Ore Recovery Research Articles

Study on Comprehensive Recovery and Utilization of a Low-grade Gold-Copper Ore in Fujian

Study on Comprehensive Recovery and Utilization of a Low-grade Gold-Copper Ore in Fujian

Effects of Sulphuric Acid and Speed Rotation on Hydrolysis of Bamboo Sawdust

Bamboo saw dust (BSD) as reducing agent in low-grade manganese ore leaching is used, and therefore the performance of cellulose hydrolysis will affect the percent recovery of Mn ore. This work is studies the effect of sulphuric acid and speed rotation on hydrolysis of BSD. The crytallinity of BSD were investigated. The parameters considered in this work are sulphuric acid concentration and the speed rotation by using 23+3 factorial design. The significant factor, main and interaction effect were investigated based on Analysis of Variance obtained from statistical software. It shows that the sulphuric acid concentration is the main factors that affects the percentercent crystallinity compared to rotation speed. It is proved by the DOE analysis and supported by the crystallinity analysis of BSD before and after hydrolysis process. The optimum condition that can be suggested to obtain low percent crystallinity are when the sulphuric acid concentration and speed rotation at 4M and 400 rpm respectively. This indicates that reducing agent can be obtained from BSD in mild condition of hydrolysis due to decrement of cellulose crystallinity

Paste Pipeline Transportation of Pumping Backfill Technology with Long Distance and High Stowing Gradient in Cold and High-Altitude Areas

A paste pipeline transportation of pumping backfill technology with long distance and high stowing gradient is proposed to solve the problem of filling slurry transportation with low concentration, the filling body poor quality, and the transportation difficulties with long distance and high stowing gradient in Heiniudong copper mine (HCM). The physical and chemical properties of the backfill material, backfill proportion test, circular pipe experiment, and backfill system analysis evaluation were studied in the laboratory and outdoor, and the application in HCM was carried out to evaluate the technology. The research results show the feasibility of considering classified tailings and binder as backfill aggregates, and the optimum proportion of cement-binder-classified tailings applied in the stope and goaf is 1 : 4 : 8 and 1 : 4 : 15, respectively, with paste rheological properties of mass fractions of both being 74%∼76% and the backfill strength of about 1.5 MPa at 28 d. Furthermore, when backfill proportions and rate of flow are 1 : 4 : 8 and 50 m3/h, the pressure loss of the pipeline is around 0.4 MPa/100 m, and the backfill pump meets the backfill requirements. On this condition, the technology is capable of obvious economic benefits with the backfill cost of only 25.56 yuan/t, remnant ore recovery rate of 80%, and new output value of 1.28 billion. It creates a precedent for the paste pumping backfill technology with long distance and high stowing gradient in cold and high-altitude areas. The technology also provides reference mining experience for similar mines.

Orogenic gold ores in three-dimensions: A case study of distinct mineralization styles at the world-class Cuiabá deposit, Brazil, using high-resolution X-ray computed tomography on gold particles

Ore studies using conventional petrographic methods have an intrinsic limitation that spatial (three-dimensional, 3-D) distribution, orientation, shape and mineral associations have to be inferred by 2-D observations and measurements. Further, developing statistically significant data sets is particularly difficult for gold ores, given the fine grain size and low gold abundance even in high-grade ores. We use high-resolution X-ray computed tomography (HRXCT) and improved processing methods to provide 3-D data of ore textures and in-situ gold particles to discern mineralization styles and improve ore recovery. This is a case study of the Archean, world-class Cuiabá orogenic gold deposit in the Quadrilátero Ferrífero province, the largest underground mine in Brazil.Selected samples represent distinct mineralization styles at the Cuiabá mine: BIF-hosted ore from the Fonte Grande Sul (FGS) orebody, and gold-bearing quartz-carbonate veins hosted in a metamafic unit from the Veio de Quartzo (VQZ) orebody. This study characterizes shape, size, orientation and mineral associations of over 16,000 gold particles. Segmentation and quantitative measurements of gold particles are particularly challenging due to their small grain size. This issue is addressed using the recently formalized partial volume and blurring (PVB) method. Gold mineralization styles differ in 3-D textural and quantitative aspects including size and spatial distribution. Analysis of the VQZ sample shows that gold particles are mainly free in the quartz-carbonate matrix, with many particles in direct contact with a diverse sulfide mineral assemblage. Gold particles in the BIF-hosted FGS ore form bands in association with high pyrrhotite zones within iron carbonate-rich bands. PVB analysis show that the long axis of measured particles in both ore types ranges from ∼25 μm to 1.4 mm, with median values around 115 μm. Larger particles dominantly are sub-equant to prolate spheroids, with smaller particles appearing to be dominantly sub-equant to equant grains. The HRXCT data reveal that the kinematic processes affect gold shape and orientation and indicate the influence of deformation at the grain scale.

