Flotation Tailings Research Articles

Review of the Failure at the Flotation Tailings Storage Facility of the “Stolice” Mine (Serbia)

A detailed analysis of the accident that occurred at the flotation tailings storage facility (TSF) at the inactive mine “Stolice” in 2014 is provided in this paper. All factors that caused the accident have been analyzed, with a review of the consequences of the accident, their accident class according to the Global Industry Standard on Tailings (GISTM), and the implemented measures to rehabilitate the TSF and the surrounding area after the accident. It has been concluded that the TSF had not been properly maintained even before the accident occurred and that the unfavorable weather conditions in Serbia in the May of that year contributed to the filtration disturbance and multiple tailings spillages from the TSF. It has been stated that the consequences according to the GISTM span from “low” to “significant”, with the group of environmental consequences having the highest rank (3). Although the accident occurred without recorded human casualties, with the damage being of a local nature, it is considered one of the most significant accidents at a TSF in Serbia in the last 20 years. The reconstructed TSF is considered stable now, with a low-to-medium risk of failure.

Study on the effect and mechanism of coarse magnetite on the flotation of fine-grained hematite

Lean hematite ore is one of the refractory iron ores in China due to its fine crystal size, deep oxidation degree and complex mineral composition. With the deep mining of mine ore, the disseminated particle size of iron minerals is finer, and the ore properties are more complex and difficult to be separated. Therefore, the research on high-efficiency beneficiation technology of fine-grained lean hematite ore is the requirement and important direction of the development of refractory iron ore beneficiation. In this paper, the carrier flotation method is used to comprehensively recover hematite, and the addition of coarse-grained magnetite can effectively improve the recovery rate of fine-grained hematite. The flotation results show that when the ratio of magnetite to hematite is 1: 1, the recovery rate of iron ore can reach more than 96.34%, and the efficient recovery is realized. Based on the results of pure mineral test, the carrier flotation test of Anshan lean hematite ore was carried out. The results show that the reverse flotation concentrate product with TFe grade of 67.88% and total iron recovery of 84.81% can be obtained, and the TFe grade of the reverse flotation tailings is 24.20%. It can be seen that compared with the traditional process, the grade of carrier flotation tailings is 2.37 percentage points lower than that of the original reverse flotation tailings; the carrier reverse flotation process can effectively improve the recovery rate of iron concentrate. In addition, the theoretical analysis of carrier flotation was carried out by contact angle, zeta potential, mineral phase microscope, SEM and XPS. It was found that the collision probability between fine mineral particles and bubbles was low, which limited the recovery of fine mineral particles in the flotation process. Fine-grained hematite can be non-selectively attached to the carrier particles, thereby affecting the adsorption of starch on it, thereby improving the recovery rate of fine particles; The carrier reverse flotation can further improve the flotation efficiency by gravity and shear force. Carrier reverse flotation provides a new idea for effective recovery of iron ore.

Adsorption-catalytic synergistic Fenton degradation of potassium butyl xanthate in flotation tailing wastewater by renewable iron-loaded sludge: Performance, kinetics and mechanism

Adsorption-catalytic synergistic Fenton degradation of potassium butyl xanthate in flotation tailing wastewater by renewable iron-loaded sludge: Performance, kinetics and mechanism

Synergistic organic manure treatment with Al/Fe/Ca-based fluoride-fixing agents promote soil formation and utilization of phosphate flotation tailings

