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By Zetti Gavelán

Lack of skilled workforce in mining industry during the current decade has been a serious concern for most mining organization globally. At the beginning of this decade the major envisioned difficulty was related to generational breach of mining professionals. By 2003, mining industry recover from a low metal prices cycle following by a worldwide mining investment boom in 2005 that triggers a talent war among mining companies that was not only limited to professional level but skilled operators as well. Despite world?s economy collapse in October 2008, impact on mining industry was small comparing to other industries and appears to show evidence of healing. Under this context, fundamentals that have been driven lack of skilled workforce for the mining industry in previous years still valid. Starting in 2007, Minera Argentina Gold SA - Barrick Veladero, subsidiary of Barrick Gold, initiated an innovative selection and training program for candidates without exposure to mining. Following our corporate social responsibility mandates, Barrick Veladero decided to change its hiring process entirely, concentrating on selection and training of local people with no experience rather than hire experienced mining operators. This paper shares Barrick Veladero experiences on its hiring program, challenges faced, and current results.

May 1, 2010

By Hongfei Li, JianLong Wang

This paper introduces the technical characteristics, transformation scheme and production practice of the project in which Copper P-S Converter was replaced by Bottom Blowing Continuous converting for the first time, and introduces the technical achievements obtained by the project. The application practice proves that bottom blowing copper continuous converting technology can completely solve the problems of fugitive off-gas pollution in the ladle transportation, large amount of flue gas and fluctuation of flue gas caused by P-S converting. Comprehensive energy consumption and operating costs of producing blister copper, safety and environmental risks are greatly reduced, and economic, environmental and social benefits are significant. It is one of the preferred technologies for upgrade and to replace P-S converting technology. INTRODUCTION At present, more than 80% of copper production capacity is still using the P-S converting technology with a history of more than 100 years. During P-S converting process, melted copper matte needs to be transported by ladle, and it needs to rotate the furnace body several times, causing a lot of serious environmental and safety risks such as SO2 flue gas emission and low-altitude SO2 pollution. In addition, the high operating cost caused by low rate of air blowing, the short life of the refractory materials and high energy consumption during P-S converting become big problems. Other problems such as large amount of flue gas, low SO2 concentration, large fluctuation of flue gas volume and SO2 concentration, and discontinuous flue gas cause high cost of subsequent flue gas treatment are also needing to be solved. In order to solve these problems during the P-S converting, metallurgical engineer in various countries has spared a lot of efforts to develop continuous blowing process for decades. The continuous converting that have been industrially applied are flash continuous converting, Mitsubishi continuous converting, Noranda continuous converting, Bottom Blowing Continuous converting and multi-lance top blowing continuous converting

Jan 1, 2019

By S. Goodman

Successful exploration for and exploitation of a mineral deposit requires an accurate understanding of the deposit architecture in 3D, and of the development of the mineralizing system through time, which requires a thorough comprehension of the structural framework of the deposit. The integration of geological, geophysical and geochemical datasets to arrive at a 4D structural framework is well-established and has been successfully applied by SRK Consulting in projects from deposit to regional scale. In some cases, we can go beyond this, applying the SEEBASETM technique, which utilizes a ?bottom-up? approach to basin analysis. SRK?s approach primarily relies on the interpretation of magnetic and gravity data, calibrated with many other datasets including mapped geology, topography, event histories, wells/drillholes and seismic data. The SEEBASETM model is a culmination of this data integration and is produced through an iterative process of structural interpretation and geophysical modeling, where an interpretation of basement composition and 4D structure is used to constrain depth-to-basement geophysical modeling, and the modeled depths are then used in conjunction with the structural interpretation to constrain contouring of the basement topography. The result is a rigorously calibrated, 3D view of the basement surface, which can be used to understand the basement and basin architecture, and thus serves as a predictive exploration model. This technique, which was developed for use in the petroleum industry, is now finding success as a predictive tool in mineral exploration.

