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By Finlayson EJ

New detailed mapping, micropalaeont- ology and isotopic dating have been successful in revising the stratigraphy of Woodlark Island. All Middle to Late Miocene intrusives and volcanic rocks have been assigned to the Okiduse Volcanics. In addition, the results demonstrate that the Tabukui beds and Wonai Hill beds (N9, late Early Miocene to earliest Middle Miocene) are a single unit. Similarly, the Nasai and Suloga Limestones are confirmed as a single stratigraphic unit as they yield similar foraminiferal ages (upper Te, Early Miocene), and both occur in the same stratigraphic position unconformably overlying basement Loluai Volcanics. Radiocarbon dating suggests a latest Pleistocene age for the upper Kiriwina Formation. Extensive Quaternary cover and low relief prevents detailed structural airphoto interpretation. In general, the structure is simple with all units on Woodlark dipping to the northwest a_4 exhibiting prominent local block faulting.

Jan 1, 1991

By M. F. Koehler

[THE LITERATURE contains relatively few thermodynamic values for interoxidic compounds, and these often are unreliable. This is especially true of data for ferrites and aluminates for which heat of formation determinations may involve extreme experimental difficulties. This report contains the results of new experimental work for obtaining the heats of formation of five ferrites (Ca2Fe2O5, CaFe2O4, MgFe2O4,, NaFeO2, and LiFeO2) and one aluminate (CaAl4O7), and a review and revision of previous similar work of this laboratory on five aluminates (Ca3A12O6,, 12/7 CaO-Al2O3, CaAl2O4, NaA1O2, and LiAlO2), The results for the 11 compounds are combined with entropy and high-temperature heat content data to derive tables of heats and free energies of formation (both from the oxides and from the elements) for the temperature range 298° 1,800° K, The relative stabilities of the compounds are discussed. The presented data all have application to ceramic reactions, and those for the calcium and sodium aluminates are applicable in considering soda-lime sintering processes for extracting alumina from silicate minerals such as day and labradorite.]

Jan 1, 1961

By A. G. Doll

"Many power-based grinding models exist, and most operators are familiar with Fred Bond's ""third theory"". Bond's model is most commonly used to describe primary and secondary grinding to product sizes above, for example, 100 µm. Operators sometimes use Bond's equation to describe grinding in situations where it is not appropriate, such as fine grinding below 50 µm. Using an alternative model would be a better choice in this situation.Bond's equation is one in a large family of models. Related equations better suited to fine grinding include the “signature plot”, Von Rittinger's model and Charles' equation. These models have a similar form to Bond's and can be fit to industrial regrind mills and laboratory tests using the simple regression tools in the charts of computer spreadsheets. Operators will find that fine grinding calculations are both easier and more accurate when using the alternative equations to fit their regrind milling surveys or when performing laboratory regrind mill scale-up for plant design.None of these models are new: Charles' equation was published in 1957 and Von Rittinger's model was proposed in 1867. A quick refresh how to apply these equations can turn these oddities from the undergraduate curriculum into useful tools for plant optimization. They can be fit to any type of mill including stirred and tumbling ball mills.IntroductionBuilding a case for fine grind modelling requires reviewing some theory before building a framework for doing fine grinding calculations. The actual calculation procedure recommended is easy and can be applied by plant metallurgists. Don’t be frightened by the theory; the end result is reasonably simple.The observation that mineral breakage is related to the absorption of energy is usually attributed to Von Rittinger (1867). Lynch and Rowland (2005) describe that the ability to empirical test and calibrate energy models of breakage didn’t exist until the widespread adoption of electric motors in the mining industry during the 1930’s and 1949’s. A series of technical papers published by the Allis Chalmers company during the 1940’s demonstrate that they were attempting to do such a calibration using Von Rittinger’s model which is usually stated as: “the energy consumed in the size reduction is proportional to the area of new surface produced.”"

