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By Sylvain Picard

"RESUMELe gisement Silidor représente la dernière découverte effectuée dans le camp minier de Rouyn-noranda (fin 1985). Située à 5 km à I' ouest-nord-ouest de Rouyn-Noranda, ce gisement aurifère doit être mis en exploitation au cours du mois d’avril 1990. 11 est constitué par plusieurs facies minéralises comprenant une tonalité, un dyke mafique, une brèche carbonatée et par une veine de quartz. L’or se présente en particules très fines associées ou non à la pyrite. La zone d’altération hematitisee se développe sur une largeur de 30m environ de chaque cote de la veine. Une première étude structurale montre que la minéralisation s’est mise en place le long d’une zone de tension, orientée NNW, résultant de I' interaction de deux zones de cisaillement dextre orientées N270° et N300°. Les réserves diluées sont évaluées jusqu’ à 700 m de profondeur, à 4,78 Mt à 5,4 g/t Au. ABSTRACTThe Silidor ore deposit, located 5km WNW of Rouyn-Noranda, is the latest discovery made in the Rouyn-Noranda mining camp. Discovered at the end of 1985, this gold deposit will begin production in the spring of 1990. The deposit comprises several facies including a tonalite, a mafic dyke, a carbonated breccia and quartz vein. Gold is present as very fine grains, associated or not with pyrite. A 30m wide hematized alteration zone is developed on both sides of the structure. A structural analysis shows that the mineralization took place along a NNW oriented tension zone that resulted from the interaction of two dextral shear zones, oriented N270 and N300 respectively. The diluted ore reserves are estimated at4,78 Mt at 5,4 g/t down to a depth of 700m."

Jan 1, 1990

By P. K. Gummer

Gold was fir t di covered on the Seabee property in 1947 by the Con olidated Mining and Smelting Company Limited. Acquired by Claude Resource Inc. in 1984, the property is now being explored by Placer Development Ltd . in joint venture with Claude Re ources. Between Augu t, 1985 and March, 1986, Placer completed 54 drill holes, extending the main o. 2 and o. 5 vein y tern to a depth of 1000 feet below surface and di covering two new satellitic gold zone 800 feet west of the o. 5 gold zone. Drill proven and probable geological re erve in four zone ( os. 2, 5, II, 14) to a vertical depth of 1000 feet, are approximately 1,120.000 tons averaging 0.32 oz. Au/ton.

Jan 1, 1986

By L. V. Kin, V. V. Lodeischikov, A. V. Byvaltsev, IRGIREDMET JSC, O. D. Khmelnitskaya, S. V. Petrov

"Irgiredmet Institute carried out the study on the development of gold recovery technology from refractory ore from a Russian deposit. The ore contained 1.1 g/t Au, 1.5% of pyrite and 2.4% of pyrrhotite. The recovery of gold achieved was 84.7% using gravity with generation of high grade recoverable gold concentrate and gravity middling. In the course of lime treatment, cyanidation of the middling followed by tailings detoxification the consumption of reagents was 13.0 kg/t CaO, 95 kg/t NaCN and 72 kg/t Ca(OCl)2 of the concentrate. This was capital intensive.In order to decrease the consumption of reagents, magnetic separation of the middling of the concentrates from gravity separation was added to this flow sheet. Pyrrotite recovered to magnetic fraction, which did not contain gold in accordance to this advanced flow sheet. The mass yield of pyrite goldbearing non-magnetic fraction decreased down to 1.5 times in relative to gravity middling. The slurry filtration prior to and after cyanidation allowed the organization of cyanide solution recycling.The results of pilot plant tests using the developed flow sheet showed that the content of pyrrhotite in the non-magnetic fraction decreased from 59.1% to 11.3%. The reagents consumption was 12 kg/t CaO, 3 kg/t NaCN and 16 kg/t Ca(OCl)2 of the pyrite concentrate. The total gold recovery using this flow sheet was 83.4%. Due to a decreased consumption of reagents, cost efficiency of the process was achieved. Based on the results of this research, a project design criteria was developed. A processing plant with the annual throughput of 2.6 million tonnes of ore is being designed."