May 1, 2003

By Richard Klimpel

"Over the last 15 years, there has been a number of plant scale sulfide mineral flotation tests organized by this author specifically aimed at understanding how flotation rate varies with reagent type and dosage. With regard to the Interaction of collector dosage, particle size, and frother type, the eight specific trends listed in the paper (Illustrated with data from two specific case studies) are qualitatively the same for almost all plants studied in some depth. The Interaction of flotation rate K and equilibrium (Infinite time) recovery R in the plant has been denoted as the R/K trade-off. The effect of collector dosage on the rate of flotation of specific particle size intervals is particularly interesting and once understood, is quite useful In understanding how industrial flotation plants respond to changing reagent type and dosage. The location and magnitude of the R/K trade-off can be different for various ores and various plant configurations, but the RIK trade-off is almost always present. The purpose of this paper is to clearly describe and illustrate the nature of this very Important flotation system Interaction so that operating engineers have a starting reference point for beginning successful plant Improvement studiesIntroductionIt is well recognized by industrial froth flotation plant operators that improving existing sulfide circuit performance is often difficult due to the self-compensating systems nature of the froth flotation system, e.g. Klimpel (1988, 1989, 1995). Extensive plant testing with existing and new reagents by this author throughout the world has clearly shown that one of the more important self-compensating mechanisms is the Interaction between collector dosage, frother type, and particle size. The term self-compensation in flotation simply means that as the plant operator consistently changes one parameter in the same direction, the nature of the froth flotation system Is such that the circuit will eventually begin to resist this controlled parameter change by automatic (uncontrolled) changes In other parameters and the overall circuit performance will go through a maximum or minimum (hence causing the operator to stop the parameter change)."

Jan 1, 1997

By M. A. T. Cocquerel, M. G. Burcher, J. G. Eacott

"This paper is based on work carried out to examine the feasibility of increasing the smelting rate of two oil-fired calcine-charged reverberatory furnaces . Various constraints prevented physical enlargement of the furnaces and so allention turned to the use of preheated combustion air or oxygen enrichment to bring about the required increase in production. As a basis for further calculation, a furnace heat balance is determined using the smelting metallurgy and combustion data for the current operation. Various other operational data are tabulated and compared with similar data from other operating copper smelters.New heat balances are derived for various levels of combustion air preheat at the current smelting rate, and the effects of air preheat on oil consumption, combustion air and waste gas volumes are tabulated. These data are used to demonstrate the level of increased throughput which would be obtained before constraints in existing gas handling equipment are encountered.A similar exercise for oxygen enrichment is set out, with oil consumption, combustion air and waste gas volumes at various enrichment levels tabulated for the current smelting rate. These data are in turn used to demon strate the smelting rate which would be achieved at a practical level of enrichment before waste heat boiler and gas handling constraints are encountered.The energy implications for both cases are summarized."

Jan 1, 1980

By Mika Muinonen

"Dewatering of nickel concentrate at the INCO Thompson integrated mine-metallurgical site takes place within the smelter plant. The eight original 60 foot diameter cable torque thickeners of which four were operational and eight Dorrco internal drum filters have been replaced with a single Outokumpu high rate thickener and a single Lasta pressure filter. The modernization has resulted in a filter cake which has a more stable and lower moisture content. Control of the entire dewatering process has been improved and costs have been reduced.Typical filter cake produced from the Dorrco filters during processing of high tonnages of Birchtree mine ore was ~20% moisture vs. ~13% moisture for concentrate produced from the Thompson mine ore. The high moistures were correlated with the amount of fine magnesium silicate gangue in the concentrate. As Birchtree ore processing rates have increased, this filter cake moisture became unacceptable because of material handling problems and the impact on furnace metallurgy. The Dorrco filters were replaced with the Lasta filter in 2003 that has consistently produced filter cake at ~11% moisture.This paper summarizes the piloting, design, installation and commissioning of the modernized dewatering circuit.INTRODUCTIONThe Thompson integrated nickel processing facility was built in 1958 and originally consisted of a mine-mill-smelter and, soon after, a refinery.Nickel concentrate is pumped from the mill as a dilute slurry directly to the smelter facility ~200 meters along an unheated pipe trestle. Originally, nickel concentrate was de-watered using up to 8 conventional 60-foot diameter thickeners and 8 internal vacuum drum filters."

