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By H R. Pein

"The Corridor Sands deposit is situated approximately 190 km north east of Maputo, Mozambique adjacent to the paleo-mouth of the Limpopo River. This deposit is one of the worldÆs largest known economically viable resource of titanium, accounting for approximately one third of the global resource. As part of the pre-feasibility and feasibility study for development of the Corridor Sands project, various geological and geotechnical investigations were undertaken, selected results of which are presented in this paper. These investigations have enabled a detailed understanding of the material characteristics of the orebody as well as the tailings materials, such as the fines, to be developed. These material characteristics have had a significant impact on the selection of the mining and disposal methods currently being implemented."

Jan 1, 2001

By Boris Albijanic, Hongjuan Chen, Bogale Tadesse

"During mechanical flotation of fine liberation sized complex rare earth ores, a large fraction of ultrafine valuable mineral particles are lost to tailings. Fine gangue mineral particles can be entrapped within the froth phase and be floated together with target mineral particles. Another common challenge in ultrafine rare earth mineral flotation is the formation of a sticky froth (excessively stable froth). Thus, the improvement of the flotation of fine and ultrafine rare earth bearing minerals through appropriate cell design and reagent selection is highly desirable. The aim of this study is to investigate the performance of a pico-bubble cavitation flotation column in improving selective recovery of ultrafine rare-earth elements (REEs) bearing minerals from complex and low-grade ores. In pico-bubble cavitation column, very small bubbles are generated which are capable of attaching to ultrafine particles and consequently modifying their surface properties. The overall effect of these processes should lead to improved flotation selectivity and recovery of ultrafine RE mineral particles. In this paper, the effects of a number of flotation operating variables on recovery and grade of REE minerals was investigated using a continuous pilot scale pico-bubble column flotation. At the flotation conditions used in this study, subtle monazite upgrade was observed while gangue Fe oxide was preferentially floated even in the presence of depressants. INTRODUCTION Rare earth elements (REEs) have unique properties such as superconductivity and catalytic activity, and are in high demand for a large number of cutting-edge technologies in the modern world including the production of mobile phones, fuel cells, displays, permanent magnets for wind power generation, high-capacity batteries, green energy devices, etc (Naumov, 2008). REEs tend to be intimately associated within many minerals, and are rarely found in high concentrations in deposits for economic extraction. Up to now, among over 250 minerals containing REEs, bastnasite (Ce), monazite (Ce, Th and La) and xenotime (Y) are the three main minerals that contain sufficient grade of REEs for commercial exploitation (Zhang and Edwards, 2012). The deposits that contain bastnäsite, monazite and xenotime usually have complicated mineral compositions, with various valuable minerals of fine embedded grain sizes in association with complex gangue minerals. In order to ensure liberation of the REEs minerals, fine grinding to less than 30 µm is necessary. This fine grinding leads to difficulty in recovering a large proportion of fine and ultrafine particles in separation processes such as flotation (Trahar and Warren, 1976). Consequently, highly liberated and ultrafine fine valuable REEs particles are lost into tailings during traditional mechanical flotation. Besides, the formation of extremely sticky and stable froth inhibits the drainage of hydrophobic particles. Thus, improving flotation of rare earth bearing minerals, particularly the ultrafine size range, is highly desirable."

