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By J. S. Stewart

THE Mackenzie River Valley proper extends from Great Slave lake to the Arctic ocean (see map, Figure 1). From Fort Providence, near Great Slave lake, to Aklavik, near the Arctic coast, the distance by river is 1,000 miles. This will give some idea of the extent of the region. Throughout the length of the river valley, the Precambrian complex rocks are overlain by unaltered sedimentary rocks several thousand feet in thickness. These sediments are dominantly marine and Palaeozoic in age. In places, the Palaeozoic strata are overlain unconformable by Cretaceous an Tertiary sediments. The strata have been folded and, in general, the inclination of the beds to the horizontal is comparatively gentle. Here, then, we have the essentials for potential oil fields covering an area over 1,000 miles in length. This is probably the largest comparatively unexplored potential petroleum region in North America, and surface evidences of oil in the form of seepages occur in several places. Many of these seeps were known to the early explorers, and the first Geological Survey report on the Mackenzie River region, containing the results of field work carried out by R. G. McConnell in 1887-88, noted the petroliferous character of the Devonian strata and the indications of petroleum(l). This region is, however, handicapped by distance from markets and sources of supply of materials; by lack of modern transport facilities; by short season for river transport, and mountainous country separating the valley from the Pacific seaboard; and by the long winter season with weather that makes outdoor work difficult. These factors have tended to discourage the enthusiasm of would-be promoters, and so development in this region has been slow In 1920, a test well drilled about fifty miles northwest of Fort Norman discovered oil in commercial quantity, and three additional wells proved a small area to be productive.

Jan 1, 1944

By T. J. Fee

Introduction The International Nickel Company have been pioneers in mine rescue work in Ontario. Following the fire in the Hollinger mine in the spring of 1928, the Company purchased ten self-contained oxygen breathing outfits, ten all-service gas masks, an oxygen high-pressure pump, life line, inhalator, ten cap-lamps with batteries, and a complete stock of spare and accessory parts. An instructor from the U.S. Bureau of Mines, Pittsburgh, came to the Creighton mine for one week and trained twenty men in the use and limitations of all equipment. Monthly training courses were commenced in the fall, and have continued to the present time. Of the original twenty men trained, ten are still on the Creighton mine rescue teams. In the meantime, and as an aftermath of the Hollinger fire, the Work-men's Compensation Board of Ontario and the Ontario Department of Mines established a central mine rescue station at Timmins. This was completed in the spring of 1930. It was then decided, in view of the great expansion in mine development in the Sudbury district, to build and equip a similar station to serve all the mines there, namely, Creighton, Frood, Garson, Falconbridge, Sudbury Basin, and Errington. An extensive survey was made and a central location selected opposite the Frood mine. The International Nickel Company donated a site, and the station was completed in the fall of 1930. The station is a two-storey brick building, 27 ft. by 47 ft., with a large garage adjoining it. The ground floor is used for equipment and for training the crews. On the second floor are living quarters for the superintendent and his family. The garage houses a closed truck, in which all necessary equipment can be taken to any desired point for training purposes or for actual mine rescue work, if and when necessary. The garage is also used as a gas room, where the men are trained in an irrespirable atmosphere. A gas-tight observation window between it and the training room enables the superintendent to see and direct the men undergoing training, weight exercisers and other devices being used.

Jan 1, 1935

By Tapio Leskinen

"In the 1960’s, when large flotation machines appeared on the market, this was a part of a world-wide trend towards the use of ever larger machine units.One factor that has played a decisive part in this development is the increased of the amounts of ore that have to be handled as the exploitation of successively poorer ores becomes necessary. The advancing application of instrumentation and automation has worked in the same direction.In 1970, a working group was set up within Outokumpu Oy, for the purpose of flotation machine development. The principal task of this group was to be the design and development of- a new type of rotor of increased efficiency, especially for the dispersion of large amounts of air- a mechanical-pneumatic flotation machine of substantially larger size that existing plant.This development project required an extensive program of theoretical investigation and experimental studies at several concentration plants of the company.The commonest flotation machine presently used in the concentration plants of Outokumpu Oy is the VK-3. It is of the mechanical-pneumatic type, and the volume of its cell tanks is 2.5 m3 per shaft. This machine was developed within the company, and its first development phase i s represented by the OKKO 3, built for the Kotalahti concentrator in 1959. It was, in its time, comparable in volume to the largest machines that were then in use anywhere in the world."