Jan 1, 2016

By Morris Wong

"An antimony-bearing gold ore was studied in a cyanide slurry containing 7 -20 ppm dissolved oxygen. 1 g/L NaCN at pH 9.5 and 48 hours retention time is required to extract 94 % of the gold from the 90 % minus 200 mesh ore by conventional cyanidation at a 7 ppm O2. As oxygen concentration was increased to 20 ppm, the conditions required to extract the same amount of gold were: 0.25 g/L NaCN, 16 hours retention time and 80% minus 200 mesh. Compared to conventional cyanidation, oxygen-enriched leaching was shown to reduce retention time by three times and cyanide consumption by four times. The kinetics of gold dissolution related to the dissolved oxygen concentration in slurry was expressed by following equation:XAu = 100 * (Kt + I0) / (1 + Kt + I0)where XAu is the percentage gold extraction; t is the retention time; K and Io are the slope and the intercept of the plot of XAu I (100 -XAu) vs. t respectively. The values of Kand Io depend upon the concentrations of cyanide and oxygen, the fineness of grind and the origin of ore sample etc.. A preliminary cost analysis was also performed to justify the use of oxygenating slurries.IntroductionGeneral:It was proposed almost 149 years ago (Elsner 1846) that oxygen is required for the dissolution of gold in cyanide solution. However, little attention were paid to oxygen application in the gold milling industry during the following one hundred years. Due to the constraints of gold price, reagent costs, refractory ores, and environmental issues in the past decade, the oxygen enhancement in the cyanidation process was introduced to the gold milling industry for increasing operation efficiency, reducing production costs, and meeting the requirement of environmental standard.The advantages of oxygen-enriched cyanidation have been reviewed by several investigators (Hague 1992, Jara et al 1992, Henderson 1989, Loroesch et al 1988, Elmore 1987, Cathro 1963). Common to all conclusions are the following advantages."

Jan 1, 1995

By Fabian Mohr

The methyl substituted binuclear cyclometallated gold(I) complexes [Au2(?-n-MeC6H3PPh2)2] [n = 4, 5 and 6] all undergo oxidative addition reactions with halogens and, in the case of the 4- and 5-methyl substituted analogues, with dibenzoyl peroxide to give metal-metal bonded digold(II) complexes [Au2X2(?-C6H3-n-MePPh2)2]. The 4- and 5-methyl substituted digold(II) compounds on warming or exposure to light undergo carbon-carbon coupling reactions to give gold(I) complexes containing the corresponding 2,2'-n,n'-biphenylyldiphenylphosphines. The 6- methyl substituted digold(II) complexes rapidly isomerise above ca. -20° C to binuclear heterovalent gold(I)/gold(III) compounds [XAu(?-2-Ph2PC6H3-6-Me)AuX{?2-(6-MeC6H3-2-PPh2)}], containing a fourmembered cyclometallated ring on a gold(III) centre. Surprisingly, no carbon-carbon coupling could be induced for any of the 6-methyl substituted compounds, even under forcing conditions. This difference in reactivity of the 6- methyl derivative, compared to that of the 5-methyl and unsubstituted analogues was studied by ab initio calculations using a series of model compounds. The results are in excellent agreement with the experimental findings and suggest that the key-step in the C-C coupling reaction is the partial opening of the Au(III)-P bond in one of the transition states. In the 6-methyl system this transition state is energetically inaccessible and hence no C-C coupling occurs.

