7.3.1.1 Copper-Smelting Processes – Sulfide Route

The general principle of copper smelting is to remove sulfur and iron (i.e. the two main impurities of copper sulfide ores) using oxygen. Figure 7.2 shows the predominance diagram for the CuFeS-SiO₂-O₂ system at a temperature of 1200 °C. Silica is added to the system (using a 2 : 1 CuFeS₂:SiO₂ molar ratio) to produce a slag phase upon ferrous oxide (FeO) formation during the O₂-enriched air blowing. At the beginning of the smelting operation, a (Cu,Fe)-rich matte and solid silica are present in the reactor. As oxygen is introduced into the system, it generates SO₂(g) and ferrous oxide. The obtention of the adequate SiO₂:FeO ratio leads to the formation of fayalite slag. When the (Cu-rich matte + slag) meta-equilibrium is reached (gray zone #1 on Figure 7.2), the smelting process is completed. The copper matte smelting can be summarized using the following overall reaction:

In theory, it is thermodynamically possible to identify operating conditions that lead to the direct production of metallic copper called direct-to-copper smelting by further oxidizing the matte (see the grey zone #2 in Figure 7.2). However, these operating conditions lead to significant copper oxidation and solid magnetite formation (spinel). Other fluxes such as CaO need to be used to lower the liquidus temperature of the slag. More importantly, copper is transferred to the slag phase. To make the process economically viable, the copper-rich slag phase needs to be treated with coke to recover the valuable metal. The Cu-rich matte is then converted into liquid copper using a similar strategy.

The primary production of copper from sulfides is performed via one of the following smelting technologies (Schlesinger et al. (2011c)):

• Submerged tuyere smelting: Oxygen-enriched air is the primary reactant, which is blown into the liquid matte via submerged tuyeres to produce SO₂ and FeO (which is subsequently fluxed into a fayalite slag synthetically produced by adding silica). The concentrate can be injected into the system via either the submerged tuyere (in which case the concentrate needs to be dry and has to meet specific particle size requirements) or from the top of the reactor as in the Noranda technology. The latter feeding strategy is more suitable for charging a scrap of various sizes into the smelter. Silica is also added to the system to flux the oxidized iron.
• Flash smelting: This technology values a solid (concentrate) -gas (O₂-enriched air) combustion-like reaction to promote the simultaneous formation of liquid matte and slag droplets, which eventually reach the matte–slag pool at the bottom of the reactor. This two-phase liquid system needs to settle in the furnace to promote efficient liquid–liquid phase separation and transfer of the copper from the slag phase to the matte. One drawback of this approach is that the concentrate feed needs to be calcined and being injected as fine particles.
• Vertical-lance smelting (ISASMELT): This technology developed in Australia is based on the use of a submerged gas-cooled lance, which blows a mixture of (oxygen-enriched + natural gas) blast into the liquid matte (see Figure 7.3). The vertical reactor is fed from the top and can tolerate various feed sizes, which ease the introduction of scrap. Magnetite, which is formed by the matte’s oxidation, is a key phase in this process. Firstly, it protects the lance from the matte as it solidifies around it. It also plays a central role in the production of SO₂(g) as it reacts with the matte phase via a solid–liquid reaction. Alvear Flores et al. (2014) presented an exhaustive study about the importance of this compact technology for recycling. It is also a central unit operation of the Umicore process detailed by Hageluken (2006).
• Direct-to-copper smelting: As explained previously, the general idea behind this smelting strategy is to severely oxidize the matte until it produces metallic copper. According to Schlesinger et al. (2011a), about 25 wt% of Cu is inevitably transferred to the slag phase. It will be recovered by reducing the copper oxide in the slag using coke.

