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By Gregg Vickell

"High tonnage, low grade gold mines such as Omai’s afford economies of scale that are unto their own insufficient to sustain profitability at today’s depressed gold price. Complimenting profitability objectives is the perpetual challenge to reduce operating costs and optimize unit operating efficiencies while being ever vigilant of issues such as environmental control.Key to achieving these objectives at Omai are throughput and availability gains, gold recovery improvements, cost cutting, a skilled workforce and reliable power generation.This paper attempts a cursory review of startup-to-date plant performance advances and future performance-driven objectives.INTRODUCTIONCommercial gold production at Omai commenced in January 1993 with annual output climbing to 250,000 opy by the end of 1994. A mill expansion completed in August 1996 increased gold production to 340,000 opy by the end of 1997. Current milling capacity is 20,000 in tpd (nom.) with actual capacity exceeding 24,000 mtpd on occasion depending on ore size and competency. Mill feed consists of hardrock (andesite, quartz diorite and rhyolite) and softrock (saprolite, laterite) at a nominal ratio of 70:30, respectively. Present day average weighted grade of the feed is 1.4 g/t gold.PROCESS DESCRIPTIONPrimary crushing, ore handling and the original 12,000 mtpd hard and soft rock SABC mill, hereafter referred to as Mill ‘A’, remains essentially as documented by Konigsmann (1 994) complete with poststartup modifications. The mill expansion, hereafter referred to as Mill ‘B’, features an 8000 mtpd hard rock SABC grinding circuit similar in configuration to Mill ‘A’ and differing only in equipment size. Mill ‘B’ includes a 9000 H.P.,reversible variable speed , single pinion drive, 32 ft. x 15 ft. SAG inill , a 7 ft. short head pebble crusher and two parallel 5400 H.P., 16.5 ft. x 30 ft. ball mills with dedicated cyclone packs consisting of D26 hydrocyclones."

Jan 1, 1999

By Mark Aylmore, Guy Deschenes

"Increasing pressure are motivating the gold mining industry to develop a substitute to cyanide for leaching gold. This effort in mainly based on the high toxicity of cyanide and on the lack of efficiency of cyanide in the treatment of mildly refractory gold ores (sulfides and pre-robbing). Consequently, in only a few years, consortia have been created to elucidate and implement a new technology. The most promising lixiviant that has been selected to replace cyanide is thiosulfate. This paper will present a critical analysis on the latest developments in thiosulfate leaching and illustrate the potential and limitations of the technology with experimental data. Finally an assessment of the cost and performance of thiosulfate leaching relative cyanidation will be evaluated.INTRODUCTIONConcerns regarding the toxicity of cyanide and inability of cyanide solution to effectively leach some carbonaceous gold ore or mild refractory gold ores have made the gold mining industry evaluate alternative lixiviants to leach gold. This concern has been heightened recently by the cyanide spill at a mine in Romania, which devastated a 250-mile stretch of the Danube River and its tributaries, and the environmental damage at Ok Tedi mine in Papua New Guinea. Some countries and locations have prohibited the application of cyanidation to recover gold, such as, California and Colorado States, Japan and Turkey etc.There are several potential non-cyanide lixiviants (Sparrow and Woodcock, 1995) for gold extraction, such as thiosulfate, thiourea and halogen. A pilot feasibility test (Marchant and Broughton, 1987) had suggested that thioureation might result in acutely toxic byproducts even after treatment. Scientists of Swedish Academic of Sciences reported that thiourea is a carcinogen. Halogen is also a pollutant to the environment. Thiosulfate leaching appears to have very good potential as an effective and less hazardous procedure for gold and silver extraction from auriferous ores. In addition to being much less toxic and relatively cheap, it forms a stable anionic complex with gold but is less reactive to other metals, for instances, copper and carbonaceous materials, in comparison with cyanide."

