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By Terry Gerritsen

Electrode control systems regulate power by physically positioning electrodes. Due to the size and weight of the electrodes, this is a slow process where power fluctuations of less than a few seconds are not fully corrected. Nickel laterite AC furnaces operating in high voltage, shielded arc mode typically incur frequent power variations of approximately +I- 20% around the power set point. To provide enhanced power control, Hatch has recently supplied, at commercial scale, a thyristor-switched reactance system called an SPLC. The SPLC incorporates predictive control software to operate furnaces at the maximum transformer rating by reducing the large power swings experienced without the SPLC. The resulting higher average power enables higher production, without increasing transformer or power plant capacity. A 60 MW production version of the SPLC has been commissioned on Falconbridge Dominicans's laterite nickel furnace, yielding excellent results with a 3 to I reduction in power fluctuations. Keywords: SPLC, AC Power Supply, Electric Furnace

Jan 1, 2004

By Ann M. Hagni

Microscopic and sub-microscopic gold commonly occurs in the sulfides pyrite (FeS2) and arsenopyrite (FeAsS). To recover refractory gold from pyrite and arsenopyrite, these sulfides are roasted to create porosity and permeability prior to cyanide leaching. A new technique of adding Ca(OH)2 and roasting, lime roasting, is being tested fro retention of contaminant sulfur and arsenic and effectiveness of gold extraction. The lime roasting of pyrite at 500°C contains anhydrite (CaS04), oldhamite (CaS), calcium sulfite (CaS03), hematite (FenOs), magnetite (Fe304), and minor amoupts of pyrrhotite (Fel.,S) and pyrite. At 600°C, 650°C and 700°C the main constituents are anhydrite, oldhamite, and hematite. At 750°C, 800°C, and 850°C anhydrite and hematite are the predominant phases with minor oldhamite. The roasted arsenopyrite at 450°C contains arsenic sulfides, calcium sulfite, hematite, and minor oldhamite, magnetite, and pyrrhotite. At 650°C 700°C 750°C and 800°C, anhydrite, oldhamite, arsenic sulfides, and hematite are present. At 850°C, the same minerals are present with only minor oldhamite. The textures indicate excellent porosity and permeability at high temperatures. At lower temperatures, rims of low porosity and permeability surround many of the pyrite and arsenopyrite particles, and the complete oxidation process from pyrite to pyrrhotite to magnetite to hematite can be observed in individual particles. Additional process mineralogy techniques employed in this study are electron spectroscopy for chemical analysis (ESCA) and cathodoluminescence microscopy (CLM) in the determination of the calcium-sulfur- oxygen phases. Automated scanning electron microscopy-energy dispersive spectroscopy was utilized to analyze particles for gold content. Reflected light microscopy (RLM) revealed arsenopyrite and pyrite encapsulated in quartz particles of the gold-bearing concentrate, gold roast calcine, and cyanide leach residue.

Jan 1, 1993

By Ann M. Hagni

Microscopic and sub-microscopic gold commonly occurs in the sulfides pyrite (FeS2) and arsenopyrite (FeAsS). To recover refractory gold from pyrite and arsenopyrite, these sulfides are roasted to create porosity and permeability prior to cyanide leaching. A new technique of adding Ca(OH)2 and roasting. lime roasting. is being tested for retention of contaminant sulfur and arsenic and effectiveness of gold extraction. The lime roasting of pyrite at 500°C contains anhydrite (CaS04). oldhamite (CaS). calcium sulfite (CaS03). hematite (Fe2O:J), magnetite (Fe304). and minor amounts of pyrrhotite (Fe1.xS) and pyrite. At 600°C. 650°C and 700°C. the main constituents are anhydrite. oldhamite. and hematite. At 750°C. 800°C. and 850°C. anhydrite and hematite are the predominant phases with minor oldhamite. The roasted arsenopyrite at 450°C contains arsenic sulfides. calcium sulfite. hematite. and minor oldhamite. magnetite. and pyrrhotite. At 650°C. 700°C, 750°C. and 800°C, anhydrite, oldhamite. arsenic sulfides, and hematite are present. At 850°C. the same minerals are present with only minor oldhamite. The textures indicate excellent porosity and permeability at high temperatures. At lower temperatures, rims of low porosity and permeability surround many of the pyrite and arsenopyrite particles, and the complete oxidation process from pyrite to pyrrhotite to magnetite to hematite can be observed in individual particles. Additional process mineralogy techniques employed in this study are electron spectroscopy for chemical analysis (ESCA) and cathodoluminescence microscopy (CIM) in the determination of the calcium-sulfur-oxygen phases. Automated scanning electron microscopy-energy dispersive spectroscopy was utilized to analyze particles for gold content. Reflected light microscopy (RIM) revealed arsenopyrite and pyrite encapsulated in quartz particles of the gold-bearing concentrate, gold roast calcine, and cyanide leach residue.

