The strong contamination of the interface is the main problem that results in low flotation efficiency of sphalerite in cyanide tailings. However, the consumption of cyanide and dissolved oxygen, as well as the concentration of ions including Zn²⁺ and SCN⁻ in the leaching solution, decreased with the use of ceramic ball medium. The conclusions obtained from SEM–EDS indicated that the use of ceramic ball medium avoided the excessive surface oxidation caused by the galvanic couple actions between the iron ball medium and the sphalerite. XPS analysis also proved that the chemical environment on the surface of sphalerite was optimized by porcelain ball medium compared with iron ball medium, avoiding the formation of Fe–OOH and Fe–O hydrophilic substances, especially [Fe(CN)₆]³⁻, thus increasing the adsorption of the collector on the surface of sphalerite. Therefore, grinding with ceramic ball medium exhibited excellent performance in terms of the cyanide process, which was approximately 5–10% higher than that obtained by grinding with iron ball media in the flotation test.

To explore the relationship between total cyanide and easily liberatable cyanide in cyanide tailings and its impact on the environment, the leaching and release characteristics of total cyanide and easily liberatable cyanide in a tailings pond were studied using a dynamic leaching experiment, and a dynamic model was established. The results show that the release concentration of total cyanide in cyanide tailings is higher than that of easily liberatable cyanide. With the increase in leaching time, the release of cyanide is more sensitive to leaching intensity. The first-order kinetic, second-order kinetic, modified Elovich, double constant, and parabolic diffusion equations were used to fit the cumulative release of cyanide. According to the kinetic results of cyanide cumulative release, in addition to the first-order kinetic equation, the other four equations revealed good fitting for the cyanide leaching process, which shows that cyanide released under simulated rainwater leaching was not simply via surface diffusion, but controlled by multiple factors.

After long-term accumulation, weathering and rainfall of metals and non-metals in tailings ponds, the chemical composition will change, and the surface heavy metals may gradually migrate to the bottom layer, thus showing a cumulative effect. At present, the distribution of cyanide and heavy metals in vertical space has not been reported. In this work, 12 sampling points were arranged for a cyanide tailings pond by the grid method in Jilin, China. The contents of cyanide and heavy metals were determined according to the standard method. The results show that the heavy metals in the tailings were mainly Cu, Pb, Zn and Mn (342.83 mg/kg, 571.09 mg/kg, 610.15 mg/kg and 796.63 mg/kg, respectively), and the concentration of heavy metals in the horizontal direction does not change significantly. It should be noted that the concentrations of zinc and manganese did not change much in the vertical direction. The concentration of copper and lead increased with the increase in depth. The concentration of cyanide at the sampling site far away from human activities was higher at 6.71 mg/kg, and the average concentration is 2.5 mg/kg. In addition, because the cyanide is unstable and affected by factors such as light and rainfall, the concentration change in the vertical direction fluctuates.

To address the marked differences in the host phases of phosphorus and vanadium, severe ore sliming, and the difficulty of synergistic recovery from the low-grade phosphorus–vanadium-associated ore in Mazharweina, Xinjiang, process mineralogical analysis, phosphorus flotation condition tests, and acid leaching-solvent extraction tests were conducted to investigate phosphorus and vanadium recovery from this ore. The results showed that the P₂O₅ and V₂O₅ grades of the raw ore were 16.71% and 0.798%, respectively. Phosphorus occurred mainly in collophane, whereas vanadium did not form independent minerals and was mainly dispersed in carriers such as mica, argillaceous silicates, and colloidal limonite. P₂O₅ and V₂O₅ showed no obvious natural enrichment in any particle-size fraction, and effective separation of P and V could not be achieved by particle-size classification. Therefore, a pre-desliming-direct phosphorus flotation process was adopted. Under the conditions of a grinding fineness of 80% passing −74 μm and an EM-C2 dosage of 1400 g/t, closed-circuit flotation comprising one roughing, one scavenging, and three cleaning stages produced a phosphate concentrate with a P₂O₅ grade of 28.32% and a recovery of 59.50%, while the P₂O₅ grade of the tailings decreased to 1.52%. Vanadium was mainly enriched in the slime and dephosphorization tailings. At a liquid-to-solid ratio of 8:1, an H₂SO₄ mass concentration of 250 g/L, and a leaching time of 8 h, the V₂O₅ leaching rate from the slime was 77.83%. After N235 extraction, Na₂CO₃ stripping, vanadium precipitation, and calcination of the leachate, a vanadium product with a V₂O₅ purity of 98.42% was obtained, with an overall recovery from the raw ore of 63.83%. The results demonstrate the feasibility of combining phosphorus flotation with hydrometallurgical treatment of the slime and dephosphorization tailings for the comprehensive recovery of phosphorus and vanadium from this complex phosphorus–vanadium ore.