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By W. Griffith

"While oxalic acid precipitation practices are commonly applied on purified rare earth elements (REEs) pregnant leaching solutions (PLS), few efforts have been reported using oxalic acid (OXA) to recover REEs directly from untreated PLS. High oxalic acid demand is the most significant issue in all these direct oxalic precipitation (DOP) practices. Preliminary research on DOP of REE from a dilute PLS originating from an acid bake followed by water leaching was conducted. At 75°C, the oxalic acid demand was between 12 and 18 times the stoichiometric requirement for complete REE precipitation. Conducting DOP at 25°C had superior REE recovery but the product purity and filterability declined. Between 75°C and 100°C, REE recovery increased with increasing temperature. When sodium hydroxide was used to adjust the pH during DOP, in the range of 0.5 to 1.5, higher pH benefited REE recovery. Using ammonium hydroxide in place of sodium hydroxide, the REE recovery was reduced by close to 5%. Sodium chloride addition to the PLS was also investigated and the results showed that at both 75°C and 25°C, the REE recovery was enhanced substantially. At 25°C, when the oxalic acid addition was controlled at 4.7 times the stoichiometric requirement, the overall REE recovery reached 95% with the addition of 240 g/L NaCl. Other reagents such as KCl, NH4Cl also showed significant enhanced REE recovery at the same oxalic acid addition. The addition of sodium sulphate appeared to largely enhance the precipitation of LREE but inhibited the formation of HREE oxalate. All results indicate the formation of rare earth double salts may play an important role in REE DOP sulphate media.INTRODUCTIONThe conventional routes for recovering rare earth elements (REEs) from a sulphuric acid pregnant leaching solution (PLS) involves an initial purification step for removal of iron, aluminium for example by precipitation through neutralization, followed by precipitation of REEs as hydroxide, oxalate or carbonate salts, or by bulk solvent extraction for REE recovery (Xia, 2013). Another route was used by Bayan Obo in their historical flowsheet, where the PLS was first subjected to double sulphate precipitation, then the remaining REEs, mostly heavy rare earth elements (HREE), were recovered by oxalate precipitation (Li, 2011)."

Jan 1, 2016

By L. Xia, B. Hart, G. Gingras, M. Furlotte

Plant and laboratory test work was performed in order to optimize a copper/zinc flotation separation circuit at Matagami concentrator, Canada. The testing identified a significant difference between the laboratory testing and the plant flotation, particularly for the sphalerite recovery in the rougher copper concentrate. Tests indicate that sphalerite recovery of the flotation concentrate in the laboratory testing was significantly higher than in the plant. The ToF-SIMS surface analysis identifies surface species responsible for the discrepancy in Zn recovery. It shows that sphalerite flotation is in response to the collector attachment and a significant enrichment of Cu on the surface of sphalerite grains. A comparative surface chemical analysis was performed on both sphalerite and chalcopyrite grains from the plant flotation and laboratory testing in order to identify surface species variability between the two testing. Two significant observations were revealed. First, there is considerably more hydrophilic iron oxyhydroxide on sphalerite surface from the plant flotation sample relative to those from the lab testing. The secondary coating likely completely or partially passivates sphalerite surfaces, reducing Cu enrichment and hence sphalerite floatability. Secondly, comparing the surface chemistry on sphalerite and chalcopyrite, the hydrophilic iron oxyhydroxide species have preference for the sphalerite grains relative to chalcopyrite. Despite numerous studies on the effect of galvanic coupling during grinding, there is limited documented evidence on the effect of the milling environment on the flotation and plant operations. This observation suggests that interaction between sulphide minerals and grinding media have a significant influence on subsequent flotation. This paper discusses the results of a detailed surface chemical evaluation of sphalerite grains from both the plant and laboratory testing in the context of grinding environments and pulp chemistry.

