The Navachab Gold mine, an orogenic lode gold deposit, is located in the Karibib region of the Pan-African (ca. 550-500) Damara belt of central Namibia. Gold mineralisation is developed within the steeply NW dipping limb of the Karibib dome. Here, ore envelopes trend along three main orientations: a) trends shallowly towards the NE (the down plunge extent), b) trends sub-vertically in and along the down plunge extent and c) trends sub-horizontally across the down plunge extent. The down plunge extent represents the bulk of the gold mineralisation, hosting the only high grade ores mined at Navachab. As such, past work primarily focused on establishing the controls to the mineralisation observed here. The sub-vertical and sub–horizontal ore trends are seen as secondary, lower grade, being hosted in the footwall. By cutting pushbacks into the footwall, in an effort to regain access to high grade pit bottom, future gold production almost exclusively relies upon optimally mining these ores. This underlines the importance to investigate and outline the mineralising controls to the secondary ore trends. This study identified the following prevailing quartz vein sets developed within the footwall, set (1) dips shallowly towards the NE (conjugate vein set), (2) steeply towards the NW (bedding parallel veins) and (3) steeply towards the SE (S2 foliation parallel). The NW and SE dipping sets contain high average gold grades, occurring at an infrequent vein density. The NE dipping veins, as a result of occurrence density alone, was highlighted as the dominant gold hosting set. Veining occurred during the late stages of the NW-SE directed, sub-horizontal shortening (D2) event and is associated with top-to-the-NW thrusting and NW-verging folds. Re-Os molybdenite dating from auriferous quartz veins indicates mineralisation occurred at 525-520 Ma. As crustal shortening amplified the Karibib dome, flexural flow developed fractures along bedding planes, providing the control to bedding parallel veins (NW dipping). With continued crustal compression the dome later experienced fold lock up associated with reduced mean rock stress and sub-horizontal extension occurred along the steeply NW dipping limb. Horizontal extensional gashes sucked in fluids to form the shallowly NE dipping conjugate vein set. These features suggest the regional D2 strain as the first-order control to quartz vein development, down plunge and within the footwall ores. To further define the secondary ores, lithological and structural controls were evaluated on a more detailed local scale. With equal amounts of biotite schist and calc-silicate host rock (bulk of the footwall lithology) material analysed, the biotite schist units were found to contain a larger volume amount of quartz veins. The mineralisation incurred is also developed at higher average gold grades compared to that of the calc-silicates, demonstrating biotite schist having the optimal rheology for quartz vein emplacement. Normal faulting and thrusting occurs widespread, at all scale levels, across the footwall. These were primarily observed along bedding foliations and secondly at higher angles cutting across foliation. The study did not constrain the extent of these, but can conclude faulting plays a very prominent role in re-distributing the secondary ores parallel to bedding along sub-vertical trend planes. Great care should be placed in properly modelling these with 3D software such as Leapfrog. The Navachab gold mineralisation came about as a result of convergent and collisional tectonics activating metamorphic dehydration of the crustal metapelites. As these fluids ascended they absorbed gold from the crust, emplaced by either a magmatic or paleo-placer source. The gold enriched hydrothermal fluids amalgamated in large scale 1ste order structures (shearing of the steep NW limb of the Karibib Dome, the Mon Repos Thrust Zone) that acted as primary active fluid path ways. In the case of Navachab the gold enriched fluid fluxed along these pathways while interacting with fluid sinks related to a physical throttle (brittle schist, folding, bedding parallel shears) and/or a chemical trap (marbles). By summarising and detailing the fluid sinks and active fluid pathways identified by this and previous works, it is strongly recommended that a mineral approach system be designed and implemented as targeting model to lead future exploration endeavours.