Kajian Perolehan Hasil Bijih Timah Berdasarkan Ukuran Butir Terhadap Variabel Magnetic Separator Skala Laboratorium

Industrial-scale Magnetic Separator is used to process tin ore by separating magnetic and nonmagnetic minerals using a belt conveyor speed and splitter opening only one variation. Based on this, this research is intended to obtain a more varied and optimal combination of variable settings in the recovery of tin ore yields. The method used is quantitative data analysis method by obtaining nonmagnetic minerals using a combination of splitter openings and belt conveyor speed based on grain size, calculating the content of tin ore in non-magnetic minerals using the Grain Counting Analysis method and calculating the composition of tin ore recovery. The research was conducted based on grain sizes of 60 mesh, 80 mesh, and 120 mesh using a combination of splitter opening sizes of 0.2 cm, 0.4 cm, and 0.6 cm and using a belt conveyor speed of 27.12 cm/s, 28.15 cm/s, and 28.40 cm/s. The sample used for each grain size is 9,000 grams with repetition for each combination of variables, namely 30 times. The results of this study obtained the highest yield of tin ore at grain size of 80 mesh as much as 1,753.7 grams in 8,920.3 grams of non-magnetic minerals with tin ore grades of 19.66% using a combination of speed 27.12 cm/s and splitter opening 0.2 cm. Based on the results of the study, the tin ore contained in the non-magnetic material processed by the Magnetic Separator, the grain size of 80 mesh is higher than that contained in the grain size of 60 mesh, while the grain size is finer than 80 mesh, the lower the recovery of tin ore contained in the non-magnetic material. Meanwhile, the larger the splitter opening and the speed of the conveyor belt, the higher the nonmagnetic minerals obtained.

Preliminary Study on the Reductive Leaching of Manganese Ores using FeSO?/Iron Powder as Reducing Agents under Room Temperature

The leaching of manganese (Mn) ore in sulphuric acid (H2SO4) under reductive conditions has been studied. The effects of leaching parameters such as ore/reductant mass ratio, acid concentration, ore particle size, solid/liquid ratio, leaching time and different reductant potential on the maximum recovery of manganese have been investigated. The optimal leaching conditions were ore/reductant mass ratio of 1:3.4, acid concentration of 10% v/v H2SO4, ore particle size of 63-200 µm, particle size of iron powder of –150 µm, solid/liquid ratio of 1:20, and leaching time of 1.5 hours at room temperature. A comparative analysis on the recovery of manganese ore was also investigated under the optimal leaching conditions for two different reductants, iron sulphate (FeSO4) and iron powder. The maximum manganese recoveries at the optimal leaching conditions in the presence of FeSO4 and iron powder are 80.6% and 95%, respectively. The results indicate that manganese can readily be leached during short time at room temperature. Furthermore, iron powder has a stronger reducing power compared to FeSO4 with respect to manganese ore recovery.

Preliminary Study on the Reductive Leaching of Manganese Ores using FeSO₄/Iron Powder as Reducing Agents under Room Temperature

The leaching of manganese (Mn) ore in sulphuric acid (H2SO4) under reductive conditions has been studied. The effects of leaching parameters such as ore/reductant mass ratio, acid concentration, ore particle size, solid/liquid ratio, leaching time and different reductant potential on the maximum recovery of manganese have been investigated. The optimal leaching conditions were ore/reductant mass ratio of 1:3.4, acid concentration of 10% v/v H2SO4, ore particle size of 63-200 µm, particle size of iron powder of –150 µm, solid/liquid ratio of 1:20, and leaching time of 1.5 hours at room temperature. A comparative analysis on the recovery of manganese ore was also investigated under the optimal leaching conditions for two different reductants, iron sulphate (FeSO4) and iron powder. The maximum manganese recoveries at the optimal leaching conditions in the presence of FeSO4 and iron powder are 80.6% and 95%, respectively. The results indicate that manganese can readily be leached during short time at room temperature. Furthermore, iron powder has a stronger reducing power compared to FeSO4 with respect to manganese ore recovery.