Jan 1, 2016

By The Staff

"IntroductionThe published information concerning treatment of uranium ores has been very limited as far as Canadian ores are concerned. Such information has dealt mainly with refining processes for concentrates or rich ores, designed to recover radium rather than uranium; these methods could not be economically applied to low-grade ores.The purpose of this paper is to describe treatment possibilities for some of the .ores that have been treated at the Mines Branch, and .to show how the character of these ores influences methods and results.What is a Uranium Ore?Prospecting for uranium has been extremely active during the past two years and a great many occurrences of radioactive minerals have been recorded. Such prospecting has been based to a large extent on the use of the Geiger counter in the fie1d. This equipment is an invaluable aid when the results are properly interpreted.Most definitely, indications of high radioactivity do not always prove the presence of a valuable orebody. First we have to determine:(1) Is the radioactivity due to uranium or to thorium. There is no immediate market for thorium.(2) Can the uranium in the ore be economically recovered.Chemical analysis is the most exact method of determining uranium content, but a fair estimate of the uranium, and also of the thorium, content can be made from the difference in gamma and beta activity. The Mines Branch guards against misinterpretation by determining both beta and gamma activities and, if more than a minor amount of thorium is indicated, then a Chemical analysis is made."

Jan 1, 1950

By L. D. Snyder, J. K. Russell

"The Lower Jurassic Iron Mask batholith intrudes Upper Triassic Nicola Group volcanic rocks in the Quesnellia terrane of southcentral British Columbia and hosts Cu-Au mineralization. The batholith comprises three mappable intrusive phases, including: Pothook gabbro to diorite, Cherry Creek monzodiorite to monzonite and Sugarloaf diorite. The Iron Mask hybrid forms a fourth mappable unit that makes up to 45% of the areal exposure in the batholith. It is highly variable in texture and composition and subdivided into three types; Type I and II have abundant (partially digested) xenoliths, many of which are derived from the Nicola Group, whereas Type III is xenolith-poor and shows extreme outcrop-scale textural diversity.Pothook and Cherry Creek rocks are similar in texture, mineralogy, whole rock and mineral composition and style of alteration. Contacts between the two can be intrusive or gradational. Rocks collected from these two suites have compositions that are permissive of a single magma batch, although simple crystal fractionation does not explain all chemical variations within and between the two suites. The Sugarloaf unit comprises a diverse group of amphibole-phyric rock types and is texturally and mineralogically distinct from Pothook and Cherry Creek suites.The Iron Mask hybrid represents progressive stages of physical and chemical interaction between Nicola Group country rocks and Pothook magma. Magmatic assimilation (selective or partial) is supparted on the basis of field relationships, petrographic observations, and REB compositional data. Thermal dehydration and partial assimilation of Nicola Group rocks by Pothook magma is shown to be a valid means of promoting early and extensive volatile saturation in the Iron Mask batholith."

Jan 1, 1995

By James L. Parks, David M. Kjos

"Significant increases in mill liner life and plant availability can be achieved by following guidelines that have evolved from over twenty years of practical experience with autogenous grinding (A0) and semi-autogenous grinding (SA0) mills. A0 and SA0 mill liner functions, performance criteria, designs, wear-resistant materials and maintenance procedures are reviewed. Operating practices that can lead to accelerated wear and breakage are discussed.IntroductionEfforts to improve the performance of liners for autogenous grinding (AG) and semi-autogenous grinding (SAG) mills began as soon as the modern, large-diameter versions of these comminution machines were first intalled in commercial plants twenty years ago . While concentrators with conventional circuits were typically running at 95010 or better mill availability, many of the new autogenous and semi-autogenous circuits were struggling to achieve 80% availability. Liner maintenance, which involves not only material and labour costs but also loss of production, was quickly identified as one of the major reasons for downtime.Today, liner designs and materials have evolved to the point where virtually maintenance-free liner life of two to three years is possible in autogenous mills and one year in the more severe operating conditions encountered in semi-autogenous mills. Of concern, however, is that hard-won lessons from the past sometimes seem to be forgotten or ignored during the design of new mills and the operation of existing mills. As liner suppliers, we are often confronted with performance problems that could have been avoided by applying well-publicized guidelines for liner design and materials selection.This paper provides an overview of AG and SAG mill liner functions, criteria for evaluating performance, designs, wear-resistant materials and maintenance procedures. Also, operating conditions that can lead to accelerated wear and liner breakage are discussed in detail."

