Application of NSR-AISC Method in Reserve Estimation and Technical Economic Evaluation of Low-Grade Polymetallic Deposits: Taking a Low-Grade Copper-Molybdenum Mine in Guangdong Province as an Example

The development and utilization of low-grade polymetallic deposits with strategic mineral resources is one of the important measures to alleviate the current high dependence on strategic mineral resources in China. However, domestic mining enterprises and most mining consulting and design institutes usually use general industrial indicators to carry out reserve estimation and technical and economic feasibility studies on low-grade polymetallic deposits, which cannot truly reflect the economic value of such deposits. The article expounds on the definitions of net return value (NSR) and on-site total maintenance cost (AISC) of common ore smelters in the evaluation of overseas mineral resources. Taking a low-grade polymetallic copper-molybdenum mine in Guangdong Province as an example, comparing the research results showed the NSR-AISC method and the general industrial index method in low-grade polymetallic deposit. There are huge differences in the results of reserve estimation;through the further introduction of Taylor’s formula and the research results on the relationship between investment intensity and production scale, a more reasonable mine life and investment scale are recommended, and a more in-depth comparative study has been carried out in the dimension of technical and economic indicators. Based on the comparative study of the above two methods in reserve estimation and the evaluation results of technical and economic indicators, the author believes that the NSR-AISC method can better reflect the true value of low-grade polymetallic ore projects, and should be popularized and applied in resource evaluation and development practice. This article further describes the application status of the NSR-AISC method for reserve estimation and the evaluation of technical economic indicators, and suggests the main points that should be paid attention to in the use of the NSR-AISC method.

Leaching kinetics of low-grade copper ore containing calcium-magnesium carbonate in ammonia-ammonium sulfate solution with persulfate

The leaching kinetics of copper from low-grade copper ore was investigated in ammonia-ammonium sulfate solution with sodium persulfate. The effect parameters of stirring speed, temperature, particle size, concentrations of ammonia, ammonium sulfate and sodium persulfate were determined. The results show that the leaching rate is nearly independent of agitation above 300 r/min and increases with the increase of temperature, concentrations of ammonia, ammonium sulfate and sodium persulfate. The EDS analysis and phase quantitative analysis of the residues indicate that bornite can be dissolved by persulfate oxidization. The leaching kinetics with activation energy of 22.91 kJ/mol was analyzed by using a new shrinking core model （SCM） in which both the interfacial transfer and diffusion across the product layer affect the leaching rate. A semi-empirical rate equation was obtained to describe the leaching process and the empirical reaction orders with respect to the concentrations of ammonia, ammonium sulfate and sodium persulfate are 0.5, 1.2 and 0.5, respectively.

Extraction of valuable metals from low-grade nickeliferous laterite ore by reduction roasting-ammonia leaching method

Nickel and cobalt were extracted from low-grade nickeliferous laterite ore using a reduction roasting-ammonia leaching method.The reduction roasting-ammonia leaching experimental tests were chiefly introduced,by which fine coal was used as a reductant.The results show that the optimum process conditions are confirmed as follows:in reduction roasting process,the mass fraction of reductant in the ore is 10%,roasting time is 120 min,roasting temperature is 1 023-1 073 K;in ammonia leaching process,the liquid-to-solid ratio is 4:1(mL/g),leaching temperature is 313 K,leaching time is 120 min,and concentration ratio of NH3 to CO2 is 90 g/L:60 g/L.Under the optimum conditions,leaching efficiencies of nickel and cobalt are 86.25% and 60.84%,respectively.Therefore,nickel and cobalt can be effectively reclaimed,and the leaching agent can be also recycled at room temperature and normal pressure.

Transformation and leaching kinetics of silicon from low-grade nickel laterite ore by pre-roasting and alkaline leaching process

The mineralogical phase transformation of a low-grade nickel laterite ore during pre-roasting process and the extraction of silicon during alkaline leaching process were investigated.The results indicate that the reaction activity of nickel ores is effectively improved by pre-roasting at650°C for2h,because of the transformation of lizardite into magnesium olivine and protoenstatite.When finely ground ore samples(44-61μm)pre-roasted firstly react with sodium hydroxide solution(60g/L)with a solid/liquid ratio of1:5at140°C for120min,the extraction of silicon can reach89.89%,and the other valuable elements of magnesium,iron and nickel are accumulated in the solid residues.The leaching kinetics of nickel laterite ore can be described successfully by the diffusion through the product layer control model.The activation energy is calculated to be11.63kJ/mol and the kinetics equation can be expressed as1-3(1-x)2/3+2(1-x)=13.53×10-2exp[-11.63/(RT)]t.