Optimization of Process Parameters for Enhanced Up-gradation of Qilla Saifullah Copper ore through Froth Floatation Technique

In this work, we focus on the up-gradation of the copper ore of Qilla Saifullah in Pakistan through the froth flotation technique. The chemical analysis of the head copper ore sample reveal the presence of 2.85% Cu, 22% Fe2O3, 52.9% SiO2, and other minor minerals. The optimum grinding time and liberation size of the copper ore have been determined as 30 minutes and +149-105 µm, respectively, for further processing. The chemical reagents are optimized in order to get a maximum grade and recovery of the copper ore. After comparisons and analysis of the results obtained, it can be concluded that the maximum grade and recovery of the copper ore are achieved at the dosage 300 (g/t) of the collector potassium amyl xanthate (C6H11KOS2), 250 g/t of pine oil, 250 g/t of a depressant (Na2SiO3), conditioning time of 10 minutes for a collector, flotation time of 6 and 10 minutes, and pH of 10 using the froth flotation technique.

Pillar recovery using new wireless blasting technology: a case study in Vazante mine, Brazil

The application of explosives has never been considered as a major influencing factor during either the designing of a mine or the selection of the mining method. However, this has changed with the launch of WebGenTM wireless blasting system, which allowed underground mines to exploit their orebody applying methods previously not possible due to limitations imposed by the use of wired detonators. The wireless blasting system is based on magnetic induction communication, and its signal is capable of overcoming hundreds of meters through rock, water, and air, to reach individual primers in the blastholes without any physical connection.A noble application of wireless detonators is being used in Vazante mine, Brazil, an underground zinc mine where ore pillars are left in the mined stope to secure stability and minimize dilution by limiting the hydraulic radius. The recovery of these pillars is financially desirable but involves extra time and costs associated with scaling, backfilling, reinstalling infrastructure, accessing areas previously blasted (less stable), drilling, charging with explosives, and subsequently, firing and mucking out the blasted material with expected high dilution. Applying 100% wireless detonators made it possible to safely preload the pillar together with production blasts before losing access to the area, a method named Temporary Rib Pillar (TRP). After all the stope is mined and the pillar accomplished its objectives, the primers are initiated without neither the need for the extra cycles described previously, nor the need to re-enter the area. Thus, it was possible to reduce the exposure of people and equipment, reduce operational cycles, and increase ore recovery, directly contributing to anticipate the ore production while guaranteeing the safety of the teams involved.

Optimum Open Pit Design for Kenticha Tantalite Mine, Southern Ethiopia

The mining sector’s share in Ethiopia’s economy is gradually increasing. Among metallic mines, Legadembi for gold and Kenticha for tantalum are the main contributors. At Kenticha, the Ethiopian Mineral Development Share Company is producing tantalite concentrate of 40-60% grade by open pit mining. The mine area is comprised of the rocks of Neoproterozoic age, pegmatite, granite, serpentinite, and talc-chlorite schist. The ore-bearing pegmatite intruding the basement serpentinite, and talc schist rocks, is asymmetric, N-S trending, and locally affected by fractures and local faults. Mining, at present, is being done using conventional methods not by developing benches. An optimum open pit design is developed and its impact on the run of mine is evaluated based on the field data, technical mine report data, borehole data, tonnage data, grade data, and mine survey. The suggested design has taken into account the mine drainage, bench design, haul road design, and mine waste dump design. Also keeping in view the future mining and production requirements. The suggested open pit mine design can minimize the dilution and improve the ore recovery.