Jan 1, 2018

By E. H. Mohr

AT the McGill concentrator of the Nevada Consolidated Copper Co., all flotation operations have been carried out in Forrester cells since November, 1926. In respect to cost of operation, the new cell has proved to be superior to the Callow, which it supplanted, and its incorporation in the flow sheet has resulted in a marked improvement in the metallurgy. Before citing the advantages of the new machine over the old, it should be explained that the Forrester has not only replaced the Callow, but it has made possible the elimination of a number of tables as well. When Callow cells were in use, table concentration preceded flotation. The ball mill pulp was elevated, by means of air lifts, to roughing Wilfley tables situated two floors directly above the flotation floor. On an intermediate floor were additional tables which cleaned the concentrate from those above. The finished table concentrate constituted approximately 60 per cent. of the total mill production, and contained about 40 per cent. of the recovered copper. The table tailings were returned to the classifiers, the overflow from which went to the flotation cells and the oversize to the ball mills. As soon as the Forrester cells were installed, this tabling operation was discontinued. The ball mill pulp now discharges into the classifiers direct, without the intermediate tabling. It was impracticable to employ the present flow sheet when Callow cells were in use, for the reason that unless the greater part of the heavy mineral was first removed, it accumulated on the mats and adversely affected the flotation operation. By tabling ahead of flotation, it was practicable to maintain the concentrator at 10 to 20 per cent. greater capacity (depending upon the pyrite content of the ore) than was possible by flotation alone. This was due to the necessity for finer grinding in the latter case. Even though the table concentrates contained only 6 to 8 per cent. copper content, the low smelting costs made it advantageous to operate at high mill capacity, at the cost of low grade of product.

Jan 1, 1927

By S B. Stewart, C J. Restarick

Over a four-hour period of steady plant operation samples were taken at regular intervals frum points within the grinding and flotation circuits. Size fractions of each composite sample were analysed for lead and zinc values and a comparison made of the behaviour of the lead, zinc and gangue minerals, size by size throughout the whole circuit. Analysis of the breakage behaviour using dynamic mathematical models gave the rates of breakage of each size fraction of each main component. Simulated grinding experiments shaved that the main grinding circuit had some reserve capacity. In the lead flotation circuit all lead values between 150 mesh (Tyler) and 5 ¦m had excellent flotation characteristics, and a major proportion had a high rate of flotation which was practically the same for all sizes. Increase in overall lead recovery would require improvement of the flotation behaviour of ultra-fine lead minerals. These were slav floating and were responsible for approximately half the lead losses in the final zinc concentrate and one thiro of the lead losses in the final tails. All sizes of zinc values between 150 mesh and 7 ¦m floated well and the performance of the zinc flotation circuit was mainly limited by the relatively poor flotation of the very fine zinc values. Approximately 60 per cent of the zinc lost to the final tails lay within the -7 ¦m fraction. The overall flotation conditions were conducive to very good selective flotation of the major proportion of the values and limitations in metallurgical performance were mainly due to the properties of the ultra-fines.

Jan 1, 1971

By AIME

Burton J. Westman-Besides decreasing the diamond size, there appear to be two other approaches open to overcome excessive diamond loss and, more particularly, the rapid diamond polish that took place in this quartz-feldspar rock. First, however, a short discussion of the diamond characteristics should be made in order to discuss the situation more clearly. The West African diamonds which we can assume were used. for this drilling have two disadvantages when drilling hard, exceptionally fine-grained rocks. These are rapid polishing and, if .the bit isn't "drilled in" carefully, a tendency to damage, resulting in excessive diamond loss. The normal wear resistance of any diamond is inversely proportional to the compressive strength of the diamond, therefore only the highest grade, of diamonds free from insipient fractures should be employed. Using upwards of 90 diamonds per carat instead of the20 to 30 per carat not only decreases the total projected area of the cutting points but increases the stress per point which, in turn, causes greater depth of penetration into the rock. Considering this depth of penetration, therefore, suggests an approach other than increasing the, number of diamonds per carat and that is to decrease the actual number of diamond points (also the carat weight per bit) so that the penetrating force on an individual diamond is increased. This penetration increases the chipping or rock shearing action.