Jan 1, 1990

During heap leaching of secondary copper sulphide ores operated at low heap temperatures and high levels of leach solution salinity, copper recovery is usually unsatisfactory, due to a slow chemical and bacteriological leaching of the covellite minerals either formed from leaching of chalcocite or naturally occurring in ores. A laboratory test program was conducted to investigate the effect of leach solution temperature in the range of 20 to 50°C on copper extraction during stirred reactor, bottle roll, and short column leaching of fresh or heap leached secondary copper sulphide ores. It was demonstrated that increasing heap temperature is an effective approach for improving recovery of copper from secondary copper sulphide ores. A review of the technologies available for heating heap leach solution or decreasing solution salinity to improve copper recovery from secondary sulphide ore indicates that an integration of copper concentrate pressure leaching with heap leaching is potentially a cost-effective approach. A copper flotation concentrate sample was tested through pressure oxidation with the use of heap leach solutions of varying strengths for the repulping of the concentrate in order to examine its effect on pressure oxidation performance. The test results indicate that use of heap leach solution of high salinity in the pressure oxidation process does not cause a negative impact on the process performance for copper extraction, sulphide oxidation, and recovery of gold (if present) from pressure leach residue, and has additional advantages of removing the bacterially inhibitory aluminum from heap leach solution and generating more acid and ferric iron in the pregnant leach solution. A flowsheet configuration for full integration of the copper concentrate pressure oxidation with heap leaching of secondary copper sulphide ore is proposed and discussed.

Jan 1, 2014

By S. Foged

Umicore and MWH jointly demonstrated that mechanical dewatering of goethite and neutralised tailings significantly increased the life of a 30 ha tailings pond and improved the operation of the pond. Umicore operates one of its zinc production plants in Balen, Belgium. As part of its production process, an excess of neutralised and goethite¬containing tailings is generated. These have historically been stored inside large lined ponds. As the storage area of the current goethite pond was nearing its capacity, Umicore studied the potential for an extension to the existing embankment. However, this extension would face intense regulatory scrutiny, because the environmental and safety impacts were considered important. Together with Umicore, MWH suggested and investigated an alternative treatment approach consisting of mechanical dewatering prior to storage. A feasibility study was prepared where wet storage was compared to dry storage. Differences in safety risks, environmental impacts and economic costs were outlined. The study demonstrated that the life of the tailings pond could be increased from 6 to 22 years, seriously reducing the environmental impacts and safety risks at no significantly higher cost than would have been incurred with the continued operation of the existing pond. The results of the study are presented in this paper.

Jan 1, 2006

By Hatch, A. Reple, W Valery

The mining industry is facing declining feed grades which require the mining and processing of larger volumes of material per tonne of product. Pre-concentration aims to remove barren material at as coarse a particle size and as early in the process as possible to reduce processing costs, energy consumption, water losses, and greenhouse gas emissions while increasing production and profitability. This may be achieved using bulk ore sorting—the separation of barren gangue from a fully loaded conveyor belt based on the grade as measured or inferred from a sensor. This paper evaluates the technical and economic viability of bulk ore sorting for the PanAust Phu Kham copper deposit in Lao PDR. Bulk ore sorting technology is described briefly along with discussion of appropriate sensor and diversion systems for the Phu Kham operation. To evaluate the economic implications, the actual mined material and costs from 2014 were used as the base-case. Sampling data was not available; therefore, the ore grade variability that would feed a bulk ore sorter was estimated using geo-statistical tools based on mine grade control data. The sorter performance when treating this material was simulated considering the accuracy of the sensor measurement and errors associated with material diversion. The impact on inferred cash flow was estimated for several different scenarios considering the additional costs for sorter operation and reject handling. The evaluation demonstrated that bulk ore sorting at Phu Kham has the potential to increase the amount of metal in product and/or reduce processing plant throughput (and costs) to improve the annual inferred cash flow for the conditions investigated. The sorter is the gatekeeper – only material above the cut-off grade (value-adding) reports to the processing plant; no waste dilutes the feed. Likewise, above cutoff grade material is not lost to waste; preventing lost revenue.

Jan 1, 2016

By K. Hardy

As part of Inco's $115 million Fluid Bed Roaster SO2 Abatement Project, a new weak acid treatment plant (WATP) was constructed to treat bleed solutions from the Copper Cliff Smelting Complex gas-cleaning circuits. The WATP produces an iron-arsenic co-precipitate, termed "basic ferric arsenate" (BFA), which requires disposal in an environmentally acceptable manner. Following analysis of many potential options, the short-listed disposal alternatives were: A) Co-disposal of the BFA?gypsum sludge with rock tailings, a portion of which would be used in mine backfill, with the remainder being sent to the tailings area, and B) Co-disposal of BFA?gypsum sludge with pyrrhotite tailings in the tailings area. Testwork showed that the addition of BFA?gypsum sludge to mine backfill samples increased the uniaxial compressive strength (UCS) of the backfill for most samples, especially those at lower binder content ratios. Extensive analysis was conducted on samples taken from the Copper Cliff tailings area, where BFA?gypsum sludge from Inco's Copper Refining Electrowinning Department has been co-deposited with pyrrhotite tailings since 1990. Pore water samples from these areas contained very low arsenic concentrations, indicating that the arsenic had remained stable in that environment. Ultimately, co-deposition with pyrrhotite tailings was chosen as the preferred option for the disposal of BFA?gypsum sludge from Inco's new WATP.