Soil utilization of phosphate ore flotation tailings (PTs) was achieved using F-fixing agents combined with organic manure treatment to address issues related to inefficient stabilization of P-F, heavy metal solidification/stabilization, poor physicochemical properties, and ecological disruption. The formulation for PTs soil utilization included 1.75 % polyaluminum sulfate (PAS), 1.75 % FeSO4, and 0.25 % CaO, with PTs content at ≥86.75 %. This approach enhanced water retention and improved nutrient and biochemical conditions. Various nutrient indexes met general planting soil requirements: organic matter 24.93 g/kg, available K 252.26 mg/kg, available Ca 2048.67 mg/kg, available Mg 246.13 mg/kg, and ammonia-nitrogen 42.47 mg/kg. The water-soluble P content of PTs-based soil decreased to 6.96 mg/kg, while available P increased to 677.99 mg/kg. Mn leaching toxicity was less than 0.016 mg/L, with a stabilization efficiency of 96.86 %. Water-soluble F in PTs-based soil was reduced to 11.133 mg/kg. This study maximized phosphorus resource utilization and prevented the migration of water-soluble P, F, and Mn to the surrounding environment. Potting experiments showed PTs-based soil was more effective than red soil and PTs-based raw materials in cabbage plantation, achieving a maximum seedling emergence rate of 98.33 %. Microbial diversity increased, and community structure improved in 40-day soil formation experiments, with PTs-based soils developing microbial communities involved in carbon and nitrogen cycling, enhancing resistance to external environmental disturbances, and promoting ecological utilization. These findings offer practical insights into the ecological utilization of large quantities of PTs and present a cost-effective approach for producing planting soils or ecological mulches from similar solid wastes.

Synthesis of sodium 1,3-propanebis(dithiocarbamate) and study of its flotation properties

Development and selection of a suitable reagent for efficient flotation is of great importance for ore beneficiation processes. Dithiocarbamates represent a class of organic compounds that are widely used as reagent-collectors in ore flotation. In this work, we conducted an analytical review of the methods currently used for obtaining this class of compounds and established the main method for obtaining dithiocarbamate salts to be the interaction of primary or secondary amines with carbon disulfide (CS2) in the presence of potassium or sodium hydroxide. The aim of the reviewed work was to synthesize sodium 1,3-propanebis(dithiocarbamate), as well as to study its flotation properties in the beneficiation of gold-bearing sulfide ores. A simple and effective method for sodium 1,3-propanebis(dithiocarbamate) synthesizing was proposed. Its structure was confirmed using physicochemical methods (infrared spectroscopy, 1H, 13C nuclear magnetic resonance spectroscopy). A set of studies found that the combined use of potassium butyl xanthate and sodium dialkyl dithiophosphate (BTF-1552) with sodium 1,3-propanebis(dithiocarbamate) led an increase in the efficiency of extraction of the target component of up to 81.63%, as well as an increase in its content in the concentrate of up to 7 g/t. In addition, the combination of sodium 1,3-propanebis(dithiocarbamate) with potassium butyl xanthate was shown to enable an extraction efficiency of 80.51%, as well as a valuable component content in the concentrate of up to 6.5 g/t with a low gold content in flotation tailings of 0.07 g/t. The proposed synthesis method and the results of flotation tests of gold-bearing ores are of interest for further research in the field of organic synthesis and practical application.

Production of low-sulphur tailings by hydrocycloning

Acid mine drainage (AMD) is an environmental problem in gold mines containing sulphur minerals. The potential for acid drainage in tailings facilities must be considered when studying their dry closure. Therefore, controlling the sulphur content in the final layer of tailings deposition has the potential to prevent AMD after closure. In this context, the rougher flotation tailings were hydrocycloning in a pilot plant in a gold mining company to assess the potential for generating modified tailings with low-sulphur content. A total of 10 samples were composited over one hour of pilot plant operation. The hydrocyclone overflow showed an average reduction of 36.2% in sulphur content and an average metallurgical recovery of 20.0%. The underflow had an average increase in sulphur content of 16.8% and 80.0% metallurgical recovery. These results demonstrated the possibility of generating modified tailings with reduced sulphur content through hydrocycloning.