Jan 1, 1950

By S. Thibeault, P. Filiatreault, B. Straight, D. Castendyk, L. Cameron

"Unmanned aerial vehicles (UAV), or drones, capable of collecting and retrieving water samples from pit lakes, tailings ponds and other mine water basins will revolutionize environmental programs by lowering sampling costs and reducing safety related risks. Costs typically include a boat and associated equipment, boat transportation, road maintenance, water safety devices, water safety training, and additional staff. Typically, the total cost can exceed $20,000 USD per sampling event (excluding lab analyses). Hazards include potential for drowning, unstable pit walls, and unconsolidated sediment surrounding tailings ponds. Hatch has developed a novel device for aerial drones which currently enables the collection of water samples from depths up to 390 ft (119 m). This system only requires a pilot, a sample technician, and a safe operations point overlooking the water body. Hatch anticipates this will reduce monitoring cost by at least 50% while improving safety, reducing staffing requirements and accelerating data acquisition. This manuscript demonstrates profiling and sample collection from a 95 ft (29 m) deep reservoir and a 377 ft (115 m) deep pit lake, and highlights the potential of this emerging technology. INTRODUCTION Working on or near water at active, closed, or abandoned mine sites is inherently dangerous. Between 1996 and 1999, the United States Mine Safety and Health Administration (MSHA) reported 17 fatalities involving drowning at US metal, non-metal, and coal mines (MSHA, 2000). Open pit mines which fill with water following mine closure, called pit lakes, present numerous risks to environmental monitoring teams including drowning, hypothermia, falls to the ground from the top of high walls, and rock avalanches resulting from steep, unstable pit walls. Some pit lakes are simply not easily accessible for ongoing monitoring. However, the routine collection of water samples from the entire water column on at least an annual basis is required to monitor the evolution of water quality over time, to verify the accuracy of numerical predictions, and to manage lake water chemistry in order to produce “treatable” influent to water treatment plants, and ultimately, produce acceptable site discharge water quality. As a product of these risks and the need for water sample, mining companies spend considerable expense on pit lake safety as part of monitoring. This includes training personnel for “on water” duties and hazards, purchasing or hiring specialized equipment and personnel (such as boats, boat operators, personal floatation devices, and mustang suits), stabilizing pit walls and maintaining safe access roads to the lake surface, employing spotters with direct communication to emergency medical services to watch sampling activities (because cell phones and radio signals can be blocked by pit walls), and utilizing redundancy to prepare for unexpected incidents and accidents. For example, one Canadian mining company requires pit lake water samplers to hold Marine Emergency Duties (MED) training to ensure the entire boat crew is sufficient and competent for the safe operation of the vessel, and uses three boats for water sampling: two on the water at all times, and one on the shoreline, ready to deploy. The costs of these safety mechanisms greatly adds to the overall expense associated with pit lake sampling. Consequently, environmental programs may need to reduce the frequency of pit lake sampling in order to keep monitoring programs within an acceptable budget"

Jan 1, 2017

By C Smith, J Cilliers, K Hadler

Unstable flotation froths, that burst before they overflow, produce no concentrate. Increasing the froth stability to improve recovery is therefore an obvious control strategy. Froth stability can be measured readily as the fraction of air entering the flotation cell that overflows as froth, rather than bursting on the surface; the ‘air recovery.’ Air rate is one of the most significant variables affecting froth stability. It is interesting that froth stability (air recovery) is maximised at a specific air rate, above or below which it decreases. It is important for metallurgists to realise that this maximum froth stability also produces the maximum mineral recovery, without compromising concentrate grade. Flotation cells must be operated at the air rate that produces the highest air recovery. The peak air recovery (PAR) to maximum mineral recovery relationship was first demonstrated by Hadler and Cilliers in 2009. Since then, the relationship between PAR and flotation performance has been widely explored; varying air rate, froth depth and particle size. Industrially, reducing the flotation bank air rate to achieve PAR yielded better mineral recoveries than other down-the-bank air rate profiles. In this paper, we present an overview of these key developments and findings in flotation optimisation based on PAR. Furthermore, using new industrial data, we show that PAR is not just applicable to rougher-scavenger cells, but also to flotation columns. Finally, we consider the future application of such an optimising control concept. CITATION:Smith, C, Hadler, K and Cilliers, J, 2018. Flotation control using froth stability: past, present and future, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 441–454 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By Masaru Tomoguchi, Tom Nishiyama, Naoto Sasamoto