Jan 1, 2006

By M. A. Bergougnou, I. Inculet, K. I. Burgess

This paper describes the electrostatic methods of separating various minerals, reviews the principles of fluidization and outlines the technique of electrostatic sorting in fluidized beds. The application of the method to the benefication of coal is emphasized.

Jan 1, 1970

By Bird. M.

As the Cadia mill was the first 40ft diameter SAG mill, commissioned in 1998, its performance, liner and discharge design attracted considerable international attention. Twelve years on, the liner design and the control philosophy are being readdressed to further improve the mill?s performance. A comprehensive series of 3D liner scans, over life of the liner, have been used to calibrate liner mass loss and progressive changes in the liner profile. This data is being used to provide true charge mass and mill filling, plus progressive trajectory profiles over life of liner. This information, along with measures of mill performance over liner life, is being used to select the optimal operating conditions as the liner wears, and hopes to inform future liner design.

Jan 1, 2012

By A. R. Sorta, D. C. Sego

ABSTRACT Temperature, solute, texture, and thixotropic effects on time domain reflectometry (TDR) measurements of water content in oil sands tailings are investigated. Results indicate that TDR water content measurements are influenced by temperature, residual bitumen, and percent clay in the tailings. A correction of 0.03–0.10/ °C must be applied to the apparent dielectric constant (Ka) to account for the effect of temperature. The Ka-volumetric water content relationships of slurries differ significantly from the commonly used Topp calibration equation. The age of the tailings and the addition of phosphogypsum did not influence TDR water content measurements.

Jan 1, 2013

By D. H. C. Wilton, W. A. Sears

"Abstract-The 682.8 ± 1.6 Ma Tickle Point Group consists mainly of rhyolite flows and tuffs with- in the Avalon Zone of the Appalachian Orogen in Newfoundland. Stratabound massive sulfide mineralization, consisting of pyrite, sphalerite, chalcopyrite and galena, occurs at the Winter Hill prospect, where a lower pyritic, Cu-rich zone and an upper Zn-Pb-rich zone are hosted by a carbonate-calc-silicate lens. At the Frenchman Head prospect, stringers of pyrite, minor sphalerite, chalcopyrite and rare galena are hosted by felsic volcanic rocks of the Tickle Point Group .Base metal ratios, metal zonations and the overall stratigraphic framework suggest that these sulfide occurrences represent volcanogenic massive to semi-massive sulfide mineralization (stringer type at Frenchman Head). As such these prospects represent the only known VMS occurrences within the Avalon Zone of Newfoundland. Pb isotope data indicate that galena from the Winter Hill prospect is less radiogenic and older than galena from Dunnage Zone occurrences, and suggest that the Winter Hill mineralization formed in an environment tectonically distinct (exotic) from the Dunnage Zone."

Jan 1, 1996

By P. E. Cavanagh

THIS REPORT is not intended to be a ?comprehensive survey of the steel industry in Europe. For the past ten years the Ontario Research Foundation has been interested in several developments in the steel industry in Europe which it was felt would be of particular interest in Canada and might possibly find application in this country. This is an account of the present status of some of these developments. The economies of steel making in Canada are somewhat different than in the United States because of the nature of the market and the relative costs and availability of such raw materials as metallurgical coke and electric power. There are cases in Europe where a great deal of effort and money has been spent in trying to make steel profitably by methods which take advantage of locally available and cheap raw materials by modifying standard practices. ln cases wl1ere the investigations have been successful and the data are available to allow translation of the results into Canadian conditions, it is felt that such information may be of great benefit in the future in Canada and at !east should be known to all those in the steel industry in this country. The field to be covered in this report only comprises the steps from concentration and treatment of iron ores to production of iron and is only concerned with those commercially operating processes well known to us and judged to be of present interest in Canada.