Copper smelters clearly offer an opportunity to partition valuable metallic elements and other impurities that constitute WPCBs in various phases that form during the process (such as the matte, the slag, the metallic melt, and the exhaust gas-dust). One fundamental aspect to understand here is the fate of each element and contaminant under these smelting operating conditions. Figure 7.4 presents a schematic representation of the partitioning of various elements/component present in WPCBs. The first group called plastics is incinerated in situ. The products of this incineration are transferred to the gas phase. The second group called volatile metals includes zinc, lead, arsenic, bismuth, tellurium, and selenium. Apart from tellurium and selenium (which end up in the liquid copper matte as they are chalcogen elements), all these elements are mostly transferred to the SO₂-rich exhaust gas. This is potentially problematic as they may exceed regulated concentrations when released to the atmosphere and become a danger for the environment. The third group is associated to the presence of glass fibers in WPCBs. It contains oxide components like Al₂O₃, SiO₂, and CaO. These oxide components are easily fluxed in the slag phase during smelting. The last group is called low-vapor-pressure metals. It is subdivided into two categories: (i) the noble metals relative to iron (such as Cu, Ag, Au, and Pt), which will be carried in the Cu-rich liquid metal, and (ii) the reactive metals (such as Al, Ti, Ni, Co, and Fe), which will mostly be oxidized and transferred to the slag phase. A converting step is performed on the copper matte to produce the blister copper, which is refined using natural gas to remove the dissolved monatomic oxygen.

Finally, it is to be noted that many authors (Khaliq et al. (2014); Cui and Zhang (2008); Ma (2019a); Zhang and Xu (2016)) have reviewed the principal industrial e-waste recycling practices involving copper-smelting unit operations. Their studies reported the Noranda process in Canada, the Aurubis’s Kayser recycling system in Germany, the Boliden Ronnskar Smelter in Sweden, and the Umicore’s precious metal refining process in Belgium (see the process flowsheet in Figure 7.5). It is to be noted that the smelting operations at Boliden and Umicore are integrated into a much larger process scheme that also involves lead-smelting operations, which will be explained in the Section 7.3.1.3.

7.3.1.2 Copper-Smelting Processes – Secondary Smelters

There are two types of secondary smelters to process copper-rich scrap, i.e. the metal smelters and the black-copper smelters (Schlesinger et al. (2011b)). Tiltable reverberatory furnaces are used to process Cu-rich (high-grade) scrap. For low-grade scrap, a more elaborate strategy (involving a reduction step using coke followed by an O₂(g) oxidizing refining step) is required. Low-grade scrap often includes shredded automobile products and dross coming from the decopperizing of lead bullion, which both contain lead. Therefore, the typical furnaces associated to the production of black copper (such as the Kaldo furnace and top submerged lance reactors like ISASMELT and AUSMELT) are often found in lead extraction operations.

7.3.1.3 Lead-Smelting Processes

Lead, like copper, is a chalcophile element which is naturally found in association with sulfur in different minerals. Lead sulfide ores such as galena (PbS) typically contain other minerals that are rich in zinc, copper, and iron (in the form of sulfates and sulfides) as well as precious metals, notably silver (Ayres et al. 2003). E-wastes such as lead-acid batteries are therefore fully compatible with this ore type as they simultaneously contain metallic lead, PbO₂, and PbSO₄ (Arnout et al. 2011). They can be mixed with the sulfide concentrate feeds.

Laubertova et al. (2017) presented an overview of the lead-smelting technologies. Two main routes can be used to process lead sulfide concentrates and lead-rich scrap, i.e. (i) the traditional sintering–smelting in blast furnaces (for Zn-lean concentrate) or using Imperial Smelting furnaces (for Zn-rich concentrate), and (ii) the direct smelting approach. Grant (2001) reported many direct lead-smelting approaches such as the Kaldo furnace (top-blown rotary converter), the QSL kiln, the ISASMELT and AUSMELT vertical reactor with submerged lance, and the Kivcet process. Contrary to conventional copper-smelting operations, which always require large amounts of concentrate to operate, several authors report that many lead-smelting operations run almost exclusively using scrap feed. This is explained by the high lead fraction that contains specific waste materials (such as lead-acid batteries) when compared to the WPCBs, which have smaller fractions of metallic components.