Jan 1, 1976

By Newell G. Alford

COAL companies are definitely becoming conscious of the need for better organization. Mechanized mining and the intricate preparation of coal have brought with them the demand for better trained and more highly skilled mining personnel. Refinement in coal mine management is increasing the demand for young graduate engineers to start in the lower grades of the reorganization and prepare themselves to meet rapidly increasing coal mining responsibilities. The interest is wide in scope and extends from self-helps, extension work, and management study at the mines to stimulation of interest in coal mining amongst undergraduate mining students, leading to their permanent employment after graduation. Data by the United States Bureau of Mines show the total annual value of coal production (bituminous and anthracite in the U.S.) through the past 22 years as being more than 38 per cent of the combined value of all ocher mineral industry products. Natural-gas production has gained rapidly on coal since 1916; petroleum exceeded coal beginning with 1934 (see Figure 1). For this 22-year period, the Committee on Student Interest in Coal Mining, in the A.I.M.E., made a survey of the mining schools in Canada and the United States to show the approximate absorption of young graduate mining engineers by the coal industry. In 1938, 232 mining engineers were graduated as compared with 42 in 1918 (a war year) and 180 in 1923. From 1923 to 1929 the graduates declined to another low of 90, but from there on there was a rapid annual increase in the graduating output, reaching a total of 232 in 1938 and continuing upward to a maximum of over 325 graduating this present month.

Jan 1, 1940

By Tarrant Elkington

Underground mine planning practices are typically manual and time consuming and will often produce sub-optimal solutions due to the application of generic rule-of-thumb assumptions or ?tried and tested? choices. The development of computer-automated tools for underground mine planning enable a wide range of scenarios to be considered rapidly in arriving at an optimised solution, subject to input assumptions. We have applied these tools to demonstrate ?optimised? solutions that challenge industry standard practices. Further, we have incorporated grade and price/cost uncertainty into these tools to gain an understanding on how the consideration of uncertainty may further alter strategy selection. Finally, we enable the tools to dynamically adjust cut-off grades in conjunction with the evolution of prices/cost.

May 1, 2010

By Damian Connelly

The metallurgical heart of the majority of copper oxide processes is the heap leaching operation. Fresh low grade sulfide ores is a far greater challenge.The success of this unit operation is so important that operators need to understand the ramifications of their decisions or non-decisions and how they could affect the downstream processing. Heap leaching is an operation that can be described as a scientific art whereby principles of science can be used to guide experienced personnel in attaining their goals. Recent advances have been made based on better understanding the ore characteristics. The success of a heap leach operation, or otherwise, is dependent upon a number of factors, notably: The type of ore to be treated The extent of testwork completed to define the process The interpretation of the testwork results Ore preparation prior to stacking Agglomeration and curing requirements Other issues are: Compliance with the ’smaller’ requirements of environmental authorities Adequate and timely drainage of solution from the base of a leaching lift The degree of saturation in the pads Heap porosity – coarser or finer ore, which is better? Higher or lower irrigation rates – which would produce a better result? Laying out irrigation pipework for ease of operation and maintenance The addition of water, raffinate during agglomeration where needed The heap leaching of sulfide ores, including chalcopyrite, via enhanced bacterial leaching is discussed. INTRODUCTION Over the last twenty years, production of extremely pure copper from solvent extraction and electrowinning (SX/EW) has become more common in a market of real declining prices for copper metal. This paper touches on a number of heap leach operations and experiences. These projects were subjected to extensive engineering design and resulted in low capital cost, efficient operations with a number of lessons learned. The key to these operations was the performance of the heaps and associated fluid collection systems rather than the plants themselves.