Jan 1, 2006

By Willemijn Husslage

Tata Steel Group At a glance ? Part of the Tata Group ?Tata Steel Group: ?Fortune 500 company ?Top 10 global steel producer ?Annual crude steel capacity of 28 million tonnes ?More than 80,000 employees ?Manufacturing operations in 26 countries across 5 continents ?Serving customers all across the globe ?Turnover 2009-10: $22.8 billion

May 1, 2011

By Mike Robotham

When RFQ?s are received they often contain detailed technical specifications which suggest a departure from the suppliers own customary approach to the design. Both client and end user are quite rightly influenced by their own experiences. The question is who is right, the client and his consultant, the supplier or perhaps a combination of these. This paper looks at some of the fundamental variances that often occur in the approach to the mechanical design of drum hoists, based on our experiences of recent years. Director of Engineering, Mike Robotham is a chartered engineer and has been with DavyMarkham for over 30 years. Having completed a craft apprenticeship he later joined the engineering department as a junior engineer. He has a wealth of experience of working on medium to large engineering projects in the field of mechanical, electrical and hydraulic equipment

May 1, 2011

By Philip A. Dirige

This paper presents the variability of responses of paste backfill specimens that have been prepared with varying binding agent and binder combinations, when subjected to uniaxial compressive strength testing. A testing program conducted for an underground operating mine to produce cost efficient paste fill recipes using four different binding agents, in various ratios and combinations, is presented to illustrate the responses. These binding agents are Normal Portland Cement Type-10 (NPC T-10), oxidation or sulphide attack resistant binder (Binder A), replacement binder B (Binder B), and replacement binder C (Binder C). The paste fill recipes were prepared at 2, 3.5 and 5% binder content at varying binder combinations, at fixed size and tested in a Material Testing System (MTS) under constant strain rate conditions following cure times of 7, 14, 28, 56, 90, 112, 180 and 240 days. The results of tests suggested that the strength data clearly shows a trend of increasing strength with increasing binder content but not necessarily cure time. It is noted that all the blends achieved dramatic increases in strength up to 28 days of cure and then either increased or decreased moderately beyond 28 days curing, depending on the binder content and binder combination. The moderate decrease is apparent for the 2% paste fill blends and in some binder combinations of the 3.5 and 5% blends, which indicated lackluster performance beyond 28 days curing. This lackluster strength performance indicated that the binder content and combination may not be sufficient for the paste fill to achieve complete hydration and acquire adequate strength performance beyond 28 days. This study indicated that economic paste fill recipes can be designed with confidence using affordable alternative binders. However, it is noted that these alternate binders cannot be used as stand alone binders but should be used in combination with NPC to facilitate hydration of the paste fill.

May 1, 2006

By Franciszek Letowski

"Recovery of precious metals associated with a sulphidic matrix has an important economic effect on complex sulphide ore treatment. Because a significant portion of the silver and gold associated with pyrite are rejected to tailings during the selective flotation of pyritic Zn-Pb-Cu ore, other ore processing routes have been investigated.Direct leaching of the ore with ferric sulphate solution followed by separation of different components of the solid suspension by flotation was applied. Such a process allows for: production of zinc sulphate solution; transformation of galena PbS to insoluble lead sulphate PbS04 and its accumulation in a non-floatable fraction together with precipitated geothite FeOOH and other insoluble gangue minerals; separation of sulphur and non-decomposed sulphides in the floatable fraction.Laboratory research carried out on the ore samples from an Alaskan pyritic deposit from the Delta region indicates that gold and silver are mostly concentrated in the floatable fraction.Total recovery of zinc from the ore attains 96.5% to 98.5% after ferric sulphate leaching. About 85% of copper was extracted from the ore during the ferric sulphate and subsequent ferric chloride leaching, and about 6% of the copper remained in the non-floatable fraction and mixed fraction during the leaching-flotation step. Extraction of lead from the nonfloatable fraction during leaching with calcium chloride solution attained 98% to 99%. Extraction of silver from the floatable fraction during leaching with ferric chloride solution reached 98%. Extraction of gold from the floatable fraction during cyanide or thiourea leaching reached 71% to 89%. The results of the laboratory research are presented."