Optimizing ash content detection and prediction in flotation tailings using a new approach to enhance feature extraction and deep learning algorithms

ABSTRACT Flotation is a crucial separation technique in coal beneficiation. However, the inability to quickly and accurately detect ash content during the flotation process hinders the development of flotation automation. This paper introduced a novel hybrid model based on convolutional neural networks, establishing the ash content in tailings. In this method, convolutional neural network is utilized to automatically extract features from tailings images. In the model training and optimization phase after inputting the augmented data, the optimal ResNet18 model is identified by comparing five model optimization methods and replacing the SoftMax classifier with the traditional SVR in ResNet18. Additionally, a PCA-SVR model for predicting tailings ash content is constructed for comparative analysis, employing six different feature combinations to validate the effectiveness of the proposed features, with optimal support vector parameters identified through random search and cross-validation algorithms. The prediction results show that the PCA-SVR model has the highest accuracy when the feature dimension is reduced to 16, and the prediction of ResNet18-SVR is more accurate compared to PCA-SVR. This indicates that the proposed method can predict the ash content more accurately than the traditional feature engineering methods and has a broad application prospect in the field of coal beneficiation.

Flotation Tailings from Cu-Au Mining (Bor, Serbia) as a Potential Secondary Raw Material for Valuable Metals Recovery

The increased exploitation of ores leads to the generation of mining waste, which has a negative impact on the environment and human health. For this reason, it is necessary to take care of it in an adequate way by applying some of the possible treatments. In addition to protecting the environment by applying appropriate treatment, there is also the possibility of making a profit by valorizing useful elements from mining waste. In order to choose the most adequate treatment, it is necessary to perform the characterization of mining waste. This paper contains a detailed characterization of the flotation tailings deposited at the Old Flotation Tailings in eastern Serbia, originating from copper ore processing. Characterization includes physico-chemical analysis, polarizing microscope analysis, X-ray Diffraction analysis (XRD) and Scanning Electron Microscopy with Energy Dispersive Spectroscopy (SEM-EDS) analysis analysis. The obtained results indicate that the investigated flotation tailings can be used as a secondary raw material for metal recovery, in this case primarily copper (whose content is about 0.24%), gold (with a content of about 0.43 ppm) and silver (with a content of about 1.7 ppm). Considering that the content of valuable elements is quite low, it is suggested to apply hydrometallurgical treatment for their recovery.

Acid mine drainage activation mechanism on lime-depressed pyrite flotation from copper sulfide ore

The lime-depressed pyrite from Cu differential flotation tailings with acid mine drainage (AMD) as a natural activator was recovered. The effect of AMD on lime-depressed pyrite flotation was investigated by a series of laboratory flotation tests and surface analytical techniques. Flotation test results indicated that AMD could effectively activate the pyrite flotation with a sodium butyl xanthate (SBX) collector, and a high-quality sulfur concentrate was obtained. Pulp ion concentration analysis results indicated that AMD facilitated desorption of Ca2+ and adsorption of Cu2+ on the depressed-pyrite surface. Adsorption measurements and contact angle analysis results confirmed that adding AMD improved the adsorption amount of SBX collector on the pyrite surface and increased the contact angle by 31°. Results of Raman spectroscopy and X-ray photoelectron spectroscopy analysis indicated that AMD treatment promoted the formation of hydrophobic species (S0 hydrophobic entity and copper sulfides) and the removal of hydrophilic calcium and iron species on the pyrite surface, which reinforced the adsorption of collector. The findings of the present research provide important theoretical basis and technical support for a cleaner production of copper sulfide ores.