Jan 1, 1991

By Frank L. Marley

Mexican silver mining operations have existed prior to the sixteenth century. Recovery processes throughout the centuries were developed at the many mining localities and have progressed through smelting, the patio amalgamation process to cyanidation and flotation. A study of the present silver mining zones is reported in which the ores are evaluated in terms of the mineralogical constituents, their state in the matrix and the effects these are noted to impose upon the efficiency of recovery. Silver mineral identifications were by quantitative ore microscopic and scanning electron microscopic energy dispersive X-ray analysis.

Jan 1, 1982

By Fei Ruan, Chang Jie, Jieyu Zhang, Jifang Xu, Kuochih Chou

"Viscosity is an important physical parameter of slag in metallurgical processes. Most studies of aluminosilicate and aluminate systems assume that aluminum occurs in tetrahedral coordination. In this case, the O/Al value was utilized to describe the slag structure in terms of the network character of aluminosilicate and aluminate systems. Composition and viscosity data from the literature and the trends were analyzed according to the O/Al value and temperature. It is shown that the O/Al value explains the trends in more than ten systems, containing more than 200 data points. In aluminosilicate system, the trend that viscosity value changed with the O/Al value which implies that alumina works as network modifier such as basic oxides or works as network former. In aluminate system, irrespective of the system and temperature, viscosity decreased steadily when O/Al value increased.IntroductionMetallurgical slags, which are made of oxides and many also contain fluorides, play an important role in metallurgical processes such as iron making, steel making and secondary refining processes. Aluminosilicate slags are applied in a wide range in metallurgical processes. A polymerizing effect of the addition of aluminum and an association of aluminum with low-field strength cations is further supported by the physical variations in aluminosilicate and aluminate systems. The behavior of alumina in silicate melts has been studied by various authors [1-6]. It is generally accepted that Al2O3 is amphoteric [7] because Al3+ ions can behave in profoundly different ways in a silicate melt. When added to a pure silica melt, Al2O3 acted as a network modifier, breaking the bridging oxygen of the pure silica network, thereby decreasing the viscosity in a similar manner to other network modifiers, such as CaO, MgO etc. to form [1/2Ca (AlO4)]4- complexes. However, when added with alkali or alkali-earth metal oxides, some of the Al3+ ions assumed tetrahedral coordination and replaced Si in the liquid network, the missing charge was compensated by Me+ or Me2+ ions. Therefore Al2O3 acted as network former [8, 9]."

Jan 1, 2011

By P. A. Metz

The Fairbanks mining district has produced in excess of 7.5 million troy ounces of placer gold and 250,000 troy ounces of lode gold since gold was discovered in the area in 1902. The Fairbanks mining district is in the northwestern part of the Yukon-Tanana Upland Schist Terrane, an area of approximately 240,000 square km that has produced in excess of 21 million troy ounces of placer gold. All lode gold production from the terrane has been from the Fairbanks mining district, and all of this production has come from high-grade gold-quartz polymetallic sulfide veins. Gold grains in the vein systems range in size from several microns to several millimeters. The veins are localized in a complex metamorphic rock assemblage which includes greenstones and felsic metavolcanic and volcaniclastic rocks, meta-exhalatives and metasediments. Gold occurs in the metamorphic host rocks in association with arsenopyrite and arsenic-antimony sulfides which occur in lenses and disseminations within the host rock. Gold grains occurring in the metamorphic host rocks range. in size from sub-micron to a few microns in diameter.