Recovery of Zn, Pb, Fe and Si from a low-grade mining ore by sulfidation roasting-beneficiation-leaching processes

To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leaching was proposed. Firstly, several factors such as pyrite dosage, roasting temperature, carbon powder dosage, holding time and particle size affecting on the flotation performance of Zn(Pb) and magnetic separation performance of Fe were simultaneously examined and the optimum process parameters were determined. A flotation concentrate, containing 17.46% Zn and 3.93% Pb, was obtained, and the Zn and Pb recoveries were 86.04% and 69.08%, respectively. The obtained flotation tailing was concentrated by a low-intensity magnetic separator. The grade of iron increased from 5.45% to 43.45% and the recovery of iron reached 64.87%. Hydrochloric acid leaching was then carried out for the magnetic separation tailing and a raw quartz concentrate containing 81.05% SiO2 was obtained. To further interpret the sulfidation mechanism of smithsonite, surface morphology and component of the sample before and after reactions were characterized by XRD and EPMA-EDS. The aim was to achieve the comprehensive utilization of the low-grade mining ore.

Hydrometallurgical leaching and kinetic modeling of low-grade manganese ore with banana peel in sulfuric acid

Manganese was leached from a low-grade manganese ore(LGMO)using banana peel as the reductant in a dilute sulfuric acid medium.The effects of banana peel amount,H2SO4 concentration,reaction temperature,and time on Mn leaching from the complex LGMO were studied.A leaching efficiency of~98%was achieved at a leaching time of 2 h,banana peel amount of 4 g,leaching temperature of 120°C,manganese ore amount of 5 g,and sulfuric acid concentration of 15vol%.The phase,microstructural,and chemical analyses of LGMO samples before and after the leaching process confirmed the successful leaching of manganese.Furthermore,the leaching process followed the shrinking core model and the leaching rate was controlled by a surface chemical reaction(1−(1−x)^1/3=kt)mechanism with an apparent activation energy of 40.19 kJ·mol^−1.

Leaching kinetics of low-grade copper ore with high-alkality gangues in ammonia-ammonium sulphate solution

The leaching kinetics of low-grade copper ore with high-alkality gangues was studied in ammonia-ammonium sulphate solution.The main parameters,such as ammonia and ammonium sulphate concentrations,particle size,solid-to-liquid ratio and reaction temperature,were chosen in the experiments.The results show that the increase of temperature,concentrations of ammonia and ammonium sulphate is propitious to the leaching rate of copper ore.The leaching rate increases with the decrease of particle size and solid-to-liquid ratio.The leaching rate is controlled by the diffusion through the ash layer and the activation energy is determined to be 25.54 kJ/mol.A semi-empirical equation was proposed to describe the leaching kinetics.

Recovery of zinc from low-grade zinc oxide ores by solvent extraction

The recovery of zinc from low-grade zinc oxide ores with solvent extraction-electrowinning technique was investigated by using D2EHPA as extractant and 260 # kerosene as diluent. The results show that it is possible to selectively leach zinc from the ores by heap leaching. The zinc concentration of leach solution in the first leaching cycle is 32.57 g/L, and in the sixteenth cycle the zinc concentration is 8.27g/L after solvent extraction. The leaching solution is subjected to solvent extraction, scrubbing and selective stripping for enrichment of zinc and removal of impurities. The pregnant zinc sulfate solution produced from the stripping cycle is suitable for zinc electrowinning. Extra-pure zinc metal was obtained in the electrowinning test under conventional conditions.

Acid equilibrium during bioleaching of alkaline low-grade sulfide copper ore

This article reports the study on acid equilibrium during bioleaching of alkaline low-grade copper sulfide ore. Adding auxiliary agents 1# （sulfur） and 2# （pyrite） makes bacterial leaching of copper and acid production carried out simultaneously because the auxiliary agents can be oxidized by bacteria and the oxidation products involve acid. The acid required for dissolving alkaline gangue during bacterial leaching is produced, and acid equilibrium is reached during the ore bio-leaching. The recovery of copper reaches more than 95%.