• Sintering–smelting: In this approach, the concentrate is first roasted using air to convert the sulfide into an oxide via the following reaction:
  

  This reaction is important as lead oxide can be easily reduced to metallic lead by a carbon-based reducing agent (which is not the case for lead sulfide). Lead oxide is not the only phase in the sintered product. It also contains melilite, spinel, calcium silicate sulfate, as well as copper sulfide (Zhao 2013). Next, the lead oxide sinters are reduced in the blast furnace using coke. Silica and lime are also charged in the blast furnace from the top (Watt et al. (2018)). The tuyeres at the bottom of the reactor blast O₂-enriched air, which react with the excess coke to produce CO(g) via the Boudouard reaction. This carbon-monoxide-rich gas will act as the main reducing agent for this reduction process (Vanparys et al. (2020)):

  

  The liquid lead (called lead bullion) obtained from this operation is collected at the bottom of the furnace along with a slag and a matte phase. In many cases, arsenic is present in the charge and will form speiss (i.e. iron arsenide) upon the reaction with iron. Figure 7.6 taken from the work of CHAIDEZ-FELIX et al. (2014) shows the phase assemblage obtained from a lab-scale lead blast furnace experiment at 1573 K. It consists of an equilibrium between lead bullion, speiss, matte, and slag. In the Umicore process presented in Figure 7.5, the lead bullion obtained from the blast furnace is sent to a refinery unit, the matte is sent to a copper smelter, while the speiss is processed in the Umicore Plant in Olen. The Imperial Smelting process works similar to the blast furnace and is used to process zinc-rich concentrates.

• Direction smelting: In this approach, the sulfide concentrate is directly reduced by coke in a furnace. Slag-forming fluxes are also added to the reactor. The following reactions simultaneously take place:
  
  
  

As explained by Grant (2001), the direct smelting of lead sulfide concentrate to obtain a low-sulfur lead bullion and a low-lead slag in a single step is difficult to implement in practice. The PbS-O₂-S₂-(SiO₂)_0.4(FeO)_0.3(CaO)_0.1 predominance diagram calculated at 1473 K and 1 atm presented in Figure 7.7 validates this affirmation.

Figure 7.7 shows that the sulfur removal of the system performed via the O2-enriched air blowing may provide the desired (liquid lead + slag) equilibrium but inadequate compositions (see the red star on Figure 7.7). In this zone, the liquid lead activity of about 0.97 implies that it dissolved a significant amount of oxygen. Moreover, Figure 7.7b shows that the liquid PbO activity is high in this zone, which would lead to important lead losses in the slag. Therefore, the oxidation alone is not enough to smelt the concentrate. The addition of coke into the system allows a significant lowering of the oxygen chemical potential. This leads to an important reduction of the Pb losses to the slag phase (see the star in the bottom left corner in Figure 7.7b). The reduction step is also not sufficient by itself as it cannot fully control the system’s sulfur chemical potential. This explains why a two-stage strategy – oxidation followed by reduction – is required (Grant, 2001). Finally, it is to be mentioned that some direct smelting furnaces can process concentrate-free feeds. Laubertova et al. (2017) reported the example of a shaft furnace battery recovery process that uses battery scrap, coke, iron scrap, and reusable slag to produce lead bullion.

7.3.1.4 Advantages and Limitations of Smelting Processes

We conclude this section by highlighting the advantages and limitation of the smelting processes to recycle electronic waste.

7.3.2.1 High-Temperature Electrolysis

In extractive metallurgy, many reactive metals can be produced using high-temperature electrolysis. This is typically performed in molten salts such as chlorides, fluorides, and chlorofluorides. Excellent review work on the subject is presented by Yan and Fray (2010). A famous example of high-temperature electrolysis is the production of aluminum via the Hall–Heroult process, which uses cryolite (Na₃AlF₆) as the electrolyte. Magnesium and lithium can also be commercially produced via high-temperature electrolysis using molten chloride electrolytes. One significant advantage of using chloride systems is that they melt at relatively low temperatures when compared to oxide-equivalent systems. It is also possible to design a molten chloride electrolyte that is not too volatile and has good ionic conductivity. Fluoride salts can dissolve oxides, which is a major advantage in primary metal production (e.g. alumina in cryolite). However, fluoride melts tend to be more aggressive and are liquid in higher temperature ranges.