Jan 1, 2017

By S. D. Hill

To help insure the continued viability of the minerals and materials sector of the U.S. economy, the Bureau of Mines is conducting research and disseminating technology that will result in more efficient minerals extraction. As part of this effort, a process is being investigated that involves an aqueous chlorine-oxygen leaching procedure to treat complex sulfide concentrates to produce ZnCl2. Subsequent fused-salt electrolysis of the ZnCI2 produces high-purity zinc metal. Previous reports described preliminary, small-scale research on the chlorine-oxygen leaching system and fused-salt electrolysis of ZnCl2. This report discusses the details of design, construction, and longterm, continuous operation of both monopolar and bipolar fused-salt cells. The effects of temperature, electrode gap, current density, electrolyte composition, and electrode configuration on current efficiency and energy consumption were determined for a ZnCl2-KC1-NaCl electrolyte system. Zinc metal was produced at the rate of about 50 lb/day in a monopolar cell which was operated continuously for 60 days with current efficiencies from 85.8 to 98.6 pct and an average power consumption of 2.1 kwhr/lb. Bipolar cells with several different electrode configurations were operated for 90 days. A current efficiency of about 70 pct and power consumption of about 1.7 kwhr/lb of zinc were achieved using a three-electrode-pair, bipolar assembly with a chlorine gas lift pump to circulate electrolyte.

Jan 1, 1981

By Joseph L. Condon

Fifteen shots of three explosives with charge weights of 0.52, 1.1, 3.1, and 5.1 kilograms were fired in a pond to study the seismic effects of detonating explosives in water. Shock pressures and bubble periods were measured with pressure transducers in the water. Vibrations were measured on land in three orthogonal directions with particle velocity gages. Statistical tests indicated that a charge weight exponent of 0.43 best scaled the data. While shock pressures and bubble periods showed considerable differences in explosive performances, the amplitudes of ground vibrations showed less difference, and one prediction equation was therefore derived to estimate vibration amplitudes from all the explosives.

Jan 1, 1970

By Thomas Varley

"The literature on flotation of oxidized ores except for patents (especially for reagents) is very sparse. The Bureau of Mines in 1936, published a circular on the subject, by Ralston and Allen. The treatment of oxides by flotation was then in its infancy, but several of the most important mining companies in the Western United States and Mexico are now operating flotation plants for treating oxidized ores. The purpose of this present circular is to review the progress made and bring the subject up to date. One of the writers of the 1916 report (Mr. Allen) has been associated with some of the most important mining companies who have put into practical operation the process of flotation of oxidized ores; among these companies are the Shhttuck Arizona Copper Co., Bisbee, Arizona (now the Shattuck Denn Co.) and the San Francisco Mines of Mexico, Ltd., San Francisco Del Oro, Chihuahua, Mexico. His experience and other sources of information have been freely drawn on in pre-paring this paper. Mr. Ralston, assistant chief metallurgist of the Bureau of Mines, contributed many valuable suggestions and criticisms.Sulphidizing Reagents for Oxidize Ores.The various reagents and methods that might possibly be used for sulphidizing were summarized by Ralston and Allen as follows:""Sulphidizing (1) by the use of hydrogen sulphide on either the dry or the wet crushed ore, (2) by the use of solutions of the various sulphides and sulpho-compounds of sodium, (3) by the use of solutions of the various sulphides and sulpho-compounds of calcium, (4) by the use of sulphur vapor, (5) by the use of a sulphureted oil, (6) with colloidal sulphur."

Jun 1, 1927

By Larissa Koroznikova, Steve Durkin, Patrick Bouwmeester, Michael Tuck

In currently used mining methods, blasting techniques often causes dilution of valuable ore and results in costly processing requirements. In the context of narrow vein mining of thin and highly concentrated orebodies there is a need of a mining method that can reduce dilution in order to remain economically viable. This research project explored the viability of a new mining technology, thermal fragmentation, in narrow vein mining. Thermal fragmentation technology uses a flame jet to produce extreme heat that spalls the surrounding rock to a strategically located drill hole, as an alternative to traditional blasting. This paper creates a net present value (NPV) model of a mining method using thermal fragmentation, as well as for an existing method used for narrow vein mining; comparisons and evaluations were made regarding the feasibility of the new technology. It was found that while overall costs for thermal fragmentation were relatively high, reductions in wages, haulage and processing costs, as well as increased productivity and ore recovery, meant that the new method would improve the financial performance of any operation. These results identify that there is an opportunity to introduce the thermal fragmentation technology to narrow vein mines within Australia, in order to lower costs and increase profit.

