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By Alex Chapman, Cam Thomas, Bryan Karney

"During the early summer of 1989, Ontario Hydro removed a concrete portion of the North Channel Dam at Crystal Falls Generating Station by blasting and excavating in marine conditions. Because of the severity of underwater blasting and because the site was in close proximity to the replacement dam, there was concern that hydraulic shock waves generated by the detonation could damage either the new dam or its control structures. This paper examines the data associated with the hydraulic shock waves generated by underwater blasting, and studies the effectiveness of an air curtain in reducing the peak pressures. In previous trials using the air curtain technique, a rule-of-thumb factor was used to estimate that only 10% of the original pressure pulse would propagate across the air curtain. However, this value had little basis in theory. A theoretical formulation of the problem by Karney in 1989 indicated a provisional value of 0.18 for this factor, but was heavily dependent on confirmation from in-field testing. Data provided by Ontario Hydro from an actual blasting procedure indicate that a more conservative value of 0.32 may be more appropriate. Thus.although the air curtain did provide considerable reduction of peak pressures, the data at Crystal Falls indicates its performance was below the anticipated level. However, due to problems with the instrumentation and some inconsistencies in the resulting data. it is recommended that this value be further refined as additional data become available."

Jan 1, 1992

By J. B. Oosthuizen

Mechanical rock breakers are used in block caving operations to reduce the size of ore fragments which are too large to pass through grizzlys at the load-haul-dumper tipping points. Several layouts of grizzly, rock breaker and ore pass are possible. The authors evaluated three different layouts at the Premier Mine Division of De Beers Consolidated Mines. The first layout consisted of a 15m surge bin before the breaker, which was located 15 m below the production level. The second layout had a 43 m longs urge bin, with the breaker located 43 m below the production level. The third layout did not have a surge bin, the breaker was located on the production level and operated at the tip. The ore throughput, availability and utilisation, delays, costs and operational experience with each system are discussed. It was found that the rock breakers did not represent a bottleneck in the ore transfer system and a surge bin did not significantly improve the production capacity. The mine has decided to make use of rock breakers situated on the production level, breaking rocks at the tip.

Jan 1, 1997

By James L. Ash

This report contains the results of a technical and economic evaluation of conceptual systems for the underground mining, rubblization, and in situ retorting of deep oil shale deposits. Study efforts were first directed towards evaluation of the applicable geology, associated minerals, rock mechanics, in situ retorting, access mining, ventilation, rubblization, safety, and environmental impact factors. Conceptual systems were synthesized and then evaluated against these factors from both a technical and limited economic viewpoint. The results of these evaluations were then analyzed and, based upon this analysis, the following two systems were selected for further study: System II: Room and Pillar-Vertical Drill and Blast System IV: Tunnel Boring-Horizontal Ring Drill and Blast Due to the large number of variables to be considered, a computer program was developed and utilized to provide data for the cost trade-off studies and economic analysis. Results of the early computer runs indicated the value of the tool available? in the computer program and led to an extension in the contract to determine the sensitivity of the systems to several of the more apparent significant variables. These results in turn led to a refinement in the design assumptions and a better understanding of the technical and economic options available to an owner-operator of a Vertical Modified In Situ Shale Oil Extraction System. This report is submitted in fulfillment of the requirements of Contract Number 50241073 by Fenix & Scisson, Inc. to the Denver Mining Research Center, United States Department of the Interior, Bureau of Mines. Work was completed as of November, 1976.