Jan 1, 2001

By K. Lillkung

The use of secondary sources has become a very attractive option in the production of platinum group metals as the demand for these metals has increased. One of the promising sources are the spent catalysts, which provide an attractive source as they have relatively high metal content compared with the ores where the metal is originally extracted from. The concentration of platinum group metals in catalytic converters is in the range of 0.1-1 mg/g as the concentration in ores is usually less than 10 ppm. Aqueous chloride solutions are claimed to be a cost effective way in which platinum group metals can be brought into solution and concentrated. In this work the effect of different parameters on the dissolution rates of Pt, Pd and Rh were investigated. The parameters investigated were the Cl- concentration, the temperature and applied potential. Chloride concentrations used were 62.5, 90, 125 and 250 g/l. Temperatures were 25, 40, 60, 80 and 94°C. Potentials were 950, 1050, 1150 and 1250 mV vs. SCE for Pt and Rh and 380, 410, 440 and 470 mV vs. SCE for Pd. Dissolution rates were calculated from the current densities determined by potentiostatic measurements. Thin wire electrodes were used in the measurements. As expected, higher temperature, higher potential and higher chloride content generally increasedthe dissolution rate. However, the effect was nonlinear and slightly different for different metals. The results show that strongly oxidizing conditions are needed in order to achieve reasonable dissolution rates. With the highest chloride content and at potential of 1250 mV vs. SCE and 94°C the dissolution rate of Pt was nearly 1 mm/h, whereas in similar conditions Rh achieved a rate of only nearly 100 µm/h. Similarly, with the highest chloride content and at potential of 470 mV vs. SCE at 94 °C the dissolution rate of Pd was nearly 1 mm/h. Keywords: Pt, Pd, Rh, redox potential

Sep 1, 2012

By Pengfu Tan

In order to understand and control the formation of dioxins and furans in iron ore sinter plants and to possibly avoid expensive end-of-pipe removal stages, thermodynamic calculations on the stability of PCDD/Fs have been combined with sinter process modeling by means of kinetic simulations and computational fluid dynamics. Three thermodynamic databases of PCDD/Fs, derived using the Group Additivity approach and two computational molecular modeling methods, MNDO and PM3, respectively, combined with the SGTE database have been used to model the PCDDIF formation and the PCDD/F isomer distributions in iron ore sintering. The results show very similar conditions for the PCDD/F formation using the different databases. Concerning the isomer distributions of toxic PCDD/Fs, the calculated data derived from the MNDO method showed optimum agreement with measured values from industrial furnaces. The thermodynamic calculations also proved that the isomer distributions found in combustion processes are equilibrium values, while the concentration ratio of PCDDs and PCDFs in the wind boxes does not represent an equilibrium state. A CFD model of iron ore sintering was developed including the kinetics of 15 relevant chemical reactions with the aim of describing the temperature fields and the gas flow patterns in the sintering bed. On the basis of these information and experimental observations reported in the literature possible pathways for the PCDD/F formation in sinter plants have been identified: PCDD/Fs are formed in the critical temperature region of 250 to 450°C below the combustion zone, are then carried downwards with the gas and condensed close to the bottom of the sinter bed. Transported to the discharge end with the solid mixture, they are again released into the gas phase when the flame front approaches the bottom. PCDD/Fs are also formed in the last wind boxes when the hot off-gases cool down and reach the critical temperature range. Consequently, methods to reduce the dioxin formation reactions are proposed including inhibitor addition to the sinter feed and off-gas quenching.

Jan 1, 2003

By M. Rafieazad, P. Fathi, M. Mohammadi, A. M. Nasiri

This study examines the impact of low-temperature heat-treatment cycles on microstructure and corrosion performance of Direct Metal Laser Sintered (DMLS) AlSi10Mg_200C alloy. It is known that the morphology and the size of eutectic silicon are the two most critical factors affecting mechanical properties of the AlSi10Mg alloy. Although high-temperature heat-treatment cycles are common practice for the strengthening of the cast Al-Si-Mg alloys, in the case of DMLS fabricated AlSi10Mg alloy, high-temperature solutionizing and low temperature annealing cycles result in mechanical properties degradation primarily due to a dramatic decrease in the solubility of Si in the super-saturated a-Al matrix of the as-printed alloy. In this study, low-temperature heat-treatment cycles that only promote precipitation of Mg2Si accompanied by the interruption of the silicon network were investigated. SEM and XRD analysis were utilized to observe the evolution of microstructure after different annealing heat-treatments at 200, 250, 300, and 350°C for 3 h. Increasing the heat-treatment temperature contributed to the Si networks interruption and an increased precipitation of Mg2Si phase. Low-temperature annealing from 200 to 350°C was found to promote the homogeneity of the microstructure characterized by a uniform distribution of eutectic Si in ??-Al matrix. Additionally, in order to investigate the impact of different heat-treatment cycles on corrosion resistivity of DMLS-AlSi10Mg_200C, the potentiodynamic polarization testing and electrochemical impedance spectroscopy were performed in a 3.5 wt.% NaCl solution. The results revealed more uniformly distributed pitting attack on the corroded surface by further increase in the annealing temperature, which was attributed to the uniformity of Si dispersion. In comparison, low temperature annealing at 200°C led to a penetrating selective corrosion attack along the melt pool boundaries, resulting in a higher corrosion rate associated with the lower absolute value of impedance of the protective passive film on the surface.