Following extensive trials on a 100 000 tpa Research and Development facility at Kwinana, Western Australia, Hismelt Corporation (wholly owned by Rio Tinto) are confident of the future commercial success of this technology. The Hlsmelt process uses iron ore fines and non-coking coals to produce molten pig iron suitable for feeding to BOF and EAF steelmaking vessels. The unique process configuration combines direct injection of raw materials into the molten iron bath, high rates of metal and slag mixing, and effective post-combustion of the bath gases to achieve efficient coal utilisation. This is the key to the strong economic performance of the process.

Jan 1, 1999

By D. Imeson, T. Grammatikopoulos, G. Pearse, M. Aghamirian, C. Mohns

"Spodumene mineral is a source of lithium for lithium ion batteries. Thus, the production of a spodumene concentrate with an approximate grade of 6% Li2O is targeted. Traditionally, spodumene is concentrated by froth flotation using fatty acid as the collector. Amine collector can also be used to float mica ahead of spodumene as required or feldspar from the spodumene tailings. Slimes have a negative effect on spodumene flotation and they are commonly separated by scrubbing and de-sliming prior to spodumene flotation. The main losses in spodumene flotation typically occur in cyclone overflow, scrubbing slimes, spodumene flotation tailings and mica concentrate. In recent years, dense media separation (DMS) is the main concentrating process in Australian spodumene operations. In these operations, DMS is efficient at generating coarse grain concentrates and rejecting large portions of silicate gangue, thereby directly reducing the milling cost. The ore should be crushed to 3/8” or coarser while further crushing is possibly necessary on the middling particles to scalp further spodumene. The DMS feed should be classified to remove fines (typically -0.5 mm particles). The final middlings and undersize fractions could then be processed by flotation. The main requirement in DMS operations is to have good spodumene liberation at relatively large grain size. This paper reviews the most significant DMS and flotation parameters in spodumene beneficiation. The treatments of different minerals in the spodumene ore are also addressed.INTRODUCTION A number of laboratory/pilot plant programs have been recently conducted at the SGS Lakefield site, sponsored by different clients. The main objective has been to develop a process flowsheet to generate a spodumene concentrate containing 6.0% or higher Li2O (~75% spodumene) suitable for producing lithium carbonate by a hydrometallurgical process. Depending on spodumene liberation size, dense media separation (DMS) can be an attractive option to produce a final high grade concentrate and to reduce the mass to be treated in the flotation circuit. In such a case, crushing would typically be performed in stages, with DMS separation to recover a spodumene concentrate and reject gangue minerals following each stage. DMS middlings and -0.5 mm fines can be processed by flotation to further recover spodumene. In the flotation process, coarse stage grinding minimizes the grinding cost and formation of slimes. In the majority of cases, a grind size (K80) of approximately 200 microns is sufficient for flotation. After primary grinding, mica, if present in an appreciable amount, should be floated to minimize interference with spodumene flotation. Desliming after scrubbing and high density conditioning are the key process operations prior to spodumene flotation. Spodumene can then be floated from silicate gangue minerals with a fatty acid or a mixture of fatty acid and petroleum sulfonate at slightly alkaline pH. Subsequently, feldspars can then be floated from the spodumene tailings."