Jan 1, 1976

By L. Lui, A. Cauchon, A. R. Laplante

"The gravity circuit at Les Mines Camchib Inc. consists of two pinched sluices to bleed a feed to one of two 76 cm (30"") Knelson concentrator. The Knelson concentrate is upgraded to smeltable grade with a smaller Knelson (18/30 cm) and a riffleless table. Their tails are column cyanided and washed prior to recycle to the grinding circuit. This paper details size-by-size circuit performance. The Knelsons recover around 90% of free gold (as determined by Mozley Laboratory Separator), even below 37 µm. The sluices are ineffective, with upgrading ratios around 2, but approaching unity in the fine sizes. Much of the effectiveness of the circuit lies in the high gold circulating loads in the grinding circuit, which partly atone for the poor sluice performance, and the ability of the Knelsons to recover free gold. Secondary upgrading recovers slightly above 99% of the primary concentrate gold, 90% of it in a 65% Au concentrate and the balance in a heavily loaded (10 to 20 oz/st) pregnant solution. INTRODUCTIONBackgroundThe operating performance of the mill of Les Mines Camchib Inc. has been described earlier (Laplante et al., 1989). The presence of significant chalcopyrite (1 % and up) makes flotation necessary prior to cyanidation, to limit cyanide consumption and minimize environmental impact. In 1982, the presence of very high gold circulating loads in the grinding circuit lead to the commissioning of a gravity circuit, schematically shown in Figure 1. Gravity recovery of gold has been found beneficial because it increases metallurgical recovery by 1-3% (Laplante et al, 1987), as only part of the flotation tail is cyanided, and economic recovery, because of the lower net smelter return of gold in the flotation concentrate (as compared with the gravity concentrate). The latter increase is estimated, with a net smelter return of 95%, at 1.5% when gold recovery by gravity is equal to 30%. In early 1988, a sampling program was initiated to characterize the behaviour of gold in the grinding/gravity circuit. This paper reports on the results of this study."

Jan 1, 1990

By Wm. F. Netzeband

"The mode of ore occurrence and the methods and costs of mining at one of the zinc-lead mines in the Tri-State zinc and lead district, Oklahoma, are presented in this paper for the information of mine operators in other districts.No. 2 mine property was acquired by lease in 1918 after some 30 holes had been drilled and a shaft had been sunk to the 305-foot level by the first lessees. After acquisition of the lease, the present operating company drilled a rich ore body on the 260-foot level and in 1920 started sinking two shafts. Production was begun late in 1921, and the mine has been operated more or less continuously to the present time.The mill has a capacity of 400 tons in 10 hours, the usual length of the mill shift in the district.The geology of Mine No. 2 is very similar to that of Mine No. 13 except that the Cherokee shale is thicker, ranging from 80 to 160 feet, and all the formations are correspondingly deeper, the Short Creek oolite horizon here being at a depth of about 300 feet. The oolite has not been found on this property, but the lime comes in at about the level at which the oolite should appear.The ore deposits of Mine No. 2 occur in the brecciated and bowlder ground above the Short Creek oolite. The main or 260-foot level is about 40 feet above the oolite horizon, and the lower or 300-foot level is immediately above it.The ore deposition is controlled by a zone of shearing and fracturing, and the mineable areas are divided into two distinct runs which have a northwest trend with a lean or practically barren area dividing them. This characteristic has controlled the mining to a large extent, for it enables the mine superintendent to locate the large pillars in lean ore and leave only small pillars in the richer ore. Figure 1 shows the outlines of the ore body and the pillars.The ore on the main level occurs disseminated throughout the jasperoid breccia as patches or lenses in the jasperoid, or cementing massive bowlders of jasperoid breccia and chert. The ore of the lower level is more of the bowlder type than that on the main level and occurs chiefly cementing the bowlders but also with some disseminated ore in the jasperoid breccia.The gangue minerals are jasperoid breccia, chert, and dolomite. Calcite is found in minor quantities. Sphalerite is the principal ore mineral; galena is found only in minor quantities around the edges of the ore body. When galena is encountered in this ore body, the limit of the ore can be expected.The exploration and estimation of ore reserves at Mine No. 2 are similar in many respects to the methods used at Mine No. 1.4After the mine was opened up and the general character of the ore deposit had been determined, drilling operations were concentrated along the trend of the shear zone. Later drilling has proved what appears to be a parallel shear zone or an offset of the zone now being mined."

May 1, 1929

The study of gas emitted from face coal samples at the coal face is part of an investigation of instantaneous outbursts of coal and gas. Normally, seam gas essentially comprises methane of which the origin in the metamorphism of coal is undoubted. The origin of other gases sometimes occurring in high proportion in seam gas, particularly carbon dioxide, is more controversial. A simple meter for rapidly assessing gas in coal was developed. Tests with the meter enabled comparisons of seam gas in face coal at some outbursting and nonoutbursting mines, and under various working conditions. The varying results of widespread testing suggest certain trends, including some prospect of recognising outbursting potential at the coal face.INTRODUCTIONThe aim of this work is identification of differences in gas condition between coals subject to instantaneous outbursts and coals which have not outburst, and between faces of coals which will and which will not subsequently outburst. Any comparative test should be conducted right at the coal face on coal about to be worked and the results of testing should be available after a short period. Moreover, the method and the comparison of results should require no special skill, and be well within the capacity of junior mine officials. Any consistent relation between gas condition and subsequent mining conditions - outbursting or not might be valuable in prediction of outbursts.INSTANTANEOUS OUTBURSTSInstantaneous outbursts (Hargraves, 1958) , are largely unpredictable and unaccountable ejections of coal (up to 5,000 tons) and gas (up to 5,000,000 cu.ft.) from a standing coal face. Usually the gas is essentially methane (CH4 ) but may be essentially carbon dioxide (C02 ), or an intermediatemixture...