Abstract Recent sea floor sulfide deposits form when seawater, heated within the oceanic crust, discharges to the sea floor. Upon mixing with cold seawater, sulfide-forming elements such as sulfur, iron, copper, and zinc are precipitated from the fluid. Actively forming sea floor massive sulfide deposits are found from different lithologic and tectonic environments varying from mid-ocean ridges to back-arc spreading centers. At a few localities, sulfide deposits are associated with turbiditic sediments that cover the axial valley of the spreading center. The southern part (Escanaba Trough) of the Gorda Ridge (NE Pacific) is one such example. At Escanaba Trough, massive sulfide deposits are associated with small sediment hills, which were uplifted by the intrusion of sills and laccoliths within the sediments. Hydrothermal deposits are dominated by pyrrhotite-rich massive sulfides, with subordinate amounts of sulfate-rich precipitates and polymetallic sulfides. Compared to deposits hosted by volcanites, Escanaba Trough sulfides contain relatively low amounts of copper and zinc. However, the average gold concentration is relatively high for a sediment-hosted deposit, and is comparable with other, Au-enriched, sea floor sulfide deposits. Despite the relatively high Au concentration in many volcanic-hosted sea floor sulfide deposits, discrete Au grains are rare. They occur mostly with sphalerite, pyrite, chalcopyrite and tetrahedrite-tennantite. Sixteen of the pyrrhotite-rich samples from Escanaba Trough were found to contain visible Au grains. They occur mostly with native Bi and various BiTe phases, and to lesser degree, with Fe-Co sulfarsenides. Transport of Au in sea floor hydrothermal systems is attributed to the presence of Au(HS)2- complex, which is destabilized when the fluid mixes with seawater. Hydrothermal fluids are generally undersaturated with respect to Au complexes and additional mechanisms, such as remobilizing earlier precipitated Au is required to explain the high Au concentrations encountered in many deposits. At Escanaba Trough the mechanism is attributed to early precipitation of Bi as melt droplets, at temperatures greater its melting temperature, as liquid Bi is capable of collecting Au even from an undersaturated fluid. Upon cooling Au is exsolved from the Bi host as native Au or maldonite (Au2Bi).

Lead isotope analyses of galena from multiple ore deposits restricted to specific tectono-stratigraphic terranes can provide information on the age and origin of the lead. In this thesis, three separate studies of lead isotopes applied to the metallogenesis of parts of the Canadian Cordillera are presented. Lead isotope data from quartz-gold vein deposits and volcanogenic and related deposits in the Insular Belt group plot in four distinct clusters on Pb-Pb plots. Each cluster corresponds to a specific deposit type and host rock category. Two parallel evolutionary trends in the lead isotopic composition exist: 1) Sicker-hosted volcanogenic deposits to Sicker-hosted veins, and 2) Karmutsen and Bonanza-hosted volcanogenic and related deposits to Karmutsen and Bonanza-hosted veins. The trends indicate a genetic relationship between host rock and isotopic composition. These observations favour a host rock source for the lead in vein deposits and, by association, a comparable source for the gold. Plutonic or abyssal direct sources of metals are not consistent with the lead isotopic data. It is suggested that the gold was extracted from the country rock, and concentrated as veins by hydrothermal activity related to Tertiary plutons. Vein deposits are isotopically distinct from volcanogenic and related deposits, providing an empirical test for distinguishing syngenetic from epigenetic deposits. Karmutsen and Bonanza-hosted deposits are more depleted in 207Pb than similar deposits in Sicker Group rocks, indicating significantly different sources for volcanic components of these two important rock units. Lead isotope data from quartz-gold veins in the Cariboo area of the Omineca Belt, and from similar veins in the adjacent Intermontane Belt indicate that these two vein types are genetically unrelated. A mid-Mesozoic model age calculated for the Cariboo gold mineralisation event indicates that all of the deposits examined are clearly epigenetic, despite reported stratiform textures at the Mosquito Creek mine. K-Ar dates from a quartz-barite vein and from regionally metamorphosed phyllite support a synmetamorphic origin for the veins, but a distal plutonic origin is not ruled out. Recent work by Godwin and Sinclair (1982) has shown that syngenetic, shale-hosted, sedimentary exhalative