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By B. Kweto, W. Legodi, C. van Buuren, R. F. Sandenbergh

Biooxidation, and especially the BIOX® process, is a viable, economical and environmentally attractive process to liberate gold hosted in refractory sulfides by oxidatively leaching the sulfides as soluble sulfates. The moderate to high cyanide consumption during the subsequent cyanidiation of the washed biooxidation residue is probably caused by the retention on the solids of labile sulfur species formed during the multi-electron oxidation process of sulfide to sulfate. In this work the nature and mode of transport of the cyanicides in the products from the mesophile BIOX® process, as well as that of the combined mesophile and thermophile bioxidation process, were considered and strategies to reduce cyanide consumption evaluated. It was found that labile sulfur species were the major cyanide consumers in the mesophile biooxidation product, forming predominantly thiocyanate during cyanidation. The transfer of labile sulfur species was only marginally decreased by more intensive and thorough washing of the biooxidation products with warm and hot water, intense stirring and warm caustic leaching. A major reduction in cyanide consumption was however achieved by using a combined mesophile and thermophile biooxidation process. Keywords: Biooxidation, refractory sulfide gold ores, cyanide consumption

Jan 1, 2020

By I. du Plessis

The metallurgical response of a difficult to treat coal and an easy to treat coal orebody were investigated in terms of the coal yield and the coal quality in a basic coal recovery flowsheet. The purpose of this paper is to show trends and derive principles for the treatment of different coal types. This investigation entailed using the LIMN simulator to simulate the metallurgical response by varying the plant operational parameters. The controllable plant variables investigated were: ?Cyclone size ?Screen and hydro-cyclone cut points ?Medium cut density ?Feed particle size distribution.

Jan 1, 2010

By J. C. M. De Souza

The Samarco pipeline, the longest in the world for iron ore slurry transportation, started operation in 1977 to transport concentrate from the Germano facilities, in the state of Minas Gerais, to the pelletizing plant, located 396 km away, in the state of Espirito Santo, on the south-east coast of Brazil. Designed for a nominal capacity of 12 million tons per year, the pipeline reached 15.5 million/year due to operational improvement and, in late 2005, it was resized to transport 16.5 million tons per year, with the installation of a booster station. In order to ensure the accurate performance of the system, a pipeline cleaning programme was carried out and, since it has been29 years from the operation start-up, precise knowledge of the pipe condition would be fundamental to guarantee the success of the project. The cleaning programme demonstrated the potential of operational gains from this initiative, and its effect on the corrosion control and future measures, which should be permanently adopted to maintain the improved capacity. The intelligent pig runs provided essential data used to establish a pipeline integrity programme that allowed Samarco to develop a pipe repair plan based on the severity and significance of the features and to implement a future pipe replacement plan.

Jan 1, 2007

By H. E. Bartlett

This paper deals with the sampling and mass measurement for ore delivered from a shaft to a processing plant and the contribution of the data from these measurements to the metal balances from shaft deliveries to final metal production. Accurate measurement of the grade and tonnage of run-of-mine ore is important for four main reasons; ? It enables the measurement of the production from different profit centres to within statistically determined confidence limits for daily, monthly, or annual averages. The profit centres could be individual shafts within a mining complex or ore treated on a toll basis ? The monthly production at shaft head is compared to the grades and tonnage determined from underground sampling and mass measurement in terms of a shaft call factor ? The sum of the production from the shafts is the input to concentrators. This input is a major part of the total input into a complex with concentrators, smelters, and refineries. In terms of the Codes of Practice for Metal Accounting the inputs are compared to outputs and inventory changes to assess the efficiencies and unaccounted losses or gains at the various stages in the flow of metal from source to market ? On a daily or daily moving average basis, the grades and tons from shafts are monitored and compared against these quantities from underground measurements. This acts as a control on off-reef mining, dilution, and other factors underground. Accurate measurement of grade of ore at the shaft head has been a challenge because of the large particle size. The conventional wisdom has been that ore can be sampled accurately only after it has been milled to give a slurry that is sampled as feed to a flotation process. However, when the ore fed to a flotation plant comes from multiple sources each source has to be sampled separately. So, relying on the grades determined using the sampling of concentrator input with cross-stream slurry samplers is not an option for determining the grade from an individual shaft. Impala has developed a system for sampling and weighing run-of-mine ore from multiple shafts. The system involves sampling the inputs to the plants using cross-belt (hammer) samplers and weighing the deliveries using in-motion railway weighing systems. Many samples are taken. Individually they have a high variance but, as a consequence of the averaging effect of large numbers and as shown by statistical analysis, the mean results are fit for the purposes of daily grade control for shafts and for monthly accounting of production from individual shafts. In the metal accounting systems at Impala, the measurement of input, including ore and toll-treated material is compared to the output in the form of final metal and tailings losses, as well as any inventory change that takes place, in order to determine a final metal balance from mill feed to product. The calculated inventory is compared to that measured, and the difference, or unaccounted- for metal, is then reported as a percentage of input. The balances obtained over several years yields balances for platinum that are below 1% imbalanced. Keywords metal balancing, platinum mines, AMIRA codes of practice, metal accounting, sampling, run-of-mine ore, statistical analysis.