Jan 1, 2018

By P. J. G. du Toit

Coal has been mined by hydraulic means in various countries, i.e., the hydraulic excavation as well as the hydraulic transportation of coal has been practiced. The hydraulic breaking (excavating) of the coal dealt with in this report is done with a monitor directing a continuous jet of water against the coal, pressures of the water fed to the monitor range from 1500 to 2500 psi at 1000 to 2100 gallons per minute Transportation of the coal from the face is by fluming if a gradient exists in excess of 2 or, when a favorable gradient does not exist, the coal is pumped from the face by various types of pumps which can also provide a means of hoisting the coal to a higher elevation as required. This report presents: 1.) preliminary design and systems specifications of a demonstration miner (referred to in this report as a hydrominer), 2.) mining configurations showing good promise for typical U.S. thick coal seams in three seam pitch categories, (0° - 7° = Flat, 8° - 250 = Moderately pitching, 26° - 90° = Steeply pitching), along with cost and performance analyses for each system, 3.) a definition of a test program that will provide sufficient engineering data for the design of a water jet cutting head adequate to cut coal in those typical U.S. thick coal seams The cost analysis data showed hydraulic methods to be economically competitive with present thick seam raining methods. It was also obvious from studies of thick seam methods that the hydromonitor coal breaking techniques are more effective in removing most of the coal with its long reach of 60 ft. (or more), is a safer mining system, and produces very low coal-dust content in the air (usually less than 0.25 to 0.5 mg/cubic meter). It is recommended that the USBM sponsor a full fledged continuation of the hydraulic mining program: (a) lab, test, (b) preliminary field test, (c) full scale mine section demonstration.

Jan 1, 1977

By Srisharan Shreedharan, Gang Huang, Hongqing Song, Sijing Cai, Pinnaduwa H. S. W. Kulatilake

"Subsidence prediction due to ore extraction and backfilling during open stoping operations for an underground iron mine in China is reported in this paper. Geologic complexities including faults and interfaces between different lithologies, and the stoping and backfilling sequence adopted from the mine plans are incorporated in building a 3-D discontinuum numerical model. The stoping was simulated in two vertically stacked horizontal layers (Fig. 6), with a total of 16 stopes. Large displacements of up to 50 cm were predicted by the model along the roof of the stopes, and a maximum surface subsidence of 22.5 cm was predicted by the model . Backfilling was found to control subsequent deformation and subsidence. The extraction of the upper orebody led to relatively higher deformations in extracting the lower orebody. Finally, the predicted subsidences at all locations along a subsidence profile on the surface are shown. INTRODUCTIONThe Luohe iron (pyrite) mine, located in Hefei city, Anhui, is an underground mine in China, being mined through the open stoping and backfilling method. The orebody lies at a depth of about 400 m and has an average thickness of 250 m (Magang Steel Company, 2012a). Discovered in the 1960s, the location has a reserve estimate of 300 million tonnes, which have been mined since 2007. The orebody has been divided by the mining company into an east and a west section to facilitate planning for ore extraction.Figure 1 shows the surface above the underground mine, which is seen to contain many human settlements including villages, pastures and other infrastructure. These locations are sensitive in that they need to be protected from potential adversities due to the underground mining, such as surface subsidence. While it may be near impossible to completely eliminate the subsidence due to ore extraction, operations such as backfilling have been adopted by the mine in an effort to substantially reduce it. Hence, this paper focuses on describing the stability of stopes in the mine and estimates future subsidence in a part of the mining region through numerical modeling."