Jan 1, 1987

By R. Schuhmann

The thermodynamic data and calculations which underlie the design of the Q-S process for coppermaking are summarized. The efficient production of crude copper and slag of low copper content in a single con¬tinuous converter requires the establishment and control of gradients in oxygen activity and other physico-chemical variables throughout the reactor. These gradients are established by stagewise introduction of oxygen and solid feed materials to provide: (a) flash oxidation of sulfides above the bath, utilizing the great bulk of the oxygen, (b) conversion of high grade matte or white metal to crude copper by bottom blowing with restricted bath turbulence, (c) close approach to equilibrium staging of converting and deconverting reactions along the length of the slag layer, from a maximum oxygen activity near the metal discharge end to a minimum at the slag discharge end, and (d) completion of slag cleansing by deconverting. In terms of the Phase Rule, the Q-S process involves a system with a minimum of 5 components (Cu, Fe, S, O, and SiO2) and several liquid, solid, and gaseous phases. Accordingly, basic data from previously reported studies on simpler systems are utilized to prepare an equilibrium diagram for the 5-component system and to generate a quantitative thermodynamic model for the Q-S process. This model shows in particular the ranges of oxygen activities (measured by equivalent CO2/CO ratios) involved in the Q-S process and the relations of oxygen activity to temperature and other important physico-chemical variables of coppermaking.

Jan 1, 1976

By J. E. Dutrizac

A mineralogical study of a weak acid leach residue from the Clarksville zinc plant of Savage Zinc Inc. was carried out. The study focused on Pb and Cd as these elements most significantly impact on the environmentally acceptable disposal of the residue. The residue consists of major amounts of bassanite, zinc ferrite and various Zn-Pb-Fe-Cd silicate species. Also present are lesser quantities of PbSO4, sphalerite, Fe2O3 quartz, Mn-Pb-Fe oxides, SrS04, (Pb,Sr)SO4 iron gel-silica gel, jarosite and Ti oxide. The principal Cd carriers are the zinc ferrite, which contains up to 9.00% Cd and likely accounts for -45% of the total Cd content of the residue, and the various silicate species which contain up to 6.97% Cd and account for about 45 % of the total Cd. The sphalerite is a minor Cd carrier containing an average of 0.35 % Cd and accounting for -5 % of the total Cd; the other carriers contribute only trace amounts to the total Cd content. The major Pb carriers are PbS04, Pb-bearing silicates and Mn-Pb-Fe oxides. Minor proportions of the total 1.09% Pb content of the residue are carried by SrS04 and (pb,Sr)S04' Trace Ph carriers include the iron gel-silica gel, Fe-Zn-AI oxide and bassanite. The residue after the toxicity characteristic leaching procedure (TCLP) is mineralogically very similar to the initial weak acid leach residue; the amounts of calcium sulphate are approximately the same in the two products. The principal Cd and Pb carriers are also the same as those noted in the initial residue, except that some dissolution and reprecipitation of Pb-Sr sulphate takes place. That the residue fails the TCLP test is likely a consequence of the superficial dissolution of Pb and Cd from species such as PbS04 or Zn-Pb-Fe-Cd silicate.

Jan 1, 1995

By Robert D. Rogers

The demand for phosphate will continue into the future, because phosphate is currently irreplaceable as a plant nutrient and in many chemical applications. Phosphate is not recycled, so the supply must be replenished through a process of mining and purification. Numerous microbial isolates. obtained from the environment were screened find those which demonstrated the greatest promise of solubilizing PO4 from the ore of interest. Screening was accomplished in a step wise manner an agar plate bioassay, shake flask studies, and semi continuous growth conditions. It was found that with the proper conditions of nutrient supply and solution pH; the bacteria and. fungus used in the semicontinuous process were able to solubilize in excess of 80% of an Idaho rock phosphate ore (RP) within nine days. A significant amount of solubilization occurred within the first six days. There was a 'reasonable correlation between the occurrence of soluble Ca and PO.4 during the processing of the RP thus providing evidence that the RP was being disintegrated by the organisms. In all cases, HPLC analysis of the solutions in which PO4 had been. extracted from RP showed that organic acids, which have been implicated as the mechanism of solubilization, were present.

Jan 1, 1991

By Thomas R. Wildeman

Experimental methods for studying vapors over solid minerals are reviewed and an especially developed method using atomic absorption is described. The method is used to study vapor pressures over Te. Ag, Au, Ag2Te, and AuTe2 from 250°C through 10S0oe. Vapor pressures observed and, concurrently. sintering properties are significantly altered by the form of the solid. The vapor pressures of the tellurium compounds reveal complex processes occurring on the surfaces of the solids. The data given also point out a method of recovering precious metals from their telluride minerals w

Jan 1, 1987

By Djona M. Gaballah I.

