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By A. R. Javadi

Pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite (ZnS) and galena (PbS) generate hydrogen peroxide (H2O2) when placed in water, with pyrite generating the most amount of H2O2. The order of these minerals for generating H2O2 in water is: pyrite > chalcopyrite > sphalerite > galena. In previous research, the methods of generation of H2O2 in the grinding mill were studied. In this study, data from experiments were included by using MODDE 9 software for six parameters composed of water, grinding media, particle size; adding collector to the grinding, and adding depressant and pH. Finally the optimal condition was achieved. According to the results, most of the recovery is related to the particle 75 μ size and when using processed water, adding 100% collector and without depressant and pH 10. The most recovery of lead in the case of the use of fresh water is in the case of 100% collector, without depressant, pH 10 and 150 μ particle size. Maximum recovery in fraction size 45 μ in the case of water process is 100% collector, without depressant and pH 10. Maximum depressant of pyrite will be the fraction size 75 μ, distilled water, without collector, 100 depressant. Keywords: Sulfide mineral, hydrogen peroxide, flotation, oxidation

Jan 1, 2020

By E. Pothas, P. R. Jochens, L. B. McRae, D. D. Howat

Discussion H. J. S. Kriek*: During 1956 Iscor became interested in the acid smelting in a blast furnace of siliceous ores containing 15 per cent silica. This interest was as a result of the successful acid smelting in Germany at a CaO/SiO2-ratio of between 0.4 and 0.8. Acid smelting can become an economic proposition due to a reduction in slag volume accompanied by a lowering of the coke rate. Paschke and Hannel1 found that a free flowing slag (i.e. viscosities of less than two poise) would not be obtained at C/S-ratios lower than 0.82 and that the addition of 3 to 4 per cent alkalies became necessary. In titanium bearing slags it was established that maximum fluidity is obtained with 10 per cent TiO2. Acid smelting resulted in increased sulphur pickup with decreasing O/S-ratio and desulphurization with soda-ash outside the blast furnace was necessary. At Iscor Mr B. B. Segal established that, for slags with (CaO+MgO)/SiO2-ratios of 0.5, 0.6 and 0.7, the addition of alkalies and TiO2 resulted in a substantial decrease in melting point and also of viscosity. At a (CaO+ MgO)/ ratio of 0.53 the addition of 10 per cent TiO2 is as effective as the addition of 3.75 per cent alkalies. This means that a blast furnace can be run without the addition of alkalies and consequently the detrimental effect of alkalies on the lining can be avoided; also that fluid slags with 5 to 15 per cent MgO and 8 to 10 per cent Al203 can be obtained with 10 per cent TiO present. (CaO+ MgO)/SiO2-ratios smaller than 0.5 the addition of titania resulted in a decrease in melting point and also in an increase in viscosity. In ceramic crucibles the slag did not become more viscous with time but thickened appreciably when held in graphite crucibles. No titanium carbonitride could be detected and only Ti20a was found. Prof Howat also did not obtain the TiC-TiN solid solution and therefore it appears that the thickening of the slag is not solely due to the formation of the titanium carbonitrides. This is in agreement with the work of Michailov and Belyakova2 who found that TiO2 decreased the viscosity of acid slags and increased the viscosity of basic slags; that titanium sesquioxide (Ti203) and titanium monoxide (TiO) increased the viscosity of both acid and basic slags and that titanium carbide, which is found in larger amounts in basic slags than in acid slags, considerably increases the viscosity of both acid and basic slags. Patzak et al3 describes the following oxygen deficient titanias, viz Ti2O, TiO, Ti203, Ti3O5, Ti4O7, Ti5O9, Ti6O11- Ti10O9,Ti7O13,Ti8O15 and Ti9O17. Until about 1959 a fair amount of data was available on the viscosity of normal slags to which titania was added2, 4, 5, 6, 7, 8, 9. At Iscor an electromagnetic viscometer based on the viscometer developed by Bockris10 was built. Progress was very slow as our main attention was centred around problems associated with the operation of the Rotor process. During 1962 and 1963 two papers containing fundamental work on the viscosities of slags in the systems CaO-SiO2-TiO211 and CaO-SiO2-TiO2 at 0 per cent A1203, 10 per cent Al203 and 20 per cent Al20312 appeared. The work of Ross and Ohno obviated the necessity of any further work by Iscor on the viscosity of blast furnace slags containing titania. Ross gave the following optimum compositions for blast furnace slags containing titania. Titania % Lime % Silica % Lime/Silica-ratio 20 30-42 38-50 0.6-1.10 30 27-36 34-43 0.63-1.06 40 24-32 28-36 0.66-1.14 The optimum temperature is 1,500°C because at this temperature the largest workable composition range exists. Higher temperatures result in narrower optimum composition ranges and in increased rates of thickening. The factors affecting the rate of thickening were also investigated by Ross and Ohno, as the main problem with slags containing titania is not finding compositions which are fluid at operating temperatures but the thickening due to a lack of oxygen. With these slags when there is insufficient oxygen to co-ordinate each silicon atom with its own individual shell of four oxygen ions, some sharing of oxygen ions must occur, resulting in the formation of large silicate ions and a rise in viscosity. The formation of oxygen deficient titania can be prevented as long as iron oxide or manganese oxide remain in the liquid slag. Iwase13 and also Viens, Camp bell and Rogers14 obtained in the experimental smelting of ilmenite ores fluid slags with more than 3 to 5 per cent manganese oxide and ferrous oxide in the slags. It appears that titaniferous ores can be smelted in blast furnaces under the following conditions: (1) Acid slag operation, i.e. at C/S-ratios from O.6-1.0. (2) Dolomite can be used as a flux as MgO is a network modifier. (3) The Al2O3-content must be less than 15 per cent as alumina promotes thickening of the slag and decreases the optimum composition range. (4) The hearth temperature must not exceed 1,500°C. According to Ross a definite amount of energy is required for each ton of iron produced and in order to increase production an increase in heat