Jan 1, 1957

By R. R. Bassermann

THE property of Howey Gold Mines, Limited is situated at Red Lake, Ontario, one hundred and seventy-five miles east of Winnipeg and eighty miles north of the Canadian National Railway line. The mine has been a continuous producer since April, 1930, and at present is hoisting 1,350 tons of ore per day. The shaft consists of three compartments (see Figure 1, sketch B). Hoisting is done by two skips in balance, one of which is interchangeable with a cage. Sinking had been carried out at three different periods. From the surface to the 550-foot level, the sets had been spaced at seven-foot intervals, from the 550-to the 1,060-foot level at six feet, and from the 1,060-to the 1,500-foot level at five feet. During the five years of operation, the shaft had become so badly worn, due to spill from the tipples and, to a certain extent, from the loading pockets, that re-timbering was considered necessary. It should be explained that no crushing is done underground because of the ad verse effect this would have on the efficiency of the sorting. The grizzly rail clearance above the ore pockets is fourteen inches and consequently the ore hoisted is extremely coarse. More than average spill from the loading pockets and tipples under such conditions is inevitable, and it is debatable whether underground crushing with consequent increased sorting costs would not be justified. The cost of complete re-timbering according to standard practice would have been prohibitive both directly and indirectly: A lengthy shut-down of the shaft, with the consequent disorganization of the staff and workmen, could not be considered. Accordingly, it became necessary to devise some method of repair that could be carried to completion in a very limited period of time. The shaft timber for the first 250 feet was not in need of repair. From the 250-foot to the 1,060-foot level the wall plates, posts, and blocking were in good condition. These had been protected by a one-inch plank lining put in at various periods during the past three years without interfering with hoisting operations. The end plates and dividers, however, could not be protected in this manner because of the lack of skip clearance. It was decided that, if the dividers in this section could be replaced, the shaft would be as satisfactory as if completely re-timbered. Several factors had to be taken into consideration in deciding on a suit-able method of repair. In the first place, the tipple plates had to be re-designed. To complete the job in a reasonable time, not only had work to be done on the tipple, but simultaneously from several points in the shaft itself. Since all power cables, pipe lines, etc., were in the manway compartment of the shaft, this compartment had to be left undisturbed.

Jan 1, 1935

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 C. Edwards

"The processes for rare earth (RE) extraction include the decomposition of RE ore concentrate minerals, and the concurrent or subsequent leaching of the rare earth elements (REEs) from the decomposed minerals. Ore beneficiation concentrate can be decomposed by, for example, acid roasting, caustic cracking, and chlorination. The REEs can be selectively extracted. The nature of the REE extraction process depends on the concentrate mineralogy and the targeted products. Based on the experience of the Saskatchewan Research Council (SRC) in RE processing, this paper reviews REE extraction from the major RE minerals, including bastnaesite, monazite, xenotime, ion-adsorption clay, allanite, cerite, and eudialyte. An update on the development of RE extraction processes is also provided.INTRODUCTIONRE processing usually involves two operation stages, physical processing and chemical processing. Physical processing, also known as ore beneficiation, separates the RE minerals from gangue minerals to produce an RE mineral concentrate. Chemical processing, which includes extraction and separation processes, converts the RE mineral concentrate to an RE compound which is either an end product or an intermediate product for the subsequent production of individual REEs or other compounds. The RE extraction process uses one or more reagents to decompose the RE minerals and to leach the REEs into solution. The RE separation process uses solvent extraction, ion-exchange, or chemical precipitation to produce either rare earth oxides (REOs) or individual REEs. The authors have previously reviewed RE ore beneficiation processes (Zhang & Edwards, 2012). This paper focuses on extraction processes for a variety of RE minerals. Table 1 gives the chemical compositions of the RE minerals mentioned in this paper. The varieties of RE extraction methods use many different reagents. These reagents are mainly inorganic acids, alkalis, salts, and chlorine gas. The commonly used acids include sulfuric acid (H2SO4), hydrochloric acid (HCl), and nitric acid (HNO3). Sodium hydroxide (NaOH) and sodium carbonate (Na2CO3) are the most common alkalis. Salts include ammonium sulphate ((NH4)2SO4), ammonium chloride (NH4Cl), and sodium chloride (NaCl). HCl is normally used to extract RE from silicate ore minerals such as allanite and gadolinite. The alkalis and H2SO4 are mainly used to leach REs from phosphate ores minerals like monazite and xenotime. The carbonate ore RE minerals such as bastnaesite can be treated using either acids or alkalis. HNO3 is mainly used to leach eudialyte and apatite. The extraction of REs from ionadsorption clay deposits uses salt solutions. Chlorine gas (Cl2) is exclusively used in the RE chlorination process which can be used to treat any RE minerals."