Jan 1, 1987

By Richard Morland

Some simple and fundamental objectives To challenge and inspire our people to embrace safety as a value ?A stand for safety as a value ?A stand for our people ?A stand by our people ?Prevent harm to our people ?Eliminate suffering for individuals, families, friends, colleagues

May 1, 2010

By Tony Robles

"A new process is disclosed for the recovery of gold and/or silver values from activated carbon wherein they are adsorbed as cyanide complexes. A unique chemistry allows the new process to use mill process water containing an inexpensive, non-hazardous compound as the eluant. The new process has an elution period of 5 hours at 98°C. The gold in the eluant can be recovered by electrowinning or zinc precipitation. The new elution process is simple, reliable, and provides an eluted carbon that may not require thermal regeneration before reuse. Retrofitting the new process to existing processes such as Zadra, Anglo, and others will not require significant capital expenditures.IntroductionThe efficient and reliable elution process for gold/silver from activated carbon is a very important process in the gold mining industry. The most common techniques for elution of gold cyanide from activated carbon are the Zadra and Anglo procedures.Specialized elution techniques like the Micron and Acetonitrile processes have not been commercialized to significant extent probably due to relatively complex unit operations and expensive reagents.Important functions of the gold recovery circuit are:Efficiently recover gold/silver from the loaded carbon.Carbon is in good condition for reuse for optimum recovery.The pregnant solution is amenable to electrowinning or zinc precipitation to recover gold/silver from the solution.The whole process to be simple and short to reduce gold inventories.Present Elution ProcessesTable 1 shows some carbon elution processes that are available today. The two main elution processes used in the industry are: (1) the Zadra process, in which hot cyanide solution is circulated through an elution column and electrowinning cells; and (2) the Anglo American Research Laboratory (Anglo) process, consisting of a pretreatment step with caustic cyanide solution, followed by elution with deionized water."

Jan 1, 2000

By Ernie Doell, Trung Nguyen

"In early 1995 Cluff Mining commissioned a newly constructed high density tailings thickener at its Cluff Lake uranium mill. The 85 ft diameter high density thickener is one of the first of its kind in North America, and incorporates the novel paste technology in its design to produce densified (in the range of 60 -65% solids) mill tailings suitable for stable deposition in the tailings management area. This paper discusses the impetus behind the installation of such an unconventional thickener, and the many trying experiences in making it operational.IntroductionCluff Mining, a wholly owned subsidiary of Cogema Resources Inc., is the operator of a uranium mine and mill complex at Cluff Lake which is located in northern Saskatchewan. Since 1980, the company has developed and mined out several different ore bodies within the Cluff Lake region for the production of uranium concentrates, commonly known as yellow cake. Based on the known ore reserves, the Cluff Lake operations are projected to last till the early 2000's.Tailings from the Cluff Lake mill have been discharged to a tailings management area consisting of a tailings pond and two facilities for primary and secondary treatment systems. The tailings pond has an impervious bentonite lined dam, and it is divided into a solids and a liquids holding areas. Until the beginning of 1995, the mill tailings were discharged in two separate lines to the solids pond; one line carried the slurry residue of 45-50% solids (w/w) from the CCD circuit, and the other discharged raffinate tails containing less than 1 % (w/w) gypsum precipitate. Supernatant in the solids pond is treated in the primary treatment system prior to discharging to the liquids pond. Figure 1 shows the site plan of Cluff Lake tailings management area.It has been estimated that by the end of 1995 the solids pond would be filled to about 60% of its design capacity, and probably would be full before the year 2000. Therefore, Cluff Mining had to consider a more efficient tailings disposal method that could extend the service life of the existing Cluff Lake tailings management area. A high density tailings thickener was believed to be the most practical and economical system for tailings disposal at Cluff Lake mine."