Jan 1, 2012

By Xiaodong Ma, Marek Pawlik

Introduction Guar gum is a natural nonionic polysaccharide. Guaran, the functional polysaccharide in guar gum is a chain of (1?4)-linked ß-D-mannopyranose units with a-D-galactopyranose units connected to the mannose backbone through (1?6) glycosidic linkages. The poly-mannose chain is randomly substituted with galactose units at a mannose-to-galactose ratio of 1.8-1.0 [1]. Cheng et al. reported that the molecular weight of natural guar gum is 1.935 • 106 [2]. Robinson et al. found that the molecular weights of five grades of commercial guar gums ranged from 4.4 • 105 to 16.5 • 105 [3]. Guar gum is frequently used as a depressant of naturally hydrophobic gangue minerals (e.g. talc) in the flotation of PGM-bearing ores, and according to Harris et al. a wide variety of modified guar gums are available for this application [4]. The Canadian potash industry utilizes natural guar gum as a blinder of water-insoluble slimes in the flotation of potash (sylvinite) ores, a process that is carried out in a saturated KCl/NaCl brine. The role of the polysaccharide is to prevent a primary amine collector from adsorbing onto these unwanted fine minerals. Hydrophobic interactions between basal cleavage planes and the polymer appear to be involved in guar gum adsorption on talc [5,6]. Liu and Laskowski stressed the importance of metal impurities on mineral surfaces that can serve as the active sites for polysaccharide adsorption [7]. This postulate was consistent with the later observation that, in addition to hydrogen bonding, the adsorption of guar gum onto talc was strongly dependent on the surface concentration of magnesium sites [8,9]. In this case the adsorption process was visualized as metal-polysaccharide complexation facilitated by the cis-configuration of the hydroxyl groups of mannose monomers. Hydrogen bonding and chemical interactions with metal sites were also suggested by Rath and Subramanian in their studies on guar gum adsorption on biotite mica [10]. In a recent study, Wang et al. demonstrated that hydrogen bonding is the main driving force behind guar gum adsorption onto talc [11]. Their data also show that the adsorption density of guar gum on alumina is about five times lower than the adsorption density on talc. Pawlik and Laskowski reported that the amount of guar gum adsorbed on illite was independent of ionic strength over a wide range of salt concentrations [12]. The adsorption density of the polysaccharide on illite did not measurably change with ionic strength from that of distilled water to about 50%-saturated KCl/NaCl potash brine (~3N KCl/NaCl). However, guar gum adsorption on dolomite dramatically decreased over the same range of ionic strengths, and no reliable explanation of that phenomenon was proposed [12]. The water structure-breaking or structure-making capabilities of background ions are usually ignored when discussing adsorption mechanisms of polymers. The concept of “structure breakers” and “structure makers” was first introduced by Gurney, and Frank and Wen who postulated that each ion is surrounded by three distinct regions of water structure [13,14]. In the first layer, water molecules are tightly bound to the ion. The second region extends farther away from the ion and is referred to as the region of structure breaking. Only at larger distances, where the ionic field is weak, water molecules form the “normal” ice-like structure. It is also now recognized that the first layer can be further divided into the primary and secondary hydration shells. Small, strongly hydrated ions reinforce the “normal” structure of water and the region of structure breaking disappears. In contrast, large less-hydrated ions disturb the ice-like structure and generate an extensive region of structure breaking. Experimentally, these two effects can readily be distinguished through viscosity measurements. Increasing concentrations of structure-breaking ions result in a decrease of the viscosity of aqueous solutions. On the other hand, increasing amounts of structure-makers significantly reduce the viscosity of water. Alkali metal cations form a sequence known as the Hofmeister series. The Hofmeister series - Cs+, K+, Na+, Li+ - orders the cations from the least hydrated to the most hydrated. Based on literature viscosity data for aqueous

Jan 1, 2005

By André Desbiens, Jocelyn Bouchard, René del Villar

"The metallurgical performance of the column flotation process is determined by the concentrate grade and recovery. Although the former can be continuously measured using an on-stream analyzer, the latter must be estimated by a material balance calculation based on seldom verified steady-state assumptions. Consequently, the automatic control and optimization of flotation columns should be performed using secondary variables having a strong influence on the metallurgical performance, such as froth depth, gas hold-up or bubble surface area flux and bias. The aim of this paper is to present a model-based automatic control strategy using froth depth and bias measurements obtained from virtual sensors developed at Laval University. Preliminary pilot plant results on two-phase (water– air) and three-phase system (water – minerals – air) will be presented and eventual industrial implementation of such a strategy as well as the possibility of achieving effective real-time optimization from bias and froth depth measurements will also be discussed.INTRODUCTIONFlotation columns have been widely studied and used in mineral processing for about twenty years. Although a great deal of work has been accomplished to understand the influence of different variables such as froth depth (H), bubble surface area flux (Sb) and bias (Jb) on the metallurgical performance, the measurement of some of these variables is still difficult to be performed – except at the laboratory level – because in most cases, no commercial device yet exists.As a result of this, it is almost impossible to achieve neither efficient control nor optimization of the process at the industrial level, even in open-loop operation. This paper aims at presenting a simple control strategy of froth depth and bias. Description of the sensors used to achieved measurements, preliminary pilot plant results for froth depth control, eventual industrial applications and possibilities for process optimization are also treated."