Jan 1, 2014

By J. O. Roux

The Impala Platinum Limited Base Metals Refinery in Springs, South Africa, refines convertor matte containing base metals and PGM?s (platinum group metals) using a hydrometallurgical process. In addition to a PGM-rich concentrate that is further refined, the Base Metals Refinery produces nickel (in powder or briquette form), copper cathodes and cobalt powder. Five leaching stages are utilized to remove the base metals and impurities from the PGM concentrate. The project described in this paper involves a radical redesign of the second stage leach autoclave with the aim of reducing maintenance costs and downtime, increasing throughput and ensuring a safer working environment. The second stage leach process at the Base Metals Refinery, produces a PGM-rich residue through the extraction of copper and any remaining base metals from the first stage leach solid residue material. Leaching is performed in an autoclave operating at a pressure of 600 kPaG and a temperature of 140°C to 150°C. Concentrated H2SO4 is added to the feed to produce a solution with a free acid content between 10 and 30 g/L and a pH of less than 2. Oxygen is passed into the 1st and 2nd compartments of the autoclave. As a result of the aggressive, highly oxidizing conditions within the autoclave, the original design utilized a heavy wall carbon steel shell with lead lining and two layers of acid bricks. The bricks acted as a wear resistant material and ensured a low surface temperature at the skin of the lead lining, whereas the lead lining acted as a corrosion barrier to protect the steel shell. These brick lined autoclaves were heavy, maintenance intensive and prone to catastrophic failure. A radical autoclave redesign was needed to reduce the high costs of manufacturing, installation and operation. Finite element modelling was used to optimize the design and various high alloy materials were investigated for the application. Based on extensive test work, the carbon steel shell was replaced with duplex SAF 2205 stainless steel in the new design. The use of this highly corrosion resistant material, eliminates the need for the lead and brick linings and reduces the wall thickness requirements substantially, resulting in an increase in the available volume for the same external shell dimensions. The vessel is welded in circular sections to form an elongated, impermeable, cylindrical body, with several agitator nozzles and compartments. The first new duplex stainless steel unit has been in continuous operation for more than 24 months and its performance has exceeded the highest expectations. The design was so successful that a patent has been registered. The major benefits of the new stainless steel design include increased throughput, improved performance, less downtime and a safer working environment. Lead has been eliminated and replaced with a more environmentally safe material, and the total lifespan of the vessel has been extended. In addition, the autoclave project has made significant developments in advancing the field of welding and optimizing the integrity of welding quality standards in South Africa.

Jan 1, 2009

By P. K. Van Manen

Anglo Platinum?s Polokwane Smelter is situated outside Polokwane, in the Limpopo Province in South Africa. A single 168 MVA six-in-line rectangular furnace smelts dry concentrates containing platinum group metals (PGMs). Furnace energy consumption since commissioning is presented, showing that energy efficiency improves as capacity utilization increases. An energy balance for the furnace is outlined, indicating the measured or calculated energy losses from the hearth and side-wall cooling, the copper coolers, the upper furnace walls and roof, and the energy losses in off-gas, matte, and slag. From the energy balance, the potential to improve energy efficiency by controlling the slag temperature, the slag level, and the off-gas volume is derived. Furthermore, the impact on the energy balance of different concentrate types is discussed, as well as the potential impact of replacing all upper waffle coolers by plate cooler panels.