Jan 6, 1969

By V. G. Pienaar, B. M. Mbiya, P. T. Slatter

[a constant b constant cc onstant c' constant D pipe diameter (m) Dshear shear diameter (m) E error function fth theoretical friction factor F function F' function K fluid consistency index (Pa sn) kv valve loss coefficient L pipe length (m) n flow behaviour index N number of data points ?P valve pressure drop (Pa) Q flow rate (m3/s) Re Reynolds number Re3 Reynolds number (Slatter) V average velocity (m/s) Vann annulus velocity (m/s) x variable X number of independent variables y variable Y number of fundamental dimensions z variable µeff effective viscosity (Pa s) ? valve degree of opening d partial derivative ? fluid density (kg/m3) [ ] function]

Jan 1, 2007

By D. A. Cowen

Toll and custom treatment of base metal concentrates have been performed for decades as mines and concentrators are developed distant from capital and, energy-intensive smelting and refining facilities. Commercial treatment terms have been developed and refined for over a century, and are negotiated taking into account a complex array of quality and economic factors. Treatment of third-party concentrates in the South African platinum industry is a comparatively recent development, brought about by the opening up of mineral rights and operations being developed that are smaller than the critical size required to sustain capitalization of smelting or refining plant. Published commercial terms have not been demonstrated to reflect the complexities and subtleties of base metal concentrate markets. This paper outlines some key elements and parameter ranges of base metal concentrate treatment contracts, and applies similar principles to typical Merensky and UG2 platinum group element (PGE) concentrates to determine potential value distribution between the mining and treatment parties in tolling arrangements.