Jan 1, 2016

By A. Chieregati

"A heterogeneity study was conducted at Erdenet Copper Mine in Erdenet, Mongolia, to determine the fundamental sampling error associated with blasthole sampling. Using a custom-designed radial sector sampler , six large blasthole samples were collected to perform the test developed by Pierre Gy to estimate the constitutional heterogeneity factor for several types of ore. The study investigated the entire sampling protocol, from mine site to analytical laboratory, and quantified the contributions to total fundamental sampling error (FSE) of the various stages of sampling and sample preparation. The total relative standard deviation (RSD) of sampling error was determined to be ±32 percent and ±49 percent for copper and molybdenum, respectively. A classical statistical analysis conducted on the data confirmed the level of RSD native to the mine. About 95 percent of the sampling error was produced from one step in the sampling protocol, that of reducing the 1-kg (2.2-lb) sample to a 50-g (0.1-lb) sample for chemical analysis. To minimize the FSE, the use of a slightly different protocol is recommended. The 1-kg sample (2.2-lb) should first be crushed to a 0.2 cm (0.08 in.) particle size. A 250-g (0.5-lb) split should then be taken and finely ground, from which a 50-g (0.1-lb) split should then be taken. These simple steps will reduce the FSE to ±7.9 percent and ±17 percent for copper and molybdenum, respectively. The modified sampling protocol effectively reduces to zero the probability of misclassification of low-grade ore. IntroductionThe Erdenet copper-molybdenumopenpit mine is located in Erdenet in Orkhon province, Mongolia. The Erdenetiin-Ovoo copper-molybdenum deposit is 1.5 to 3 km (0.9 to 1.8 miles) wide and about 25 km (16 miles) long (Erdenet Mining Corp., 2016). Chalcopyrite, bornite covellite and chalcocite are among the major minerals found in the deposit. The mine processes approximately 26 Mt/a (29 million stpy) of ore."

Jan 11, 2017

By A Batnasan

The main techniques for removing phosphorus as an impurity element in iron ores are physical separation (eg magnetic separation, flotation) and chemical separation (acid leaching, alkaline leaching and bioleaching) processes. When phosphorus occurs in iron ores as phosphate minerals (eg apatite), physical separation techniques are available to remove the gangue minerals from iron ores. On the other hand, chemical separation techniques are suitable for reducing phosphorus levels when phosphorus is associated with iron oxides and hydroxide minerals (eg hematite and goethite) in iron ores. In this study, the dephosphorisation behaviour of high‐phosphorus iron ore was investigated using a combined alkaline roasting and water washing route. Roasting experiments were conducted in a laboratory electric muffle furnace using various concentrations of sodium hydroxide (0–50 g/kg-ore) at different temperatures (100–800°C), varying prolonged roasting times (0–2 h), and different particle size fractions (-45 – -250 μm) of iron ore. Subsequent water washing experiments were employed at various pulp densities (30–1000 g/L) and different times (30 s – 60 min) at ambient temperature. Hematite and goethite were the main minerals in the high-phosphorus iron ore identified using X-ray diffraction (XRD) analysis. The iron ore assaying 63.40 mass % Fe (total), 0.240 mass % P, 4.53 mass % SiO2, and 4.87 mass % Al2O3 were determined using X-ray fluorescence (XRF) analysis. The results showed that the dephosphorisation rate of the iron ore was reached 40 per cent under roasting with 50 g/kg-ore NaOH at 300°C for 2 hours followed by water washing for 30 s with a pulp density of 300 g/L, reducing phosphorus level from 0.240 mass % to 0.142 mass %. At the same time, the iron grade was reduced from 63.4 per cent to 59.5 per cent, caused by an iron loss. Results suggest that the phase transformation of goethite induces the dephosphorisation of high-phosphorus iron ore under alkaline roasting conditions

Aug 22, 2022

By Pengfu Tan

In order to understand and control the formation of dioxins and furans in iron ore sinter plants and to possibly avoid expensive end-of-pipe removal stages, thermodynamic calculations on the stability of PCDD/Fs have been combined with sinter process modeling by means of kinetic simulations and computational fluid dynamics. Three thermodynamic databases of PCDD/Fs, derived using the Group Additivity approach and two computational molecular modeling methods, MNDO and PM3, respectively, combined with the SGTE database have been used to model the PCDDIF formation and the PCDD/F isomer distributions in iron ore sintering. The results show very similar conditions for the PCDD/F formation using the different databases. Concerning the isomer distributions of toxic PCDD/Fs, the calculated data derived from the MNDO method showed optimum agreement with measured values from industrial furnaces. The thermodynamic calculations also proved that the isomer distributions found in combustion processes are equilibrium values, while the concentration ratio of PCDDs and PCDFs in the wind boxes does not represent an equilibrium state. A CFD model of iron ore sintering was developed including the kinetics of 15 relevant chemical reactions with the aim of describing the temperature fields and the gas flow patterns in the sintering bed. On the basis of these information and experimental observations reported in the literature possible pathways for the PCDD/F formation in sinter plants have been identified: PCDD/Fs are formed in the critical temperature region of 250 to 450 °C below the combustion zone, are then carried downwards with the gas and condensed close to the bottom of the sinter bed. Transported to the discharge end with the solid mixture, they are again released into the gas phase when the flame front approaches the bottom. PCDD/Fs are also formed in the last wind boxes when the hot off-gases cool down and reach the critical temperature range. Consequently, methods to reduce the dioxin formation reactions are proposed including inhibitor addition to the sinter feed and off-gas quenching.

