titanium and co-precipitated amounts of tin (16.5%). More than 98.2% precipitation efficiency (for titanium) was observed.

In order to get a pure titanium dioxide, the residue was re-dissolved and re-precipitated at a temperature of 90 °C and a pH of 0.5 for 3 hours. The slurry was filtered, washed and dried at 110 °C for 2 hours. Titanium oxide (0.201 Kg) having purity of 97.9% was collected. The titanium free liquor was taken for the recovery of tin. Tin (Sn) was precipitated using sodium hydroxide solution (pH 1.2) at room temperature (25±3) °C for 2 hours. The precipitated tin hydroxide was filtered, washed and dried at 110 °C for 2 hours. About 0.076 Kg of tin oxide was collected with a purity of 98.8%.

I n the subsequent study, the impurities including iron, aluminium, zinc and chromium was removed using sodium hydroxide solution (pH 4.5). After removing impurities, copper and nickel were recovered using sodium hydroxide solution at a pH of 5.5 and 9 respectively. About 0.076 Kg of copper hydroxide and 0.32 Kg of nickel hydroxide were recovered.

Recovery of precious metals: For recovering precious metals, the residue, R_A (0.54 Kg) was taken and agitated with about 1.25 liters of aqua regia at a temperature of 80 °C for 4 hours, to leach out precious metals. The slurry was then cooled, filtered and washed. The residue, R_B (about 0.48 Kg) and leach liquor, L_B (about 2.8 liters) were collected. Leach liquor, L_B was taken for precipitation of gold and palladium, while the residue, R_B was taken for silver (Ag) recovery. The chemical analysis of the leach liquor, L_B is presented in Table 3. More than 90% leaching efficiency of both gold (Au) and palladium (Pd) was observed.

Table 3. Chemical analysis of aqua regia leach liquor-2

Recovery of Gold (Au) and palladium (Pd): About 0.3 Kg of urea was added to the leach liquor, L_B and agitated. The gold (Au) was precipitated by adding 0.5 g of sodium metabisulphite. The precipitated gold was taken out by filtration followed by washing and drying. The dry weight of gold powder was found to be 0.29 g.

The gold free filtrate was taken for palladium recovery. I nitially, palladium was cemented by using zinc dust at a concentration of 8 g/L, the cemented palladium was re-dissolved in aqua regia and re-precipitated using sodium formate at a concentration of 22 g/L and a temperature, 95 °C. The precipitated palladium was filtered, washed and dried and finally melted at 1600 °C. The weight of palladium metal was 0.75 g.

The residue, R_B was roasted at 600 °C for three hours. The roasted residue was agitated with nitric acid (0.4 liter) and water (1.0 liter) at 80 °C for 2 hours. The slurry was filtered, washed and dried at 110 °C for two hours. The weight of the final residue was found to be 0.32 Kg. To the filtrate (1.28 liter), saturated sodium chloride solution was added till the complete precipitation of silver. The precipitated silver chloride was washed, dried and melted at 1000 °C. The weight of silver metal obtained was 14 g. The chemical analysis of the final residue was presented in Table 4. Table 4: Chemical analysis of the final residue

Example-2: Another batch of 2 Kg sieved sample was taken for the study. Slurry is made by dissolving the sample in water (about 10 liters). The slurry is then used for primary leaching with hydrochloric acid (about 5.4 liters) and 50% w/v hydrogen peroxide (about 132 ml).

The mixture is agitated for five hours at a temperature ranging from 70 to 80 °C. After predetermined time, the slurry was filtered to obtain leach liquor, L_2A (about 15.7 liters) and residue, R₂A (about 0.52 Kg). The analysis of the leach liquor, L_2A is represented in table 2. The leach liquor was taken for selective precipitation of metals present in the leach liquor. Whereas, the residue, R₂A was taken for recovering precious metals like gold, silver, and palladium.

Table la Chemical analysis of the leach liq uor, L.2A

The leach liquor, L_2A was taken for selective precipitation of metals. From the leach liquor (L_2A), barium was recovered as barium sulphate by treating with sulphuric acid (1.6 time stoichiometric amount of Ba in the liquor) at room temperature (25±3) °C. More than 99% precipitation efficiency was observed and about 0.565 Kg of barium sulphate was collected with 97.8% purity.

In the subsequent step, barium free liquor was treated with sodium hydroxide solution at 90 °C and a pH of 0.5 for atleast 3 hours. The slurry was then filtered to separate residue containing titanium and co-precipitated amounts of tin (16.5%). More than 98.2% precipitation efficiency (for titanium) was observed. In order to get a pure titanium dioxide, the residue was re-dissolved and re-precipitated at a temperature of 90 °C and a pH of 0.5 for 3 hours. The slurry was filtered, washed and dried at 110 °C for 2 hours. Titanium oxide (0.203 Kg) having purity of 97.9% was collected. The titanium free liquor was taken for the recovery of tin. Tin (Sn) was precipitated using sodium hydroxide solution (pH 1.2) at room temperature (25±3) °C for 2 hours. The precipitated tin hydroxide was filtered, washed and dried at 110 °C for 2 hours. About 0.09 Kg of tin oxide was collected with a purity of 98.8%.

I n the subsequent study, the impurities including iron, aluminium, zinc and chromium was removed using sodium hydroxide solution (pH 4.5). After removing impurities, copper and nickel were recovered using sodium hydroxide solution at a pH of 5.5 and 9 respectively.