Jan 1, 2019

By Zhuo Chen, Jun Zhou, Jiemin Zhou

Samples of the reacting suspension particles were taken from the reaction shaft of a commercial flash furnace at different heights. The processes of matte and slag formation and morphology of the samples in the shaft were observed and analyzed. Variations of matte and slag composition in the shaft were investigated. By analyzing the samples taken from the settler at different locations, settler processes were studied such as slag formation, agglomeration and separation of matte-slag phases, Fe3O4 reduction, copper-in-slag reduction etc. In accordance with the characteristic of matte and slag formation the flash furnace can be divided into 6 functional regions, and a mechanism model called as “Micro bath process” is proposed for matte and slag formation in a commercial flash furnace. INTRODUCTION During flash smelting, concentrate particles scattered in the reaction shaft complete their oxidation reactions in suspension. The droplets produced fall into the settler beneath with a small amount of un-melted solid particles, agglomerate and separate into matte and slag layers which are tapped out from their tap holes separately. With respect to the extent of slagging reaction in the reaction shaft, there isn’t any commonly accepted understanding in literature (Kimura, Ojima, Mori & Ishii, 1986; Kemori & Kondo et al.,1988; Kemori, Denholm & Kurokawa, 1989; Fagerlund &Jalkanen, 2000). Some assume that the slagging reaction in the shaft can be ignored and some propose that all slag is produced in the shaft. With regards to the characteristics of the physical-chemical changes of concentrate particle in the reaction shaft, although there are some research results reported from pilot scale and commercial scale experiments (Kemori, Ojima & Kondo, 1988; Hagni, Vierrether & Sohn, 1988; Jorgensen, Taylor, Dickinson & Stevenson, 1992; Swinbourne, Simak & Yazawa, 2002; Kivinen & Tuominen, 2005), they are generally limited themselves to theoretical analysis and calculations. No studies are published concerned with the processes of matte and slag generation in the shaft, quantitative analysis of the processes, variations of matte and slag composition and morphologies of the melts along the axis of the shaft, etc.

Jan 1, 2019

By Sangjian Zhang, Zhu Xiao, Lairong Xiao, Rushi Chen, Mei Fang, Zhou Li

The copper-graphite metal matrix composites were fabricated by two different production methods: directly sintering with copper and graphite (Cu-Gr composite), and in-situ sintering with cuprous copper and graphite (Cu2O-Gr composite). During the in-situ reaction sintering, a redox reaction occurs at the interface of the Cu2O particles and Gr particles, resulting in the improved physical and mechanical properties of Cu2O-Gr composites. Compared with Cu-Gr composite, Cu2O-Gr composite has a higher compression strength at any given graphite content due to the improved interface bonding and uniformly distributed fine graphite particles in the matrix. Increasing graphite content decreases the mechanical properties of the copper-graphite composites. The compression strength of Cu2O-Gr composite is 50-100 MPa higher than that of Cu-Gr composite. INTRODUCTION Copper-graphite composites have been widely used in industry due to their low density, relatively high strength and hardness, good thermal and electrical conductivity, arc ablation resistance and self- lubrication ability (Grandin & Wiklund, 2018; Han & Xu, 1999; Huang et al., 2017). Electric sliding contact friction pair in the high-speed railway system and motor brush in electric contact components have a very high demand on electrical conductivity and self-lubrication ability of materials. The electric devices need to bear both mechanical load and high electric current. Copper-graphite composites with good mechanical, electric and self-lubrication properties are the best candidates for application in multifunctional electric sliding contact frication components. The relative sliding speed of friction pair is developing towards 100 m/s or even faster, which can withstand the electric flow of more than 800 A, and it is often accompanied with vibration and impact (Guo & Ding, 2008). Conventional wear-resistant materials can no longer meet the current requirements and this promotes the development of copper- graphite composites with excellent comprehensive properties (Ding, Chen, Zhu, Zhang, & Zhou, 2009; Yasar, Canakci, & Arslan, 2007). There are many ways to fabricate copper-graphite composites, mainly including powder metallurgy method, liquid immersion infiltration method, rapid sintering method and electrodeposition method (Kalin & Poljanec, 2018; Kim, Lee, & Han, 2014; Liu et al., 2018; Wang et al., 2018). Each method has its own advantages and disadvantages, and the powder metallurgy method is the most commonly used one due to its simple process, high economic benefit, and industrial production applicability. The common steps of the powder metallurgy method are as follows (Samal, Parihar, & Chaira, 2013): firstly mixing the copper (copper alloy) powders and graphite with a certain ratio, following by die-forming, and then sintering at the high temperature with or without pressure. The advantage of copper-graphite composites is that the graphite distributed in the copper matrix uniformly and it can form a carbon diaphragm preventing the wear of the composite (Fu & He, 2010). However, the bonding strength of the interface between the copper matrix and the graphite is relative low as the combination between them is mechanical interlocking but not metallurgical bonding (Ma et al., 2008). It is meaningful to develop new fabrication technology of copper-graphite composites with better interface bonding, which could provide guidance for the design of copper-graphite composites with excellent comprehensive properties.