Jan 1, 2003

By J L. Mauk

The Favona low-sulfidation epithermal deposit in the Hauraki Goldfield is less than one kilometre southeast of the world-class Waihi deposit. The veins are hosted by Miocene andesite of the Waipupu Formation, have a total strike length of over 900 m, a vertical extent of at least 350 m and are on average 1-3 m wide. Individual veins have been traced for 200 m with vein widths up to 25 m. The dominant vein textures are massive, crustiform and colloform banded veins and vein breccias, with subsidiary textures including comb, zonal, flamboyant, lattice-bladed, ghost-bladed, microbotryoidal, feathery, ripple and dune, and mosaic. Quartz is the dominant gangue mineral in the veins at Favona. Accessory gangue minerals include pyrite, chlorite, adularia, illite, calcite, and apatite. Electrum is the only gold-bearing mineral and is associated with acanthite, tetrahedrite, pyrite, sphalerite, chalcopyrite, and rare galena. Whole rock geochemical analyses indicate that gold grades correlate with textures/vein types. The highest gold and silver grades are associated with vein breccias and banded veins, with the former containing 28 g/t Au and 136 g/t Ag on average. To date, veins with significant base metal sulfide minerals, and massive and crystalline quartz veins lack significant precious metals (Au and Ag), except for one banded base metal sulfide-rich vein at Favona. An overall zonation of the veins exists, with base metal sulfide minerals dominant at depth (1050 m RL) of the system are characterised by massive quartz veins.Fluid inclusions from a total of 31 quartz, amethyst, and platy calcite veins reveal homogenisation temperatures of ~190-250¦C, with final ice melting temperatures between 0.0 and -0.8. Ice melting temperatures correspond to apparent salinities of 18O value of the water that formed these veins is approximately -4.3 per mil, which is similar to meteoric water-dominated, near neutral chloride fluids in geothermal systems in the Taupo Volcanic Zone.

Jan 1, 2002

By H. M. Smith

The production of oil during the early development of new oil fields is usually accompanied by large quantities of natural gas. Since in most fields gas under pressure is the main source of the energy causing the oil to move through the sands into the wells and thence to the surface, coincident production of natural gas will continue as long as present-day producing methods are used. Generally the use of this gas is restricted, since markets are not readily available. Consequently, a large portion of the gas produced is often wasted by being blown into the air. The magnitude of gas waste is well brought out in the following statement by Miller: Only a few outstanding examples need be cited to show the enormous production of gas that accompanies the discovery of new sources. In April, 1927, the Ventura field produced approximately 210,000,000 cubic feet of gas a day; and during the summer of 1927 Alamitos Heights reached a peak gas production of 102,000,000 cubic feet a day; on December 1, 1928, as a result of the summer's deeper drilling in the Long Beach field, surplus gas production amounted to 125,000,000 cubic feet a day. It is estimated that in January, 1929, the Santa Fe Springs field will have a daily gas production of 387,000,000 cubic feet from the recently discovered Buckbee oil zone. Of this amount 274,000,000 cubic feet a day will be the surplus above the amount required for field and plant fuel, taken out by utility corporations, and that used in repressuring the Meyer sand which lies above the Buckbee zone. Another example of the waste of large volumes of gas is in the Turner Valley field of Canada, 6 where 250,000,000 cubic feet of gas is being burned daily in flambeaus, as there is no present outlet except at prohibitive cost.