In the context of recycling electronic waste, the use of high-temperature electrolysis would normally require a chlorination step to oxidize their metallic components. This step could be done via a roasting strategy. During the molten chloride electrolysis, Cl₂(g) is evolved at the anode and can be recycled in the chlorination/roasting step or used in situ to chlorinate metallic components. The main environmental concern with this process is the accidental release of chlorine in the atmosphere as a result of a process failure. The material selection to building the equipment for handling chlorination operations is also critical and challenging. Applications of molten chloride electrolysis to recover rare-earth elements from electronic waste were recently reviewed by Xi et al. (2020). They reported Abbasalizadeh et al. (2016)’s work in which AlCl₃ is used as the chlorination agent to transfer Nd and Dy (via an exchange reaction) into the eutectic LiCl-KCl-NaCl molten chloride. Nd and Dy’s transfer in the salt is possible as their respective chloride (NdCl₃ and DyCl₃) are more thermodynamically stable than AlCl₃.

Interestingly, this chlorination strategy will not transfer iron and boron in the molten chloride since their chlorides are less stable than AlCl₃. As a result, the electrodeposited rare-earth metals are free of iron. Another strategy to chlorinate the rare-earth elements from magnet scrap is to use a molten MgCl₂-KCl bath (Hua et al. (2014)). In this case, the electrolysis leads to an Mg-Nd alloy’s electrodeposition, which is inevitable because of the small reduction potential differences between these two elements. The proposed flowsheet of Hua et al. (2014) to process NdFeB scrap is presented in Figure 7.9.

7.3.2.2 Low-Temperature Electrolysis

Even though low-temperature electrolytic operations using aqueous solution should not be considered as pyrometallurgical processes, their technological importance in refining primary metals and in recycling e-waste motivated an overview of these processes in this section. Low-temperature electrolytic processes typically use sulfuric, hydrochloric, and boric acid solutions (Jin and Zhang (2020)) as well as ionic liquids. It is interesting to note that conventional hydrometallurgical processes to recover valuable metals from electronic waste use similar aqueous solution to perform the chemical leaching (Su 2020), i.e. the first step of the extraction. The subsequent purification and recovery steps in extractive hydrometallurgy are based on completely different strategies such as precipitation and cementation. Here are some recent examples of low-temperature electrolytic processes to recycle electronic waste.

• Copper recovery from e-waste: There is rich literature on the recovery of copper, which is the main metallic element contained in WPCBs using low-temperature electrolytic processes. In their work, Fogarasi et al. (2013) presented the two main electrolytic routes that can be taken to perform the copper recovery (Figure 7.10). In route A (sulfuric acid electrolyte), copper’s direct electrochemical oxidation is performed, followed by the electrodeposition of pure copper at the cathode. In route B, the Fe³⁺/Fe²⁺ redox pair is used to achieve a faster dissolution of copper (Cu). Fogarasi et al. (2013) reported that route A leads to the extraction of a 98% purity of copper for the energy consumption of 1.06 kWh per kg, while route B produces a 99% purity of copper energy consumption of 1.75 kWh per kg. Here are other examples of copper extraction based on similar processes.

Song et al. (2021) were able to electrochemically extract 95 wt% of the copper dissolved in a multimetal sulfuric acid leaching solution of waste liquid crystal display panels that contained indium, copper, aluminum, and molybdenum. Barragan et al. (2020) proposed a new strategy for the efficient recovery of highly pure copper and antimony (in the form of Sb₂O₃) from WPCB by leaching the waste in hydrochloric acid solution that also contained ferric chloride (to accelerate the copper dissolution), followed by a precipitation step and an electrowinning process. Zhang et al. (2018) proposed a greener ionic liquid electrolyte that contains N-butyl sulfonate pyridinium hydrosulfate, i.e. [BSO₃HPy]HSO₄, to replace H₂SO₄ in a slurry electrolytic system made of a graphite anode and a titanium cathode. The optimal replacement of 10% H₂SO₄ by this ionic liquid allowed to obtain a copper recovery of 90.94% with a purity of 81.69%. Finally, it is worth mentioning the technoeconomic assessment for the recovery of metals from WEEE reported by Diaz and Lister (2018), which compared the low-temperature electrolytic recycling technology to the black copper-smelting operations (which also include an electrolytic step to produce pure copper). As explained previously, the black copper-smelting route of low-grade scrap starts with its melting and subsequent reduction using coke. An oxidizing step is then performed to remove less noble elements like iron, lead, zinc, and tin. The more noble metals (Ni, Ag, Au, and platinum groups) are collected in the anode mud of the electrorefining process. In their study, Diaz and Lister (2018) noted a competitive alternative to process electronic waste via the low-temperature electrolytic recycling technology alone when compared to black copper smelting. A significantly lower capital investment can be achieved with the electrochemical process (i.e. 2.9 kg e-waste per dollar of capital investment) compared to the black copper-smelting process (i.e. 1.3 kg per dollar of capital investment). This is not a surprising conclusion and validates the conventional industrial practices that use leaching and electrolytic refining to extract copper from copper oxide ores. In this case, the main difference is the much higher quantity of impurities in the e-waste (compared to the oxide concentrate), which will require larger electrolyte purification units.