Effect of dissolved metal ions from mineral surfaces on the surface wettability of phosphate ore by flotation

The dissolved metal ions have a significant effect on the wettability of mineral surface in flotation slurry. In this work, the composition and characteristic of phosphate ore raw (R), flotation concentrate (FC), and flotation tailing (FT), the dissolution of metal ions from mineral surface and its effect on mineral surface wettability were investigated. The results of the dissolution test indicated that the concentrations of Ca2+ and Mg2+ are high (FT > R > FC), while Fe3+ and Al3+ were low in the slurry. H3PO4, as the slurry pH regulator, has the weakest dissolved ability of metal ions from mineral surface compared with HCl, HNO3 and H2SO4. The surface tension tests revealed that metal ions will combine with NaOL in the slurry, causing the consumption of NaOL. The contact Angle tests performed that some of NaOL adsorption sites on the FC and FT surface were covered the hydrolysates of Fe3+ and Al3+, which inhibited the hydrophobicity of the FC and FT surfaces. Ca2+ and Mg2+ consumed part of NaOL in the slurry, inhibiting the hydrophobicity of the FT surface. However, the adsorption of Ca2+ and Mg2+ on the FC surface increased the adsorption of NaOL, which enhanced the hydrophobicity of the FC surface.

Linking ores flotation results to the Kinsenda copper deposit mineralogical features: enhancing the recovery of copper retained in flotation tailings

ABSTRACT Since the Kinsenda concentrator commissioning in 2015, the selective flotation of mixed copper ores (67% as sulphides and 39% as oxides) from the Kinsenda deposit (DR Congo results in a recovery yield of 57% of copper as oxides versus 98% as sulphides.. Talings from the flotation of sulphide minerals contain 1.14% as total Cu including 0.38% as oxides. This research adopted a geometallurgical approach for enhancing the recovery of not floated copper. Consequently, it thoroughly studied the Kinsenda copper deposit mineralogy to explain causes of poor results of the industrial flotation of copper and conducted chemical and granulochemical analyses of tailings to describe and comprehend the copper behaviour during flotation. The effects from particles size in tailings, and the reagents’ suite components and their dosages on the copper flotation efficiency were investigated. Flotation results were significantly improved when tailings contain 70% particles with size smaller than 106 μm were floated using a reagents suite including 3600 g per t NaHS (sulphidiser), 360 g per t SNPX (collector) and 360 g per t AM810 (promoter). Indeed, the concentrate's grade in total copper reached 23.75%, with a recovery yield of 91.07% and 15.41% in oxidised copper, with a recovery yield of 91.26%.

Concentration of Silver from Recycling of Fine Powder of Wasted Videogame Printed Circuit Boards through Reverse Froth Flotation and Magnetic Separation Processes

Electronic waste stream grows day by day; printed circuit boards are a kind of solid waste that accounts for 6% of electronic waste. When these are discarded, they can cause soil, water, and air contamination; however, if recycled, these can be considered as a secondary source of metals. Physical comminution of printed circuit boards generates particles with sizes smaller than 250 µm, which are typically not included in the recycling process because they are considered as dust and unvaluable; nevertheless, precious and base metals can be found in these particles. The concentration of metals like silver, among others, from these particles can be achieved by reverse froth flotation in a flotation column followed by magnetic separation of the tails products. A mass balance of the flotation column feed, concentrate, and tails indicates that using a pulp modified with 5 ppm methyl isobutyl carbinol plus 5 g/ton oleic acid (both biodegradable reagents), the concentration of the products improved, resulting in recoveries of 86.13 and 13.87% in the concentrate and tails zones, respectively, with a grade of 74.4% in the tails flow. Magnetic separation of the tails product increases slightly the concentration of silver, reaching a silver grade of 74.5%, a recovery amount similar to those obtained employing complex and environmentally unfriendly processes.