Oct 1, 2003

By M. Sutherland

"Increased efforts to improve mine safety have led to various successful forms of ground control in Canadian underground mines. At Inca Limited's T-3 Mine in Thompson, Manitoba, a perimeter blasting method has been implemented in jumbo drilled development headings.The technique utilizes an increased number of perimeter blast holes in the drill pattern while decreasing the linear charge density of explosives in any individual perimeter hole. The decrease in perimeter charge density means less blast damage to the surrounding structure.The benefits realized include increased safety through less loose ground and substantially less scaling time, less overbreak and improved long-term drift stability. This paper highlights the development and implementation of the perimeter blasting program.IntroductionThompson Mine Thompson Mine is a 6000 ton per day underground nickel mine located just outside the city of Thompson, 400 miles north of Winnipeg. The mine is serviced by the T-I shaft and by the T-3 shaft. The shafts are 10 000 ft apart and are connected on several levels with the deepest operating levels at T-1 and T-3 being 4000 ft and 2400 ft respectively.The T-3 Mine presently produces 52% of the total Thompson Mine production and 65070 of the 3000 ft of mine development each month. Current development of the 5 million ton 1, C orebody below 2400 level at T-3 will maintain these production and development ratios.Depending on equipment requirements, development headings at T-3 Mine vary in size. The 1, C ramp, at 17 1/2 ft wide by 10 1/2 ft high, is presently the largest development heading at T-3. Fifty- two per cent of the development footage is done with four Gardner Denver two boom electric /hydraulic jumbos and one Gardner Denver one boom electric /hydraulic jumbo. The remaining 48% is driven with hand-held jack-leg drills."

Jan 1, 1990

By Mari Yamamoto

Controlled thermolysis of gold(I) complex, [Au(C13H27COO)(PPh3)], at 180 °C for 5 h under an N2 atmosphere affords the new type of monodispersed spherical gold nanoparticles capped by myristate and a small amount of PPh3 ligands (average diameter 23±1.5 nm), in spite of no use of solvent, reducing agent and stabilizer. Much attention has been paid to the preparation and structure determination of transition metal nanoparticles, because of their widespread use in technological applications.1 In general, the chemical reduction of metal salts and organometallic complexes2 by the use of reducing agent is carried out to prepare gold nanoparticles. In some cases, a large amount of solvent and stabilizers such as surfactants,3 polymers4 and ligands5 are used to prevent agglomeration of particles and control the particle size. For instance, Brust et al. reported a two-phase reduction of AuCl4 - by NaBH4 in the presence of dodecanethiol as stabilizer to prepare gold nanoparticles.6 On the other hand, O?Brien et al. recently reported an one-phase preparation of gold nanoparticles by the reduction of AuCl4 - by NaBH4 in the mixture of tri-n-octylphosphine oxide and octadecylamine which act as both solvent and stabilizer, but this procedure also needs reducing agent.7

Oct 1, 2003

By Rick Valenta

Is the Labrador Central Mineral Belt part of the IOCGU Club? ?I sent the club a wire stating, PLEASE ACCEPT MY RESIGNATION. I DON'T WANT TO BELONG TO ANY CLUB THAT WILL ACCEPT ME AS A MEMBER.? GrouchoMarx Acknowledgements ?Ian Dunlop, Matt Lennox-King, Mark O?Dea, Dave Barbour, Derek Wilton, Lucas Marshall, Roland Butler, Mark King, Sally Howson,

May 1, 2006

By Ho-Sung Yoon, Jyothi Rajesh Kumar, Jin-Young Lee

"Global population is a key factor in future energy requirements, especially considering the important elements of global energy resources, electricity production, the accessibility/non-accessibility of electricity, greenhouse gas emissions and the states of nuclear reactors. Everyday human requirements depend on industrial products, and industries need a considerable amount of electricity to run. Nuclear fuel is inexpensive and bolsters industrial growth. As uranium reserves are being consumed rapidly, an alternative source of energy in the nuclear industry is thorium. Here, primary energy supplies by various sources from 1993 to 2011 as well as estimates of future needs to 2020 are discussed. This study finds that primary energy sources are mainly fossil fuels, renewables sources, and nuclear and hydro energy. In addition, growth depends on increases mainly in renewable sources.Recovery of thorium from its ores by using hydrometallurgical routes were described and discussed. The final part of this paper was presented a case study of the thorium extraction from Korean monazite ore leach liquors by using Prieme JM-T diluted in A150 diluent as an extractant system. And possible separation of the thorium from rare earths (REs) and base metals were calculated by measuring of separation factors and finally, the recovery of thorium from loaded organic of Prieme JM-T was established.THORIUM IS FUTURE NUCLEAR FUEL?Global energy consumption is increasing everyday with the high population growth, high industrialization of numerous countries and the higher level of good and services consumption that follows it. If we look at specific sectors of the economy, oil consumption grew by 0.8 million barrels per day, globally, whereas the increase was 0.4 % for natural gas and coal 1.8 % for nuclear power and just under 2.0 % for hydroelectric power (Asheley 2009). Use of renewable sources of energy for power generation and transportation increased by 3% in 2014 (BP 2015)). Thus alternative source of energy must be identified and developed to meet this increasing demand. Thorium is a potential one for the nuclear industry."