Jan 1, 1996

By T. R. Raponi

"The Williams Mine, Canada's largest gold mine, processes over 6,000 t/d of ore utilizing SAG/ball milling, leaching and CIP, carbon stripping, electrowinning and refining to produce gold dare bullion.Grinding accounts for over 50% of milling costs with grinding media, power and mill liners making up over 75% of grinding costs. Since 1991, several changes in these areas have been made which have significantly reduced costs, while fineness of grind and mill recovery have both been increased.This paper discusses the changes made in the grinding circuit over the past few years and assesses their impact on operating costs and mill operations.IntroductionThe Williams Mine is located in the Hemlo area of northwestern Ontario, about 40 km east of Marathon. It is operated by the Williams Operation Corporation, which is equally owned by Teck Corp. and Homestake Canada Inc. Canada's largest gold mine, the Williams Mine commenced operation in October 1985 at a rate of 3, 000 t/d. Mill throughputs have typically exceeded design, averaging 3, 397 t/d in 1986, and 3,380 t/d in 1987. In August 1988, production was increased to 6,000 t/d. Throughput averaged 4,212 t/d in 1988. Production increased over the next few years averaging 5,477 t/d in 1989, 6,248 t/d in 1990, 5,987 t/d in 1991, and 6,283 t/d in 1992. With a recovery of over 95.5%, production for the first three quarters in 1993 is averaging 6,408 t/d.GeologyThe main ore body is a steeply dipping conformable zone of microcline-quartz-muscovite rock. It occupies the contact zone between hanging wall metasedimentary rocks and foot wall felsic quartz feldspar porpyritic rocks of possible volcanic or subvolcanic intrusive origin. The ore body is enveloped by muscovite schists which are thickest on the footwall side. The schist is enriched with tourmaline, pyrite, and green mica containing vanadium and barium. The ore body contains pyrite and barite which occurs mostly in laminae and bands to veinlets and breccia fillings. The gold occurs as free (native) grains, 20 to 30 µm in size at grain junctions of the microcline and quartz. Visible gold is rare in the main zone. There are commonly two or more discernable layers of economic value which are separated by weakly gold mineralized material which usually consists of muscovite schist, barite rich zones or sulphide poor-microcline-quartz rock and in the upper and western fringes by biotite rich schist with feldspathic fragmental characteristics. Molybdenite is a common constituent as very fine disseminated grains in the microcline-quartz rock and is most prominent with gold enrichment. Stibnite and realgar are two of the more common accessory sulphides."

Jan 1, 1994

By V. Lakshmanan, Dan Mackie, J. Melnbardis

"Crystalline Flake Graphite is a strategic mineral. Since 1954 nearly all North American consumption has been imported from offshore. The Bouthillier Project in Quebec, Canada, is Orrwell Energy's bid to enter the market, with plans to begin production of a 500 TPD operation in 1984.Milling of flake graphite is considered easy, yet more problems have been encountered in flowsheet and plant design than in most other minerals. A fresh approach to existing technology was taken, giving results that are highly satisfactory. The key to graphite liberation without flake deterioration is found to be in autogeneous grinding. In addition to producing marketable coarse graphite by such autogeneous grinding, flotation and tabling, super purity fines can be produced using repeated regrinding, elutriation and flotation. This paper discusses the design of a 500 TPD plant based upon the results obtained from bench and pilot plant studies. INTROOUCTIONWhen the Black Oonald mine at Calabogie Ontario ceased production in 1954 nearly all North American crystalline flake graphite had to be imported from offshore. Orebodies in the United States were too small or too low in grade to be exploited economically.The Houthillier Township near Mont Laurier in Quebec (Figure 1) was a known graphite-bearing area The Orrwell property in that area was delineated in 1957 Some test work including minor footage of diamond drilling was carried out by Italia Copper and the MHM Syndicate prior to 1981 when Orrwell acquired the property"

Jan 1, 1984

By Steve Massman, Edward Wipf, Gary Beerkircher

"As mine size and tonnage continues to increase, material handling/processing equipment has also increased in size with 400 MT haul trucks, 2000 KW crushers, 12 M SAG/AG mills, 8.5 M ball mills, 3000 M3 flotation cells and 12.5 MW HPGR’s all with planned designs to increase in size further. Related equipment including vibrating screens have also increased in size with screens seeing feed rates of 3000 MTPH. Open circuit SAG Pre-Crush and closed circuit HPGR flowsheets are becoming more common with large screening requirements. Large multi-slope (banana) or horizontal screens are commonly used for these applications, however as the size of these screens has increased, the number and size of the brute force/overhead exciters has increased but with limitations. An example is a SAG discharge screen that will see surges of material from the SAG mill which overload the screen, reducing the screen stroke or throw causing a large reduction in screen efficiency with excessive fines & moisture in the screen oversize feed to the pebble crushing circuit. This paper will present some actual examples of the use of Two-Mass screens as a solution to these overloading problems and the advantages of Two-Mass screens in other high tonnage applications.Brute Force Drive Compared to Two Mass Drive Vibrating ScreensThe standard single mass screen as depicted by figure 1 is generally recognized in the mining industry. It is characterized by the screen body – labeled as the single mass (Mt) that is supported by isolation springs (Kiso). The stroke needed to process material is obtained by incorporating large eccentric wheels and/or drive boxes (Fb) directly mounted to the screen body via a drive tube."