"In Hanaoka Mine, which was one of the major ""black ore"" (Kuroko) mines in the north part of Japan, mining has finished in 1994, and only mineral processing plant is available now. Recently this mineral processing plant has been applied to the soil washing. Soils contaminated with heavy metals are excavated and transported to the plant. The soils are separated to cleaned soil and metal enrichment through the process which consists of scrubbing, classification, extraction, insolubilization, flotation, and so on. The process is flexibly designed for each contaminated soil, in order to the products can be recycled as cleaned soils or metal sources. . Soils contaminated with arsenic and lead were treated in 1999, which were entrusted from Environmental Agency of Japan.IntroductionDowa Mining has had the mine site named ""Hanaoka Mine"" and ""Kosaka Mine"" in the north part of Japan. Those mines consisted of some ore deposits and produced ""black ore"" (Kuroko) and ""yellow ore"" (Oko), which are complex sulfide ore. The ""black ore"" mainly consists of sphalerite, galena and barite with accessory chalcopyrite, tetrahedrite and Au, Ag minerals. The ""yellow ore"" mainly consists of powdery and clayey ore containing chalcopyrite and. pyrite, with a little sphalerite and galena. Each mine site had mineral processing plant respectively, ""Hanaoka Mine"" had two and ""Kosaka Mine"" had one plant. Those plants were operated by flotation process to separate each mineral, and produced concentrates of copper, lead, zinc, pyrite, barite and so on.Unfortunately, the mine had closed one by one because of change in economical circumstances and the lack of ore reserves. Mill plants were also closed and dismantled, but the last ""Matsumine mill plant"" of Hanaoka Mine was remained to be utilized usefully. In the circumstance like that, we came across with soil remediation business. Recent reports say that subsurface contamination in Japan is being grown substantially and that it can be a serious matter beyond the imagination of our society. So nowadays we have a tendency to pay attention to the environmental issues including subsurface one."

Jan 1, 2000

By S Servant

In the mine of this study, cemented hydraulic backfill is placed under the form of a slurry at a percentage of solid by weight varying from 50 to 65 per cent. This slurry settles and cures in the stope. During this process an important amount of water must be drained from the stope. Field observations in the case studied here show that the geometry of the drains system has a significant impact on the drainage of the stopes. The mode of filling produces a backfill showing an evident stratification typical of a sedimentation process. In situ observations on this material show that this stratification translates into a very strong heterogeneity in the physical properties of the cemented backfill. This heterogeneity should affect the hydraulic conductivity of the backfill in the stope. In order to study this aspect, finite element simulations were performed. This paper presents the results of these simulations, where the impact of this stratification on hydraulic conductivity is related to the performance of the backfill drainage system for various geometry. This preliminary study is aimed at providing a better understanding of the role of the drains in the dewatering process of the backfilled stope in the case of this mine.

Jan 1, 1998

By S. P. Holt, Robert S. Sanford

"INTRODUCTION The wartime demand for iron ore and alloy steel on the west coast revived interest in iron reserves in western United States and Alaska. Magnetite was known to have been the most abundant metallic mineral at some of the old copper mines on Prince of Wales Island. In the summer of 1942, the Bureau of Mines began a program of preliminary investigation followed by core-drilling two magnetite deposits on the island.The Poor Man iron deposit on Kasaan Peninsula, Prince of Wales Island, Southeastern Alaska, was examined in August 1942 by Robert L. Thorne, associate mining engineer of the Bureau of Mines, and John C. Reed and B. N. Goddard, geologists of the Federal Geological Survey in September 1942 the area was mapped geologically and a magnetic survey was made by the Geological Survey. In October and November 1942, a more thorough preliminary examination, which included cleaning out and sampling old adits and trenches, was made by thesenior author, Robert L. Thorne, and Aner W. Erickson, engineers of the Bureau. J. G. Shepherd, a consultant of the Bureau, spent two weeks in March and April 1943 investigating the Poor Man and other deposits on Kasaan Peninsula. Core drilling was completed in June 1943.L. A. Warner, of the Geological Survey, examined all cores and interpreted structural data revealed by the drilling."