Jan 1, 1931

By G. A. Calboreanu

The deterioration rate of any metal or alloy is increased in a erosion-corrosion application because the passivated surface layer is removed by the solid particles contained in the pumping fluid and the material surface becomes corrosion active again. Numerous attempts were made to engineer a stainless steel with higher hardness and unaffected corrosion performance. Alloys of special interest in such an area are the duplex stainless steels particularly the CD4MCu material which has the capability of being hardened due to its ferrite-austenite microstructure. Unfortunately, the CD4MCu with higher hardness comprises different intermetallic compounds which will increase the corrosion rate. The paper is analyzing the erosion-corrosion performance of a new type CD4MCu with 32% higher hardness and superior wear performance. The corrosion performance of this new alloy is better than the corrosion performance of the common CD4MCu material which the industry could supply. The corrosion effectiveness was evaluated with the Huey test, the potentio-dynamic anodic polarization test and the weight loss test whereas the erosion resistance was established with a low stress ball-on-disk type apparatus. The Wilfley hard CD4MCu material had 39% lower weight loss and 32% smaller volume loss than the common CD4MCu alloy The hardness range of the centrifugal pump parts poured and processed with the Wilfley hard CD4MCu is 320-350 HBN versus 250-255 HBN hardness of the common CD4MCu. After approximately 18 months of experimentation and field testing it was clearly established that the life time of the Wilfley hard CD4MCu is higher than the common CD4MCu and consequently centrifugal pumps were supplied to different customers in North and South America. India and more recently Australia. Sincere thanks to all the A.R. Wlifley & Sons Inc. and Western Foundries Inc. employees who supported and help to implement this work. special thanks to George Wilfley. Dave Cobb. Bill Becker Sr. and Wayne Taylor

Jan 1, 1997

By J. R. Thoenen

Occurrences of titanium, zircon, and other heavy minerals are widespread in Florida. Deposits concentrated as "heavy sands" by wave and wind action have been mined an or near Atlantic coastal beaches since 1916. An important inland concentration at Trail Ridge; .believed, to have formed along an ancient share line, was disclosed by preliminary investigations (1947) by the Bureau of Mines. This report describes investigations to determine the extent and grade of "heavy-sand" concentrations in other inland stabilized dune areas in the central and northeastern parts of the peninsular region of the State (fig. 1). Portions of the central highlands south of Trail Ridge and terraces and dunes in the coastal lowlands region, east and northeast of the ridge were investigated. Holes were drilled in 10 areas (see table 1 and fig. 2) along State and County roads and on private property. They were spaced 250 to 2,000 feet. apart to obtain a cross-sectional picture of "heavy-sand" content in the dune areas. These sections were generally placed in east-west directions normal to the general trend Of the peninsula to intercept one or more of the ancient shore lines in the various areas (figs. 3 to 13, incl.). A total of 433 drive piped and 333 jet holes were drilled, and 7,293 bulk and core samples were taken. Intermediate holes were placed at closer intervals when field analysis indicated 1 percent or more of heavy minerals present. All of the holes drilled and the analysis of samples that contain 1 percent or more of heavy minerals are listed in table 2. Profiles (figs. 14 to 20, inclusive) show the more favorable concentrations, which are irregular in outline and erratic in mineral content. The highest concentrations averaged 1.0 to more than 2.0 percent heavy-mineral content, but a few individual hole samples exceeded 3.0 percent in grade.

Jan 1, 1949

By Muhlis Nezihi Saridede, Burak Birol

"Plastic wastes become a huge problem day after day due to the growth of plastic industry. Hence, plastic wastes are used as a raw material for other industries to eliminate this problem. Polyethylene Terephthalate (PET), which is a commonly used for plastic bottles, may be used as a reductant for iron production because of its C and H content. In this study, composite pellets containing iron ore concentrate, flux and reductant were prepared. Crystallized PET bottles and coke were used as reductant after ground. The prepared pellets were reduced at 1400°C for 30 min. and iron nuggets were produced successfully. Usage of waste PET purely as reductant in composite pellets gives a higher metallization degree and proper carbon content than coke or PET and coke mixture.IntroductionPlastics can be seen everywhere, especially at packaging industry in our common life. The increase of plastic usage ends up with an important problem, the increase of plastic wastes. Many studies have been carried out about usage of plastics as reductant in blast furnace such as Ogaki et al.[1], Heo and Baek [2], Saridede [3], Minoru et al.[4], etc. Besides, Kato et al. [5], Nomura and Kato [ 6] and Rongfang et al. [7] have pointed out the use of plastics in coke ovens regarding to its calorific content. In EAF steelmaking, a mixture of coal/coke and plastics can be used as an energy resource instead of only coal or coke. [8] Also, the usage of PET and PC in self-reducing composite pellets increased the reduction rate. [12]PET (Polyethylene terephthalate ), which is commonly used for soft drink plastic bottles, is an important municipal waste due to its C and H content. Its molecular formula is (C10H8O 4)n. Demir and Dincer showed that PET can be crystallized at 250°C to have the ability to be ground to a size of -100 µm. [11]"