Selective chlorination of unroasted and roasted spent hydrorefming catalysts was studied between 300 and 600°C using the following gas mixtures : C12+CO, C12+N2 and CI2+air. Such materials contain Co, Ni and Mo compounds supported by alumina and/ or silica. They also contain variable amounts of V compounds, C, S, etc., accumulated during the life time of the catalyst. Direct oxychlorination of the C+S rich sample allows higher extraction rates than those obtained by the carbochlorination of the same roasted one. It seems that the roasting step doesn't favor the economic metals' recovery. Results of the chlorination of C+S poor spent catalysts, in the presence of CO, in the chlorinating gas mixture, or of carbon added to the sample, were comparable. Cobalt and nickel chlorides were .extracted from the chlorination residues by their dissolution with acidified water, whereas molybdenum and vanadium chlorinated compounds were recovered by selective condensation from the vapor phase. The recovery rates of valuable metals as well as the selectivity were higher for the chlorination of spent catalysts than results obtained by the hydrometallurgical processing of the same sample. Results of the scale-up of the oxychlorination of the unroasted sample confirmed those of the laboratory scale. The global recovery rates were about 88 pct of cobalt, 93 pct of molybdenum and 76 pct of vanadium. The carbochlorination was applied to the recovery of W and Ti from Denox catalyst that is a carbon and sulfur free one. Recovery rates higher than 99 pct at temperatures lower than 500 °C could be obtained. This may indicate that the chlorination process can be applied to the treatment of various industrial wastes or by-product containing transition and/or refractory metal compounds A flow - sheet for the pretreatment of these materials is proposed.

Jan 1, 1995

By M. Jortikka

The smelting process of the old reverberatory furnace including its capacity has been improved over the years in different ways eg. with oxy-fuel burners, parallel reactors with the reverb or with various modifications of the reverb furnace itself. The future reverb modifications with old concepts are difficult or impossible as the increased copper losses in slag call for separate slag treatment and particularly the process modifications do not solve the environmental issues but instead mean unavoidable high capital and operating costs in gas cleaning and sulphur recovery. During the last decades improvement in process performance and increase in capacity of copper flash smelters has been achieved easily by substituting the pre-heated process air by highly oxygen enriched air. The flash furnaces of today, however, are inherently designed to operate with high oxygen enrichment. Further process improvement with a view to production costs and capacity increase as well as to the more stringent environmental requirements is possible with even higher oxygen enrichment. This, apart from the advantages it generates, requires improvements in and careful design of the feed system, concentrate burner, furnace cooling and the off-gas line as well as adaption of the flash converting instead of the conventional Peirce-Smith converters. A logical answer even for the revamping of the reverb smelters in the future is the total replacement of the reverb furnaces with flash smelting together with flash converting. This paper discusses the possibilities for future process improvements of the existing flash smelters taking into account the process concepts, equipment modifications and capital aspects not forgetting the environmental issues. The revamping of the PASAR smelter in Philippines is introduced as an example of a recent success story.