The role of microbubble dose in combined microflotation of fine particles

Flotation of small particles is one of the global challenges facing the mineral raw materials processing industry. Large amounts of non-ferrous and rare metals are lost in the flotation tailings in the form of mineral particles below 15 µm in size as a result of the low effectiveness of their capture by coarse bubbles generated in conventional flotation machines. The method of combined microflotation, developed in recent years, uses conventional coarse bubbles (CB) and microbubbles (MB) produced in the stand-alone generator of air-in-water microdispersion, which serves as the flotation carriers. Depending on the MB dose, the effect of their application may be positive or negative. The theoretical analysis of various mechanisms of particle transfer onto the surface of coarse bubbles and further into the froth layer allowed to obtain the formula for the optimal MB dose f=d<sub>d</sub>/2d<sub>p</sub>r<sub>p</sub>, where d<sub>d</sub> is MB size; d<sub>p</sub> and r<sub>p</sub> respectively are the size and the density of particles. Experiments performed on the copper ore flotation tailings at the Atalaya Mining (Spain) and Chaarat Kapan (Armenia) concentrators showed that, besides the optimal MB dose in the range of 1-3 ml/g, there is another optimal MB dose in the range of 10-20 ml/g, where the copper recovery increases by several percent compared to the reference test (f = 0). The deep minimum in copper recovery is observed in the area between the optimal MB doses, which is by several percent lower than the value in the reference test.

Enhanced recovery of low-grade copper ore and associated precious metals from iron tailings: A case study in China

In the processing of copper and precious metal recovery in a mine in Chengde, Hebei Province, the decrease in copper mineral content and finer grain size distribution of the selected ores have caused challenges in the existing milling process. This has led to difficulties in achieving significant dissociation of chalcopyrite, resulting in decreased separation efficiency of copper and associated precious metals. Particularly, there is a notable loss of precious metals in the flotation tailings. Using BPMA and other detection equipment, the occurrence state, distribution patterns, and liberation degree of copper and associated precious metals in iron tailings were investigated. Subsequent flotation experiments examined the impact of grinding regimes on mineral liberation and reagent regimes on mineral enrichment. Subsequently, flotation experiments were conducted to examine the influence of grinding regimes on mineral liberation and reagent regimes on mineral enrichment. The research findings revealed that copper and precious metal contents in iron tailings are extremely low and finely disseminated within gangue minerals, posing challenges for liberation. Under the optimal conditions determined from flotation experiments, a laboratory closed-circuit process consisting of one rougher, one scavenger, and two cleaner stages yielded a copper concentrate with a grade of 11.33 %, a recovery efficiency of 66.95 %, and containing 9.42 g/t gold, 52.72 g/t silver, 2.78 g/t platinum, and 8.46 g/t palladium. The SEM-EDS analysis further emphasizes the necessity to tailor grinding and flotation processes to optimize the liberation and subsequent recovery of these valuable minerals, avoiding excessive grinding of gangue minerals which could negatively impact flotation performance.

Processing the Fluorocarbon-containing Waste from Electrolytic Aluminum Production

The characteristics of fine fluorocarbon-containing waste from aluminum production are given and the influence of changes in aluminum production technology that have occurred in recent years on the formation structure, composition and properties is analyzed. The currently proposed methods for processing these wastes are reviewed. An assessment of the method for processing fine fluorocarbon-containing waste from aluminum production was carried out on the example of coal foam flotation tailings using the method of causticization with lime milk to produce a product containing synthetic fluorite, used as a mineralizer for the preparation of straight cement. The influence of operating parameters (temperature, process time, dilution) on the completeness of causticization of raw materials has been determined. The greatest influence on the extraction of fluorine into fluorite and sodium into solution is exerted by the liquid to solid ratio (L:S), at that, the yield of fluorite reaching 220 kg per ton of supplied raw material. It has been found that the high alkalinity of the solution allows sodium hydroxide to be extracted from it and used after causticization as a gas cleaning solution.