Jan 1, 2012

This paper examines the new water-hydraulic (hydropowered) drill rigs that have been developed and proven in the last decade primarily in tabular, shallow-dipping, narrow-reef hard-rock South African mines. These include drill rigs for flat-end tunnel development, narrow inclined raises and winzes, large inclines and declines and longhole-based developments such as ore passes and vent raises. The hydropower technology rigs offer advantages both in safety and performance over traditional compressed-air-powered, hand-held development, and are generally more cost-effective than imported oil electro-hydraulic drill rigs. Furthermore, hydropowered rigs are inherently energy-efficient. The benefits of hydropower are discussed and a ?comprehensive development model? is presented which demonstrates that ore-bodies can be accessed in significantly shorter times. ?Localized? half level hydropower systems are introduced and a range of ancillary equipment available is listed.

Jan 1, 2012

By J. Guerrero, O. Pavez, N. Toro, A. Nazer

In the 19th century, mining in northern Chile was of great economic importance. Today, the ruinsof the copper smelters and abundant fragments of refractory bricks from their furnaces remain. Therefractory materials have not been previously studied. the origin, physical and chemical properties, and compressive strength of 10 different samples of historical refractory bricks collected on-site were investigated. The characterization results showed that the bricks are silico-aluminous, unlike the magnesium-chromium bricks used by current foundries. It was determined that 90% of them were produced in the UK and 10% were manufactured in Chile. The phases present are mullite, quartz, cristobalite, and tridymite, and the chemical composition corresponds mainly to SiO2. The compressive strength of the historical bricks is similart to those currently used,. These results may contribute to the knowledge of the 19th century mining and metallurgical historical heritage of Chile and the UK.

Jul 5, 2023

By M. Paricheh, M. Rahmanpour

"The cost efficiency and high reliability of semi-mobile or combined in-pit crushing-conveying (IPCC) and truck systems, compared to conventional truck-shovel systems alone, make them more attractive for use in modern mining operations. Semi-mobile systems, which are the most common systems in open-pit mining, combine the advantages of both systems such that fewer trucks are required, operating costs are lowered, and the operation is environmentally sound. One major aspect of utilizing this system is determining the location of the in-pit crusher and the timing of relocations. Facility location models are rarely used in the mining sector, but these models may be used for solving the in-pit crusher location problem. To this end, the main parameters affecting the IPCC location in open pit mines are reviewed and examples given of how they are applied to a dynamic location-relocation problem. Subsequently, the model is implemented for Sungun copper mine in Iran and the number and the exact time of relocations of the IPCC units determined. IntroductionNowadays, compared to the last century, metal prices are more volatile, average and cut-off grades have decreased, and stripping ratios increased. From a mining engineering point of view, many ore deposits with simple access and low capital and operating costs have been exhausted. The remaining deposits are located at depth, with no easy access, which requires high initial capital costs (Osanloo, 2012). Mining at great depths and high tonnages is possible only through open pit mining. Today, more than 80% of open pit mines in the world use a shovel-truck system for loading-hauling operations (Osanloo, 2010).With increasing depth, haulage distances increase and the number of loads per truck decreases. As a result, fuel, tyre, and depreciation costs per ton increase. Owing to these factors, open pit mining at great depth (300 to 1000 m) using shovel-truck systems faces some technical and economic problems. In-pit crushing-conveying (IPCC) systems have been known in the mining industry for many decades. The idea was introduced in 1956 in Germany. Today, the cost efficiency and high reliability of IPCC make it more appealing than conventional shovel-truck operation, especially in longer life projects with lengthy transportation distances and high production rates (Koehler, 2003). Since 1956, different aspects of this system have been reviewed by many researchers. Barua and Lanergan (1985) developed a computer program that compares the tonnage aspects of various conveyor layouts. They believed that to maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must first minimize the cost of the flatter slope, which is expressed in terms of additional stripping or tied-up ore. Dos Santos and Stanisic (1987) evaluated the design of a high-angle conveyor at the Majdanpek copper mine. Then, Sturgul (1987) applied GPSS (general purpose simulation system) to find the best location of an in-pit crusher."