Jan 1, 2009

By J. Kempken, T. Schreiter, R. Degel, H. Schmieden

After a short summary of the historical milestones of the submerged arc furnace, the paper will highlight different applications and their specific smelting furnace requirements in the non-ferrous area. Additionally, aspects of rectangular furnace design and the features of the copper cooling system for the side wall will be emphasized. Finally a state-of-the-art design tool as well as a new invention to improve slag cleaning technology will be featured, which shows great potentials especially for the PGM and copper related industry. Keywords: submerged arc furnaces, aerro alloys, non-ferrous, copper, slag cleaning, PGM, rectangular furnace, DC technology

Jan 1, 2006

By L. A. Cisternas, A. Lopez-Valdivieso, Y. L. Botero, L. Cortes

In the processing of sulfidic ores, pyrite is usually depressed and ends up in the tailings. However, pyrite is the most common species responsible for acid mine drainage (AMD) due to its spontaneous oxidation. This AMD affects the tailing chemical stability by dissolving some toxic or heavy metals species such as As, Cd, Co, Pb, and Zn, which can migrate into the surrounding ecosystems. Therefore, the strategies to reduce pyrite’s environmental impacts are a crucial point to produce more chemically stable tailings. Consequently, numerous studies have been made to find procedures that could inhibit or restrain AMD. This work aims to develop a procedure for designing conceptual flotation circuits that can produce tailings with low content of pyrite and to assess this type of circuit with traditional flotation circuits. The method consists of applying optimal flotation design where three products are considered: metal concentrate, pyrite concentrate, and tailing with low content of pyrite. Two approaches are considered; in the first, a flotation circuit is designed to generate a concentrate rich in all sulfide minerals and to obtain a tail with low pyrite content. Then, a second flotation circuit is designed for the selective separation of the bulk into metal concentrate and pyrite concentrate. This is called fractional separation design. In the second approach, a flotation circuit design is realised that generate the three products simultaneously. This is called integrated separation design. Both fractional separation design and integrated separation design are assessed in this paper. This methodology is focused on cleaner production (waste prevention and resource minimisation) rather than end-of-pipe technology (treat existing waste). The methodology was applied to a copper sulfide ore from a local company in the Antofagasta region (Chile). This ore contains chalcocite, covellite, iron oxides, and pyrite. Keywords: Flotation, tailing, pyrite, cleaner production

Jan 1, 2020

By F. T. Cawood

This article offers a personal and independent view on the threats to the sustainability of the South African minerals industry. It argues that if the six identified threats of insufficient spending on research and development, general standard of education, nationalization, inadequate infrastructure, Aids prevalence and labour inefficiency are not acted upon as a matter of urgency, the future of the mining industry, along with its important contribution to the economy, is at risk.

Jan 1, 2011

By A. D. Pienaar

The aqueous chemistry of niobium and tantalum compounds with hydrofluoric acid is a mature field of investigation. In contrast, very little attention has been paid to reactions of these compounds in non-aqueous environments. Similarly, reactions of elemental fluorine with niobium and tantalum compounds have received little more than cursory attention in literature. This work discusses the isothermal reaction kinetics derived from thermogravimetry, using a less redundant approach than is currently the standard. We also discuss the most likely kinetic models for this reaction.

Jan 1, 2012

By P. Saari

The Talvivaara deposits are located in Eastern Finland, within the Kainuu Schist Belt. These polymetallic sulphide deposits are of relatively low grade, but bioheapleaching enables economically profitable nickel extraction from the ore. Zinc, cobalt, and copper are also produced. Talvivaara?s production process begins with large-scale open-pit mining and four crushing stages, followed by stacking of the ore in bioleaching heaps. The ore is first leached for 13?14 months on the primary leach pad, after which the leached ore is reclaimed, conveyed, and re-stacked on the secondary heap pad. After secondary leaching, the barren ore will remain on the secondary heaps permanently. In the metals recovery process, the metals are precipitated from the pregnant leaching solution. Construction of the process was begun in spring of 2007 and first metal sulphides were produced at the plant in October 2008. The development of the bioheapleaching process was begun in a 17 000t on-site pilot heap operated during 2005?2008 and has continued in the industrial heap. At industrial scale, the development measures have focused on improving the permeability and aeration of the heap. Challenges with crushing and aerations systems at the beginning of industrial-scale leaching delayed the increase in metal recovery, but after the first two operational years the leaching results have improved significantly and the process is performing as targeted.