Co-extraction of valuable metals and kinetics analysis in chlorination process of low-grade nickel-copper sulfide ore

To efficiently co-extract Ni and Cu from low-grade nickel-copper sulfide ore,chlorination roasting with NH;Cl followed by a water leaching process was investigated.The results show that 98.4%Ni and 98.5%Cu can be synchronously extracted when the ore particle size is 75-80μm,the roasting time is 2 h,the mass ratio of NH;Cl to ore is 1.6:1 and the roasting temperature is 550°C.The evolution behavior of various minerals was elucidated using X-ray diffraction(XRD)coupled with scanning electron microscopy(SEM).The kinetics of the chlorination process based on the differential thermal and thermogravimetric analysis(DTA-TG)data was analyzed by Kissinger method and Flynn-Wall-Ozawa(FWO)method.The chlorination process of low-grade nickel-copper sulfide ore mainly contains two stages:the decomposition of NH;Cl and the chlorination of ore.The maximum apparent activation energies(Ea)at two stages are determined to be 114.8 and 144.6 kJ/mol,respectively.The condensed product of exhaust gas is determined to be ammonium chloride,which can be recycled as the reactant again,making the process economic and clean.

Experimental study on preferential solution flow during dump leaching of low-grade ores

The phenomenon of preferential solution flow during dump leaching of low-grade ores was studied.The formative mechanism of preferential solution flow was investigated through analyzing the relationship between permeability and ore diameter,and the relationship between surface tension and ore diameter.The preferential solution flow happened within the fine ore area when the dump was unsaturated.And it could happen within the coarse ore area when the dump became saturated.The results of experiment show that the outflow of coarse ore area increases sharply with higher applied rate.The outflow of fine ore area is greater than that of coarse ore area when the applied rate is below 3.2 L/min,and the preferential solution flow happens in fine ore area.But the preferential solution flow happens in coarse ore area when the applied rate is higher than 3.2 L/min.The result of the experiment is consistent with the mechanism analyzing.

Improved recovery of a low-grade refractory gold ore using flotation–preoxidation–cyanidation methods

In this work, different flotation–preoxidation–cyanidation methods are considered for treating a lowgrade refractory gold ore. On the one hand, the results of selective flotation show that 22% and 31.1%of total Sb and As, respectively, remained in the final tailings and only about 28% of the total Au remained for further cyanidation processes. On the other hand, in bulk method of flotation the maximum Au recovery of 90.6% achieved after 60 min of flotation at the grind size with K80 of 146 micron. In addition, the bulk flotation method resulted in the concentrate with low concentrations of Sb and As elements. To improve the recovery of low-grade refractory gold ores, flotation should be followed by roasting, biological, or pressure oxidation processes so that the gold could be liberated prior to cyanidation processes. It is also found that the pressure oxidation pre-treatment of the concentrates prior to cyanidation may yield high gold recoveries of over than 83%. In these processes, recoveries are controlled by the temperature and the oxygen partial pressure in the solvent. However, by utilizing the bio-oxidation technique, the oxidation of sulfur to sulfate cannot be completed and, consequently, the gold recovery may be limited to only 72.2%.

Study on hydrometallurgical process and kinetics of manganese extraction from low-grade manganese carbonate ores

In order to utilize low-grade manganese ore resources effectively, a hydrometallurgical process was developed for manganese extraction in dilute sulfuric acid medium, and the kinetics of leaching manga- nese was also investigated. At room temperature, manganese from low-grade manganese carbonate ores was extracted by sulfuric acid leaching without reductants. During the extracting process, single-factor analysis method was used to evaluate the effects of grinding fineness, sulfuric acid concentration, liquid-to-solid ratio, agitation rate and leaching time on the leaching efficiencies of Mn and Fe. The optimal leaching conditions are determined as coarse particles of below 2 mm size （without ball-milling）, sulfuric acid concentration of 0.86 mol/L, liquid-to-solid ratio of 5：1, agitation rate of 150 r/rain and leaching for 180 min at room temperature. Under the optimal conditions, the leaching efficiencies of Mn and Fe are 96.21g and 13.35%, respectively. In addition, through the experiments at different temper- atures, it is found that the leaching process follows the shrinking core model under the conditions of changing acid concentration and intermittent reaction device. Moreover, the apparent activations of effective diffusion and chemical reaction in the kinetic model are calculated to be 18.83 and 27.15 kJ/mol, respectively.