Oct 1, 1946

By S. R. Hutchins

Several refractory sulfide and carbonaceous gold ores were examined to determine whether microbial pretreatment could be used to enhance gold recovery. Test samples were biologically leached using unique. thermophilic microorganisms to solubilize iron sulfide minerals. Gold recovery from sulfide ores and concentrates was enhanced following pretreatment with the thermophiles; responses varied widely for the individual materials. Although bioleaching had little effect on the preg-robbing characteristics of the carbonaceous gold ores, cyanidation of the bioleached residues yielded higher gold recoveries than cyanidation of the native ore. By combining bioleaching with abiotic processes designed to inhibit preg robbing, gold recoveries of 83-98% were obtained from carbonaceous ores yielding 10-50% gold recovery by direct cyanidation.

Jan 1, 1987

Gold recovery from refractory ores can be very difficult and complicated depending on the causes of the refractoriness. Typically, a fine wind and/or the oxidation of the gold-bearing sulfide minerals are required to liberate gold from the host rock matrix. Mineral separation techniques, particularly surface¬wettability based separation (i.e., flotation) and magnetic separation, can be used to pre-concentrate the gold and sulfide minerals, or to remove mineral species that are harmful either to the oxidation of the gold-bearing sulfides or to the cyanidation following the oxidation. In this paper, several processes are described, including oil-agglomeration flotation recovery of fine-grained gold and sulfide, reverse flotation of carbonate minerals prior to pressure or bio oxidation, reverse flotation of elemental sulfur after pressure or bio oxidation but prior to cyanide leaching, and magnetic concentration of gold-bearing iron oxides. The integration of these techniques with gold hydrometallurgical processes may make the economic recovery of gold from an otherwise uneconomic refractory ore possible.

Jan 1, 2011

By T. Karlkvist, H. R. Kota, F. Andreas, B. I. Pålsson

"Separation of different calcium minerals have long been an interesting and challenging problem. In this investigation calcium mineral separation is examined by: microflotation, zeta potential measurement and adsorption, using novel collectors having two functional groups instead of one. In theory, by varying the distance between the functional groups, it could be possible to preferentially target one calcium mineral by matching the spatial distance between the sites on the mineral surface. In this investigation two new surfactants have been tested to estimate their ability to float apatite and/or calcite. Preliminary findings show that an increase in distance between the functional groups favors one mineral over the other, and this might be due to differences in the mineral surface structure. INTRODUCTION Apatite is a strategically important mineral which has both great value for industrial use as a precursor to phosphoric acid; and in the agricultural industry as an important source of phosphorous in fertilizers. This has led both the European Union and the United States to recognize it as an important resource for the future. Currently there is no lack of apatite in the world, but it is important to develop the beneficiation to ensure future supply. A common problem in apatite flotation is the selectivity towards calcite, which it is commonly associated with. Although calcite is an important industrial mineral its abundance makes it not usually worth to extract from apatite ores unless it is easy to separate1,2. Both apatite and calcite are semi soluble minerals with calcium as one of their constituents. The most common way to separate apatite and calcite is to use flotation. A process that involves making the mineral particles to extract from the system hydrophobic by adsorbing a collector on the mineral surface. One way to study collector adsorption on a mineral surface is by zeta potential measurements. The zeta potential is the charge at a slip plane a short distance from the surface. By measuring the zeta potential with and without the presence of collector it is possible to investigate if the collector is present at the surface or not, a change is indicative of collector adsorption. The zeta potential of fluorite, apatite and calcite without collector can be seen in Figure 1. The small difference in zeta potential between apatite and calcite shows the similarity in their surface properties and gives an indication of their separation difficulty. One additional problem with semi soluble minerals is their tendency to slightly dissolve in the pulp and thereby change the ionic composition of the solution. This makes separation and research problematic as the freed ions may readsorb on the mineral surface and interfere with the collector adsorption on the mineral surfaces3–5."