• Noble metal and rare-earth recovery: The acidic sulfate electrolytic solution typically used to electrorefine copper is not strong enough to dissolve noble metals, which end up in the anode mud. Because of that, a stronger electrolyte would be required to extract them via a subsequent electrolytic process. Lister et al. (2014) reported such a two-stage electrorecycling strategy (Figure 7.11). In the first stage of the process, iron, nickel, tin, copper, silver, and rare-earth elements are dissolved in an acidic sulfate solution. Copper and silver are then plated at the cathode. Rare-earth elements dissolved in the solution could be potentially recovered at this stage by precipitation using sulfate double salts. In stage 2, pallidum and gold could be dissolved in an HCl solution (via the anodic production of Cl₂) and finally plated at the cathode.

It is also to be noted that the reduction potential of copper and gold is significantly different (i.e. +0.34 V vs +1.83 V relative to the standard hydrogen electrode). Therefore, it is interesting to evaluate the possibility to electrodeposit each metal from a single electrolytic solution selectively. This is a particularly attractive strategy if one can find a more environment-friendly electrolyte than cyanide solution that can efficiently dissolve gold such as ammoniacal thiosulfate solutions. Kasper et al. (2018) studied the electrochemical behavior of gold and copper simultaneously present in such a solution. They performed electrowinning tests using graphite electrodes to conclude that it is feasible to use such an electrolyte to extract copper and gold at two distinct potentials. Unfortunately, the purity of the individual metal deposits was not reported in their study.

7.3.3.1 Roasting

Roasting is a pyrometallurgical treatment that can be used during the recycling process of electronic waste. It consists of heating a system to a specific temperature in some reactants’ presence (which can be solid, liquid, or gas). Solid–gas reactions are often promoted in these reactors. In the case of electronic waste, Panda et al. (2020) recently proposed a low-temperature chlorination roasting step of pyrolyzed WPCB using ammonia chloride (NH₄Cl) in air at temperatures between 200–325 °C (Figure 7.12).

The thermal decomposition of NH₄Cl leads to the production of H₂(g) and HCl(g). The latter gaseous compound acts as the chlorination agent via a chemical reaction of the following type (e.g. copper):

This approach provides a recovery of 93% for Cu and 100% for Ni, Zn, and Pb using a roasting temperature of 300 °C in the form of metallic chlorides. These metallic chlorides could in turn be processed using a molten chloride electrolytic method as detailed previously. Roasting strategies have also been used in the recycling of permanent magnets. Yoon et al. (2014) used oxidation roasting at 600 °C to convert Nd into Nd₂O₃, which was then processed via leaching.

7.3.3.2 Molten Salt Oxidation Treatment

Molten salt oxidation treatment is another pyrometallurgical strategy that can be used to recycle electronic waste. It relies on the introduction of electronic waste in a molten salt bath (which can be a eutectic carbonate or hydroxide mixture). As explained by Flandinet et al. (2012), these molten salts are specifically selected for their ability to dissolve solid undesirable e-waste components such as oxides, glasses, and plastics, as well as gaseous compounds (CO₂ and halogenated compounds). It is also important to prevent the dissolution of metals in the molten salt to recover them more easily, it is not always desired to generate a more severe oxidizing environment (using air injection in the reactor for example) as it could ultimately lead to the transfer of valuable metals into the salt. In their work, Flandinet et al. (2012) used a eutectic KOH-NaOH molten salt at 300 °C to treat WPCBs. They were able to virtually recover all the metallic fraction of the WPCBs. This approach also prevented the release in the gas phase of most of the halogenated compounds, which were trapped in the molten salt.