Jan 1, 2012

By G. Deschênes

Located in North West Tanzania, Buzwagi is an underground narrow-vein gold mine owned by Barrick Gold. The plant was commissioned in 2000. The reserves are estimated to be more than 25 years. The ore, containing gold, silver and copper mineralization in sulphides, is processed at 3,300 tpd. This paper presents the results of an investigation to determine the performance of an ammonia-cyanide leaching approach to recover gold from two samples. Coarse gold was recovered by gravity prior to leaching. The gravity tailings had the following composition; 1.97 g/t Au and 0.31% copper for the medium cyanide soluble copper (MCSC) sample and 2.80 g/t Au and 0.53% Cu for the high cyanide soluble copper (HCSC) sample. Experiments were conducted for 72 hours in stirred tank reactors with monitoring and control of parameters. The straight cyanidation produced gold extractions of 76.1% for the MCSC sample and 84.6% for the HCSC sample. Associated cyanide additions were 6.1 kg/t and 11.5 kg/t respectively. The gold extraction of the MCSC sample was improved to 88.8% using 350 ppm NaCN and 700 ppm ammonia. The associated cyanide and ammonia additions were 2.5 kg/t and 1.7 kg/t respectively. For the HCSC sample, 82.5% gold extraction was obtained using 500 ppm NaCN and 600 ppm ammonia. The cyanide addition and ammonia additions were reduced to 3.8 kg/t and 1.0 kg/t respectively. No significant decrease in gold extraction was noticed when the cyanide and ammonia were reduced to 150 ppm and 500 ppm respectively. In the presence of ammonia, a significant amount of cyanide was oxidized to cyanate by the cupric ion. In the absence of ammonia, the formation of copper cyanide was the predominant reaction responsible for consumption of cyanide.