Jan 1, 2019

By Rob Sloan, Simon Yu, Rob Robison, Mark Schmidt

"In 2000-2009, Highland Valley Copper (HVC) tested new generation 660 mm cyclones supplied by various manufacturers in an aim to replace existing first generation 762 mm cyclones without success. Due to plant infrastructure limitations, new generation larger (838 and 800 mm) cyclones were not feasible. In 2010, an approach of testing hybrid 800 mm diameter cyclones was initiated by HVC. The hybrid cyclone consists of an 800 mm cyclone upper head section and the use of existing 762 mm cyclone conical sections. Three hybrid 800 mm cyclones were engineered and fabricated, targeting a throughput of 850-950 mtph and were installed in the E line cyclone cluster. Since starting operation on September 20, 2011, three hybrid 800 mm cyclones have been running at combined throughput ranges of 750 to 1050 mtph. Based on plant Daily Met Reports, the test hybrid cyclones demonstrated a decrease in misplacement of coarse copper bearing material to the cyclone overflow and better cyclone efficiency when compared to the existing 762 mm cyclones at the same throughput. The hybrid 800 mm cyclones also demonstrated significantly higher capacity per unit when compared to the existing cyclones. Based on the improvement demonstrated in the full plant trials, HVC decided to install the new hybrid 800 mm cyclones in all five grinding lines starting in July, 2012. A more detailed discussion of the full plant trials and initial results of the first complete retrofit of E line will be presented. INTRODUCTIONHighland Valley Copper Partnership (HVC) is Canada’s largest base metals mine and one of the world’s largest producers of Cu-in-concentrate. The mine, located in the central interior of British Columbia, is owned (97.5% interest) and operated by Teck Resources Ltd. In 2011, HVC produced about 210 million lbs. of Cu and 8 million lbs. of Mo. HVC’s Mill Modernization Project (MOP), which was budgeted for $475 M, is currently under intensive construction and is scheduled for commissioning in Q1 2014. The following description was valid at the time of writing this paper.The operation has been expanded a number of times since the startup of the original Lornex mill in 1972. The plant started up with two 9.75 m (32 foot) diameter primary SAG mills. It was expanded in 1981 with the addition of a 10.36 m (34 foot) diameter SAG mill. In 1988, the Highmont concentrator grinding mills and flotation section were moved and attached to the Lornex mill site. This added two10.36 m (34 foot) diameter AG primary mills to the Highland mill. This brought the total number of primary mills to five referred to as Line A, B, C, D and E: three SAG mills (A, B and C) each with two ball mills in parallel and two AG mills (D and E) each with one ball mill. Each AG mill line is equipped with one short head pebble crusher. Ball mill circuits are classified using 8 clusters of ten Linatex 762 mm cyclones to reach a target P80 330 micron (equivalent 43% +100 mesh), before the cyclone overflow product advances to the bulk flotation circuit."

Jan 1, 2013

By Davis Myers

OVERVIEW ?A summary of population demographics data ?Potential impact of Baby-Boomer, Generation X, Millennial demographics in the mining workforce ?Planning with constrained resources Age Cohorts ?Baby Boomers: born 1946 -1965 ?Booming economy, TV, cold war, JFK, Vietnam ?Generation X: born 1965-1980 ?There was always space travel ?PC?s, Black Monday, Challenger, Desert Storm, fall of Berlin Wall ?Millennials: born 1980-2000 ?There was always the internet ?IM, Facebook, Sept 11, Columbine

