Bastnaesite is one of the most important rare earth element (REE)-bearing minerals; however, its high chemical stability limits REE extraction by direct HCl leaching, necessitating the use of conventional, energy-intensive NaOH cracking. To overcome these processing challenges, microwave-assisted NaOH roasting followed by HCl leaching was developed to enhance REE recovery from Dong Pao bastnaesite concentrate and compared with direct HCl leaching and conventional NaOH roasting. Direct HCl leaching achieved 72.4 % REE recovery under harsh conditions (12 M HCl, 90 °C, 8 h, and 10 mL/g). Conventional NaOH roasting required treatment at 600 °C for 160 min to achieve 98.5 % REE extraction following HCl leaching. In contrast, microwave-assisted NaOH roasting for just 90 s significantly improved subsequent leaching by promoting mineral transformation and increasing mineral reactivity. Process optimization identified a NaOH-to-concentrate ratio of 0.2 (w/w), HCl concentration of 3.7 M, temperature of 90 °C, and liquid-to-solid ratio of 10 mL/g as the optimum conditions, achieving 98.3 % REE extraction. Characterization results confirmed bastnaesite decomposition and formation of Na-containing products during roasting, while kinetic analysis indicated ash layer diffusion-controlled LREE leaching and diffusion-dominated mixed-control HREE dissolution, with activation energies of 8.49 and 20.13 kJ/mol, respectively. Compared with direct HCl leaching, the proposed process reduced HCl consumption by 73.3 % and energy requirements by 40.6 %, while increasing REE production from 215.6 to 292.7 g per kg of concentrate. Compared with conventional roasting, microwave-assisted roasting enabled a comparable REE extraction while reducing the pretreatment time from 160 min to 90 s and energy consumption by 91.1 %.
This work provides the first systematic quantification of how frother chemistry influences air recovery, an important indicator of froth stability and overall flotation performance. Air recovery research has so far for only focused on froth physics for a given flotation reagent suite, thus leaving a gap in our understanding of the links between flotation physics and chemistry. This research offers interesting opportunities for bridging those gaps and enabling advanced optimisation strategies. Two distinct frother types, an alcohol and a polyglycol, and a blend of these, were tested for a sulphide mineral system in a bench-scale continuous flotation tank, allowing air recovery to be measured at a range of superficial gas velocities. Concentrate and tailing samples were taken to quantify the impact of these changes on the metallurgical performance. Results demonstrated that for all chemistries tested, air recovery increased with frother concentration before reaching a plateau. Crucially, this plateau occurred at concentrations consistently higher than the CCC99 . Pyrite recovery exhibited a similar trend with superficial gas velocity, reaching a maximum under conditions corresponding to peak air recovery. A strong positive relationship was observed between air recovery and pyrite recovery for all frother chemistries. However, the magnitude of this relationship depended on frother chemistry. For instance, while DF 250 generated the highest air recovery, the MIBC-DF 250 blend consistently produced the highest pyrite recovery, demonstrating that air recovery alone does not fully determine flotation performance. These findings not only demonstrate a relationship between frother type and air recovery but also challenge the conventional practice of dosing frother solely to control pulp bubble size, with important implications for metallurgical recovery.
Recovering fine spodumene by flotation remains challenging because bubble–particle attachment is inefficient. Selectivity between spodumene and K-feldspar is particularly difficult because the two aluminosilicates present closely similar surfaces. We build on selective separation of spodumene from quartz using Ca(II)-activated flocculation–flotation with anionic polyacrylamide (PAM) and sodium oleate (NaOl). The current study compares sodium oleate with a glycine-based collector (BASF DP-OMC-1396) and evaluates polymer-assisted flocculation-flotation. Flotation is investigated across increasing complexity: synthetic binaries of spodumene and K-feldspar; a ternary spodumene–K-feldspar–quartz mixture; and real ore from Mt Cattlin, Western Australia. Ca(II)-activated NaOl is effective in a narrow pH and dosage window. The glycine-based collector performs strongly in Ca(II)-free spodumene–K-feldspar mixtures at 500 g/t; in the ternary system, which contains more spodumene (30 wt%), comparable upgrading requires 1500 g/t. Expressed per unit of spodumene in the feed, the two dosages are similar (∼4350 and ∼ 5000 g/t of spodumene), indicating that collector demand scales with spodumene content rather than that selectivity is lost. Polymer-assisted flotation with anionic PAM–NaOl increases recovery in the synthetic systems but lowers grade, a response consistent with the flotation of mixed-mineral aggregates, since Ca(II)-activated anionic PAM is known to aggregate both spodumene and K-feldspar. In the Mt Cattlin ore, flocculation-flotation increases mass pull without improving selectivity. Alkaline attrition scrubbing at pH 12.5, which combines slime rejection with alkaline conditioning of the mineral surfaces, enables the glycine-based collector to deliver the only genuine upgrading observed on the ore (∼16–24 wt% spodumene, extending to ∼ 70 % recovery at higher collector dosage).