Jan 1, 2017

By I. Silin, H. Wotruba, A. Liono, I. Volobayev, Y. Rubinstein, D. Gürsel

Old, abandoned or in situ processed and storaged tailings can be a potential source of remained valuable metals and minerals. By reprocessing of tailings, it is possible not only to achieve economical profit from recovery of valuable metals, but also to reduce environmental damage caused by tailings such as generation of acid mine drainage from sulfide minerals, mobilization of heavy metals, soils and groundwater contamination. In addition, storage facilities can be successfully redesigned with modern geotechnical requirements and monitoring to avoid a dam collapse of abandoned tailings. In this case study, the processing of low grade abandoned tailings was investigated. Copper (0.15%), cobalt (0.032%) and gold (0.35 ppm) were determined as main valuable and recoverable metals. Cobalt occurs as cobaltite and cobaltiferrous pyrite, while chalcopyrite represents copper. The fine fraction of tailings below 32 μm is enriched with cobalt and copper. Preliminary test-works with initial tailings sample were aimed to recover copper, cobalt and gold. The previously tested magnetic and gravity separation could not recover the metals in a pre-concentration step. Flotation tests of the tailings sample had shown a suitability for flotation without previous regrinding of the tailings. The valuable minerals are first proposed to be recovered by collective flotation, followed by a selective circuit with production of a low grade copper concentrate and a cobalt bearing pyrite concentrate. Based on various tested collectors, amounts of sulfidisation agents and activators, a suitable reagents regime was determined and used for up-scaling by locked cycled test. Keywords: reprocessing, tailings, copper, cobalt, gold, flotation, cobaltite, cobaltiferrous pyrite

Jan 1, 2020

By G. Chisakuta, P. Bwando, G. Mitshabu

By the end of 2022, the south-eastern Copper Belt region of the Democratic Republic of Congo was host to more than 40 small, medium, and relatively large sized copper solvent extraction sites. To date, solvent extraction reagents supplied in this region have never exceeded 50% of ketoxime content. We generated laboratory extraction and strip isotherms and used these with the BASF IsocalcTM modelling software to assess the benefit of using higher ketoxime blends on the main metallurgical performance indicators such as copper recovery and net transfer. Pure ketoxime extractants combine a relatively low strip point and a decent loading capacity, which are most effective as the pH of the pregnant leach solution gets closer to 2.

Jan 1, 2023

By K. J. Craig, M. A. Swart, Q. G. Reynolds, L. Hockaday

A solar reflector soiling study was carried out at a ferromanganese smelter in South Africa to assess the soiling rates at different locations around the plant. Several meteorological parameters were monitored to give insight into the conditions that lead to increased soiling. Mineralogical characterization of dust samples collected from the reflectors and the atmosphere revealed that only a certain size fraction is of importance with regard to soiling, and that the dust can be attributed to both raw materials and smelter products. Proximity to the dust source was the primary driver for increased soiling. The site that experienced the most soiling was very close to raw material heaps; this was deemed an outlier and was excluded from the summary statistics. The secondary driver for increased soiling was location relative to the smelter dust sources and the wind’s direction and speed. The reflector set at the best location experienced 13.1% less soiling than the set at the ‘worst’ (but still feasible) location, represented by an averaged mean daily reflectance loss of 0.0186. The study revealed that while there are periods of intense soiling at this particular site, proper planning of reflector location in relation to the smelter dust sources can significantly mitigate the soiling rate.

Jun 5, 2023

By N. F. Dawson

Recovery of a chromite concentrate from UG2 flotation tailings streams, of a grade suitable for further pyrometallurgical treatment for the production of ferrochrome, has certain specific challenges not only by virtue of the decoupled nature of the primary PGM recovery process and secondary gravity based chromite recovery circuits, but also because of intrinsic variations in the nature of the ore and its mineral composition. Variability in the ore, together with upstream process dynamics, creates a challenging environment in which to achieve optimal recovery of a gravity concentrate with accurate grade control. This paper examines certain inter-relational aspects between the ore gangue and chromite mineral types, the influence of the primary PGM extraction circuit and secondary chromite extraction activity as applied specifically to UG2 ore against a background of actual plant experiences and circuit configurations applied in order achieve a suitable quality of chromite concentrate.