Jan 1, 1962

By Steve Kral

When the price of gold cracked the US$9.60/g(US$300/oz) mark last year, the gold-mining industry began to crank up again. This followed several years of gold prices that had languished in the mid-200s. Existing mines increased production and mines that had lain dormant for many years were getting second looks. The United States still produces a large amount of gold. But not much of the new production is coming from that country. Instead, strict environmental regulations still have companies wary of trying to invest the millions needed to start a new mine or expand existing ones. And after years of studies and negotiations between legislators, environmental interests and the mining industry, the United States still has not crafted a new mining law.

Jan 1, 2003

By Joram M. Amir

Even when they are cased! [ ] In bored piles, as well as in drilled shafts, and in caissons too, defects will occur! [ ]

Jan 1, 2001

By W. C. Joughin

Problems with the stability and performance of rockpasses in deep-level mines are common. This paper presents the results of an investigation into the recorded performance of rockpasses in deep-level gold mines. Records relevant to the geological environment in which the passes are excavated, their excavation details, and their historical performance are not at all well documented on mines. In general, the quality of data available from mines is poor. The investigation showed that very large spans frequently developed in passes during use, and that rock mass quality, the orientation of the pass with respect to the dip of the strata, and the magnitude of the stress acting normal to the axis of the pass were the main factors of influence. From the historical information obtained, the risks of instability in rockpasses are identified.

Jan 1, 2005

Effective gas management in coal mines, as well as proper ventilation design, is very important for maintaining the safety of underground coal miners. Advances in numerical modeling techniques have enabled evaluations of the coal mining environment using advanced grid designs and computed gas distributions through detailed mathematical models. These models help engineers to see the unseen areas of the mine through visualization of the data and ultimately help improve worker safety. This paper summarizes the modeling efforts conducted by NIOSH in the U.S. and CSIRO in Australia to address various gas management issues encountered in longwall mines. The modeling studies involved two of the most widely used grid-based techniques, namely computational fluid dynamics (CFD) and numerical reservoir modeling. This paper discusses the application of these techniques for gob inertization and gas control for spontaneous combustion, pre-mining degasification of coalbeds, gob gas venthole design and performance evaluation in longwalls, and gas management during development mining.

Jan 1, 2009

By Finlayson EJ

Epithermal gold is hosted by late Middle Miocene Okiduse Volcanics on Woodlark Island. Dating of associated alteration minerals indicate mineralisation to be broadly contemporaneous with volcanism (12.2 - 12.5 m.y.) and is probably related to subduction of the Solomon Sea plate along the Trobriand Trench. Mineralisation is polymetallic and structurally controlled. Pyrite, sphalerite, galena and minor chalcopyrite occur in steeply dipping fracture zones and/or hydrothermal breccia bodies associated with mixed clay-pyrite alteration envelopes. These envelopes occur within a terrane of regionally propylitised high-K, calc-alkaline andesites and related epizonal subvolcanics. Gangue minerals include quartz, calcite, dolomite and minor anhydrite. High level advanced argillic alteration and siliceous cap rocks are not present but may have been eroded off the top of the system. Fluid inclusion and stable isotope data suggest that the carbonates and paragenetically associated auriferous base metal sulphides probably deposited from an isotopically evolved (+8%. 180) reduced hydrothermal fluid sourced from underlying sedimentary rocks. The fluid was predominantly of low salinity (0-4 wt % NaCl equivalent) and had a temperature of about 270¦C at the time of boiling and mineral deposition.

Jan 1, 1987

By S. J. Duda

Procedures for detecting and locating seismic sources are adapted to the case of entrapped miners. During an emergency the entrapped miner is likely to generate seismic signals, by knocking against the walls of the shaft. With a seismic array station deployed near the disaster area, the seismic signals can be recorded. Two algorithms are presented for the location of the source from the arrival times of a seismic phase: The first algorithm produces one, most probable location, under the assumption of constant velocity of propagation of seismic waves in the medium. The second algorithm is similarly applicable, however without any restriction on the seismic velocities in the medium between the source and the array. While the first algorithm uses an analytic approach, the second one is based on a trial and error procedure. The second algorithm gives locations with arbitrarily high accuracies, depending only on the accuracy of arrival time measurements and on the compatibility of the true velocity distribution and the assumed velocity model. The algorithm is shown not to require under certain conditions direct .arrival time measurements, thus eliminating a potential source of errors in the location of entrapped miner.