About 0.079 Kg of copper hydroxide and 0.32 Kg of nickel hydroxide were recovered.

Recovery of precious metals: For recovering precious metals, the residue, R₂A (0.52 Kg) was taken and agitated with about 1.25 liters of aqua regia at a temperature of 80 °C for 4 hours, to leach out precious metals. The slurry was then cooled, filtered and washed. The residue, R₂B (about 0.45 Kg) and leach liquor, L₂B (about 2.7 liters) were collected.

Leach liquor, L₂B was taken for precipitation of gold and palladium, while the residue, R_2B was taken for silver (Ag) recovery. The chemical analysis of the leach liquor, L_2B is presented in Table 2a. More than 90% leaching efficiency of both gold (Au) and palladium (Pd) was observed.

Table 2a: Chemical analysis of aqua regia leach liq uor, L

Recovery of Gold (Au) and palladium (Pd): About 0.312 Kg of urea was added to the leach liquor, L_2B and agitated. The gold (Au) was precipitated by adding 0.5 g of sodium metabisulphite. The precipitated gold was taken out by filtration followed by washing and drying. The dry weight of gold powder was found to be 0.29 g.

The gold free filtrate was taken for palladium recovery. I nitially, palladium was cemented by using zinc dust at a concentration of 8 g/L, the cemented palladium was re-dissolved in aqua regia and re-precipitated using sodium formate at a concentration of 22 g/L and a temperature, 95 °C. The precipitated palladium was filtered, washed and dried and finally melted at 1600 °C. The weight of palladium metal was 0.78 g.

The residue, R₂B was roasted at 600 °C for three hours. The roasted residue was agitated with nitric acid (0.4 liter) and water (1.0 liter) at 80 °C for 2 hours. The slurry was filtered, washed and dried at 110 °C for two hours. The weight of the final residue was found to be 0.31 Kg. To the filtrate (1.26 liter), saturated sodium chloride solution was added till the complete precipitation of silver. The precipitated silver chloride was washed, dried and melted at 1000 °C. The weight of silver metal obtained was 14.3 g. The chemical analysis of the final residue was presented in Table 3a.

Table 3a. Chemical analysis of the final residue

Elements Cu Ni Pb Zn Au Ag Pd Sn Fe Al Mn Cr Ba Ti

% BDL BDL BDL BDL 0.003 0.16 0.006 BDL BDL BDL BDL BDL BDL BDL

Claims

A process for recovering metals from waste monolithic ceramic capacitors comprising the steps of:

pulverizing of the waste monolithic ceramic capacitors to obtain particles of preferred size;

sieving of the pulverized capacitors;

preparing an aqueous slurry of sieved powder obtained in step b);

leaching of the aqueous slurry of step c) with hydrochloric acid and hydrogen peroxide by agitating it for 4-6 hours at a preferred range of temperature to obtain a leached slurry;

filtering the leached slurry obtained in step d) through a filter press to obtain a leached liquor containing metals including alkaline earth metals and transition metals and a residue containing precious metals including gold, silver, and palladium;

treating the leached liquor of step e) with sulphuric acid for selective precipitation of Barium as barium sulphate;

treating the barium free liquor of step f) with sodium hydroxide solution for selective precipitation of transition metals at their respective pH ranges;

agitating the residue of step e) with aqua regia at a preferred temperature range for 3-4 hours to obtain a silver rich precipitate and a supernatant containing gold and palladium;

treating the supernatant of step h) with urea and sodium metabisulphite and filtering the solution to obtain a palladium containing liquor and gold precipitate;

cementing the palladium from the liquor of step i) with zinc dust and treating the cemented palladium with aqua regia and sodium formate at a preferred range of temperature to obtain palladium powder; and k) melting the palladium powder of step j) at a temperature ranging from 1200- 1800 °C to obtain pure palladium.

2. The process of recovering valuable metals as claimed in claim 1, wherein the preferred size of particles is less than 300 μιη.

3. The process of recovering valuable metals as claimed in claim 1, wherein the preferred range of temperature for leaching of slurry is from 70° to 80 °C.

4. The process of recovering valuable metals as claimed in claim 1, wherein the preferred range of temperature for agitating of residue is from 75° to 95 °C.

5. The process of recovering valuable metals as claimed in claim 1, wherein the preferred range of temperature for treating cemented Palladium is from 90° to 100 °C.

6. The process of recovering valuable metals as claimed in claim 1, wherein the preferred range of temperature for melting the Palladium powder is from 1500- 1600 °C.

7. The process of recovering valuable metals as claimed in claim 1, wherein the transition metals are titanium, copper, nickel, zinc, iron, manganese, and chromium.

8. The process of recovering valuable metals as claimed in claim 1, wherein the precipitate of step h) is roasted at a temperature ranging from 550-650 °C for 2- 4 hours.

9. The process of recovering precious metals as claimed in claim 9, wherein the roasted precipitate is agitated with a mixture of nitric acid and water at a temperature ranging from 75-85 °C for 1-3 hours and filtered to obtain resid and silver rich filtrate.

10. The process of recovering precious metals as claimed in claim 10, wherein the silver is precipitated as silver chloride by treating the filtrate with sodium chloride solution.

The process of recovering precious metals as claimed in claim 1, wherein the precipitation efficiency of Barium is 99 %.

12. The process of recovering precious metals as claimed in claim 1, wherein the gold precipitate is purified using state of the art process.