deposits in the autochthonous part of the Canadian Cordillera contain lead which has evolved in a high U/Pb environment. This 'shale' curve evolution model applies to deposits which have an upper crustal (host-rock) lead source. Ainsworth-Bluebell, Carmi and Slocan camps, and lead associated with the Moyie intrusions, all contain lead which plots substantially and variably below the 'shale' curve. Their departures from this curve provide evidence for a second, uranium poor, possibly lower crustal lead source, for which a growth curve, referred to as the Bluebell curve, can be constructed. The lead data are interpreted within the framework provided by these two growth curves. Mixing of lead between these two lead reservoirs is proposed to explain the linear array of data from Slocan and Carmi camps. Mixing lines, joining points of equal time on the two growth curves, provide a method for interpreting lead data from these deposits. Introduction of relatively unradiogenic lead into the upper crust via magmas which originated in the lower crust is invoked to explain the mixing. The three studies considered here illustrate the differences in lead isotopic characteristics of different tectono-stratigraphic terranes and show that the development of local models for the interpretation of common lead isotope data has application to exploration. / Science, Faculty of / Earth, Ocean and Atmospheric Sciences, Department of / Graduate

The Neoarchean greenstone belts of the Canadian Superior Province host world-class Au and base metal (Cu-Zn-Pb) massive sulfide deposits with distinct geological features, including a wide range of different host rocks and crustal settings. The range of settings is reflected in the trace metal signatures of their ores. This study examines the trace element geochemistry of pyrite from 55 different Archean volcanogenic massive sulfide (VMS) deposits in Canada to test the relationship to their host rocks, the deposit sizes and their grades. The database includes 258 samples of pyrite from 47 deposits in the Abitibi Greenstone Belt (AGB), together with 30 samples from 8 deposits in the Western Superior (Sturgeon Lake, Uchi, Benny, and Manitouwadge belts) and 45 samples from 6 deposits in the Slave Province (Hackett River, Amooga Booga, and High Lake belts). We used statistical methods to characterize the trace element geochemistry of pyrite in grab samples from the deposits, as well as larger samples representing many thousand of tonnes of ore from monthly concentrates. The study focused on pyrite mineral separates comparing samples from different deposits and different ore types within individual deposits. The analysis shows the trace element geochemistry of pyrite is a useful fingerprint of the different mineralizing systems, with trace element enrichments and depletions reflecting different source rocks, inferred temperatures of ore formation, and the scales of the hydrothermal systems. A comparison of the Abitibi samples to other deposits in the Superior Province shows distinct trace element signatures between primitive and more evolved crustal settings of different age. Similar results are found among 102 samples of pyrite from 30 deposits in Proterozoic and Phanerozoic belts across Canada. District-scale variations in pyrite chemistry mainly reflect host rock and correlate different bulk Cu/(Cu+Zn) grade ratios of the deposits. Pyrite samples from Cu-rich deposits are enriched in Cu, Bi, Co, Ni, Se, Te and Mo; whereas pyrite samples from Zn-rich deposits are enriched in Pb, Ag, Cd, In, Ga, Sn, As, Sb, Hg and Tl. The same patterns are observed in Cu-rich versus Zn-rich zones of individual deposits. Statistical analyses reveal pyrite samples from VMS deposits in the AGB that are associated with primitive mafic-ultramafic tholeiitic rocks (e.g., Potter-Doal and Genex from Timmins, and East Sullivan and Dunraine from Val d'Or camps) are enriched in Cu (>5000 ppm), Co (>1500 ppm), Se (>4000 ppm), and Ni (>250 ppm), whereas pyrite from deposits associated with tholeiitic to calc-alkaline felsic rocks (e.g., Abcourt-Barvue from the Amos-Barraute camp) are commonly enriched in Pb, Ag, Au, Cd, In, Sn, As, Sb, Hg, Tl (10s to 100s of ppm). These variations closely match primary trace element abundances in unaltered volcanic rocks compiled from over 4000 high-quality analyses of samples from the Superior Province. Whole-rock data for rhyolite confirm high concentrations of Pb, Ag, Bi, Te, Cd, In, Ga, Sn, Hg, and Tl compared to basalt and komatiite, which have higher Cu, Co, Ni, and Se. The variation in trace element concentrations in pyrite is remarkably consistent for different deposits. We