Jan 1, 2003

By Zachary Mayer

Agenda ? Kidd Mine ? Background ? Problem & Opportunity ? Challenges associated to depth - Production Rates - Productive Work Time - Operational Delays - Transfer Levels - Mucking Time - Impact on Hoist ? LHD Automation ? Sandvik AutoMine Lite ? Expected Benefits Kidd Mine - Background ? Located in Northern Ontario, Canada ? Underground base metal mine operating since 1966 ? Total 148 Mt mined (2.31% Cu, 6.08% Zn, 75 gpt Ag) ? 2.35 Mt/Y from stopes as deep as 9500 ft (2.9km) below surface ? Mining method: blasthole open stoping ? Mine/Mill: 1263 employees and contractors ? 2012 National John T. Ryan Safety Award

Aug 1, 2013

By Dave Rose, Jacques McMullen

"The Detour Lake mine is located 200 km northeast of Timmins, Ontario, Canada.The mill processes a free milling gold ore using both a gravity circuit and leaching with a state-of-the-art carbon-in-pulp circuit. The grinding circuit is a conventional SABC circuit with a tonnage of 55,000 tpd at 95 microns with the addition of pre-crushing upstream of the SAG mill.The Detour Lake mine has one of the largest gyratory crusher in the world (60"" x 113"") and over the years, many improvements have been made to this crusher with the result of improving plant throughput. This paper will present how choke feeding the gyratory crusher has been found to be one of the main influences on crusher product particle size and plant throughput. It will explain how the effect of choke feeding the gyratory was measured and from that, the various changes made to improve choke feeding of the gyratory crusher.IntroductionOre is hauled from the open pit mine and dumped directly into a single 60"" x 113"" primary gyratory crusher. The live capacity of both the feed and discharge hoppers of the gyratory crusher are designed for two (2) truck loads each, assuming a nominal payload of 300 tonnes per truck. The surge bin under the primary crusher feeds an apron feeder that controls the ore feed rate at approximately 4,200 tph onto a 227 metres long x 1.52 metres wide belt conveyor with a nominal capacity of 4,200 tph.The primary crusher is operated at an open side setting of 172 millimetres to produce a F80 feed to the secondary crushing circuit of 135 millimetres."

Jan 1, 2017

By K. K. Murray, D. G. Gelderland

"The pressure oxidation autoclave at the Goldcorp’s Red Lake Gold Mines (RLGM) was commissioned in 1991. Since commissioning, the circuit has been well maintained and its capacity increased to mechanical and gaseous oxygen limitations. With advances in available technology, opportunities to produce more sulphide concentrate and a changing business environment, an optimization project has been initiated at RLGM. A multi-phase approach has been developed to optimize both the operations and economics of the circuit. This paper describes the history and future of the autoclave circuit at RLGM.AcknowledmentsWe would like to thank the operations and maintenance staff at Red Lake Gold Mines for their support and openness of mind to enact changes to enhance the operation. We would also like to acknowledge the members of the integrated team from Red Lake maintence, metallurgy, operations and corporate technical services that did the hard work to push past the status quo in order to advance productivity.OpportunityThe Red Lake Gold Mines (RLGM) pressure oxidation (POx) circuit has been in operation for over 25 years processing the high sulphide content gold bearing minerals. Increased sulphide concentrate production from improved flotation recovery and increasing gold sulphide association at depth has led to the POx circuit becoming a throughput bottleneck. The opportunity exists to process increased tonnages of sulphide concentrates and increase gold production. In order to realize this value, it will be necessary to increase the throughput capacity of the POx circuit.BackgroundThe POx autoclave at the RLGM, commissioned in 1991, was Canada’s first gold autoclave. It was installed to replace the two-stage fluid bed roaster that was built when the Campbell mine and mill were constructed in the late Forties. The circuit was originally designed to process 12.5 t/d sulphide sulphur in about 50 t/d of Campbell flotation concentrate. Following the merger of the Campbell and Red Lake Mines in 2006, the POx circuit started processing flotation concentrate from both Campbell and Red Lake Mills. In recent years, the circuit has been processing about 18 t/d sulphide sulphur in 75-85 t/d flotation concentrate."