Jan 1, 2016

By C. Peña, C. A. Zaror, M. Bustamante, M. Vega

The goal of this paper is to present the carbon footprint of Chilean cathodic copper, as a function of the electricity generation mix, ore grade, and process efficiency. Ore grades and acid leach-solvent extraction-electrowinning (LX-SX-EW) yields in the range 0.5-1.0% and 60-80%, respectively, were considered here. Additionally, electricity supply from two Chilean suppliers, SING (1.0 ton CO2/MWh) and SIC (0.4 ton CO2/MWh) was assessed. Simulation results show that oxide plants located in northern Chile, served by the SING network, feature carbon footprint within 4.3 - 11.4 MWh I ton cathodic copper, as compared with 3.5 - 9.3 MWh I ton copper in plants served by SIC. The footprint was found to increase as ore wade and process yields decrease.

Jan 1, 2011

By S. Kamado

Development of new SF6 alternative cover gas with low global warming potential is strongly desired for melting and casting of magnesium alloys. We have tried to apply several kinds of HFC and HFE gases with low potential from 30 to 950 as protection gases to AZ91D alloy melt in the temperature range of 650 to 800°C. As a result, protective film consisted of MgF2 and MgO is formed on the melt surface and ignition-proof effect is sufficiently obtained at all the temperature ranges even when flow rates of mixed gases containing CO2 as a carrier gas are almost the same as those using SF6 gas. Reduction of global warming potential exceeds 98% at 650°C and 89% at 800°C in comparison with a mixed gas of CO2 and SF6.

Jan 1, 2006

By S. K. Shaha, W. Kasprzak, D. L. Chen, J. Friedman, F. Czerwinski

The role of solutionizing conditions within the T6 heat treatment of the cast Al-Si-Cu-Mg alloy with micro-additions of Zr-Ti-V was assessed in terms of their influence on microstructure and mechanical properties. As a solution treatment, a single step process at 485 and 510oC and a two-step process at 510 and 525oC for varying time intervals from 0.5 to 9 h, were examined. In all cases, solution treatment was followed by isothermal aging at 150oC for 100 h. The microstructural analysis showed that the solution treatment changed the eutectic silicon morphology from needle-like, present after casting, to spherical shape in the T6 state and that the extent of changes was strongly influenced by temperature and time. The Cu- and Mg-rich phases completely dissolved during solution treatment while the Fe-containing phase, Al8Si6Mg3Fe, remained partially in the microstructure. The Ti, V and Zr containing phases, Al13Si2Ti3Zr and Al3Si26FeTiV10, present in as-cast state, were partially dissolved during solution treatment, reducing their size by 10 - 48%. The studied alloy reached the maximum ultimate tensile strength of 335 MPa and yield stress of 303 MPa for two-step solutionizing with a total time of 4.5 h. It is concluded that the two-step solution treatment resulted in higher tensile properties than the single-step process.

Jan 1, 2015

By Chris M. Healey

"Abstract- The Jasper Gold mine in northern Saskatchewan produced a total mill feed of 140 127 tonnes at an average grade of 19.0 g/t Au. Both diamond drilling and chip samples from under­ ground workings returned extremely high values, up to 12 851 g/t (375 oz/I) Au over 0.10 meter from a chip sample. In the presence of such extremely high grades, the accurate estimation of reserves is a major problem.ln the first part of this study, production data from the mine were used to evaluate the results of a variety of reserve estimation techniques in this environment. In two out of three cases studied, indicator kriging with assays cut to 342.9 g/t (10 oz/I) gave extremely good results. In the third case, however , indicator kriging gave unacceptable results. Instead, lognormal kriging of uncut data gave the best results.Secondly, from the same database, a number of drill-hole spacings were simulated, and reserves estimated from the simulations. As may be expected, the results of this analysis show increasingly better results with tighter drill-hole spacing. However, even with a 5 m by 5 m spacing the variation from actual grade was quite large. This is not a systematic under- or over-estimation of grade, rather a random scatter of results .The results of this study show very clearly the problems of deriving good estimates of reserves when extremely high-grade samples are encountered. There is a high degree of randomness to this database, so reserve estimation in this situation will remain a high-risk operation."