Jan 1, 2011

By S Maritsa, E K. Chiw, S M. Masuka, ika, T Chisahwira

The challenges facing the iron and steel industry on a global scale include depletion of resources, excessive energy consumption and CO2 emissions. The demand for iron products in Zimbabwe is increasing due to massive infrastructural development like the expansion and improvement of mines, transportation systems and water reticulation systems. In Zimbabwe, cast iron is mainly produced from recycling scrap metals. However, there is currently a shortage of scrap, which has resulted in the vandalism of state-owned properties such as bus terminals and railway wagons and lines. Nevertheless, Zimbabwe is rich in limonite ores, which are idle, and it is the aim of this research that this ore be effectively used to produce feed for the blast furnace. The Blast furnace is still the dominant route for producing cast iron where sintering and coking processes are the main routes for preparing the Blast furnace feed. Sintering limonite ore fines produce a product of poor grade because of the high value of loss on ignition that has a negative impact on Blast furnace efficiency. Furthermore, the sinter plant and the coke making plant are a source of huge environmental pollutants. Since low-grade Zimbabwean coal fines are also available in abundance, they were used in the preparation of the composite materials ensuring maximum utilisation of this resource. Results showed that the limonite ore-carbon composite pellets sample with the binary basicity ratio (CaO/SiO2) of about 0.27 and coal addition of 2.5 wt per cent achieved above the minimum physical quality requirements for pellet of drop number of at least 4, a dry compression strength greater than 2.2 kg/pellet and an indurated compression strength greater than 250 kg/pellet. Limestone was used as the binder avoiding the addition of more impurities from bentonite since the limonite ore has a high SiO2 content of about 8.78 wt per cent.

Sep 18, 2023

By J. F. Strunk Jr., K. M. Maestas, Y. Chai, R. Legrand, R. L. Henry

"Groundwater with elevated sulfate and metals was treated using a sulfate-reducing pilot PRB consisting of a series of six 55-gallon drums filled with ZVI (Fe0), wood chips, brewer’s grain, limestone, and emulsified vegetable oil (EVO). Sulfate was reduced from >3000 to <250 mg/L and Cd, Cu, Mn, Ni, and Zn concentrations declined up to 3 orders of magnitude in 2 months. After 6 months of operation at a 7.5-day retention time, sulfate broke through. The system was then rejuvenated by injecting EVO in each drum and operated for 5 more months at a 9-day retention time. Sulfate-reducing bacteria in the pilot system grew by 3 to 4 orders of magnitude in 2 months. Some cementation of the PRB media was observed in the first 2 drums after processing 25 pore volumes of water, due to preferential upstream precipitation of sulfides; this can be avoided through operational modifications. Resulting permeability losses were estimated at 0.1% to 1% per year. Absorption/desorption was tested at bench scale to evaluate the use of various oil-soaked sorbents as a slow-release source of dissolved carbon. The tests show that oil sorbents can increase the useful life of a PRB between oil refills to over 100 pore volumes. INTRODUCTION High sulfate concentrations, low pH, and dissolved metal exceedances are common characteristics of mining-influenced water (MIW). Above-ground such MIW is commonly treated using biochemical reactors (BCR). In a BCR, water generally flows vertically downward through a layer of biomass, such as wood chips, manure, etc. The dissolved organics resulting from the decay of this material are oxidized first by microbes using oxygen; when oxygen is exhausted, other bacteria can use nitrate, and then sulfate to oxidize organics. Sulfate-reducing bacteria (SRB) reduce sulfate to sulfide, which quickly reacts with metals to form largely insoluble and stable metal sulfides. These reactions can be summarized as follows:"