Jan 1, 2012

By K. Muehlenbachs, J. U. Stemler

"Abstract - The Boston deposit is a recently discovered Archean lode-gold deposit located approximately 700 km northeast of Yellowknife, in the southern portion of the Hope Bay volcanic belt. Primary H2O-CO2-NaCl fluid inclusions associated with gold mineralization at Boston are of low salinity (4.6 ± 1.4 eq. wt.% NaCl, n = 182) and contain pure CO2. Homogenization temperatures of liquid-phase inclusions average 270° ± 33°C (n = 141), but some fluid inclusion assemblages contain coexisting liquid- and vapor-phase inclusions that homogenize near critical conditions, suggesting phase separation. In these groups, liquid-phase inclusions homogenized at 299° ± 25°C (n = 24), and vapor-phase inclusions homogenized at 338° ± 28°C (n = 18). These temperatures overlap within error, and a median temperature of ~320°C is taken as representative of trapping conditions within the veins. An approximate pressure of 2-3 kbar was estimated from fluid inclusion isochores at this temperature. Stable isotopic compositions of carbonate minerals from the mineralized zones range from -4.7‰ to -3.1‰ in d13C, and from 11.2‰ to 12.8‰ in d18O; quartz ranges from d18O = 12.8‰ to 14.9‰; and pyrite ranges from d34S = 2.1‰ to 3.5‰. The d18OH2O, d13CCO2, and d34SH2S values of the fluid calculated from these minerals at 320°C are 6‰-9‰, -2.4‰ ± 0.6‰, and 1.5‰ ± 0.5‰, respectively, and are intermediate between metamorphic and magmatic compositions. A unique source for the fluids cannot, therefore, be constrained from these data. Petrographic and geochemical data for the Boston deposit are consistent with data from other Archean lode-gold deposits worldwide, implying that the Boston deposit formed in a similar environment and from similar types of fluids. These results suggest that the Hope Bay volcanic belt may represent a major new mesothermal gold belt in this part of Nunavut. Sommaire - Le gîte Boston a été récemment découvert à environ 700 km au nord-est de Yellowknife dans la partie sud de la ceinture volcanique de Hope Bay, il s’agit d’un gîte d’or filonien d’âge Archéen. Les inclusions fluides primaires de H2O-CO2-NaCl associées à la minéralisation aurifère du gîte Boston présentent une faible salinité (4.6 ± 1.4 eq. pds.% NaCl, n = 182) et contiennent du CO2 pur. La température d’homogénéisation des inclusions à phase liquide est en moyenne de 270° ± 33°C (n = 141), mais certains assemblages d’inclusions contiennent des phases liquides et gazeuses coexistantes qui s’homogénéisent dans des conditions quasi critiques, suggérant qu’elles résultent d’une séparation de phase. Dans ces groupes, les inclusions de phase liquide s’homogénéisent à 299° ± 25°C (n = 24) et les inclusions de phase gazeuse s’homogénéisent à 338° ± 28°C (n = 18). Ces températures se recoupent à l’intérieur des limites de l'erreur de mesure, et nous considérons qu’une température médiane de ~320°C est représentative des conditions de piégeage dans ces veines. Les isochores permettent d’estimer une pression d’environ 2-3 kbar à cette température. La composition des isotopes stables des carbonates dans les zones minéralisées s’étend de -4.7‰ to -3.1‰ pour d13C et pour de 11.2‰ à 12.8‰ pour d18O; les valeurs de d18O pour le quartz varient de 12.8‰ à 14.9; et les valeurs de d34S dans la pyrite vont de 2.1‰ à 3.5‰. Les valeurs de d18OH2O, d13CCO2, et d34SH2S du fluide calculé à partir de ces minéraux à une température de 320°C est de 6-9‰, -2.4‰ ± 0.6‰, et 1.5‰ ± 0.5‰ respectivement, reflétant des conditions intermédiaires entre des compositions métamorphiques et magmatiques. Il n’est donc pas possible d’identifier une source unique pour ces fluides à partir de ces données. Les données pétrographiques et géochimiques du gîte Boston sont similaires à celles d’autres gîtes d’or filoniens ailleurs dans le monde, ce qui implique que le gîte Boston s’est formé dans un environnement similaire à partir des mêmes types de fluides. Ces résultats suggèrent que la ceinture volcanique de Hope"

Jan 1, 2006

By P. K. Kaiser, C. D. Martin, F. Amann

"This article focuses on the uniaxial compressive strength (UCS) of homogeneous and heterogeneous rocks. Critical factors impacting the UCS of rocks are reviewed and then separate assessments of UCS data from homogeneous and heterogeneous rocks are presented. The variability (coefficient of variation) of UCS test results for homogeneous rocks is found to be generally <25% while for heterogeneous rocks the variability is >35%. Failure mode variation is found to be the cause of the variability of UCS test results in heterogeneous rocks. Empirical engineering design methods that use UCS as an input parameter are discussed and suggestions are provided for selecting appropriate UCS values.INTRODUCTIONThe UCS is one of the most commonly used rock engineering parameters whether for rock mass classification or for strength determination. The mean UCS and its variability are often assumed to represent a reliable rock material property. For the reasons discussed in this article, UCS is rarely representative of the intact rock UCS. The UCS test simply records the collapse load during uniaxial loading of a cylindrical specimen. As suggested by many, the UCS is an index rather than a unique engineering parameter. The UCS is thus a proxy for rock strength which depends on the loading rate (e.g., Bieniawski 1967), specimen geometry (e.g., Hudson et al., 1971), specimen size (e.g., Bieniawski 1968), and many other factors. Furthermore, the UCS is not the same as the Hoek-Brown strength criterion parameter sci.There are differences in opinion on what is commonly considered to be ‘intact rock’. When sampling ‘intact’ pieces of core from drillholes crossing blocks of rock bound by block forming discontinuities, samples will be ‘intact’ but some samples will be veined, others will be damaged, and others will be neither. According to Hoek (1983), intact rock pieces are “the unfractured blocks occurring between structural discontinuities in a typical rock mass” and the sci parameter for intact rock is determined using the fitting procedure proposed by Hoek and Brown (1997) based on the data requirements outlined by Hoek (1983). Studies (e.g., Bewick and Kaiser 2014; Kaiser and Kim 2014) have shown that independently determined UCS and sci values for the same rock type are not equivalent. Thus, UCS data cannot be used as a replacement for sci."

Jan 1, 2015