Jan 1, 1973

By K. -H. Lux, Z. Hou

ABSTRACT: It is well known in underground mining and waste disposal that the development of excavation damaged zones (EDZ), and thus, hydraulic pathways, is closely related to the excess in the corresponding damage boundary (= long-term strength) of the rock mass and the resulting damage. By integrating short-term and long-term strength, dilatancy, damage and healing into the newly extended Hou/Lux constitutive model for rock salt, it is possible to treat the tertiary creep, to identify the EDZ around underground facilities and to predict the time-dependent development of the EDZ in rock salt. Under in-situ conditions, the EDZ of a 37-year-old drift at the Sondershausen salt mine and the EDZ around a drift dam in rock salt are numerically examined. The numerical analysis gives satisfactory results. RÉSUMÉ: Dans les exploitations minières et le stockage souterrain des déchets, le développement des zones endommagées et fracturées (EDZ – excavation damaged zone), créant des chemins préférentiels pour écoulement, est étroitement lié au contexte géotechnique et à la méthode d’excavation. En intégrant la résistance court et long terme, la dilatance, l’endommagement et le colmatage, le nouveau modèle rhéologique Hou/Lux pour le sel gemme est capable de traiter le fluage tertiaire, d’évaluer l’extension de l’EDZ et de prédire l’évolution de EDZ dans temps autour des ouvrages creusés dans le sel gemme. L’EDZ d’une galerie datée de 37 ans dans la mine de sel de Sondershausen et les EDZ autour de barrage des galeries dans un sel gemme ont été modélisées. Les résultats numériques fournissent des résultats satisfaisants. ZUSAMENFASSUNG: Bekanntlich ist die Entstehung von Auflockerungszonen und damit hydraulischen Wegsamkeiten in der unmittelbaren Umgebung um bergmännische Hohlräume mit der Überschreitung der entsprechenden Schädigungsgrenze eng verbunden. Unter Berücksichtigung der Bruchfestigkeit, Schädigungsgrenze, Dilatanz, Schädigung und Verheilung kann das Stoffmodell Hou/Lux das Verhalten des tertiären Kriechens beschreiben, die Auflockerungszone im Salzgebirge ermitteln und ihre zeitabhängige Ausdehnung vorausberechnen. Mit diesem Stoffmodell werden die Auflockerungszonen einer 37jährigen Strecke in der Salzgrube Sondershausen und die Auflockerungszonen um ein exemplarisches Abdichtungsbauwerk im Salzgebirge numerisch untersucht. Dabei werden zufriedenstellende Berechnungsergebnisse erzielt.

Jan 1, 2003

By Eric S. Lindquist

High groundwater levels combined with variable soil and rock profiles present challenges to the design and construction of deep excavation support systems. Increasingly, secant piling techniques are the preferred technical solution for shoring in these cases. Recent developments in drilling equipment, tooling and procedures allow cost-effective construction of deep, overlapped pile systems to tight tolerances in extremely difficult ground conditions. This paper summarizes critical design considerations for secant pile shoring systems, detailing recent advances in installation procedures and verification techniques that allow secant piles to be used for an expanding range of project conditions. Technically- and commercially-viable secant pile shoring designs require that stringent drilling tolerances be achieved in order to successfully provide combined structural support and groundwater cutoff systems, particularly for unreinforced circular secant pile shafts acting in ring compression. Modern drill rigs and tooling can advance fully cased shafts in geotechnical conditions which range from cohesionless soils below the groundwater table to hard rock, within a single drillhole. Down-hole instrumentation provides verification that tolerance requirements are met and allow secant pile shoring to be used at sites previously considered unacceptably risky. In California, secant pile shoring systems are being used for excavation depths of over 100 ft (30 m) on projects with challenging site conditions. Two case studies are presented detailing severe geotechnical conditions and excavation depths which represent the regional state of practice in secant piled excavation support. These projects provide real world examples of the design, construction and construction verification techniques that permit secant pile shoring to be used in more demanding geotechnical conditions and to greater depths than ever before.