Jan 1, 2010

By K. Karaman

This study investigated the effect of surface roughness on Schmidt rebound hardness (RL). Four different test surfaces of rock samples were studied: natural, ground, cut surfaces, and core samples. There was significant variability of standard deviation based on the RL on the natural surface, which indicated high roughness of the rock surface, whereas surface polishing caused a significant decrease in standard deviation. ISRM and ASTM methods were compared to estimate unconfined compressive strength (UCS) for different testing surfaces. RL obtained from the cut surface was found to be more reliable than those obtained from other testing surfaces for the prediction of UCS; however, regression and ANOVA analyses revealed that the ISRM method gave a more accurate UCS estimation of rocks with highly rough surfaces. It was also shown that RL values obtained from a cut surface were significantly higher than those obtained from core samples. Therefore, a comparison between RL values obtained from core samples and cut surfaces was made based on previous studies. This study statistically showed that estimated UCS values are not statistically significant if Schmidt rebound tests are not performed on similar surfaces. In addition, the mineral grain sizes of the studied rocks, different testing surfaces compared with those in literature, and standard deviation from RL are evaluated and discussed. The Schmidt hammer technique is a rapid, inexpensive, and straightforward method for estimating UCS for preliminary assessment; however, roughness of the surface should be eliminated if variations are shown in the surface rebound hardness.

Apr 14, 2024

By M. J. U. T. Van Wijngaarden

SYNOPSIS The corrosion characteristics of a new steel, 14Cr, which is a modification of the well-known 3CR12, in aqueous sulphate solution were studied through polarization diagrams and a computer analysis based on the polarization¬resistance technique. It was found that the corrosion rate of 14Cr steel is independent of pH for pH values above 6, but, below a pH value of 4, the corrosion rate increases sharply with decreasing pH. In air-saturated solutions of neutral and mildly alkaline pH, the corrosion rate is approximately 60 µm per year. The 14Cr steel shows a twentyfold improvement in corrosion resistance over that of 3CR12 in sulphate solutions at pH 2,0. This is considered to be due to the beneficial effect of its higher chromium and nickel contents.

Jan 1, 1985

By K. Woodward, Y. Potvin, B. McFadyen, M. Grenon

Open stoping has become a popular mining method in hard rock mines, not only due to the safety of the method as a non-entry approach, but also because of the high extraction rate and low costs. At mine sites, stope performance is evaluated by calculating stope overbreak using the stability chart. However, limitations of the stability chart regarding the precision of the predictions, nonconsideration of factors such as the influence of blasting, and the exclusion of underbreak have led to non-optimal designs. The capabilities of today's computers have increased the amount of data being collected and the power of models being built. This article presents a step towards a new stope design approach where stope overbreak and underbreak are measured and georeferenced using octrees at an approximately cubic metre resolution and predicted using multivariate statistical models (partial least square, linear discriminant analysis, and random forest). Results show that overbreak and underbreak location along the design surface and their magnitude are predicted with good precision using a random forest model. These predictions are used to build the expected geometry of the open stope. The resolution of the data and the use of multivariate analysis has enabled the prediction of variation in stope performance along the design surface, going well beyond the simple qualitative per stope face prediction provided by a traditional stability chart approach.