Jan 1, 1994

By Chu Yong Cheng

Manganese is one of the main impurities in the recovery of base metals from their ores. This is especially true when laterite ores are processed using the pressure acid leach (PAL) process. The use of intermediate precipitation, solids/liquid separation and re-leach in the three Western Australian nickel laterite plants to separate nickel and cobalt from manganese and other impurities makes these processes complicated and costly in capital and operation. The research work carried out by the solvent extraction (SX) group at the A J Parker Cooperative Research Centre for Hydrometallurgy (CSIRO Minerals) has led to the development of direct solvent extraction processes to separate nickel and cobalt from manganese and other impurities. The simplicity of the process flowsheets and the expected savings in capital and operating costs are the major advantages of the new direct solvent extraction processes over the existing processes. By using a new synergistic organic system, Versatic 10 / dccyl-4-pyridinccarboxylatc ester, with a pilot plant leach solution from BHP¬Billiton (after iron precipitation), the metal values (Ni and Co) together with zinc and copper were separated from manganese and other major impurities (Mg, Ca and Cl) in the first SX circuit. The co-extracted manganese, magnesium and calcium were easily scrubbed out. The extraction and stripping kinetics of the metals with the new synergistic organic system were very fast and the synergistic system was found to be chemically stable. pH-extraction isotherms of metals with D2EHPA showed that the extraction order as a function of pH50 was Zn2+ > Ca2+ > Mn2+ > Cu2+ > Co2+> Ni2+> Mg2+. Semi-continuous test work with a cobalt pilot plant leach solution from Peko Rehabilitation Project Pty Ltd and a synthetic solution to simulate a concentrated laterite leach solution from BHP-Billiton showed that manganese, together with calcium, copper and zinc can be separated from nickel, cobalt and magnesium. This led to the development of another direct solvent extraction process to separate manganese and other impurities from nickel and cobalt and to recover nickel and cobalt from leach solutions. A project was initiated by HiTec Energy Ltd to recover manganese from waste laterite leach solutions using SX technology and to further purify it to produce high purity electrolytic manganese dioxide (EMD) for the high quality alkaline battery industry. Batch tests were conducted to select a suitable organic system for the extraction of manganese in terms of selectivity and extraction and stripping kinetics. A total of eight organic systems were tested using waste laterite processing solutions. A Versatie 10 / synergist system showed very promising results. By using the synergistic system, high manganese recovery would be obtained, fewer SX stages would be required and a higher A:O ratio could be used in extraction, resulting in a smaller organic inventory and lower capital costs. The risk of gypsum formation is also minimised with the synergistic system. Keywords: Solvent Extraction, Manganese, Laterite, Synergist

Jan 1, 2004

By G. A. Irons, D. Colbert

Several electric arc furnace (EAF) shops have increased the zinc and lead grades of EAF dust by reinjecting it into their furnaces; however, the kinetics of this operation are poorly understood. To understand the kinetics, EAF dust was injected into 70 kg heats of iron or steel in an induction furnace. Supplemental additions of carbon and oxygen to overcome the heat load on the system were also investigated. In all cases, the dust was chemically transformed, and the zinc content increased from 16% Zn to 60 - 80% Zn. About 50% of the zinc in the EAF dust dissolved in the iron or steel. The zinc oxide in the dust is thought to be reduced by carbon and the zinc vapour subsequently dissolved in the iron. A kinetic model for the dissolution process is proposed, and the implications for EAF dust recycling are discussed.

Jan 1, 2000

By Zhang Libo, Li Wei, Peng Jinhui, Guo Shenghui, L. Jing

Partially stabilized zirconia ceramics (PSZ) were prepared by microwave sintering using natural baddeleyite which was obtained by flotation of baddeleyite ore as starting materials. Because of using natural baddeleyite as raw and the process producing no waste, it realized shorter process, energy savmg, environmentally-friendly and non-damage processing. In this work natural baddeleyite stabilized with 3.8wt.% CaO were sintered at 1300°C for 60min in a multi-mode 2.45GHz microwave furnace. Microwave sintering could fast heating and sintering. Besides that, microwave sintering could heat materials wholly, low the sintering temperature, improve the materials properties and save energy. X-ray spectrum displayed the product composed by the tetragonal phase and monoclinic phase that also showed the product was PSZ. Fine and uniform microstructures were observed in the product through scanning electron microscopy. PSZ which was with high density was confirmed by the method of the Archimedes technique.

Jan 1, 2011

By Qiu Dingfan

The Slurry Electrowinning Process (SEP) is a new hydrometallurgical method and clean technology, Sulfide ores can be treated by the SEP without S02 emission and the sulfur from the sulfides can be recovered in element form which is easy to stockpile and can be transported at low cost. In a special slurry electrowinning cell one can achieve at the same time the leaching of elements from the feed and the electrowinning of metals from the electrolyte, Because the anode reactions are utilized for leaching of some elements the consumption of the energy is decreased, SEP also has the advantage of separation some elements, It is especially suitable for treatment of complex concentrates containing Pb, Cu, Au, Ag and S, which could hardly be accepted by any smelter. In this work, the behavior of different metals in this process has been studied and it was found that the separation of metals in some minerals is possible through the SEP, Beijing General Research Institute of Mining and Metallurgy (BGRIMM) has studied the possibility of the SEP to treat bismuth concentrates containing S, Be and F in laboratory and pilot-test scale as well as in commercial plants, The commercial plant was set up in 1997, It has been in operation since that time, According to the results of laboratory experiments and pilot- plant test, a new plant to treat complex gold minerals was set up in 1998 and is in operation in 1999