Reductive recovery of manganese from low-grade manganese dioxide ore using toxic nitrocellulose acid wastewater as reductant

The hydrometallurgical strategy of extracting Mn from low-grade Mn ores has attracted attention for the production of electrolytic manganese metal(EMM). In this work, the reductive dissolution of low-grade Mn O2 ores using toxic nitrocellulose acidic wastewater(NAW) as a reductant was investigated for the first time. Under the optimized conditions of an Mn O2 ore dosage of 100 g·L-1, an ore particle size of-200 mesh, concentrated H2 SO4-to-NAW volume ratio of 0.12, reaction temperature of 90°C, stirring speed at 160 r·min-1, and a contact time of 120 min, the reductive leaching efficiency of Mn and the total organic carbon(TOC) removal efficiency of NAW reached 97.4% and 98.5%, respectively. The residual TOC of 31.6 mg·L-1 did not adversely affect the preparation of EMM. The current process offers a feasible route for the concurrent realization of the reductive leaching of Mn and the treatment of toxic wastewater via a simple one-step process.

Bio-leaching of low-grade large porphyry chalcopyrite-containing ore

In order to economically recover the copper from the low grade copper ores, a bioleaching solvent extraction electrowinning plant with a design capacity of 2 000 t cathode copper per year in Dexing Copper Mine was operated in September, 1997. The results during the 10 month period of the industrial copper dump leaching have been obtained and the approaches have been carried out to enhance the copper dissolution from the waste rocks. The overall operation flowsheet is depicted. The problems confronted in the process and possible way for improving are discussed.

New Technology “Flotation to Form Agglomerates and Magnetic Separation” Allows Great Breakthrough for World Low-Grade Light Rare Earth Ores

Rare earth resources are relatively scarce worldwide, but their global consumption is increasing year-by-year. At present, China has about 36% of the global rare earth reserves, but provides 90% of the world＇s supply, which has generally met world demand and promoted the development of the world economy. In order to continuously and stably supply rare earths to international markets, the Chinese Government has financially supported the Institute of Multipurpose Utilization of Mineral Resources within the China Geological Survey to study the utilization of low-grade rare earth ores. Following many years of experimental research, the project has developed a new technology entitled ＂Flotation to Form Agglomerates and then Magnetic Separation＂, which will bring a technological revolution to the world＇s light rare earth ore dressing.

Review on development of low-grade scheelite recovery from molybdenum tailings in Luanchuan, China: A case study of Luoyang Yulu Mining Company

Luanchuan area is one of the regions with the largest scale of scheelite beneficiation and the largest output of scheelite concentrate in the world. After years of innovation and progress, the beneficiation technology of low-grade scheelite associated with molybdenum tailings in Luanchuan area is becoming more and more perfect. In this study, the development process of low-grade scheelite recycling technology in Luanchuan area was reviewed, including raw ore properties, beneficiation process, flotation equipment and flotation reagents. Meanwhile, taking Luoyang Yulu Mining Co., Ltd. as an example, the effects of various technical transformations such as the optimization of the beneficiation process, the column-machine combined process, and the high-efficiency flotation reagents were elucidated in detail. However, the recycling technology of low-grade scheelite of Luanchuan area is still possible to be improved. As a result, coupled with the latest research progress, the development direction of low-grade scheelite beneficiation in Luanchuan area was also prospected. It is of great significance to further improve the recovery efficiency of low-grade scheelite resources in Luanchuan area and this can provide technical reference for other scheelite plants.