Jan 1, 2011

Considerable improvements in diamond drilling equipment have been made since the early 1950s. A recent drilling programme, however, at Williamtown, near Newcastle, carried out by the Joint Coal Board on behalf of the Electricity Commission of New South Wales, emphasized deficiencies in existing equipment when particularly diflkult coring conditions were encountered.Spurred by unacceptably low recoveries in coal seams in this area, the Joint Coal Board devised and developed, with the assistance of the Electricitv Commission, a new typc core lifter and a special 26 ft triple tube core barrel, designed to operate in conjunction with it in bad coring conditions.The core lifter, whieh is propped open clear of incoming core during drilling, is activated by fluid impulses generated inside the core barrel. It can be used with double tube barrels when drilling firm rocks whichdo not change diameter by compaction inside the barrel.Softer rocks. such as clayey types, require the use of a rriple tube barrel.Results from testing carried out to date in difficult areas in New South Wales coalfields have shown that greatly improved results can be obtained with the new equipment. Although developed essentially for coaldrilling, the new core lifter and barrel could be adapted to other fields of diamond drilling. Even in areas where no great coring difficulty is encountered the cost advantage deriving from the greatly inereased life of the core lifter and the adaptor in which it is housed could justify its use.

Jan 1, 1967

To improve the ventilation of large-opening under-ground stone mines by using stoppings made from stone left in place. Background Underground stone mines typically use large room-and-pillar openings with little ventilation control. Stone miners may be exposed to contaminants that include silica dust, diesel particulate matter (DPM), exhaust gases, nuisance dust, and face shot (blasting) fume gases. The health risks from breathing these fine particles and toxic organic materials are well documented. Although contaminants such as dust or diesel exhaust can be some-what controlled at the source, the escaping contaminants are often diluted and removed from work areas using ventilation with minimal controls. In many cases, current mine ventilation practices use trial-and-error techniques to improve ventilation. The growing realization of the importance of a healthy work environment along with increased regulations will require many stone mines to implement new ways to improve ventilation. In-place stone stoppings are one option. Mine air flows from one point to another because of a pressure differential between those points. This pressure differential is established using a mechanical fan in conjunction with separations (stoppings) between high-pressure (intake) and low-pressure (return) airways. The separation of intakes and returns is an important concept that is used in most other types of mines. Normally, air moves through underground stone mines with minimal control technology. Fans are placed in locations where they are thought to be most beneficial, and intakes are not separated from returns. In a few smaller underground stone mines, proper ventilation can be achieved with a few strategically placed fans. However, as these mines increase in size, moving adequate quantities of air through-out the mine becomes harder, and separation of intakes and returns becomes a problem. Building traditional mine stoppings to separate the intake and return airways in entries with a cross-section area often exceeding 800 ft2 is difficult. The stoppings must be able to withstand both the dynamic pressure from face shots and the static pressures created by the fans. Construction and maintenance difficulties, as well as high costs, are the main factors preventing these traditional mine stoppings from gaining wide acceptance by the stone mining industry.

Dec 1, 2002

By V. Kecojevic

Environmental and economic costs related to equipment fuel consumption and carbon dioxide (CO2) emission present a substantial challenge to the mining industry. Haul trucks are an integral part of the overall surface mining system and they consume a substantial quantity of fuel. Consequently, they produce a significant amount of CO2. The objective of this research is to (i) establish a mathematical relationship among a truck?s fuel consumption, power and engine load factors and (ii) determine the amount of a truck?s CO2 emission and the associated cost that may arise from potential CO2 legislation. In order to achieve these objectives we have considered original equipment manufacturer (OEM) haul trucks, commonly used in surface mining operations. The research presented here may be used by mining professionals to help determine the cost and environmental burden of the trucks? application, and efficiently manage energy consumption.

Jan 1, 2011

Splashing and wave formation has been studied experimentally by using one-seventh water model of CLU (Creusot-Loire Uddeholm) reactor with two different tuyere configurations by using paraffin and water as simulated slag and simulated metal respectively. Experimental results have been discussed in terms of total splash volume, relative slag volume in splash (dispersed-slag phase holdup in splash), wave frequency and wave amplitude with respect to gas injection rate and tuyere configuration at bath heights between 0.25m and 0.38m. Gas flow rates ranged from 0.00599m3/s to 0.01312 m3/s. The results of splash measurements showed that splash volume increased with increasing gas flow rate and decreasing bath heights for both tuyere arrangements. However, off-centre configuration resulted in relatively higher amounts of splash. The slag volume in splash (Dph in splash) was also found to increase with decreasing bath heights and increasing gas injection rates. The high speed photography revealed that the frequency and wave amplitude increased with gas injection rate. The experimental results also showed that the frequency and vertical displacement of the rotating wave increased when the tuyere orientation became off-centered. This might be used to explain why off-center configuration resulted in generating higher splash volumes than the center configuration.