Jan 1, 2003

By Michael Sheedy

Recoflo® ion exchange technology utilizes fine mesh resin beads, a fully packed resin bed, and counter-current regeneration. These unique features help to improve exchange kinetics, reduce regenerant consumption, and increase the concentration of strip solutions. Commercially, Recoflo technology has been extensively used in the following areas: the separation of strong mineral acids from dissolved metal salts, the purification, removal and recycle of dissolved metals in the surface finishing industries, and the production of high purity water. In the mining and metallurgical industries Recoflo ion exchange systems have been installed to treat a copper refinery bleed stream at Falconbridge's Kidd Creek plant and to treat a nickel and cobalt bearing ammonium sulfate effluent stream at another primary metal producer's facility. In the first application an APU® Recoflo system is used to separate the sulfuric acid at 200g/L from the dissolved copper and nickel sulfate salts at 80 g/L. The absorbed sulfuric acid is recovered from the resin by elution with water and is then returned to the refinery electrolyte circuit. In the second application a chelating ion exchange resin is used to remove nickel and cobalt at 0.01 - 3.0 g/L from process streams containing 120 - 240 g/L ammonium sulfate. This Recoflo system has been designed to reduce the nickel and cobalt levels to 0.001 - 0.002 g/L while recovering cobalt and nickel at concentrations up to 30 g/L and 15 g/L respectively. Case studies for both these metallurgical applications and other Recoflo processes are presented in this paper.

Jan 1, 1998

By G. Mawire

The extraction of scandium from its oxide with a task specific ionic liquid [Hbet][Tf2N] offers potential advantages of lower energy consumption, recycling of the extractant and reduced environmental impact compared to conventional methods. In this paper, the influence of reaction temperature, solid to liquid ratio and water to the ionic liquid ratio on the extraction of scandium from a synthetic scandium oxide (Sc2O3) system was studied. The influence of investigated parameters on leaching was very significant as approximately 98% scandium was extracted from its oxide after 72 h. Furthermore, the ionic liquid also selectively extracted 47% scandium after 48 h from a mixture of pure metal oxides containing oxides of tantalum, niobium, and scandium. The research demonstrated the potential of [Hbet][Tf2N] to selectively extract scandium from columbite processing tailings and other scandium bearing ores like Ixiolite, Heftetjernite and Tantalite which predominately contain tantalum and niobium.

By Liang Li

Thermodynamic analysis and experiments were conducted in order to verify the feasibility of preparing crude titanium tetrachloride (TiCl4) via the carbochlorination of low-grade titanium slag in molten salt. Titanium slag, assaying 74.6 wt% TiO2 with high calcium and magnesium oxide impurities, was treated by an optimized carbochlorination process in NaCl molten salt. These impurities in the titanium slag were chloridized simultaneously, and chlorination products FeCl2, MnCl2, MgCl2, CaCl2 and CrCl3 were collected in the furnace slag. XRD and SEM/ EDS analysis of residue shown that SiO2 and Al2O3 in titanium slag were difficult to chlorinate completely. Theoretical calculations and industrial-scale experimental studies reveal the content of TiCl4 in the products was more than 98.8 wt% and thus proved the feasibility of utilizing low-grade titanium slag for TiCl4 preparation by molten salt chlorination technology.

By U. Lauwers

The influence of different waste materials in sinter blends on the quality and mineralogical composition of a sinter is described. As a result of partial preservation of the original textures of waste materials in the final product the degree of assimilation of the original material in the sinter product can be ascertained. Recycling of waste materials such as dolo char , arc-furnace slag, scrap fmes, mill scale and BOF-mud via the sintering process a saving on costs of raw materials such as flux and fuels, and a reduction on pollution can be achieved. A comparison of a typical sinter with a sinter produced through the utilization of waste materials showed a distinct correlation between certain types of waste material and some metallurgical properties. Dolochar in a sinter blend acts as part of the fuel and flux that are normally added to a sinter mix. Generally dolochar improves the quality of the sinter and decreases the consumption of coke. Other waste materials adversely influence the quality of the sinter. The glass and calcium ferrite content of a sinter when dolochar is used in the blend is similar to that of a sinter produced with only coke as a fuel in the sinter mix. The glass content is lower and the calcium ferrite content is higher, in a sinter produced from dolo char in the blend compared to a sinter produced with additions of other waste materials. Changes in the mineralogical composition are related directly to the metallurgical properties of the sinter product.

Jan 1, 1991