Jan 1, 2012

By B. Geelhoed

Gy's formula for the variance of the fundamental sampling error (FSE) is based on a model that assumes independent selections of particles. A generalization of this model, which can theoretically lead to more realistic representation of actual sampling processes, was, for the first time, experimentally verified by repeated sampling from a well-mixed batch consisting of particles of zirconium silicate of two different sizes. The predictions for the sampling variance made in Gy?s theory and in the more general model were compared with the experimentally observed sampling variance. During the experiment performed here, the variance prediction made in the theory of Gy ranges from 0.96 × 10?7(g2/g2) to 0.95 × 10?7(g2/g2), which is approximately a factor 60 lower than the experimentally observed variance (5.76 × 10?6(g2/g2)). The difference between the Gy prediction and the experimentally observed variance cannot be fully explained by the uncertainty in the experimentally obtained variance due to the finite number of degrees of freedom. Therefore, the difference is likely to be a significant effect. This shows that, under the experimental conditions, Gy?s formula for the variance of the FSE probably undervalues the sampling variance, unless an empirical correction factor for grouping and segregation is applied in the theory of Gy. The new theory resulted in a variance prediction of 6.00 × 10?6(g2/g2), which is much closer to the observed variance than the Gy prediction. This indicates that the generalization of Gy?s theory offers an outlook for providing a better variance prediction than the Gy predictor in the case of dependencies between particle selections. The results are discussed. Keywords: sampling, variance, parameter for the dependent selection of particles, Gy, experiment.

Jan 1, 2009

By S. Konetschnik, E. Dreyer, A. Filzwieser

"More intense furnace operations require effective cooling in order to achieve low refractory wear and good furnace lifetimes, making cooling technology an important aspect of furnace operations. However, the use of water - today’s standard cooling medium - has some disadvantages, as it can cause problems both during furnace start-up and operation, namely hydration problems, corrosion, and explosions. Apart from these aspects, the energy that is removed by extensive cooling cannot be recovered and is normally lost in the cooling towers.With METTOP’s new patented cooling technology it is possible to overcome the disadvantages of water by using an alternative cooling medium, namely ionic liquids (ILs). The main advantage of ILs is the fact that there is no risk of explosion should this liquid come in contact with molten metal. Furthermore, ILs can be used at higher temperatures than water - up to 200°C. This relatively high cooling medium temperature avoids hydration and corrosion problems and allows recovery of the energy by combining the ILTEC system and the organic Rankine Cycle (ORC) technology, for example.Due to the removal of the explosion risk, much higher safety standards can be achieved, particularly in critical areas such as tap-holes"

Jan 1, 2014

A large body of work exists discussing the impact of grinding media on mineral flotation. Generally, the work indicates that a change to a less electrochemically active grinding environment has positive benefits on downstream processing. There is a fear, however, that these benefits may not be realized when treating low sulphide containing ores, typified by UG2 style deposits. A series of experiments was developed to test the flotation response of UG2 ores using the Magotteaux Mill® to determine if measurable differences in the pulp chemistry could be discerned using a range of grinding media types. Further, the flotation responses of the contained sulphides (chalcopyrite, pentlandite and pyrrhotite) were determined. The data collected provided strong evidence that the grinding environment can significantly influence the pulp chemistry and flotation characteristics of all sulphide minerals, even when the ore contains less than 0.5 per cent total sulphide. The results show that the change to an inert media type produced a substantial improvement in the flotation rate of all sulphide minerals, as well as the PGMs.