note that randomly sampled pyrite from almost any part of a deposit with a bulk enrichment in a particular element shows notable enrichment in that element compared to pyrite from other deposits. Pyrite from a deposit with a bulk enrichment in Te, for example (Quemont in the Noranda camp), will almost certainly contain more Te than pyrite from other Te-poor deposits. We test this observation among 47 deposits for 15 different elements. Pyrite samples from Au-rich VMS deposits (e.g., Horne, Quemont, Bousquet #2, and Dumagami) have anomalous Au (>6 ppm) and Te (>70 ppm). Co-enrichment in other elements such as Bi, Se, In and Sn may reflect a common felsic magmatic source. Other trace element enrichments appear to reflect the scale of the hydrothermal system (e.g., depth and extent of leaching). For example, pyrite samples from several large-tonnage deposits (Kidd Creek, Horne #5 Zone, and Geco) have high Sn concentrations (from 450 to 15000 ppm) possibly reflecting the large volumes of felsic rock from which the Sn was extracted. In other deposits, co-enrichment of Sn with Bi (>100 ppm) and In (>10 ppm) suggest a magmatic contribution to the ore fluids Principal Components Analysis (PCA) combined with hierarchal clustering confirms systematic trace element variability in pyrite from deposits with different host rocks and bulk Cu/(Cu+Zn) ratios. However, pyrite from deposits in different terranes seems to record major differences in the crustal compositions of those terranes. For example, pyrite samples from bimodal-felsic deposits show the same trace element signatures (i.e., enrichments in Ag, As, Sb, and Hg) in the AGB and in the Western Superior. In contrast, pyrite samples from deposits in the Slave craton tend to show a distinct enrichment in Pb, U and Th that may be related to the more mature and thicker crust in the Slave compared to the AGB. Other deposit types (magmatic Cu vein deposits, orogenic Au deposits) also show dramatically different pyrite compositions. Pyrite concentrates from magmatic Cu vein deposits in Chibougamau are enriched in Cu, Co, Ni, Te, As, Sb compared to VMS in the AGB, and samples from orogenic Au deposits in Timmins and Val d'Or are enriched in Au and Mo and depleted in Pb, Bi, As, and Sb compared to VMS. These differences highlight the potential application of the trace element signatures of pyrite during exploration for different deposit types in the same region. Trace element signatures of pyrite in grab samples compared favourably to much larger bulk samples from the same deposits (e.g., monthly concentrates and mine tailings) giving some confidence that the much smaller samples can provide a reliable first-order fingerprint of the deposits as a whole. LA-ICP-MS analyses of individual pyrite grains also agreed well with bulk analyses of pyrite over a wide range of trace element concentrations (10s to 100s of ppm).

Elevated manganese (Mn) concentrations have been measured in groundwater within the Roanoke River watershed, Virginia. Concentrations of Mn often exceed the secondary drinking water standard. A historic belt of Mn ores, the James River-Roanoke River Manganese District (JRRRMD), occurs in the eastern part of the watershed. The project objectives were to 1) evaluate the formation of the JRRRMD ore deposits and 2) analyze existing groundwater chemistry data to evaluate sources and processes that control groundwater Mn. Analysis of ore minerals, morphologies, and chemistry provides support that the ore deposits are supergene in origin, consistent with previous work. Spatial correlations between Mn ore locations and stream terrace deposits support a model of ore formation in which Mn-oxides were precipitated near discharge zones as anoxic groundwater mixed with oxic groundwater. Terrace deposits present at elevations higher than modern streams suggests that topography has been inverted, allowing ores to be found at higher elevations than what is typically associated with ores formed in discharge zones. Analysis of groundwater chemistry data shows positive correlations between Mn, calcium and bicarbonate concentrations in groundwater, suggesting that carbonate-bearing lithologies are probable sources of Mn to groundwater. Regionally, groundwater flows toward the Roanoke River where the flowpath terminus is marked by elevated Mn. The inverse correlation of Mn with dissolved oxygen suggests that reducing conditions that develop along flowpaths allow for Mn to persist in groundwater. Overall, results suggest that the same processes that allowed for formation of the JRRRM ore deposits continue to occur today. / Master of Science

No description available.