Lin et al. (2017) used a different approach based on the use of a molten (Li,Na,K)₂CO₃ eutectic carbonate salt reactor operated at temperatures between 550 and 700 °C. They also performed air injection to promote the oxidation reactions. Overall, they were able to recover 95% of the copper available in the WPCB even though they expected more copper oxidation (which should have lowered its recovery efficiency). Other molten salts have been studied in the literature, such as the NaOH-Na₂CO₃-NaNO₃ salt (Liu et al. (2016)) and LiCl–KCl eutectic mixtures (see for example Riedewald and Sousa-Gallagher (2015)) (Figure 7.13).

7.3.3.3 Distillation

As discussed previously, some metals (like cadmium, zinc, and lead) contained in e-waste are more volatile than others (like copper and tin). One can take advantage of this property to separate and purify some metallic fractions by distillation. Figure 7.14 shows the evolution of the vapor pressure (expressed in log₁₀(Patm)) of some pure metals calculated using the FactSage software. Lead chloride is also presented in Figure 7.14 to see the impact of chlorination on the lead volatility. The boiling point at 1 atm for Cd (766.5 °C), Zn (908.3 °C) and PbCl₂ (950.8 °C) is also clearly visible in Figure 7.14 (i.e. the temperature at which the vapor pressure reaches 1 atm). A threshold vapor pressure of 10⁻⁶ atm is also drawn in Figure 7.14. It represents a typical vapor pressure above which volatilization becomes experimentally non-negligible in open systems.

Zhan and Xu (2008) explored the possibility to remove zinc from Cu-rich fractions via vacuum separation at 1123 K. The application of a vacuum lowers the total pressure of the system, which increases the volatility of all the metals at a given temperature. Using such an approach, they were able to purify copper particle obtained from WPCBs from and original purity of 90.68 wt% to a final purity of 99.84 wt% by volatilizing zinc. These authors presented a series of experimental studies on the purification of copper-rich particle using vacuum separation as reported by Ma (2019a).

7.3.3.4 Pyrolysis

Pyrolysis is another important recycling operation, which allows an efficient and more environment-friendly separation of organic and volatile components from the metallic fraction of WPCB. It is defined as the thermal degradation of organic materials under vacuum or inert conditions to produce value-added products such as oil, gas, and carbon black (Xiong et al. 2020; Jadhao et al. 2020; Huang and Lo 2020). The use of conventional or microwave pyrolysis in recycling electronic plastics such as high-impact polystyrene (HIPS) and acrylonitrilebutadiene styrene (ABS) widely emerged in the last decade due to the several advantages it offers: (i) it closes the loop of waste plastic stream, (ii) it produces value-added products at low operating costs and temperature, and (iii) it potentially prevents the release of toxic gases when compared to combustion and incineration processes(Czégény et al., 2012; Miskolczi et al. 2008).

The pioneering work of de Marco et al. (2008) provides fundamental data about the pyrolysis products of electronic wastes. According to these authors, the solids that remain after the pyrolysis are composed of metals and ceramics mixed with about 3–34 wt% of char-like carbonaceous compounds. The collected condensable gases produce a low-viscosity liquid composed of aromatic compounds (in the case of WPCB). The noncondensable pyrolysis gases are composed of light hydrocarbon molecules as well as significant amounts of CO and CO₂. It is to be noted that one major challenge that limits the use of e-waste pyrolysis byproducts is the presence of halogenated flame retardants, mostly BFRs such as tetrabomobisphenol A (TBBPA). These brominated compounds contaminate the collected liquids and gaseous products by producing bromophenols and release toxic emissions such as hydrogen bromide (HBr) (Das et al. 2020; Ma and Kamo 2019).

Pyrolysis is typically an endothermic process that requires energy to proceed. Conventional pyrolysis methods of e-waste plastics or WPCB include fixed-bed reactors (Hall and Williams 2007), fluidized-bed reactors (Zhang et al., 2018; Guo et al. 2010), and rotating reactors (Ma et al. 2018). The performance of all these pyrolysis strategies is well documented in the literature. The use of microwave irradiation to perform the pyrolysis of WPCB is an interesting alternative to these conventional methods because of the following advantages (Huang et al., 2016): (i) it is a volumetric heating method (and therefore more energy efficient), (ii) it is a selective-heating method; (iii) it is a high-heating-rate method, and (iv) it is an easy-to-control (start/stop) method. Zhang et al. (2021) recently proposed a microwave-assisted catalytic process to pyrolyze WPCB.