Jan 1, 2009

By P. R. Jochens, D. D. Howat, C. W. Wolhuter

Prof D. D. Howat (Member): I am a man with a grievance this afternoon. The other authors of the paper entered into a conspiracy to ensure that I presented the paper, and finally even had the audacity to say that this was the least that could be expected from me as I had done so little of the other work. Looking at the distinguished audience here that includes some figures who have forgotten more about cement that some of us have ever learned, I realise how right the other authors were. I feel very much cast in the role of the fool who was pushed in where angels fear to tread. I hesitate to say, Sir, that I would not use the word angels in any definition of the authors of this paper. In presenting the paper, may I try to put forward a few simple points about cement that are probably as little known to the majority of the audience as they were to me-at least until very recently. Concrete, made by mixing Portland cement, sand, crushed rock and water is the most widely used constructional material in the world. Generally the production of actual Portland Cement is about two-thirds that of steel and by the time this is formed into concrete the total is about five times the world tonnage of steel produced. A very rough approximation suggests that world consumption of concrete is about one ton per person per annum. Apart from fresh water mankind consumes no other material in such quantities. What mankind originally required was a mortar to be used for joining blocks of stone or bricks. Such mortar was originally made from limestone burned in a kiln and then mixed with water. Production of mortar of this type began over 5,000 years ago. Some samples from pyramids in Egypt have shown this mortar to be in good condition although over 4,500 years old. The progress of technological development was very slow until 1824 when an English bricklayer, Joseph Aspidin, patented a new mortar material which he made by burning limestone and clay together in his kitchen stove. Aspdin called his product 'Portland Cement' because, in colour, it resembled a well-known building stone quarried in the Isle of Portland on the south coast of England. Aspid's same types of raw materials are used today in modern cement production and the product is still referred to as Portland cement. A lime-containing material such as chalk or shell and a clay-type material such as shale, slate or clay itself, are mixed in the required proportions, ground and fed to a kiln where the minimum burning temperature is about 1,500°C. The raw materials react to form hard chunks of a material known as 'clinker'. About 5 per cent of gypsum is added to the clinker and the mixture is then ground to a very fine powder, the function of the gypsum being to control the rate of setting of the concrete. The key oxides in cement production are lime, silica, alumina and iron oxide. The possible reactions occurring during and the products of the burning process are shown in Table I.

Jan 12, 1968

By L. G. Austin

The paper summarizes the current state of development of simulation models for SAG and FAG mills, and gives the results of recent investigations of the physical processes occurring in autogenous grinding. The breakage process is treated as the sum of three regions of breakage actions: normal breakage caused by nipping of particles between media (steel balls or pebbles); abnormal breakage caused by media when the particle or lump is too big in relation to the media to be readily nipped; and self-breakage resulting from the chip¬ping fracture and abrasion of the tumbling action of rock lumps. Each region of breakage action has associated specific rates of breakage and primary progeny fragment distributions. A simplified form of the model was used to predict the performance of an 8 m diameter SAG mill with L/D = 0.5 grinding a copper ore, and predicted maximum capacity and minimum kWh/ton at about 6% ball load at 25% total filling. Two FAG mills of L/D = 2 were necessary to give the same capacity, and the simulations indicated a lower kWh/ton for these mills. For this ore, both systems were technically feasible.

Jan 1, 1987

By N. C. Joughin

A simplified economic model is derived for the financial return from a new gold mining project. The model is used to analyze the effects of changes in the mining rate, capital costs, working costs, value of the reef and the stoping width. The most effective avenues for improving the return from a new project are indicated. It is shown that those technological innovations which would improve the return significantly, and therefore make it possible to open more mines, are the use of smaller mines, means for improving the mining rate to capital costs ratio for smaller mines, planning and control of mines by computer simulation, an improved mine valuation procedure and means for mining at a decreased stope tramming width.

Jan 1, 2014

The peak in Sulphur price late 2008 caused again intensified investigation into the decomposition of sulphates and spend acid. One option is the recycling of sulphur from spend acid, which already started more than 40 years ago in the field of spend acid from organic processes in refineries and acrylic glass manufacturing. Since this acid couldn't be neutralized and discarded due to high organic contents, another source of SO2 to produce sulfuric acid is the decomposition of metal sulfates. The decomposition of metal sulfates was investigated at the next peak in Sulphur price in the 80?s. The first major scale pilot tests and studies of gypsum decomposition were carried out for clients in Florida. The decomposition of gypsum was quite often mainly influenced by the strategic decision of some countries to get own and independent acid production. In other cases the industry doesn?t have enough space to dump the gypsum. Therefore the company was also forced to convert the gypsum into ?saleable? products to avoid these residues. Other projects were made for the recycling of iron-sulfates from the pigment industry. The motivations for these projects were not only given by acid price. For the German pigment industry it was no longer permitted in the 80?s to dump the iron sulphate residues into the sea. These regulations forced the industry to develop new technology, which allows converting of the FeSO4-residues into iron oxide and acid. Therefore economical and mainly environmental factors influenced the decision of the companies to develop and to build these plants. Beside the above mention iron- and calcium-sulphates many other process exist, which produce metal sulphates as an intermediate product. These cases are not forced by environmental or strategic factors. In cases like Ni, Co or Mn-decomposition the main product of the plant is treated in the decomposition plant and later on used as ferrous-alloys. These final products are higher in price compared to e.g. CaO, which exclude the high-energy consumption for the endothermic chemical decomposition reaction. The reaction of SO2 generation based on acid or sulfate decomposition takes place in the range of 1000°C. The energy demand of the endothermic reaction is considerable and leads due to the high fuel-combustion off-gas stream to a diluted SO2 gas. Energy prices are usually increasing at the same time like sulphur. Sulphuric Acid based on decomposition is therefore more expensive than as acid based on sulphur combustion. The paper describes fluid bed technology and possible process concepts for decomposition plants. Additionally the thermodynamic background of the decomposition process and the main influence factors are provided to receive a better understanding of this technology.