Jan 1, 2017

By Troy Holm, Matthew L. Fuller, Frederick H. Earnest

Virtual Reality Animation Of The Proposed El Dorado Mine In El Salvador Assists Regulators In Reviewing An EIA

Jan 1, 2005

I had hoped when the time came for me to address this meeting that gold mining would be in a much happier position than we find it to-day.As it is, and because of the demands of national security, we find this great and valuable primary industry relegated to the background and in danger of temporary extinction.These notes refer mainly to gold mining in Western Australia, whence comes the major portion of Australia's production, and their purpose is to record a brief review of the industry immediately prior to the general disorganization that is following the call-up of men from the mines.During this period of reduced production the loss of comparable statistical data will make it impossible to follow the progress of the industry, and, since the period may be long, it occurred to me that information might be derived from an examination of the trend of recent production that would indicate what may be expected of the industry when full production is resumed in the future.Since the revival of gold mining, which started in 1931, most of the old-established mines have reorganized and, in some cases, completely renewed and enlarged their equipment in order to achieve maximum production, and old mines hitherto unpayable have been re-opened and equipped. At the same time large sums have been spent in scientific prospecting and in search for new mines, and it is safe to say that at the end of 1939, when gold production reached its peak, practically every known mine in Western Australia that could be worked profitably was in production.During this period the number of men employed in the industry rose from 5,961 in 1931 to approximately 15,000 in 1939. From the outbreak of war in 1939 to the end of 1941 the output from the mines was maintained at full pressure, maximum tonnage beingrecorded in 1940, although a reduction in the grade of ore treated caused a slight falling off in gold production.During 1941 there was a. fall in both tonnage and grade. The decline in tonnage was due partly to the closing of the Lancefield and Mount Magnet mines, and partly to the shortage of labour brought about by the response of the goldfields to the call for men for the fighting services, the average number of men in the industry in this year being reduced to 12,929.With the entry of the United States into the war in December last and the pooling of the resources of the allied nations, the need for gold as a purchasing medium passed, and with its passing there arose the demand for the transfer of the man-power engaged in gold mining to other fields of industry directly connected with the war effort.There was a ready response by those connected with the industry who...

Jan 1, 1942

By Hassan Ghaffari

? HPGR Process/Overview ? HPGR vs. SAG ? The Manufacturers ? Mine Operations using HPGR ? Projects by Wardrop/Tetra Tech ? Economic Benefits ? Questions Main Operating Parameters ? Press Force (N/mm2) ? Roll Speed (m/s) ? Moisture Content (%) ? Feed Particle Size Distribution Specific throughput M* = throughput/D*L*V (ts/hm3) Flake thickness/strength in balance with the compressing force (floating roll situation)

Aug 1, 2013

By J. Allaire, P. N. Muteb

The Meadowbank mine is operated by Agnico-Eagle Mines Ltd. and is located in the Kivalliq District of Nunavut in northern Canada. The process plant operates presently at a throughput of 11,500 tpd with gold grade varying between 2.9 and 3.6 g/t. The ore supplied to the process plant come from two different open pits and it expected to come from a new open pit early 2014. The ore is mid-hard materials so the feed size ore have a bigger impact on the throughput. The average gold recovery is 94.0% with about 15-30% of gold recover from the gravity circuit. The Meadowbank mine commissioning started in February 2010 utilizing a flowsheet consisting of one stage of crushing followed by primary and secondary grinding. The commercial production began in March 2010. During the commissioning several difficulties were observed to reach the designed throughput of 8,500 tpd. It was evaluated to insert a second stage of crushing. The insertion of the secondary crushing have two objectives, first to reach easily the process plant design throughput as well as to evaluated the mill capacity increasing. The process plant team evaluated various expansion alternatives eventually deciding to install a secondary crusher, increased the SAG and Ball mills motors capacities and maintain the rest of the process unchanged. Early 2011, with the improvement in the gold markets, decision was made to install the secondary crusher between the primary crusher and the primary grinding. The process plant continued to operate on the initial design until early June 2011. This paper presents the SAG mill throughput increasing from the designed tonnage of 8,500 tpd to 2012 targeted tonnage of 11,500 tpd, secondary crusher commissioning and the difficulties to operate in Canada Artic (Nunavut).