Metallogeny of the Baiyangping Lead-Zinc Polymetallic Ore Concentration Area, Northern Lanping Basin of Yunnan Province, China

The Lanping Basin in the Nujiang-Lancangjiang-Jinshajiang （the Sanjiang） area of northeastern margin of the Tibetan Plateau is an important part of eastern Tethyan metallogenic domain. This basin hosts a number of large unique sediment-hosted Pb-Zn polymetallic deposits or ore districts, such as the Baiyangping ore concentration area which is one of the representative ore district. The Baiyangping ore concentration area can be divided into the east and west ore belts, which were formed in a folded tectogene of the India-Asia continental coUisional setting and was controlled by a large reverse fault. Field observations reveal that the Mesozoic and Cenozoic sedimentary strata were outcropped in the mining area, and that the orebodies are obviously controlled by faults and hosted in sandstone and carbonate rocks. However, the oreforming elements in the east ore belt are mainly Pb-Zn -Sr-Ag, while Pb-Zn-Ag-Cu-Co elements are dominant in the west ore belt. Comparative analysis of the C-O-Sr-S-Pb isotopic compositions suggest that both ore belts had a homogeneous carbon source, and the carbon in hydrothermal calcite is derived from the dissolution of carbonate rock strata; the ore- forming fluids were originated from formation water and precipitate water, which belonged to basin brine fluid system; sulfur was from organic thermal chemical sulfate reduction and biological sulfate reduction; the metal mineralization material was from sedimentary strata and basement, but the difference of the material source of the basement and the strata and the superimposed mineralization of the west ore belt resulted in the difference of metallogenic elements between the eastern and western metallogenic belts. The Pb-Zn mineralization age of both ore belts was contemporary and formed in the same metaliogenetic event. Both thrust formed at the same time and occurred at the Early Oligocene, which is consistent with the age constrained by field geological relationship.

Study on experiment and mechanism of thermal dissolved sulfuration of low grade lead-zinc oxide ore in lanpin

The thermal dissolved sulfuration technology is brought forward and performed based on the characteristic of low grade lead-zinc oxide ore in lanpin. Using sulfur as the sulphidizing agent in the experiment, the oxides in the sandstone and ignimbrite are changed into sulfides. The disproportionation reaction of sulfur in a solution is confirmed as 4S＋3H2O=2S^2-＋S2O3^2--＋6H^＋. The dynamics process is studied and the first-order reaction rate equation -1n（1-a）=ktt is obtained. The effects of the reactive products, stirring speed, dosage of sulfuration agent, value of pH and sulphidizing temperature on the sulfuration of oxide ore are investigated. The results indicate that the reactive apparent activation energy is 100.8 kJ/mol and the sulfuration ratio of lead-zinc oxide ore reaches 60% under the conditions of pH 5.9-7.5, the sulfuration temperature of 130 ℃, sulfuration time of 180 min and the stirring speed of 800 r/min.

Recycle of Wastewater from Lead-Zinc Sulfide Ore Flo-tation Process by Ozone/BAC Technology

Lead-zinc sulphide ore contains lead sulphide (galena), and zinc sulphide (sphalerite). In the first flotation stage, galena is rendered hydrophobic with an organic collector such as xanthate, while sphalerite is kept from floating by depressants, and in the second flotation stage, activator was used to activated zinc flotation. Since the organic regent used are different in the two flotation stage, wastewater from the second zinc flotation stage can’t be directly recycled to the first lead flotation stage. Wastewater from flotation process for concentrating lead-zinc sulphide ore often containing organic compounds such as diethyldithiocarbamate(DDTC), xanthate, terpenic oil(2# oil) and thionocarbamate esters (Z-200), are environmentally hazardous. Their removal from contaminated water and the reuse of the water is one of the main challenges facing lead-zinc sulphide ore processing plants. In this study, synthetic wastewater containing DDTC, xanthate, 2# oil and Z-200 at concentrations ranging from 21 to 42 mg/L was fed into an Ozone/Biological activated carbon (BAC) reactor. Analyses of the effluent indicated a chemical oxygen demand (COD) removal over 86.21% and Total organic carbon (TOC) removal over 90.00% were achieved under Hydraulic retention time (HRT) of 4h and O3 feeding concentration of 33.3mg/L. The effluent was further recycled to the lab scale lead concentrating process and no significant difference was found in compare with fresh water. Furthermore, lead-zinc sulphide mineral concentrating process was carried out at lab scale. The produced wastewater was treated by Ozone/BAC reactor at O3 feeding concentration of 16.7mg/L and HRT of 4h. The effluent analysis showed that TOC removal was 74.58%. This effluent was recycled to the lab scale lead-zinc sulphide mineral concentrating process and the recovery of lead was not affected. The results showed that by using Ozone/BAC technology, the lead-zinc sulphide mineral processing wastewater could be recycled.