Effects of micro- and nano-bubbles on the flotation separation of unburned carbon from coal fly ash

ABSTRACT Nanobubble technology has found extensive use in enhancing the flotation efficiency of various minerals. The utilization of nanobubbles in flotation encompasses bulk nanobubbles, surface nanobubbles, and microbubbles; however, a comprehensive comparison of their respective roles is lacking. Furthermore, there is a dearth of studies on employing nanobubble systems in the flotation of coal fly ash. The coal fly ash characters were initially studied to reveal its ultrafine size (d50: 24.80 μm) and porous structure that includes a multitude of hydrophilic functional groups. These attributes contribute to the inadequate hydrophobicity of unburned carbon within the ash. Subsequently, the effects of different collector types, frother types, and air flow rates on the efficiency of carbon removal using conventional flotation methods were examined. Comparative analysis showed that conventional flotation using diesel oil outperformed that which employed kerosene as the collector. Additionally, higher carbon removal efficiency was achieved using MIBC, with a longer carbon chain length, as the frother, as opposed to sec-Octyl alcohol. Furthermore, a direct relationship between air flow rate and coal fly ash flotation performance was observed – higher air flow rates corresponded to improved flotation outcomes. Finally, a comparison between conventional flotation and nanobubble-assisted flotation (utilizing bulk nanobubbles, surface nanobubbles, and micro-nanobubbles) was conducted. Remarkably, micro-nanobubble flotation yielded the most promising results, with the lowest loss-on-ignition in flotation tailings (3.15%) and the highest removal efficiency of unburned carbon (85.01%). These findings satisfied the loss-on-ignition requirements of Grade I fly ash standards.

Experimental research on combined magnetic separation and flotation for coal gasification fine slag

ABSTRACT To address the challenging issue of achieving high-precision carbon-ash separation in coal gasification fine slag, magnetic separation and flotation techniques were employed to study their respective separation effect. Optimization tests were conducted to determine the optimal magnetic field gradient and magnetic field intensity, and the ash content of magnetic separation concentrates was 72.18%, with a reduction of 7.28% compared to the sample. Additionally, the magnetic separation tailings exhibited an ash content of 93.99%. Through flotation tests, the optimum conditions for kerosene and composite collector was established. The ash content of flotation concentrates decreased by 16.39% compared to the sample, and the ash content of flotation tailings was 95.83%. A comparative analysis revealed that the combined separation process outperforms single separation methods in reducing the ash content of the concentrates, and realize the further quality improvement and separation effect. Detailed analysis of magnetic characteristics and XRD, XPS analysis studies on the sample and the products, indicated that the magnetic separation effectively pre-removes a certain amount of magnetic high ash minerals. Furthermore, the collector treatment significantly increased hydrophobicity difference between residual carbon and ash matter, which ultimately enhancing the flotation effect.

Gold recovery from cyanidation residue by chloride leaching and carbon adsorption – Preliminary results from CICL process

There is a vast amount of globally underutilized low-grade mine tailings and leach residues, including those from primary processing of gold. In this research, the target is to recover the remaining gold (10.9 g/t) from weathered refractory iron-rich residue that had previously been subject to autoclave oxidation, subsequent cyanidation in a conventional carbon-in-leach (CIL) circuit as well as storage at tailings area. Chloride leaching has been considered as one of the most promising cyanide-free gold leaching methods and it has shown positive outcomes in treating primary gold ores, concentrates, and flotation tailings. Therefore, in the current study, the iron-rich residue investigated was subjected to chloride leaching combined with simultaneous carbon adsorption. The investigated parameters included leaching time (2–8 h), chloride concentration ([Cl−] = 0.2–5 M), type and concentration of oxidant ([Cu2+]/[Fe3+] = 0.1–1 M), as well as type and concentration of activated carbon (14–25 g/L), whereas S/L ratio (100 g/L), acidity (pH = 1), and temperature (90 °C) were kept constant. Leaching results indicate that up to 40% of the remaining gold could still be recovered from the investigated residue with optimized chloride leaching. According to the results, the most important parameter for gold recovery was the leaching time. Moreover, of the studied oxidants, cupric ions were shown to contribute more to gold recovery when compared to ferric ions (35% vs. 24% at [Cu2+]/[Fe3+] = 0.1 M). Nevertheless, an increase of cupric concentration from 0.1 M (low-concentrated) to 0.5 M, resulted in only a slight increase in gold recovery (from 36% to 40%), whereas no further improvement in gold recovery was achieved with a 1 M cupric concentration. Two studied activated carbon products showed equal effectiveness in gold adsorption. In-situ carbon adsorption was shown to occur effectively in chloride media, as all dissolved gold could be detected in the activated carbon, and the concentration of remaining gold in the pregnant leach solution was minimal (< 0.02 mg/L). These findings indicate that low-concentrated chloride leaching of leach residues from industrial gold processes can allow an enhanced recovery of gold from previously mined and treated raw materials.