Mineralogical Prediction on the Flotation Behavior of Copper and Molybdenum Minerals from Blended Cu–Mo Ores in Seawater

The copper ore in Chilean copper porphyry deposits is often associated with molybdenum minerals. This copper–molybdenum (Cu–Mo) sulfide ore is generally mined from various locations in the mining site; thus, the mineral composition, oxidation degree, mineral particle size, and grade vary. Therefore, in the mining operation, it is common to blend the ores mined from various spots and then process them using flotation. In this study, the floatability of five types of Cu–Mo ores and the blending of these ores in seawater was investigated. The oxidation degree of these Cu–Mo ores was evaluated, and the correlation between flotation recovery and oxidation degree is presented. Furthermore, the flotation kinetics of each Cu–Mo ore were calculated based on a mineralogical analysis using mineral liberation analysis (MLA). A mineralogical prediction model was proposed to estimate the flotation behavior of blended Cu–Mo ore as a function of the flotation behavior of each Cu–Mo ore. The flotation results show that the recovery of copper and molybdenum decreased with the increasing copper oxidization degree. In addition, the recovery of blended ore can be predicted via the flotation rate equation, using the maximum recovery (Rmax) and flotation rate coefficient (k) determined from the flotation rate analysis of each ore before blending. It was found that Rmax and k of the respective minerals slightly decreased with increasing the degree of copper oxidation. Moreover, Rmax varied greatly depending on the mineral species. The total copper and molybdenum recovery were strongly affected by the degree of copper oxidation as the mineral fraction in the ore varied greatly depending upon the degree of oxidation.

Laboratory study on three-dimensional characteristics of gravity flow during longitudinal sublevel caving

Ore recovery and dilution in caving mines are strongly determined by the flow characteristics of broken rock, especially under inclined boundary conditions. A physical draw model was designed and used to investigate the gravity flow characterization of granular materials during longitudinal sublevel caving. The three-dimensional (3D) shapes of flow axis, isolated extraction zone (IEZ) and ore remnants under inclined walls were obtained using the discharged quantity method and 3Dmine software, which indicated that stope boundary conditions and orebody dip have an impact on the characteristics of gravity flow. The experimental results also showed that the gravity flow characteristics of 3D draw experiments under the inclined walls were significantly different from those of the two-dimensional (2D) draw experiments. The flow axis is a line segment and lies in the same plane and only the width of the IEZ can be displayed in 2D experiment. However, the flow axis is a space curve segment in 3D draw experiment. Since being cut off by the front and inclined upper walls simultaneously, the 3D IEZ is more complicated than the 2D IEZ and both the width and thickness of the draw body are displayed simultaneously in 3D state. The shape and location of ore remnants in 3D state are also different from that in 2D state. Therefore, the 3D IEZ is more consistent with the actual situation than 2D IEZ for sublevel caving stope. The 3D gravity flow presented may be used to determine the design of the extraction layout and improve ore recovery during longitudinal sublevel caving.

World mineral loss and possibility to increase ore recovery ratio in mining production

ABSTRACT Based on the data of global mineral production and ore recovery in mining and mineral processing, this paper presents the annual production of 25 minerals and calculates the annual losses of 20 minerals from the year 1920 to 2018. The results indicate that the annual production of 23 minerals has increased by between 1 and 930 times since 1920, similar to the increase in the GDP per capita since 1960. At the same time, a vast amount of minerals has been lost in mining and mineral processing since 1920. Hence, this paper discusses the possibility of increasing ore recovery ratio.

Investigation of Backfilling Step Effects on Stope Stability

Cemented rock fill (CRF) is commonly used in cut-and-fill stoping operations in underground mining. This allows for the maximum recovery of ore. Backfilling can improve stope stability in underground workings and then improve ground stability of the whole mine site. However, backfilling step scenarios vary from site to site. This paper presents the investigation of five different backfilling step scenarios and their impacts on the stability of stopes at four different mining levels. A comprehensive comparison of displacements, major principal stress, and Stress Concentration Factor (SCF) was conducted. The results show that different backfilling step scenarios have little influence on the final displacement for displacement in the stopes. Among the five backfilling scenarios, the major principal stress and stress concentration factor (SCF) have almost the same final results. The backfilling scenario SCN-1 is the optimum option among these five backfilling scenarios. It can immediately prevent the increase of the displacement and reduce the sidewall stress concentration, thereby preventing possible failures. Using the same strength of CRF can achieve the same effects among the four mining levels. Applying backfilling CRF of the same strength at different mining depths is acceptable and feasible to improve the stability of the stopes.