Jun 5, 2023

By S. Bradshaw, A. Mabentsela

"Physical and numerical modelling methods were followed to identify mechanism(s) for tundish slag entrainment in a bare tundish and one with a flow control device (FCD).The physical and numerical models made use of water and paraffin to model steel and slag respectively. Observations from the physical model showed that the steel-slag interface remains immobile in both cases. Entrained paraffin formed droplets approximately 1 mm in diameter. Results from both models (numerical and physical) showed that in both cases (bare and FCD case), areas of high entrained slag concentration exist near the inlet region. The entrained slag concentration decreases towards the tundish endwalls.Flow patterns and velocities tangential to the steel-slag interface from the numerical model showed that slag entrainment in both the bare tundish and tundish with a FCD possibly takes place via two mechanisms. First, the slag moves across the steel-slag interface via mass transfer; secondly small velocities tangential to the interface at depths greater than 10 mm below the interface carry the already ‘entrained’ slag into the bulk steel phase. These tangential flow patterns are dominant in the inlet region, hence the high concentration of entrained slag in this region. IntroductionThe tundish serves as the last metallurgical vessel through which steel passes before solidifying in the moulds. It is in the tundish that the last traces of non-metallic inclusions should be removed, otherwise the inclusions carry over to the moulds and cause defects in the steel product. In the tundish, the slag is less dense than the melt and thus resides on the top of the melt. The slag provides a sink into which the non-metallic inclusions float to and dissolve. The slag also protects the melt from air and heat loss. However, when slag is entrained in the melt as a result of increased turbulence or shearing at the steel-slag interface, slag can become a source of nonmetallic inclusions.Numerous studies on flow patterns in the tundish have been carried out since the introduction of the tundish. These studies were aimed at improving the flow characteristics in the tundish (Chattopadhyay, Isac, and Guthri, 2010; Cloete, 2014; Jha, Dash, and Kumar, 2001; Jha, Rao, and Dewan, 2008;"

Jan 1, 2017

By K. C. Sole, M. E. Mitshiabo

Although surfactants have proven to increase metal extraction by improving solution flow in porous media, these reagents have not been widely adopted in heap leaching operations because of potential incompatibilities with downstream processes that are subject to surface properties. The BASF leaching aid, LixTRA 118, was previously found to have no significant effect on post-leach operations. This study evaluated its performance in ores of different particle size distributions and the impact on acid consumption and metal recovery. Column leach studies, carried out on a copper oxide ore and mixed oxide–sulfide ore, showed that the effect of LixTRA 118 was more pronounced in ores with wider particle size distributions than in those with more uniformly sized particles. Solution percolation rates increased by up to 18% in the presence of leaching aid, and increased copper recovery and acid consumption. Improved wetting facilitated the kinetics of acid diffusion at macroand micropore level. There was no significant effect on iron recovery

Jan 1, 2023

By S. Swanepoel, A. V. Cherkaev, Q. G. Reynolds, M. Erwee

Ferrochrome (FeCr) is a vital ingredient in stainless steel production and is commonly produced by smelting chromite ores in submerged arc furnaces. Silicon (Si) is a componrnt of the FeCr alloy from the smelting process. Being both a contaminant and an indicator of the state of the process, its content needs to be kept within a narrow range. The complex chemistry of the smelting process and interactions between various factors make Si prediction by fundamental models infeasible. A data-driven approach offers an alternative by formulating the model based on historical data. This paper presents a systematic development of a data-driven model for predicting Si content. The model includes dimensionality reduction, regularized linear regression, and a boosting method to reduce the variability of the linear model residuals. It shows a good performance on testing data (R2 = 0.63). The most significant predictors, as determined by linear model analysis and permutation testing, are previous Si content, carbon and titanium in the alloy, calcium oxide in the slag, resistance between electrodes, and electrode slips. Further analysis using thermodynamic data and models, links these predictors to electrode control and slag chemistry. This analysis lays the foundation for implementing Si content control on a ferrochrome smelter.