Thesis (Ph.D.)--University of the Witwatersrand, Faculty of Science, 1987

I denna uppsats studerar vi kommutativitet i polynomringar i en variabel med icke-kommutativ multiplikation över en domän, så kallade Ore-utvidgningar. Vi Ore-utvidgar en polynomring i en variabel med koefficienter tillhörande en kropp. Vi studerar centralisatorer i denna Ore-utvidgning. Centralisatorn av ett icke-konstant element visar sig vara kommutativa och fria moduler. Vi undersöker även när centralisatorerna som en algebra är genererade av ett element. Några nya exempel av sådana centralisatorer presenteras. Vi undersöker också en ny metod för att bestämma om centralisatorn som en algebra är genererad av ett element. / In this thesis, we study commutativity in polynomial rings in one variable with noncommutative multiplication over a domain, which is called Ore extensions. We look at the Ore extension of a polynomial ring in one variable with coefficients belonging to a field. We study centralizers in this Ore extension. The centralizer of a nonconstant element turns out to be commutative and a free module. We also investigate when the centralizers as an algebra are generated by one element. Some new examples of such centralizers are presented. We also look into a new method to determine if the centralizers as an algebra are generated by one element.

Extensões de Ore são anéis de polinômios, denotados por R[x o &], nos quais a variável x e os elementos de R não comutam necessariamente. Algebras de Hopf fracas são algebras que tamb em são coálgebras e satisfazem um conjunto de axiomas de compatibilidade entre essas estruturas. Neste trabalho investigamos extensões de Ore cujo anel base e uma algebra de Hopf fraca. Mais especi camente, dada uma algebra de Hopf fraca R, estudamos sob quais condições R[x o &] e uma algebra de Hopf fraca com uma estrutura que estende a estrutura de R. Sob certas hipóteses, obtemos condições necessárias e su cientes para que a extensão de Ore seja uma álgebra de Hopf fraca, obtendo assim um resultado que generaliza um teorema de Panov para o contexto de algebras de Hopf fracas. / Ore extensions are polynomial rings, denoted by R[x o &], in which the variable x and the elements of R do not commute necessarily. Weak Hopf algebras are algebras which are also coalgebras and satisfy a set of axioms of compatibility betweem these structures. In this work, we investigate Ore extensions whose base ring is a weak Hopf algebra. More speci cally, if R is a weak Hopf algebra then we study under what conditions R[xo &] is a weak Hopf algebra extending the structure of R. Under certain hypotheses, we obtain necessary and su cient conditions for an Ore extension to be a weak Hopf algebra, obtaining a result that generalizes a Panov's theorem to the setting of weak Hopf algebras.