Jan 1, 2016

By R. F. Coulter

THE Flin Flon concentrating and metallurgical plants (Figure 1) of the Hudson Bay Mining and Smelting Co., Limited, lie adjacent to the town of Flin Flon, on the Manitoba-Saskatchewan border, 400 miles northwest of Winnipeg. Since the start of milling operations in 1930, over 50 million tons of ore from the Flin Flon mine have been treated in the concentrator. During the past few years, newly discovered orebodies have been brought into production and their ores processed at the company's main metallurgical plants. All these properties are within a 75-mile radius of Flin Flon and have aided substantially in augmenting plant metal production as well as retarding depletion of ore reserves in the Flin Flon mine. •Current satellite producers are the Schist Lake, Coronation, and Chisel Lake mines. The Schist Lake mine, located 3 ½ miles southeast of the main plant, is a copper-zinc producer. It was brought into production in 1954. The

Jan 1, 1962

By P Vass, C Butler, J Condon, B Hooper

DeGrussa Copper Mine, resource status 10.2 Mt at 5.7 per cent Cu, 2.1 g/t Au (Sandfire Resources, 2013), commenced concentrate production through its 1.5 Mt/a conventional copper flotation processing plant in September 2012. Geometallurgical assessment during the feasibility study identified potential process-critical ore factors that required further investigation to determine potential impact on processing. Ore characterisation and metallurgical processing performance test work of ore types ensued and subsequently, the first quantitative DeGrussa geometallurgy model was born. Now one year into steady state production, the geometallurgical program has evolved significantly from its infancy and provides valuable insight into ore characterisation and behaviour in the processing plant. This paper details geometallurgical investigations undertaken at DeGrussa to date; from feasibility metallurgical test work through to geometallurgical modelling and management practices, as well as reconciliation between observations at feasibility study level and subsequent observations in production.CITATION:Condon, J, Butler, C, Hooper, B and Vass, P, 2013. DeGrussa Copper Mine, Western Australia - feasibility to first production: a case study in developing a working geometallurgy model from first principles, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 59-66 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

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Jan 1, 2011

By Lj. D. Andric

Development of modern devices for mechanical and mechanochemical materials activation has been dated since the beginning of last century. Rapid development of high-energy mills that are different in type, size, design and performances, has been occurred during the last four decades. Because of complete understanding of mechanical processing, entire knowledge of theoretical principles of high-energy mills is necessary. These principles are main causales of process that are happened in materials during the mechanical activation. This paper is occupied by theoretical principles of operation of high-energy mills. Conventional high energy mills with balls, high-energy planetary mills and high-energy vibrational mills are particularly considered in the paper.

Jan 1, 2014

By K. E. Waters, J. E. Nesset, F. Rosenblum, R. Langlois, J. A. Finch

"Some sulphide mineral mixtures when exposed to air and a warm and humid environment can exhibit a self-generating temperature rise, referred to as self-heating or spontaneous heating. The observed self-heating derives from the oxidation reactions of the active sulphides. In previous publications by the authors, it was proposed that the self-heating of sulphides proceed in three distinct stages; namely, Stage A, Stage B and Stage C. In Stage A (<100OC) the sulphides, in the presence of moisture are oxidized, generating heat through a series of sequential reactions and forming elemental sulphur as one of the products. In Stage B (>100OC) the elemental sulphur produced in Stage A is further oxidized to generate additional heat and SO2. Stage C (>~380OC), sometimes referred to as thermal runaway or roasting, is not addressed in this paper. This communication focuses on the type and role of elemental sulphur found in Stages A and B, and the sequence of reactions that is occurring in Stage A. The sequence is divided into two phases, the first based on galvanic interaction between the sulphide minerals that generates an acidic environment but little heat, and the second involving the production of H2S which rapidly oxidizes to produce heat and elemental sulphur. To date, data suggest that the primary oxidation product formed in Stage A is the stable orthorhombic a-S8 form of elemental sulphur. Above 96OC the a-S8 transforms to a monoclinic ß-S8 structure which is less stable and will readily oxidize to generate heat and SO2. Above 159OC a proportion of the ß-S8 sulphur will transition to a polymeric Sµ form of sulphur which can readily volatilize. INTRODUCTION Some particulate and granular materials, when exposed to appropriate atmospheric and physical conditions, can exhibit a self-induced temperature rise resulting from accumulated heat due to internal exothermic oxidation reactions. This phenomenon is commonly referred to as self-heating or spontaneous heating. Under favorable physical and environmental conditions, these reactions can elevate the temperature of a given material from ambient to the point of auto-ignition. Self-heating behaviour should not be equated to pyrophoric behaviour which describes a group of materials (e.g. certain finely powdered metals) that will ignite spontaneously in air below 55OC. There are numerous materials such as coals, hay, forestry wood products, food products to name a few that can exhibit self-heating behaviour. This publication continues the focus of the authors on the self-heating of sulphide mineral mixtures. Many of the metals of commercial use found in nature are in the form of these sulphides which must be mined, crushed, ground and concentrated prior to smelting or hydrometallurgical processing in order to produce primary metal. The issues of sulphide self-heating therefore pertain to the mining and processing of metals."