Nov 1, 2011

By J. E. Sepulveda, R. Morrow

"In full scale mineral grinding operations, the steel balls normally utilized as auxiliary grinding media get to be consumed at different specific rates, as determined by 3 basic mechanisms: abrasion/corrosion (wear), impact ball breakage and impact surface spalling, the latter relating to the removal of rather small pieces of steel coming off the exposed surface of the balls due to repetitive impact with other balls in the charge or the surrounding mill liners. In conventional ball milling applications, the clearly predominant mechanism is gradual wear by abrasion/corrosion. In the case of SAG mills, where the make-up balls may be as large as 6”f diameter, breakage and spalling mechanisms are significant contributors to total ball consumption, although not to the extent of overriding wear mechanisms.Ball wear has been well characterized by the Linear Wear Theory, as presented at the last CMP 2014. Impact breakage has been customarily monitored, at the ball manufacturing site, via Drop Ball Tests (DBT) consisting of repeatedly dropping balls from a standard height and accounting for the number of broken balls and the weight loss (spalling) of the remaining unbroken balls. When properly conducted, DBT’s should provide meaningful indications of the breakage and spalling resistance of large diameter balls. The current publication is focused on the interpretation and projection of DBT results to full industrial scale, taking into account all 3 mechanisms of grinding media consumption in SAG mills.INTRODUCTIONA thorough, quantitative understanding of the independent contribution of all active grinding media consumption mechanisms is of the highest practical relevance, particularly in semiautogenous (SAG) milling operations where ball breakage and spalling can not be ignored or neglected, as it has been normally acceptable for conventional ball milling processes. The high severity impact conditions currently imposed in typical SAG mill environments create serious challenges to the manufacturers and suppliers of large diameter (5”–6”) SAG balls, whose products must be capable of surviving such harsh operational conditions with minimum breakage or spalling."

Jan 1, 2015

By T. Backers

In fracture mechanics modeling the rock mass is seen as the combination of rock material and fractures, which are evident on different scales. The activation and propagation of fractures dominates the stability and performance issues as well as fluid flow potential. Once fractures have started to grow into each other and create through going separations, stability of rock engineering constructions may be violated. Further, as fracture mechanics inherently incorporated a scaling law in the form of the participating fracture length, scaling issues can be more accurately mirrored. This contribution highlights some examples of application of a fracture mechanics based code on some rock mechanics issues such as numerical laboratory experiments, borehole stability or the creation of fluid pathways due to underground constructions. The results show explicitly the creation of fractures and connection of existing ones, and hence the safety and performance issues. It has to be concluded that the basis for frequent application of fracture mechanics on specialized challenges in Geomechanics is laid, and the needs for future developments are defined.

May 1, 2009

By C. O’Hara, S. Roy, K. Swedburg, J. Katchen

"New Gold Inc.’s New Afton concentrator commenced operation in June 2012, producing a copper-gold-silver concentrate. During the first two years of operation, various optimization efforts have improved mill throughput, recoveries and concentrate grade. Daily mill throughput has increased 20% above the design rate. This was achieved through more consistent instantaneous feed rate and size distribution from the mine reclaim stockpile, higher pebble crusher reliability, changes to the SAG and ball mill grinding media sizes, larger SAG grate openings, reduced SAG discharge screen-deck aperture size and an increase in the maximum ball mill power draw limit. Metal recoveries at New Afton are primarily limited by liberation in the coarser size fractions; therefore, improving the grind size at a given mill throughput has been the main means of improving recovery. The average rougher flotation feed F80 size was reduced from 175µm in Q1 2013 to 165µm in Q1 2014, while the average operating mill throughput increased by 25%. Other flotation circuit changes focusing on recovery improvement include the addition of froth crowders to the rougher cells and the use of a stronger frother. The average concentrate grade increased from 22% Cu in Q1 2013 to 25% Cu in Q1 2014. This was achieved by redirecting the flash flotation concentrate stream to the regrind circuit and by the addition of lime to the cleaner circuit.This paper describes the above improvements made at the New Afton mill. The paper describes the objectives, timing and end result of each area of improvement.IntroductionThe New Afton Mine is situated approximately 10km west of Kamloops, British Columbia, 340km north-east of Vancouver. The mine has a moderate climate, access to transportation and electrical infrastructure, nearby fresh water supply, and close proximity to industrial and mining service businesses. The population of Kamloops is around 90,000 with most mine employees living in or around the city."

Jan 1, 2015

By K. D. Hester

"1.0 INTRODUCTIONThe uranium industry started in Canada with the reopening of the gravity concentrator at Eldorado's Port Radium mine in the North West Territories in 1942. In 1948, uranium exploration and processing was opened to the private sector, and the Federal government offered special pricing incentives to the industry. Eldorado explored the Lake Athabaska deposits in northern Saskatchewan, and in 1953 opened the Beaverlodge mine. Other mines in the area soon followed. In 1953 also, the deposits in the Blind River - Elliot Lake area and in the Bancroft area of Ontario were discovered. Mines and mills were quickly developed in these areas.Canadian uranium production peaked at 15,900 short tons of U3O8 in 1959. In that year, uranium exports at $300,000,000 exceeded those of all other minerals in value. On October 1, 1959, Canadian uranium operations employed 13,900 people.Up to 1956, it was a seller's market. By 1958, world production of uranium was far ahead of requirements. Accordingly, in November of 1959, the USAEC announced it would not renew it's options for continued delivery of uranium from Canadian mines. This led to the ""stretchout"" of existing uranium contracts, and many mines were shut down (see Table 1) .In the early 1970's, a commercial uranium market seemed to be developing. In Ontario, at Elliot Lake and Bancroft, mines reopened. But competition was coming from the Athabaska basin of Saskatchewan. Currently it is these high grade, open pit deposits of northern Saskatchewan which are in the forefront of Canadian uranium developments. Canadian developments since 1942 reflect the political and market forces, as well as metallurgical progress and changes in environmental standards. This paper attempts to show these process changes in broad perspective."