Behaviors and Overlying Strata Failure Law for Underground Filling of a Gently Inclined Medium-Thick Phosphate Deposit

In this study, the No. 6 pit in the eastern mining area of the Jinning phosphate mine in China was taken as the research background. In order to reduce the cost of filling, an improved pillarless sublevel caving method is proposed. This method greatly improves the ore recovery rate by adding a recovery route. In addition, the combination of similar material simulation experiments and numerical simulations (discrete element and universal distinct element code) revealed the deformation and failure laws of the surrounding roof rock and the characteristics of the surface subsidence. The results indicate the following. (1) The similar simulation experimental results indicate that the deformation of the overlying rock layer originated from the direct roof of the goaf and gradually developed into the deep part of the rock layer. An irregular stepped caving zone formed in the goaf. The maximum surface subsidence was located above the phosphorus orebody, and it gradually decreased toward both sides. As the stope approached propulsion, the location of the maximum subsidence gradually moved toward the propulsion direction. (2) The numerical results revealed that the displacement of the overlying strata was nonlinear, and it decreased with increasing roof height. A support pressure concentration area was formed within a certain range of the stope roof. The numerical simulation results are basically consistent with the similar simulation experimental results.

A Stochastic Model Updating Gold Reserve Estimation by Using Monte Carlo Simulation

This paper elucidates a new idea and concept for exploration of the gold ore deposits. The cyanidation method is traditionally used for gold extraction. However, this method is laborious, time-consuming, costly, and depends upon the availability of the processing units. In this work, an attempt is made in order to update the gold exploration method by the Monte Carlo-based simulation. An excellent approach always requires a high quality of the datasets for a good model. A total of 48 incomplete datasets are collected from the Shoghore district, Chitral area of Khyber, Pakhtunkhwa, Pakistan. The cyanidation leaching test is carried out in order to measure the percentage of the gold ore deposits. In this work, the mean, median, mode, and successive iteration substitute methods are employed in such a way that they can compute the datasets with missing attributes. The multiple regression analysis is used to find a correlation between the potential of hydrogen ion concentration (pH), solid content (in %), NaCN concentration (in ppm), leaching time (in Hr), particle size (in µm), and measured percentage of gold recovery (in %). Moreover, the normal Archimedes and exponential distributions are employed in order to forecast the uncertainty in the measured gold ore deposits. The performance of the model reveals that the Monte Carlo approach is more authentic for the probability estimation of gold ore recovery. The sensitivity analysis reveals that pH is the most influential parameter in the estimation of the gold ore deposits. This stochastic approach can be considered as a foundation to foretell the probabilistic exploration of the new gold deposits.

Numerical analysis of the stability of arched sill mats made of cemented backfill

Sill mats made of cemented backfill are important structures largely used in underground mines to provide a safer working space, increase ore recovery and improve ground stability conditions. When large span sill mats are exposed at the base, flexure can take place and tension appears at the base center of the sill mats. As cemented backfill and other geomaterials usually have a strength smaller in tension than in compression, a common practice in field is to add reinforcement structures built with bolts and metal grids. The construction of these structures is expensive and time-consuming. In this paper, the use of arched sill mats is for the first time proposed as a new solution to reduce the occurrence of tension and increase the stability. Binder consumption and metal reinforcement can thus be reduced, resulting more economic design of cemented backfill sill mats. A series of numerical simulations have been performed to analyze the efficiency of arched sill mats in reducing the minimum required cohesion. For the stope geometry considered in this study, the numerical results show that arched sill mats exhibit an improved stability compared to traditional rectangular sill mats. The minimum required cohesion of sill mats increases as the sill mat span increases but decreases as the curvature of the sill mats increases. The overlying uncemented backfill has effect to decrease the stability of sill mats. The minimum required cohesion of sill mats increases with the thickness of the uncemented backfill above the sill mats. These results remain valid as long as the stopes are wide enough, and the prevailing failure mode of sill mats is tension by flexure. More work is thus necessary to establish a criterion for the application of this new concept to increase the stability of cemented backfill sill mats against the failure of tension by flexure.