Jan 1, 1993

By R. J. Mitchell

"High narrow backfills of significant strike length approach the stress conditions associated with a free standing wall but this lower bound design approach is not considered efficient for earth structures. This paper considers the design of high narrow backfills using upper bound limit analysis, retaining wall theory with backfill arching and centrifuge model tests. Complex toe support situations pertaining to operational factors at Falconbridge's Lockerby Mine are investigated. It is shown that the susceptibility of high narrow fills to toppling is much lower than anticipated from the retaining wall analysis and that toppling can be avoided by a suitable toe support. It is also concluded that adhesion between the fill and an adjacent pillar can markedly increase the stability of the fill. INTRODUCTIONFalconbridge's Lockerby Mine pillar projects will expose cemented tailings backfills to heights (H) up to 55 m. The exposed backfills may have strike lengths (L) greater than the heights but are relatively narrow, having widths (w) of up to 12 m. The bulk cement content of the fills varies from 8.3% (12.1 T:C) to 3.0% (33.1 T:C). Exposed faces may have up to a metre of overhang due to stope overblasting, as shown in Figure 1. These fills have a solids specific gravity of close to 3.0 and were placed at a pulp density of about 68%. The tailings have a fairly uniform grain size with 32% less than 0.075 mm (200 sieve) and about 12% slimes (less than 0.02 mm sizes). The fill exposed by early pillar removal will be backed by adjacent, near vertical, ore pillars. Future pillar removal may expose similar fills which are not supported by an adjacent ore pillar but which will be backed by either an open void or by uncemented backfill. Furthermore, these fills may be undercut when the pillar slot is driven, further reducing stability (case 3, Fig. I) or a wedge could be left to help support the fill (case 2, Fig . 1). This paper presents limit equilibrium analyses for these various backfill conditions in a 22: 1 T:C fill and com- pares the results to data from centrifuge model studies at a scale factor of A= 157 as indicated in Figure I. Prototype fills have a 6 m deep cone plug of 16:1 T:C which is accounted for in the comparison of centrifuge model test with analytical results."

Jan 1, 1989

By Sarah Bélair, Denis Beltrami, Deka Moussa Ragueh, Valérie Weigel

"Scandium is a highly valuable element belonging to the rare earth elements (REE) but it has different chemical and physical properties from those of lanthanides. This metal is generally used in aluminum-scandium alloys and in fuel cells but the development of its market is slowed down by its high price (more than 1000 US$/kg for a 99.0% grade product). Indeed, the low scandium content in the continental crust (an average of 22 ppm) and the difficulties to extract it selectively from solutions containing high impurity concentrations increase the cost of the recovery processes. Nowadays, scandium is primarily recovered as a byproduct.ERAMET is developing a hydrometallurgical process in Gabon (Mabounié) in which the scandium content in the deposit reaches 10 to 100 ppm. The Maboumine project consists in the extraction and purification of uranium, niobium, tantalum and REE from Mabounié ore. In the frame of this hydrometallurgical project, ERAMET had also studied the feasibility to recover scandium as a byproduct. During the ore leaching, uranium and rare earth elements including scandium are solubilized into the pregnant leach solution (PLS). Among the various methods investigated to recover scandium from the PLS, liquid-liquid extraction using Primene JM-T as extractant was chosen and tested first at laboratory scale and then at pilot scale. Selective stripping of the loaded organic was carried out by using phosphoric acid solution, allowing the selective recovery of scandium from thorium and zirconium partially extracted from the PLS by Primene JM-T under the process conditions."