Jan 1, 1987

By Phillip Cancilla, Daniel Roy, Frank Rosenblum

"The manual moisture determination methods employed today by mineral processing plants falls short in providing timely information required for automation control. The costs associated with transporting and handling concentrates still represent a major portion of the overall treatment price. When considering the cash flow of a mining operation, that is governed by both the smelter contract, with moisture penalties and the quantity and quality of the concentrates shipped, an efficient method of on-line moisture content would be a welcome tool.A novel on-line determination system for ore concentrate moisture content would replace the normally tedious classic manual procedure. Since the introduction of microelectronic-based control systems, operators have strived to reduce the treatment costs to the minimum therefore a representative and timely determination of on-line moisture content becomes vital for control set points and timely feedback. Reliable sensors have long been on the “wish list” of mineral processors since the problem has always been that you can only control what you can measure. Today, the task of moisture determination is still done in the classical technique of loss in weight by uncontrolled procedures. These same methods were introduced in the earliest base metal concentrators. Generally, it is acceptable to have ore concentrate moisture content vary within a range of 7 to 9%. Many times, delays in manually achieving reliable feedback on the moisture content results in the moisture varying from 5 to 12% before corrective actions can be made.This paper first reviews the traditional and widely available methods for determining moisture content in granular materials by applying physical principals and properties to measure moisture content. All methods are in some form affected when employed on mineral ore concentrates. This paper describes a unique and promising on-line moisture sensor in two mineral processing applications, which not only automates the tedious tasks but also results in reliable moisture feedback that can be used in the optimization of the de-watering process equipment such as pressure or vacuum filters and fuel-fired driers. Finally, two measurement applications will be presented to indicate the usefulness and to summarize the measurement requirements for the proposed method of employing drag force and mechanical properties of the material itself to determine the moisture content."

Jan 1, 2001

By Thomas Deschamps

The main objective of this work was to investigate the effect of binder addition on the geochemical properties of surface paste tailings. Nine leaching columns were prepared following different disposal configurations: non-cemented, non-cemented with cemented layers, and cemented .The proportions of cement by total paste weight were between 0 and 1% for a whole column; the maximum amount of cement in a single layer was 2%. The material used was a sulphide-rich tailings coming from a Canadian hard rock polymetallic mine. The columns have been flushed once a week during 30 weeks. The recovered leachates were analyzed for pH, Eh, conductivity, and a wide range of elements. The drainage rates were monitored during the initial disposal as well as after each column flush. The results show that binder addition can improve the whole environmental performances for some of the studied disposal configurations. The disposal configuration consisting of intercalation of three 2% cement paste layers within non-cemented layers (0.66 wt.% of cement for the whole column), and the configuration with 2% cement added only in the two bottom layers (0,44 wt.% of cement for the whole column), showed the best results in terms of water quality and drainage rates.

May 1, 2007

By E. T. C. Spooner, Ioannou S. E.