O cenário atual das commodities minerais tem obrigado as empresas de mineração a otimizar suas operações e buscar soluções inovadoras para manter a lucratividade desejada de suas minas, controlar os riscos envolvidos e garantir a sobrevivência do negócio. Neste contexto, com a menor disponibilidade de minérios de alto teor, os depósitos minerais remanescentes exigem a lavra e o beneficiamento de grande quantidade de massa por tonelada de metal produzido. Apesar de apresentarem um teor médio decrescente, estes depósitos frequentemente contêm regiões com minério de teor mais elevado incorporados às porções de baixo teor ou mesmo na massa estéril. Por melhor que seja o conhecimento das reservas, a lavra seletiva dessa massa de teor mais elevado é praticamente inviável operacionalmente. A forma mais eficiente de aproveitar estas reservas é por meio do processo de pré-concentração de minério por ore sorting (seleção/classificação de minério). O objetivo deste processo é remover, quanto antes possível, o material com teor não econômico. A consequência imediata é uma potencial redução dos custos de processamento e do consumo de água e energia para a mesma massa de metal produzido. Existem várias tecnologias aplicáveis em pré-concentração; este estudo é focado especificamente na apresentação e discussão dos impactos causados no empreendimento mineiro de Phu Kham (PanAust), a partir da instalação e utilização do equipamento de pré-concentração, bulk ore sorter. Este equipamento faz uso de um sensor de ressonância magnética, e tem por objetivo a separação de grandes volumes de estéril direto no transportador de correias totalmente carregado (in-pit ou correia de alimentação da usina), baseando-se nos teores medidos ou inferidos a partir de seu sensor. A pré-concentração com bulk ore sorter é fundamentalmente utilizada para aumentar a produtividade baseada na capacidade fixa da usina e atenuar os prejuízos causados pelas incertezas nas definições de teores da mina. Isto é, a partir da maior precisão no controle do teor, pode-se reduzir a diluição e a perda de minério oriundas da operacionalização e explotação das reservas minerais, controlar o teor de corte e o teor médio na alimentação da usina ou ainda separar os diversos tipos de minérios para serem tratados em diferentes rotas no processo de beneficiamento. Este trabalho estuda a viabilidade técnica e econômica da aplicação do bulk ore sorter na mina de Phu Kham, no Laos. Inicialmente a tecnologia de sorting é apresentada, juntamente com os sensores disponíveis e o sistema diversor. Os resultados econômicos foram avaliados a partir dos dados de produção da mina no ano de 2014. Sem a possibilidade da realização de estudos de amostragem, a variabilidade do minério foi estimada a partir da variância de dispersão, uma ferramenta geoestatística. A estimativa do desempenho da separação inclui os erros associados a leitura do sensor e do sistema diversor. Avaliações econômicas foram realizadas para diferentes cenários e demonstraram que o bulk ore sorter tem o potencial de aumentar a produção metálica e o lucro. Este sistema de pré-concentração permite a que apenas o material com valor agregado alimente a planta de beneficiamento. / The current scenario of mineral commodities has forced the mining companies to optimize their operations and look for innovative solutions to maintain the desired profitability of their mines, control the risks involved and ensure the survival of the business. In this context, with the depletion of high grade deposits, the remaining lower grade deposits require the mining and processing of larger volumes of material per tonne of product. These low grade deposits often contain a large proportion of barren gangue. The aim of pre-concentration is to remove this barren material at as coarse a particle size and as early in the process as possible. This has the potential to reduce processing costs, energy and water consumption and possibly ore transport requirements. There are several technologies that may be applicable for pre-concentration; this work looks specifically at bulk ore sorting. This involves the separation of a large volume of barren gangue from a fully loaded conveyor belt (in-pit or plant feed belt) based on the grade as measured or inferred from a sensor measurement. Bulk ore sorting may be used to increase the production rate through fixed plant capacity, or reduce the required size of downstream processing equipment. Uneconomic or marginal reserves may be upgraded making them economic to treat and improving the resource utilisation. Bulk ore sorting could also be used to reduce dilution and ore loss in mining operations by improving grade control, or to separate ore types to treat via different process routes. Thus, bulk ore sorting has the potential to improve the profitability and reduce the environmental impact of mining operations and may be applied in different ways. This study investigates three examples of how bulk ore sorting may be employed and the associated economic benefits. The technical and economic viability of bulk ore sorting is evaluated for a copper deposit in Laos. Bulk ore sorting technology is described briefly along with discussion of appropriate sensor and diversion systems for the copper deposit considered. The economic implications were evaluated using the actual mined material and costs from 2014 as the base-case. In the absence of sampling data, ore grade variability was estimated using geo-statistical tools based on mine grade control data. A process model of the sorter was developed considering the accuracy of the sensor measurement and errors associated with material diversion. This was used to simulate the separation performance. Economic evaluations were conducted for a number of different scenarios and demonstrated that bulk ore sorting has the potential to increase the amount of metal in product and annual profit for the operation considered. This is because it is cheaper to sort and reject below cut-off grade material contained in the plant feed than to treat this material through the processing plant. The sorter acts as a gatekeeper - only above cut-off grade material (value-adding) reports to the processing plant.