Jan 1, 2016

By W. Petruk

"D. Doyle How large is the beam to which you referred? Are we considering Watts, Griffis the probe of a 1 micron sized particle?& McOuatReply Yes. The beam would be about 1 micron. We could enlarge or reduce it depending on the current in-put.D. Doyle Then chemically bound material we might not see. If you are restricted to 1 micron then you would not be able to separate things smaller than this. Is this not so?Reply We won't be able to analyse it but we will see it as the scanning electron microscope has a resolution down to about 0.1 microns. We should be able to get the general composition of such a grain plus its surrounding grains. We feel we can then make a reasonable estimate if it were as small as 0.1 microns. An accurate analysis would be achieved with grains of 3 and 4 microns.D. Doyle Would this have an application in the hydrocarbon field?Reply The lowest atomic number would be sodium. We wouldn't be able to go below that atomic number. Anything with sodiumor a larger atomic number would enable us to analyse the carbon.M. Fairwether Do you have any policy relating to priorities. For instance areCominco companies going to be limited in the amout of samples permitted to give everyone a chance because quite frankly our company, and I'm sure all base metals companies can hardly wait for the installation of this equipment. I can forsee two years back-log of work in the first month.Reply Quite frankly, I haven't looked at this. At the present time the work is under the START program but I suppose we will have to look at possible changes. Anything under the START program has first line priority.B. Thorndycraft Will the briquetting of samples be automated as well? Pinson MiningReply No, not at this point. Our polish section laboratory is first-class and can turn out 18 polish sections a day which is more thar we can handle.B. Thorndycraft With the capability of this machine to operate in a hands-off manner perhaps it could start to look for the 5 micron gold particle in tailings?Reply I have such a program using a wave-length spectrometer rather than the energy dispersive. The wave-length spectometer is about 10 times as sensitive and would scan the section much faster. A one inch diameter polished section would be scanned in about 7 hours for anything 1 micron or larger."

Jan 1, 1984

By J. S. Roberts

"Campbell mines has been in operation since 1948. Part of the process consists of roasting of a flotation concentrate which contains arsenopyrite.At present operation level the milling rate averages about 850 t.p.d. and about 48 t.p.d. of concentrate is roasted. The gas emissions at present contain between 3 and 3.5 t.p.d. of arsenic trioxide, about2 - 2.5 t.p.d. of particulate matter and between 19 - 25 t.p.d. of sulphur dioxide. The particulate is extremely fine calcine dust and it assays about 4 oz. Of gold per ton.Until 1974 the gas was discharged untreated through a 125' stack.UntreatedThe decision was made in 1972 to erect a gas treatment plant, construction ran through 1973 and operation began in 1974. The cost of the installation was about 1.3 million dollars.On investigation the process technology, it was apparent that the plant operated by giant Yellowknife for the same purpose was the model we should study. I would like to acknowledge the fine cooperation we had from giant, the plant that emerged is very similar to theirs and incorporates some improvements suggested by them and based on their operating experience.The plant is designed to remove arsenic and solid dust and it has no provisions to remove sulphur dioxide.We depend upon the 125' stack to assure adequate dilution. Because of the nearby airport the stack cannot exceed 125'. A stack burner is provided that can ""y used to heat the stack gas and carry it higher into the atmosphere when there is an inversion condition.The removal of arsenic from hot gas is fairly simple. The sublimation point is about 425°f, a few degrees above this point the arsenic is a gas and a few degrees below it condenses into a solid form. The arsenic could have been removed by simply dropping the temperature below the condensation point and collecting all the solids in a baghouse. This would have removed the possibility of recovering the gold in the calcine dust and would have produced an impure and more bulky arsenic material to be placed in safe storage. The decision was made to collect the solid dust in a hot electrostatic precipitator operating above the sublimation point and subsequently lower the temperature below the condensation point and collect the arsenic in a baghouse."

