A process using a thermal reduction for producing the battery grade lithium hydroxide from wasted black powder generated by cathode active materials manufacturing.

Recent lithium consumption is doubled in a decade due to the Li-ion battery (LIB) demand for electric vehicles, the energy storage system, etc. The LIBs market capacity is expected to be in strong demand due to the political drive by many nations. Wasted black powders (WBP) are generated from the manufacturing of the cathode active material and spent LIBs. The recycling market capacity is also expected to expand rapidly. This study is to propose a thermal reduction technique for recovering Li selectively. The WBP, containing 7.4 % Li, 62.1 % Ni, 4.5 % Co, and 0.3 % Al, was reduced in a vertical tube furnace using a 10 % H2 gas as a reducing agent at 750 ºC for 1 h, and 94.3 % of Li was recovered from a water leaching, while other metal values, including Ni and Co remained in the residue. A leach solution was treated in a series of crystallisations, filtering, and washing. An intermediate product was produced and re-dissolved in hot water at 80 ºC for 0.5 h to minimise Li2CO3 content into a solution. A final solution was crystallised repeatedly to produce the final product. A 99.5 % of LiOH·H2O was characterised and passed the impurity specification by the manufacturer as a marketable product. The proposed process is relatively simple to utilise to scale up for bulk production, and it can also be contributed to the battery recycling industry as the spent LIBs are expected to overabundance within the near future. A brief cost evaluation confirms the process feasibility, particularly, for the company that produces cathode active material (CAM) and generates WBP in their own supply chain.

A study on pyro-hydrometallurgical process for selective recovery of Pb, Sn and Sb from lead dross.

This study is to propose a pyro-hydrometallurgical process for recovering Pb, Sn, and Sb from lead dross (LD), incorporating stages of roasting, leaching, and precipitation. The LD, containing 67.2% of Pb, 4.0% of Sn, and 1.4% of Sb, was first roasted at 750 °C for 2 h to oxidise the sulphide metals. Approximately 90% of Pb was oxidised from the first roasting. The LD was second roasted by mixing with 95% H2SO4 for sulphatising at 300 °C for 3 h to break the complex oxide structure of the oxyplumboromeite (Pb2Sb2O7). After the two-step roasting process, over 99% of Pb was oxidised and Sb was separated. The calcine obtained was desulphurised by 2 M Na2CO3 solution for insoluble PbSO4 to PbCO3 for selective leaching. The residue was then leached in 2.5 M HNO3 at 50 °C for 3 h and over 99% of Pb dissolved into the solution while Sn and Sb remained in the solid residue. The Pb containing solution was neutralised at pH 8 using 2 M Na2CO3 and over 99% Pb was precipitated as PbCO3 and Pb hydroxides. A residue containing Sn and Sb was leached in 7 M NaOH at 95 for 1 h and over 99% Sn was leached selectively. Sn in the solution was precipitated at pH 7 using 2 M H2SO4 as SnO2. Sb was recovered as Sb2O3 in solid reside from Sn leaching. The overall recovery rates of Pb, Sn, and Sb from the LD were 99.5%, 95.4%, and 86.3%, respectively. The proposed process is expected to contribute to recycling Pb and other metal values from LD by minimising hazardous waste emissions.

Recovery of tungsten and cobalt from cemented tungsten carbide wastes using carbonate roasting and water leaching.

Tungsten and cobalt can be recovered from cemented tungsten carbide-Co (WC-Co) scraps using a novel process based on roasting of the scraps with Na2CO3 and water leaching of the calcine containing Na2WO4. The process was evaluated at different conditions testing several parameters, including Na2CO3/WC molar ratio 1:1-2:1, roasting temperature in the range 400-900°C, roasting time at 1-3 h, leaching temperature at ambient - 90°C and leaching time 1-3 h. Thermodynamic modeling of the roasting and leaching stages using HSC and Stabcal software allowed an understanding of the speciation of the products at different conditions. Subsequent experiments to determine an optimized conditions of both roasting and water leaching were conducted aiming to yield recoveries of both W and Co in solution over 95%. Roasting of the scraps with Na2CO3 at 800°C in 3 h using Na2CO3/WC molar ratio of 1:1 will produce a Na2WO4 which can be easily water leached. Over 95% W was recovered after water leaching the calcines at 60°C in 3 h. Over 94% Co was recovered as soluble CoSO4 using 1 M H2SO4 with 1.5 M H2SO3 added as a reductant. The double step of CoCO3 precipitation using 1.5 M Na2CO3 and conversion to Co(OH)2 using 8 M NaOH allows a high purity product (>99.5% purity) to be produced with most of the major impurities as Cr and Fe remaining in the residual Na2CO3 and NaOH liquors. The levels of these major impurities are 0.05% Cr and 0.3% Fe in the final Co hydroxide product.Implications: This is a new process to treat W-Co wastes and recover W and Co. To the best of our knowledge, no study was done before in this area.

Processing of ferromanganese fumes into high-purity manganese sulphate monohydrate.