Jan 1, 2017

By M. Massinaei, M. R. Yarahmadi, S. Zeidabadi, M. Oliazadeh

"Although the majority of new flotation column installations continue to be for cleaning or recleaning, an increasing number of flotation projects are specifying columns for roughing or scavenging duties. NICICO has recently installed flotation columns operating in rougher, cleaner and re-cleaner circuits at three new copper concentrators; Miduk, Sungun and Sarcheshmeh. Miduk is probably one of the unique operations that use flotation columns in a rougher circuit. The Miduk rougher circuit comprises five CISA parallel columns. The rougher concentrate constitutes 88% of final concentrate and reports to concentrate thickener. The rougher tailings are scavenged with ten 50 m3 mechanical cells. The scavenger concentrate is then reground and cleaned in three 3.2 m columns in parallel. The cleaner tailings are scavenged in six 50 m37mechanical cells. The cleaner tailings along with the rougher scavenger constitute final tailings.In Sarcheshmeh, the cleaner concentrate is re-cleaned in a column. The column tailings are sent back to the cleaning circuit and the concentrate is reported to the concentrate thickener.This paper reviews performance details of the flotation columns utilized in rougher circuits at Miduk and compares their metallurgical performances in an industrial scale with columns at cleaning circuits for copper flotation. Practical limitations for applying the flotation columns in rougher circuits will be discussed in details.INTRODUCTIONThe use of flotation columns has grown rapidly since the first successful commercial application at Mines Gaspe, where columns were used in a cleaning circuit to float molybdenite. Early applications of columns were restricted to replacing or adding final cleaning stages in existing circuits. Columns can achieve superior cleaning as compared to conventional cells due to their ability to support a large froth depth and the use of water for washing froth. Wash water travels down through the froth and crosses the pulp/froth interface, replaces the entrained particles reporting to the froth from pulp and improve the concentrate grade."

Jan 1, 2011

By Mike Sapiro

"Heath Steele Mines is a base metal producer situated near Newcastle, N.B. presently treating 3500 TPD of complex Cu-Pb-Zn ore. The ore consists of massive fine grained sulfides averaging I. 1% Cu, 1.6% Pb and 4.5% Zn.The original mill started in 1956 at 750 TPD and was successively expanded through 1500 to 3000 TPD in 1969. In 1972 an expansion to 4000 TPD was initiated. It was planned to obtain the maximum benefit from equipment and manpower by adding some process equipment, instrumentation, an on stream analyser, and a process control system. The benefits from this approach could be greater than normally expected, because of the complex metallurgy of this ore.IntroductionThe crushing plant is of conventional design with two 5 ½ ft, Symons cone crushers with the crushed ore screened to -7/8 inch on a 61 x 16 1 Al I is Chalmers vibrating screen and stored in four 1000 Ion bins. There are two para I lei grinding circuits, each one consisting of the following: one Rod Mill in open circuit, one Ball Mill in closed circuit with two 20"" Krebs cyclones, and one Ball Mill in closed circuit with three 15"" Krebs cyclones, Al I the Mills are 9' x 121 Al I is Chalmers Mills. The secondary cyclone overflow from each grinding circuit is thickened and combined to form flotation feed. The final product is 65% - 400 Mesh. Two regrind mills have been installed as part of the expansion."

Jan 1, 1977

By Donald Strickland, Peter Kresin, Dave Hall

"Have your parents or grandparents ever tried to share with you their wisdom and you ignored it because it did not relate to you‘? I have been ignorant and inevitably grew older and wished I had listened. The result: learning the hard way. As British Astronomer Fred Hoyle once said. “It is the true nature of mankind to learn from mistakes not from example.“ Hopefully we will continue to learn from the examples shared by our fellow mineral processors in these conference proceedings. It is the intent of this paper to share wisdom from the Musselwhite Mine that you may apply to your own operation.This paper is divided into three sections:1. Feasibility3. Operation3 . DesignKey points of the Feasibility and Design stages are summarized. followed by an analysis of actual operating statistics. An attempt is made to relate these statistics back to the feasibility and design with emphasis on areas of success and those with potential for improvement."

Jan 1, 1999

By Claude Dufresne

The technical development of the niobium-bearing products is to a great extent influenced by the striving of the respective industries to maintain or strengthen their positions in all fields of application against competing materials. South East Asia, China, Eastern Europe and CIS are slowly becoming significant consumers of ferroniobium. As a result, the world niobium market has been stable despite the slowdown in steel production in 2001. This paper will describe present producers, capacities and the current market conditions.

May 1, 2003