Jan 1, 1976

By Pierre Sauvé, Guy Perrault, Armand Savoie, Pierre Trudel

"La Mine Doyon est une entreprise détenue à parts égales par Cambior Inc. et par Minerais Lac Ltée qui en est l'opérateur. Les statistiques de production au 1"" janvier 1990 indiquent que, depuis le début de son exploitation (1980), 10.3 Mt de minerai d'une teneur moyenne de 5,9 g Au/t ont été usinées. Les réserves sont estimées a 10,4 Mt à une teneur moyenne de 6,6 g Au/t jusqu' à une profondeur de-450m.Trois zones minéralisées sont présentement en exploitation. Ces zones sont caractérisées par la présence de veines de quartz-pyrite accompagnées localement de chalcopyrite. Les roches encaissantes sont les schistes a séricite (zone# 1), les roches volcaniclastiques de composition felsique (zone #2) et intermédiaire (zone #2, zone ouest) ainsi que les roches intrusives appartenant au pluton de Mooshla (zone ouest). Toutes ces roches ont subi un métamorphisme régional dont l'intensité correspond à la limite supérieure du facies schiste vert ( -425°C}.Les zones minéralisées sont situées à I' intérieur des zones de déformation et d'altération les plus intenses. L'étude des structures mineures nous indique que la région a subi plusieurs épisodes de déformation ductile a fragile qui sont antérieurs, contemporains et postérieurs al a minéralisation. L'analyse structurale indique également que la déformation majeure est associée à une importante phase de compression orientée N-S a NNE-SSW.L’ensemble des observations suggère que la minéralisation et I' altération associée sont de type épigénétique. La mise en place du minerai se serait effectuée à la fin de l'épisode principal de métamorphisme et de la déformation."

Jan 1, 1990

By G. W. Smith

"The adsorption of dehydroabietylamine acetate on quartz, hematite, rutile and baddeleyite has been deter-mined for concentrations ranging from 0.3 to 5,600 µ. moles/I and for pH values ranging from 2.5 to 10.5.The surface tension, UV and IR spectra, equivalent conductance and ion concentrations for dehydroabiety lamine acetate solutions have been determined.Contact-angle and flotability studies have established that good collection may be obtained with this amine but that selectivity depends on the suitable use of depressants for some of the minerals."

Jan 1, 1971

By Guangdong Institute of Resources Comprehensive Utilization, Y. H. Tang, H. X. Hu, T. S. Dong, Z. H. Zhang

"A world class rare metals ore deposit in Africa belongs to a massive granite type poly-metallic pyrochlore-niobium resource. Successful project development in realizing the resource is vital to Africa’s economy and pyrochlore processing technology.The ore properties are quite complex due to the presence of varieties of valuable minerals, fine dissemination and intimate association of minerals and the relative abundance of easy-floatable minerals. The separation of pyrochlore from zircon, calcite and mica is generally considered difficult. Whilst studies have previously been conducted by several independent institutions overseas, no satisfactory process has yet been developed to date for the ore deposit—possibly due to the process designed being too complex; high cost of regents; lower performance.Based on process mineralogy study and extensive testwork, the authors have in recent years developed a new technology—“De-slime—Iron-rejection—Zircon (with calcite) and Calcite reverse flotation—Pyrochlore flotation”. The flow-sheet is characterized by ?1prior removal of slimes and iron minerals; ?2 Zircon (with calcite) and Calcite reverse flotation; and ?3Pyrochlore flotation applying new type of chelating collector GYX3 developed by GZRINM in conjunction with selected regulators and depressants.The process involved in the new technology is simple to implement and facilitates stable operation; the flotation reagent regime is easily adaptable; the cost of main reagent is lower largely because it is easy to manufacture and is environmental friendly. Laboratory test results demonstrated that, for raw ore assaying 0.35% Nb2O5, 0.03% Ta2O5 and 0.45% ZrO2, pyrochlore concentrate grade of >35% Nb2O5 could be obtained with >74% recovery. Meanwhile, the Ta2O5 content in pyrochlore concentrate could be recycled, and ZrO2 recovery >50% could be achieved in zircon concentrate assaying ZrO2>63%—significantly improving the technical-economic indicator. These results were verified in pilot trials of 1.44t/d throughput. The paper presents the new technology in whole, summarizes the key factors affecting its performance and discusses in simple terms the mechanism of separation."

Jan 1, 2016