"Abstract- The Dixie Zone Au-Ag deposit is a low sulfidation (adularia-sericite) epithermal system located 1.4 km from the producing Ken Snyder mine (7.0 M oz Au equivalent reserve/resource), Midas Mining District, north-central Nevada. This WNW striking, steep south dipping, fault system is hosted by mid-Miocene (17 Ma to 14 Ma) felsic volcanic rocks that were erupted from a vent(s) to the northeast. The system was mineralized by approximately 250 ± 20°C (n = 86), near-neutral pH, ascending fluids of predominantly meteoric origin [volatile corrected salinity = 0.52 ± 0.35 eq.wt% NaCl (n = 86)]. Bulk analytical fluid inclusion gas chromatography identified a mineralizing fluid volatile composition of 99.68 ± 0.18 mole % H2O, 0.16 ± 0.09 mole % CO2, 0.11 ± 0.05 mole percent N2, 0.05 ± 0.04 mole % CH4, and 0.01 ± 0.01 mole % CnHn+i (n = 7). Sulfur species were not analyzed due to absorption within the chromatography system; however, a H2S concentration of 16 ppm to 160 ppm molar is estimated. The volatile content contributes approximately 50% of the microthermometrically measured freezing point depression. Cl- concentrations up to approximately 9000 ppm suggest an additional, probably magmatic, fluid component.Ascending fluids were channelled along approximately NW striking faults related to the same primary extension (17 Ma to 14 Ma) responsible for Nevada basin and range physiography. Boiling induced by CO2 effervescence led to precipitation of Au and Ag over a 150 m vertical interval. At the time, this interval was about 460 m below the water table (approximately 38 bars), but because of erosion, it is now within 90 m of the surface. A Ag:Au ratio of 25:1 has been calculated for the Dixie Zone."

Jan 1, 2000

By B. H. Kim

It is hypothesized that the standard approach adopted to characterize the ground and to determine the rock strength properties at depth are flawed and generally tend to underestimate the rock strength and possibly the stiffness, particularly when the rock mass is characterized as massive to moderately jointed or blocky to very blocky. It is therefore necessary to conclusively established that the hypothesis, that the rock strength in highly confined rocks is greater than generally anticipated, is actually valid and generally applicable, and then to develop new procedures to establish rock parameters that can be reliably used for the design of deep excavation.

May 1, 2009

By Charles Kocsis

Over the last few years a multi-faceted feasibility study has been evaluating the introduction of hydrogen fuel cell powered equipment into underground mines to replaced diesel engine powered equipment. The analysis has shown that under normal operating conditions fuel cells could at the very least be an environmental and health benefit by eliminating combustion engines and their unwanted by-products. However, they could also generate significant reductions in the amount of ventilation a mine needs to supply when compared to dieselized operations; this would further decrease a mine?s energy consumption, its associated greenhouse gas impact and also reduce a mine?s significant ventilation associated costs. Fuel cell powered equipment also has the advantage of lower noise and heat production compared to their equivalent diesel powered counterpart. The degree to which fuel cells can generate ventilation benefits in underground mines depends upon a number of operational parameters and mine specific qualifiers such as current ventilation control or management, mine depth, inherent dust conditions and minimum velocity requirements. However another major consideration in the introduction of fuel cells in underground mines will be the safety requirements associated with diluting and removing the potentially explosive hydrogen in the event of a gas leak. In outside surface applications, it is possible to use dispersion, buoyancy, time, and lack of ignition sources as mitigating factors in deriving a low risk of an explosion. However, in underground operations these factors change, dispersion and buoyancy are limited; time may not be available and in non-coal mines ignition sources are not typically controlled. When a hydrogen leak occurs, the availability of sufficient ventilation to dilute and effectively remove the gas will be critical, consequently ventilation may be the prime risk controlling factor. In this paper, the authors evaluate the potential benefits of replacing diesel engines with fuel cells in powered production equipment, discuss the mitigating qualifiers that could limit ventilation savings and evaluate solutions to retain and maintain an additional ventilation capacity in the event of an emergency situation such as a hydrogen leak from the fuel cell stack or its distribution system.

May 1, 2007

By F. Chiesa

The permanent mould process, which is a favoured process for producing premium quality aluminium casting, is rarely used to produce magnesium alloy parts. This may be attributed different factors such as: the reservation with which the automotive industry considers using magnesium in structural castings; also the relative hot shortness of most magnesium alloys when compared to aluminium A356/357 is another factor. A third reason stems from the fact that the "heat content" of magnesium is much less than that of aluminium, resulting in the difficulty, for a relatively slow process such as permanent mould (versus pressure die casting), to reach a mould dynamic equilibrium temperature sufficiently high to avoid premature freezing and misruns. The present work will address this problem by carrying out controlled cycling of a laboratory mould, poured with aluminium and magnesium alloys. The influence of insulation of the mould and artificial electric heating on the run-in mould temperature will be determined experimentally. A thermal model will be sought for these thermal devices that will allow to predict the thermal behaviour of moulds of different geometry before a single casting is poured. The model will be tested on an industrial casting in a workshop environment.

Jan 1, 2006