A novel process for recovering manganese (Mn) values from two types of ferromanganese (FeMn) fumes (wastes from smelters) was developed incorporating several stages of leaching, roasting, and precipitation. Fumes containing high K (as K2O) were first leached in water at ambient conditions to recover K, purified to remove metal impurities, and precipitated as 99.9% pure K2SO4. The residues containing Mn and remaining impurities were then leached in sulfuric acid at 80 °C using oxalic acid as reductant yielding Mn sulphate liquors. The leach liquors were then purified to remove Fe at 95°C as jarosite at pH 2.8 and Fe(III) hydroxides at pH 5.5. Base and transition metal impurities were removed as sulfides. K, Na, Ca, and Mg were separated from Mn in a novel step based on precipitation of Mn hydroxide from the Mn sulphate liquors. Subsequent redissolution of Mn hydroxide in sulfuric acid under reducing conditions using oxalic acid produced a high purity Mn sulphate liquors containing <5 ppm of these impurities. After evaporation by boiling manganese sulphate monohydrate (MSM) was crystallized, achieving purity >99.5%. Another fume containing oil contaminants having low K was subjected to sulfation roasting to produce Mn sulphate. The calcine was water leached to produce Mn sulfate liquors and treated in a similar process to make high purity MSM. Novelty aspects of the proposed process include the recovery of K as a by-product and separation of Mn2+ from other impurities by its selective precipitation to produce a high purity Mn sulphate solution for MSM recovery. Implications: Ferro-manganese smelters discharge wastes (fumes) that are collected by bag filters or electrostatic precipitators and then store them on the ground with coverage to prevent their dispersion into the surrounding atmosphere. The work presented propose a process to reprocess these wastes into high purity manganese sulphate monohydrate (MSM). Processing of wastes containing high potassium as KOH also yields high purity potassium sulfate used as fertilizers. Some smelters spray oil onto the fume dumps to prevent material being blown onto the surrounding area. The oil-contaminated fumes have to be first roasted to burn off the oil, before they can be processed. The process developed therefore converts a hazardous waste into high purity MSM and potassium sulphate used in many applications.

Recovering valuable metals from recycled photovoltaic modules.

Recovering valuable metals such as Si, Ag, Cu, and Al has become a pressing issue as end-of-life photovoltaic modules need to be recycled in the near future to meet legislative requirements in most countries. Of major interest is the recovery and recycling of high-purity silicon (> 99.9%) for the production of wafers and semiconductors. The value of Si in crystalline-type photovoltaic modules is estimated to be -$95/kW at the 2012 metal price. At the current installed capacity of 30 GW/yr, the metal value in the PV modules represents valuable resources that should be recovered in the future. The recycling of end-of-life photovoltaic modules would supply > 88,000 and 207,000 tpa Si by 2040 and 2050, respectively. This represents more than 50% of the required Si for module fabrication. Experimental testwork on crystalline Si modules could recover a > 99.98%-grade Si product by HNO3/NaOH leaching to remove Al, Ag, and Ti and other metal ions from the doped Si. A further pyrometallurgical smelting at 1520 degrees C using CaO-CaF2-SiO2 slag mixture to scavenge the residual metals after acid leaching could finally produce > 99.998%-grade Si. A process based on HNO3/NaOH leaching and subsequent smelting is proposed for recycling Si from rejected or recycled photovoltaic modules. Implications: The photovoltaic industry is considering options of recycling PV modules to recover metals such as Si, Ag, Cu, Al, and others used in the manufacturing of the PV cells. This is to retain its "green" image and to comply with current legislations in several countries. An evaluation of potential resources made available from PV wastes and the technologies used for processing these materials is therefore of significant importance to the industry. Of interest are the costs of processing and the potential revenues gained from recycling, which should determine the viability of economic recycling of PV modules in the future.

Selective leaching of valuable metals from waste printed circuit boards.

This study was carried out to recover valuable metals from the printed circuit boards (PCBs) of waste computers. PCB samples were crushed to smaller than 1 mm by a shredder and initially separated into 30% conducting and 70% nonconducting materials by an electrostatic separator. The conducting materials, which contained the valuable metals, were then used as the feed material for magnetic separation, where it was found that 42% of the conducting materials were magnetic and 58% were nonmagnetic. Leaching of the nonmagnetic component using 2 M H2SO4 and 0.2 M H2O2 at 85 degrees C for 12 hr resulted in greater than 95% extraction of Cu, Fe, Zn, Ni, and Al. Au and Ag were extracted at 40 degrees C with a leaching solution of 0.2 M (NH4)2S2O3, 0.02 M CuSO4, and 0.4 M NH4OH, which resulted in recovery of more than 95% of the Au within 48 hr and 100% of the Ag within 24 hr. The residues were next reacted with a 2 M NaCl solution to leach out Pb, which took place within 2 hr at room temperature.

Kinetics analysis of growth and lactic acid production in pH-controlled batch cultures of Lactobacillus casei KH-1 using yeast extract/corn steep liquor/glucose medium.

This study was performed to determine the optimal conditions of yeast extract, corn steep liquor and glucose concentration for the growth and lactic acid production of Lactobacillus casei KH-1 and to assess the effect of these conditions using a response surface methodology. A Box-Behnken design was used as an experimental design for the allocation of treatment combination as 17 pH-controlled batch cultures. The growth and product parameters were estimated by Gombertz, Leudeking and Piret models from experimental data, and analyzed statistically with response surfaces. The effects of yeast extract, corn steep liquor and glucose were significant for the maximum specific growth rate, mu(max) and the maximum biomass concentration, X(ma). The interaction of corn steep liquor and glucose indicated that the positive or negative effect of glucose on mu(max) in corn steep liquor below or above 2.1% could be explained by the glucose-dependent availability of a nutrient on mu(max) Although the experiment was achieved in pH-controlled batch culture for L. casei KH-1, the growth- and non-growth-associated production rate parameters, a and b, were significant in the response surface model. The growth and lactic acid production of L. casei KH-1 were strongly affected by glucose and the importance of the media composition was demonstrated. The estimated optimal conditions of the growth and lactic acid production of L. casei KH-1 were 1.276% and 0.697% for yeast extract, 3.505% and 1.708% for corn steep liquor, and 2.390% and 2.215% for glucose, respectively.