Jan 1, 2016

By M. Szumny, P. Mertuszka, K. Fuławka

Multi-face destress blasting is the basic method of reducing rock bursts hazard in Polish copper mines. Blasting operations of this type are carried out regularly, twice a day in order to minimize the geomechanical threat while maintaining high extraction efficiency. Unfortunately, the assessment of the effectiveness of this type of work until recently was based solely on determining whether a given destress blasting triggered a seismic tremor or not. Information on the characteristics of seismic waves generated by blasting has so far been omitted in analyzes of this type. The lack of reliable knowledge about the generated seismic effect makes the introduction of constructive corrective actions impossible, which in turn significantly reduces the triggering potential of conducted destress blasting. This study presents the results of comprehensive work aimed at increasing the effectiveness of active rock burst prevention in Lubin underground copper mine. For this purpose, an innovative system for continuous monitoring of seismic waves induced by multi-face blasting was developed, and a new method for seismic assessment of the effectiveness of the destress activities was developed. Based on the monitoring data and the comprehensive analysis of waveforms, the effectiveness of individual blasting operations was determined, and on this basis, modifications to the drilling and blasting patterns aimed at increasing the generated seismic effect were proposed. The use of innovative methods of monitoring and data analysis made it possible to increase the effectiveness of the conducted activities, in terms of the magnitude of seismic vibrations, up to 250%. It should be emphasized that the introduced modifications were aimed at increasing the effectiveness of blasting operations while maintaining the current level of costs. As a result, the proposed solution turned out to be highly effective, significantly affecting the safety of the work carried out in Polish copper mines. Thus the proposed approach is planned to be implemented in other Polish copper mines belonging to KGHM in the near future

Jan 1, 2024

By Paul T. Allsman, William H. King

This paper describes the mineral deposits and discusses mining and milling methods and related operating conditions in an important gold , silver , lead , and zinc producing area in Ouray , San Juan , and San Miguel Counties , Colo . , embracing 250 square miles . The area is extremely rigged , ranging in altitude from 8,000 to 14,000 feet . White men first penetrated the area in 1860 , when a group known as "Baker's party " reached the Animas River at the present site of Silverton . Difficulty with the Ute Indians and severe winter conditions forced the group to leave and discouraged prospecting until 1870 when a second party reached the area and succeeded in producing 27 tons of ore yielding $ 150 a ton in gold . In 1874 , a treaty with the Indians opened the way for settlement , and mining was established . The total production of recovered gold , silver , copper , lead , and zinc from 1873 to 1947 in the three counties is valued at $347,057,047 . The greatest amount of the total production of metals has come from quartz veins in a series of volcanic flows and tuffs covering the area and comprising a blanket 4,000 to 6,000 feet thick . The igneous rocks were laid down upon an old , erosion surface and are underlain by Paleozoic and Mesozoic sedimentary rocks and pre - Cambrian metamorphic rocks . The entire series has been intruded in late Tertiary time by volcanic dikes and larger intrusive masses . Because of the great differences in relief generally , mining is con-" ducted through long adits to a depth of nearly 3,000 feet below surface in places . Walls of the veins generally stand well , and shrinkage stoping is the most common method of mining . As virtually all of the ore is sulfide below the zone of oxidation , flotation is the common method of milling . Generally , where gold is important , provision is made for its removal by jigs . Bulk concentration of the sulfides is practiced at smaller mills , but the larger operations make a selective lead and zinc concentrate or a lead , zinc , and copper . concentrate . ·

Mar 1, 1950

By V. L. Brant

The estimated total energy demand for the U. S. is 60 quadrillion BTU for 1970 and is expected to more than double by the year 2000. Today the energy crunch is here. It is evidenced by brownouts to conserve electrical energy, and the shortage of gas and oil. Many such headlines, including suggestions for energy rationing, are in the news today. Paradoxically we have an abun¬dant source of energy in the form of coal. Energy consumers, however, are geared to use fuel in other forms. It is the intent of this presentation to discuss processes under development for conversion of the energy contained in these vast coal reserves to more usable forms. Coal conversion to synthetic fuels should be of particular interest to those engaged in mining. Coal consumption in a single conversion plant is expected to range from 10,000 to 40,000 tons per day. Plants will be located at the mine mouth and will operate continuously 24 hours per day, seven days per week. Further, many such plants will be needed., For example, looking at natural gas only, the supply/demand gap is estimated to reach 13 trillion cubic feet per year by 1985. Assuming that gas from coal conversion were to fill only 101, of this gap, 15 plants producing 250 MMSCFD would be required. These plants would consume about 90 million T/Y of coal or 135 million T/Y of lignite. It will indeed be a challenge to mine, suitably prepare, and transport coal at these projected volumes without interruption. Many ideas have been conceived for better utilization of coal and these are in various stages of development. Those supported by public money are covered by the Annual Report of the Office of Coal Research (OCR), Department of Interior. The two processes to be discussed, the Consolidation Coal Company&apos;s (Consol) C02 Acceptor Process (1)(2)(3) and The Pittsburg and Midway Coal Mining Company&apos;s (PAMCO) Solvent Refined Coal Process (4)(5) are both sponsored by OCR.

Jan 1, 1970