Differences in Etheramines Froth Properties and the Effects on Iron Ore Flotation. Part II: Three-Phase Systems

ABSTRACT Part I of this investigation, addressing two-phase systems, showed that etherdiamine produces more stable and dryer foams, whereas smaller bubbles are formed with faster kinetics by ethermonoamine. These remarkable differences in foam stability, attributed to the distinct chain length and consequent hydrophilic–lipophilic balance, influence froth properties, and may affect the selectivity of reverse cationic flotation of iron ores. The investigations conducted in this paper evaluated the effect of froth properties on the selective flotation recovery of a Brazilian iron ore from the Iron Quadrangle. The results showed that etherdiamine presented faster kinetics of froth formation but the recovery of water and solids in the froth was lower than that achieved with ethermonoamine. The water recovery was closely related to nonselective transport of solid particles in the froth by entrainment. The etherdiamine ability to form more stable and dryer froth reduced the entrainment of fine iron-bearing particles in the froth (tailings), improving the metal recovery in the concentrate (bulk), as the iron loss was directly proportional to the water recovery in the froth and the reagent presented lower entrainment factor than the ethermonoamine. In addition, the drier froth from etherdiamine is responsible for greater quartz true flotation, rendering this collector a more selective effect on flotation in spite the more powerful effect of ethermonoamine.

A Method for Determining the Thicknesses of the Reserved Protective Layers in Complex Goafs Using a Joint-Modeling Method of GTS and Flac3D

In this study, a C-ALS underground cavity scanner was used to detect the shapes of mining goafs. In addition, GTS software was adopted to establish a three-dimensional geological model based on the status of the stopes, geological data, and mechanical parameters of each rock mass and to analyze the roof areas of the goafs. In regard to the morphology of the study area, based on a thin plate theory and the obtained field sampling data, a formula was established for the thicknesses of the reserved protective layers in the goafs. In addition, a formula for the thicknesses of the protective layers in the curved gobs was obtained. The thickness formula of the protective layers was then successfully verified. The detection results showed that the roof shapes of the goafs in the Yuanjiacun Iron Mine were mainly arc-shaped, and the spans of the goafs were generally less than 20 m. The stability of the arc-shaped roofs was found to be greater than that of the plate-shaped roofs. Therefore, by reducing the thicknesses of the protective layers in mining goafs, the ore recovery rates can be increased on the basis of safe production conditions. The formula of the thickness of the security layers obtained through the thin plate theory was revised based on the statistical results of the roof shapes of the goafs and then combined using GTS and FLAC3D. The modeling method successfully verified the stability of the mined-out areas. It was found that the verification results were good, and the revised formula was able to improve the recovery rate of the ore under the conditions of meeting safe production standards. Also, it was found that the revised formula could be used in the present situation. At the same time, it was also determined that the complexity of the rock masses obstructed the full identification of the joints and fissures in the present orebodies. Therefore, it is necessary to incorporate C-ALS underground cavity scanners to regularly observe the shapes of the goafs in order to ensure that stability and safety standards are maintained.

Comparison of Biomass and Coal in the Recovery Process of silicon in an Electric Arc Furnace

Silicon recovery of silica ore ( $$ {\text{SiO}}_{2} $$ ) has been studied with two types of carbon materials, charcoal as a biomass and coal as a fossil fuel, at elevated temperatures between 1800 and 2000°C in an electric arc furnace. The effects of porosity and electrical resistance of the carbon materials were investigated. To this end, recovery of silicon and ferrosilicon production were tested separately by charcoal and coal, and the products were investigated both qualitatively and quantitatively. A higher electrical resistance of charcoal was found in comparison with coal, with increased efficiency of the furnace and decreased electric energy consumption (per ton of product). The efficiency of the furnace using charcoal and coal was 92.13% and 77.4%, respectively. In addition, the higher porosity of charcoal facilitates the flow of SiO gas through the carbon material leading to a higher reactivity and reducing the electric energy consumption for each ton of FeSi.