Other scientific advances and progress in the field of e-waste pyrolysis in recent years mainly focus on three distinct aspects, which are:

• Fundamental description of the degradation mechanisms and kinetics of WPCBs: The reaction mechanisms and pathways for the degradation of WPCB and e-waste plastics (as well as their corresponding kinetics) have been extensively studied in the literature (see for example Yao et al. (2020);Alenezi and Al-Fadhli (2018); Quan et al. (2009); Cai and Shao-hong (2006)). Thermal analysis methods such as the thermogravimetric analysis (TGA) and the differential scanning calorimetry (DSC) have been used to identify reaction mechanisms at the laboratory-milligram scale. Larger-scale experiments typically performed in tubular furnaces are also frequently used in the literature to analyze and characterize the different pyrolysis products (i.e. the solid residue, the condensate, and the volatile fraction).

  Studies on the pyrolysis pathways of WPCB reported that hydrogen bromide is produced during the initial step of the process, followed by the depolymerization of epoxy resins, random rupture, and reactions of free radical to form the different pyrolysis products (Gao et al. 2021; Gao and Xu 2019; Ma and Kamo, 2018). The effect of different process parameters (such as the heating rate, the carrier gas, and the operating temperature) on the degradation mechanisms and the composition of the pyrolysis products have also been reported by many authors (Diaz et al. 2018; Evangelopoulos et al. 2017; XIE et al. 2016).

  Interestingly, the presence of copper is of prime importance during the pyrolysis of WPCB as it acts as a catalyst to promote some chemical reactions, especially in the presence of brominated compounds (Gao et al. 2021; Liu et al. 2018; Ma and Kamo 2018). More specifically, copper changes the pyrolysis mechanism of WPCB by promoting the conversion of organic bromides to Br₂ and HBr, while reducing the apparent activation energy (E_a) of the overall pyrolysis reaction (Gao et al. 2021). Ma and Kamo (2018) also noted that the simultaneous addition of iron and nickel particles in a two-stage pyrolysis bed reactor leads to a synergistic effect on the product yields of WPCB. Their presence promotes the formation of phenol and other aromatic compounds, such as benzene and toluene.

• Recovery of metals from the solid residue: Pyrolysis of WPCB also proved to be a promising technique to efficiently separate the volatile organic fraction from the valuable metallic components. The solid residue that contains these valuable metals (along with oxides and carbonaceous compounds) is then processed via hydrometallurgical or electrometallurgical methods to selectively recover each metal. Jadhao et al. (2020) proposed a new approach to recover the metallic fraction of WPCB using a chemical-free green ultrasonication technology of the solid residue obtained after a mild-temperature pyrolysis process performed between 200 and 600 °C. Many other approaches have been proposed to improve individual metal recovery from solid pyrolysis residue. It includes: (i) low-temperature alkaline smelting coupled with liquid–liquid phase separation (Chen et al. (2020), (ii) combination of physical separation methods such as sieving, gravity separation, and magnetic separation (Wang et al. 2020), and (iii) the use of higher pyrolysis temperatures (i.e. 850 °C) to obtain distinct copper-rich, lead-tin-rich, and slag-rich residues to promote more efficient partitioning of rare-earth elements (Khanna et al. 2018).
• Capture and stabilization of brominated compounds: Copyrolysis of e-waste plastics or WPCB in the presence of alkali and alkaline earth-based additives (such as CaO and CaCO₃) significantly enhance the quality of the pyrolytic liquid products (Gao and Xu 2019; Xie et al. 2017; Kumagai et al. 2017; Jung et al. 2012; Zuo et al., 2011; Bhaskar et al. 2004). These additives stabilize the hydrogen bromide emissions in the form of brominated salt compounds via a reaction such as the following:
  

  As a result, bromine is not transferred to the gas phase, which prevents the contamination of the collected pyrolytic oil and gas. There are still several important scientific elements to understand regarding the identification and thermal stability of the brominated compounds formed during the pyrolysis, which may impact the recovery of valuable metals.