Mar 14, 2024

By DeVerle P. Harris, David E. Euresty

"Potential metal supply refers to metals to be produced from as yet unknown deposits. Naturally, to measure potential supply requires, among other things, inference to the number of deposits, their tonnage, and their grade Thus, probability distributions were constructed for each 400-square mile cell in Northern Sonora, Mexico. A computer system was developed that simulated the occurrence of deposits for each cell, exploration for the deposits, and their exploitation. Evaluation was performed for each deposit by discounted cash flow analysis.An objective of this study was to explore the effect of the transportation network and required road or railroad construction upon potential supply. Consequently, the evaluation system was designed to consider the spatial location of a simulated deposit to the existing transportation network and the market destination. The construction of linkages to connect the deposit to existing transportation arteries as well as the carrying of the concentrates to the intended market were simulated. in terms of cost. The selection of the path took into account the existing terrain. This was done by digitizing topographical maps and representing the topography of the area by an orthogonal grid of elevations. The optimum route was selected by dynamic programming, meeting a prescribed gradient constraint. The potential supply of each cell was determined, allowing for the infrastructure effect."

Jan 1, 2014

By van der Merwem J. N.

"Coal pillars are prone to scaling over time, and are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces. Databases of failed and stable pillars cannot, therefore, be regarded as definitive.More than a decade has passed since the first attempt to predict the lifespan of pillars. This paper presents a review based on the new database and pillar strength equations. The methodology for the analysis is based on the fact that pillars scale. The ultimate safety factor (defined as the ratio between strength and load at the time of failure) is derived, followed by an estimate of the rate of pillar scaling, which results in an estimate of the expected time of failure.It is suggested that the predicted lifespan should not be used as an absolute indicator of expected pillar life, but rather as an index – the pillar life index (PLI) – to complement the safety factor and related probability of failure when evaluating pillar stability. A pillar with a predicted lifespan of less than 500 years would be regarded as being in danger of imminent collapse, while a lifespan of at least 1 000 years should be required for long-term purposes.The differences between this analysis and that published in 2003 constitute a strong case for regular reviews of all the empirically based stability indicators, namely the safety factor, probability of failure, and the PLI. Reviews at 10 year intervals appear to be required.The constants in the proposed method are valid only for the Witbank No. 1, 2 and 4 seams and the Highveld No. 2 and 4 seams.IntroductionIt is well known that over time, coal pillars are prone to scaling, and thus are progressively reduced in size and consequently subjected to increased load and reduced strength. The effective factor of safety therefore reduces and it is quite possible for failure to occur after a long time interval. In fact, scrutiny of the database of failed pillars indicates that failure occurred after periods ranging from 2 to 49 years, and not immediately after mining.The factor of safety for pillars at the time of failure is not the same as that immediately after mining. The concept of the safety factor is based on the pillar strength obtained by statistical analysis of pillar dimensions immediately, or within a month after mining, when the surveys are conducted. Strictly speaking, the safety factor as commonly used is valid only immediately after mining, and it continues to reduce thereafter. The ideal situation would be to base analyses of pillar strength on pillar dimensions at the time of failure, but that is simply not possible because the real dimensions cannot be measured at that time."

Jan 1, 2016

By Y. J. Cao

The purpose of this study was to investigate the flotation behaviour of unburned carbon in a cyclonic-static microbubble flotation column (FCSMC). The ash sample, collected from a power station in Guangdong province of China, was characterized by size analysis, X-ray diffraction, contact angle measurements, and X-ray fluorescence. The effect of the column flotation operating variables on the removal of unburned carbon from the fly ash was systematically studied. The feasibility of separating unburned carbon and ash was determined from the removal rate of unburned carbon (RUC) and loss on ignition (LOI). Within the range studied, the optimum diesel oil dosage was 1200 g/t, abies oil dosage was 600 g/t, pulp density was 20 per cent, superficial gas velocity was 1.4 cm/s, and circulating pressure was 0.20 MPa. The results indicate that the FCSMC technique is effective in removing the unburned carbon from the fly ash, which can be attributed to the generation of microbubbles and the continuous cyclonic circulation method. Under the optimized conditions, a cleaning ash with 2.13 per cent LOI and 94.21 per cent RUC was obtained.

Jan 1, 2012