Jan 1, 2004

Tumbling mills and some theoretical and practical aspects of milling are discussed. Attention is directed towards the difficulty of relating the dynamical analyses of the motions of particles within mills to the existing treatments of breakage processes. This synthesis may be advanced by a study of milling that utilizes the matrix analysis of breakage processes. Application of the matrix analysis to results obtained on (i) grinding in open-circuit a zinc ore in a conical pebble mill and in a modified conical mill, (ii) "grinding of cement clinker in a compound mill, and (iii) to some drop-weight grindability tests is discussed. It is shown that the 'IT-breakage model adequately describes the breakage processes of each of the tests. (The proportion broken in each breakage cycle is given by 'IT and is independent of particle size.)The size distributions of the products are accurately predicted in the cases considered. Matrix analyses of the breakage of coals, various ores and a cement clinker have been made. Appendices give an elementary account of the matrix methods and give the equations fully describing reverse order closed circuit operations. The matrix analyses of breakage processes and the survey are used to examine briefly closed circuit grinding.INTRODUCTIONTumbling mills are slow-speed crushing and grinding machines which have been developed from some fairly simple types into many widely differing forms. Many of these versatile mills have proved extremely useful in ore dressing and mineral beneficiation, whereas others have been used in the cement industry and less extensively in the coal industry.The performance of even the simplest tumbling mill is very imperfectly understood, whilst the more difficult problems associated with the balancing of complex milling circuits and with the establishment of the optimum operating conditions for the whole of a complex circuit are even less well apprehended.Our knowledge of the performance of tumbling mills is rather sharply divided into two categories: the first consists...

Jan 1, 1957

By M. Kolahdoozan, H. Noori

"Molybdenite flotation at the Sarcheshmeh Copper Complex has been studied as a function of pulp potential controlled using two types of sodium sulfide (mineral and chemical origin). Air and either plant or laboratory nitrogen was used as a flotation carrier. Potential measurements were taken in the rougher cells as it was proved to be vital for molybdenite flotation. Flotation test with reducing potentials between -525 and -575 mv gave the highest molybdenum recovery (93%) and the lowest copper recovery of between 4 to 15% can be achieved. These potentials did not change considerably by adding increasing amounts of sodium sulfide. At more oxidizing potentials, i.e. more than -450 mv, the recovery of copper increases while the recovery of Molybdenite drops.INTRODUCTIONThe average copper ore feed grade to the Sarcheshmeh flotation circuit is typically set at about 1.2% Cu. to be delivered to processing unit is set at 1.2%. There is no control of molybdenum (Mo) which usually varies between 0.01 and 0.04%. During concentration, a Cu-Mo concentrate is produced by one stage of roughing, one stage of scavenging and two stages of cleaning. Final concentrate contains 0.5% of Mo and 32% of Cu while the particle size is 90% under 42 microns (325 mesh). Because of the high grade of Mo in this concentrate and economic interests, a molybdenite flotation circuit plant was built in 1984. This plant was designed to treat 1432 tons of Cu-Mo concentrate (on dry basis) and consists of one stage of roughing followed by seven stages of cleaning. After dewatering the concentrate contains 53.7% Mo which is dried and shipped overseas.The majority of molybdenite production worldwide comes as byproduct of copper porphyry ores. The Mo grade typically varies between 0.01 and 0.06% and accompanies chalcopyrite and chalcocite. Due to the hydrophobic nature of molybdenite, it can be floated along with copper sulfides. Thus the overall concentrate of Cu and Mo contains 0.2-1% of molybdenum as well as 20-42% of copper. To separate the Mo, Cu and Fe minerals, depressants are used. The most important of which are sodium sulfide, sodium hydrosulfide, Anomal D, sodium cyanide, ferropottasium and ferro-cyanide. The first three regeants have been extensively used for successfully treating porphyry ores [1]. In the molybdenum plant of the Sarcheshmeh Copper Complex only sodium sulfide has been used to date."

Jan 1, 2006