Jan 1, 2009

By H. Moslemi

Metal sulphides are usually semiconductor and their electrical conductivity is low; therefore they are sources of electrons and promote electrochemical reactions. This phenomenon is of significant importance in sulphide flotation. The formation of a hydrophobic surface on the sulphide minerals is the result of various electrochemical reactions. Potential difference in the mineral-liquid interface plays an important role in the kinetics of the reactions, and consequently, on the flotation behaviour of sulphides. The value of the potential difference is influenced by different oxidation-reduction reactions at the interface area, and is measured by a two-electrode system in which an electrode is applied as reference electrode. In this study, we attempted to measure open-circuit potentials of pyrite, pyrrhotite, and steel electrodes with respect to the reference electrode. To verify the results, two reference electrodes, saturated calomel and gold, were used. Time-potential relations with respect to both reference electrodes show that the potential of both pyrite and pyrrhotite increases with time; however, this trend decreases for steel except when at pH 12. The results show a potential drop by an increase in pH value, except for steel, for which the opposite holds. This paper discusses galvanic interactions between minerals and grinding media in the grinding circuits, as well as among other minerals in the flotation processes.

Jan 1, 2012

The peak in Sulphur price late 2008 caused again intensified investigation into the decomposition of sulphates and spend acid. One option is the recycling of sulphur from spend acid, which already started more than 40 years ago in the field of spend acid from organic processes in refineries and acrylic glass manufacturing. Since this acid couldn't be neutralized and discarded due to high organic contents, another source of SO2 to produce sulfuric acid is the decomposition of metal sulfates. The decomposition of metal sulfates was investigated at the next peak in Sulphur price in the 80?s. The first major scale pilot tests and studies of gypsum decomposition were carried out for clients in Florida. The decomposition of gypsum was quite often mainly influenced by the strategic decision of some countries to get own and independent acid production. In other cases the industry doesn?t have enough space to dump the gypsum. Therefore the company was also forced to convert the gypsum into ?saleable? products to avoid these residues. Other projects were made for the recycling of iron-sulfates from the pigment industry. The motivations for these projects were not only given by acid price. For the German pigment industry it was no longer permitted in the 80?s to dump the iron sulphate residues into the sea. These regulations forced the industry to develop new technology, which allows converting of the FeSO4-residues into iron oxide and acid. Therefore economical and mainly environmental factors influenced the decision of the companies to develop and to build these plants. Beside the above mention iron- and calcium-sulphates many other process exist, which produce metal sulphates as an intermediate product. These cases are not forced by environmental or strategic factors. In cases like Ni, Co or Mn-decomposition the main product of the plant is treated in the decomposition plant and later on used as ferrous-alloys. These final products are higher in price compared to e.g. CaO, which exclude the high-energy consumption for the endothermic chemical decomposition reaction. The reaction of SO2 generation based on acid or sulfate decomposition takes place in the range of 1000°C. The energy demand of the endothermic reaction is considerable and leads due to the high fuel-combustion off-gas stream to a diluted SO2 gas. Energy prices are usually increasing at the same time like sulphur. Sulphuric Acid based on decomposition is therefore more expensive than as acid based on sulphur combustion. The paper describes fluid bed technology and possible process concepts for decomposition plants. Additionally the thermodynamic background of the decomposition process and the main influence factors are provided to receive a better understanding of this technology.

Jan 1, 2009