Jan 1, 2018

By K. J. Heale, G. G. Richards, D. J. Rioux, R. L. Stephens

Slag fuming has remained a key process for the recovery of Zn, Pb, and minor elements in the integrated flow sheet at Teck Cominco&apos;s Trail Operations for more than 75 years. Since the start-up of the KIVCETTM process and the #3 slag fuming furnace in 1997, a range of process intensification projects have been completed to improve metals recoveries and throughput based on improved process control, mechanical modifications, and, most recently, the implementation of oxygen enrichment of the blast air. This latter project, initially named &apos;nitrogen depletion&apos; and now &apos;nitrogen oxygen performance enhancement&apos; (NOPE), is the primary focus of this paper. NOPE reduces the amount of heat required to heat the inert nitrogen gas, and therefore more of the energy input can be used for reduction. It also frees up boiler capacity because the inert nitrogen gas does not then have to be cooled down. A dynamic METSIMTM model was developed to determine the opportunities to improve the process capacity and heat balance to both assist with justifying the project and to verify the observed furnace behaviour. Several process safety management issues were also addressed and a few key minor mechanical modifications were made to allow the project to proceed to commercial implementation at 25% blast air oxygen enrichment. Short term plant trials have confirmed the model predictions. The first three months of a long term, continuous trial showed a 6% increase in zinc fuming rate from average furnace performance despite the implementation of a KIVCETTM fluxing change to assist with minor element recovery that was expected to negatively impact the zinc fuming rate by 3 to 5%. This encouraging result is also in line with the model predictions.

Jan 1, 2008

By B. Das

The processing of banded iron ore by direct anionic and reverse cationic flotation to produce pellet grade concentrate is investigated. This low-grade iron ore, known as banded hematite quartzite (BHQ) contains 41.8% Fe, 41.5% SiO2, 1.2% Al2O3, and 0.5% LOI. The flotation behavior of BHQ is studied by varying the effect of concentration of collector, frother, dispersant and pH. Oleic acid and dodecylamine are used as the anionic and cationic collectors respectively. The parameters that gave the best results in cell flotation are utilized in 100 mm diameter flotation column. All the results are compared with regard to the iron grade and recovery in concentrate. It is observed that, anionic collector with dispersant dosage of 220 g/ton, at pH ~8.5 is selective and yielded a high-grade iron concentrate with low silica as impurity. Also, it is possible to achieve an iron concentrate with 64% Fe at 30-37% weight recovery. The characterization studies of the BHQ ore using X-ray diffraction (XRD) technique and liberation analysis with QEMSCAN® are carried out. The sample contains 58.8% hematite and 37.7% quartz and the liberation of hematite particles are found to be 90.3% in -25 µm size. An alternative statistical modeling approach called response surface methodology was also undertaken to optimize the operating parameters in order to maximize the iron grade and recovery. The predicted results by the model are in agreement with the experimental values.

Sep 1, 2012