Beneficiation process reengineering under the background of green development-- take the full utilization of a low-grade lead-zinc ore as an example

A zero-waste mine is an advanced mine development and resource utilization strategy that aims to encourage the reduction of mine solid waste at source and the utilization of all its components. A low-grade lead-zinc ore in Guangxi, China, contains valuable components like galena, sphalerite, pyrite, barite, etc. and the gangue mineral is primarily dolomite. This paper proposes a flow of “waste pre-discarding by dense medium cyclone - sequential priority flotation of lead and zinc recovery - flotation and gravity combined process for barite recovery” to realize full component utilization. In this technical approach, 38.81% of the waste rock that can be utilized as concrete aggregate was removed from the feed ore through pre-discarding prior to grinding, after that recovers lead and zinc by the sequential priority flotation method with yield of Pb concentrate and Zn concentrate at 7.61%. In order to more efficiently recover barite and pyrite from lead-zinc flotation tailings, the gravity equipment of reflux classifier (RC) was used, which obtained barite concentrate with a yield of 7.97% and pyrite concentrate with a yield of 2.28%. The remaining yield of 43.34% tailings can be used for filling after the recovery of barite concentrate. Utilizing waste rock, lead, zinc, barite, and other resources might help the mine achieve the dynamic balance of “mining-discarding-beneficiation-filling” and realize zero tailings discharge.

A technology for gold extraction from tailings resulting from polymetallic ore beneficiation

We propose a technology for processing of copper-zinc ore flotation tailings with the purpose of extracting gold using ultra-low (10–30 mg/dm3 ) concentrations of NaCN. The material composition of the starting materials was examined using the methods of inductively coupled plasma, gold assay atomic absorption and X-ray diffraction analysis. The NaCN concentration in solution was determined by titrimetric and photometric methods; pH levels were determined by potentiometric analysis. The laboratory studies established optimal conditions for cyanidation of tailings resulting from flotation beneficiation of copper-zinc ores, which underwent preliminary lime treatment: mass fraction of the 0.071-mm mesh minus – 70.5%; duration – 8 h; NaCN consumption – 0.3 kg/t (at a concentration of 30 mg/dm3 ); L:S =1:1. Under these conditions, the gold extraction rate amounted to 32.0–33.6%. Pre-grinding of flotation tailings to the size of 90% of the 30 mm mesh minus led to an increase in gold extraction of up to 41.5–44.7% at an NaCN consumption of 0.6 kg/t of the same concentration and L:S =1.5:1. The results obtained served as the basis for experimental cyanidation of waste tailings according to two schemes, i.e., using tailings of initial coarseness and pre-ground tailings. A good agreement was achieved between the laboratory and experimental results. At present, the use of pre-ground tailings seems unreasonable economically, due to high costs of NaCN (0.6 kg/t vs 0.3 kg/t) and active chlorine (5.6 kg/t vs 1.2 kg/t). According to pilot tests on a flotation tailing sample weighing 67 t (with an Au content of 1.35 g/t), the gold extraction level was 31.9% at an NaCN consumption rate of 0.135 kg/t. As a result, we propose a technology of gold extraction from flotation tailings of copper-zinc ore based on the application of ultra-low concentrations of NaCN. The expected profit can amount to 1276.74 mln RUB per year, with the economic efficiency of 88% and the payback period of 1.4 years).