Jan 1, 2013

By Chandrakala Kari, S. Mohan Rao, Gajanan Kapure, K. S. Raju

High carbon ferrochrome production process is energy intensive, consumes approximately 3300-3400 kWh per ton of metal produced. This amounts to ~ 34% of the total ferrochrome production cost. The power costs in India are almost 4 times greater than the leading ferrochrome producing countries like South Africa. Hence there is a need to look for alternative sources of energy. In ferrochrome making process large volumes of high concentration CO gas (~ 85%V) is developed at ~400-5000C, which if combusted within furnace represents high potential for significant electrical energy savings. In a typical heat in the state-of-the-art submerged arc furnace, about 34 kmol of CO gas is generated per ton of ferrochrome metal, which, if post combusted to CO2, represents 7620 MJ of energy. Post combustion of the evolving CO (and H2) gas to CO2 can be achieved by providing additional oxygen before the generated CO gas leaves the furnace. In this paper, a concept of post combustion in submerged arc furnace as potential source of significant electrical energy saving in ferro-chrome making process has been explored. The prospects and limitations of the same are discussed. The value of this paper lies in its attempt to explore the post combustion application to submerged arc furnace with an objective of significant reduction in electrical energy consumption in ferrochrome production process.

Jan 1, 2007

By Murad Y. Abu-Farsakh

This paper presents the analyses of load-transfer (t-z) behavior of twelve prestressed concrete (PSC) instrumented test piles that were driven in different locations of Louisiana. The load transfer method is an efficient and practical tool for the analysis of single driven pile to predict the ultimate load. The test piles were instrumented with vibrating wire strain gauges in order to calculate the shear transfer for different soil layers during the static load test. The static load test was performed on all the test piles after 10 to 14 days of end of driving. Empirical load-transfer models were proposed previously by different researchers (Fleming 1992, Hirayama 1990, Krasinski 2012, Bohn et al. 2017) after analyzing the instrumented test pile results. For this study, load transfer of each soil layer was measured and compared with the previously developed existing model. The stiffness parameter (M) for the hyperbolic equation was calibrated in order to best match the measurements. The good agreement of the proposed curves was then confirmed by applying them to assess the overall load-settlement curve of other pile load test results available in the literature. INTRODUCTION The load transfer mechanism (stress-strain relationship) for an axially loaded pile usually comprises of three loading mechanisms: axial deformation in the pile (z), soil skin friction along the shaft (qs), and soil end-bearing (qb). The development of soil skin friction with axial deformation is referred as the t-z curve, and the development of soil end-bearing with axial deformation is referred as the q-z curve. Figure 1 presents the load transfer mechanisms in an axially loaded pile and the corresponding spring mass models for t-z and q-z curves (API, 2002). In order to construct the load transfer curve in clay under vertical load, the axial load transfer (t-z) curve and end-bearing load transfer (q-z) for that particular soil should be assessed separately. The axial load transfer (t-z) method is probably the most widely used technique to study the problem of single axially loaded piles. Therefore, the load transfer through the shaft of the pile and are getting more and more important compared to the only the pile capacity analyses. The load transferred from shaft to the surrounding clay soil is a function of the width of the pile and the surface roughness of the pile and clay properties. The development of a representative procedure allows the assessment of the t-z curve in soil (sand and/or clay) that leads to the prediction of a nonlinear vertical load-settlement response at the pile head.

Jan 1, 2018