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By C. Yanqui

"Discontinuum Mechanics is based on two foundations: the idealization of the internal structure of substances, and the principle of the arithmetic mean value. For a linear discontinuum, the methods of Crystallography are used to describe their structure, in which nodes stands for the position of the homologous points such as molecules, grains, or any other motif. The principle of the mean value states that the value of any discrete function at a node is equal to the arithmetic average of the values of this function at the neighboring nodes. This principle leads to an elliptical linear differential equation of the second order, with coefficients given by the statistical second moments of the Cartesian components of the metric vector. As long as those coefficients are components of a symmetric second-order tensor, there are three mutually perpendicular directions, in which the variances reach their extreme values and the covariances are zero, showing that the mean value principle transforms the discontinuum into a statistically equivalent orthotropic continuum.The application of Discontinuum Mechanics to linear rock masses is simple and direct. It is sufficient to define the motif as the plane of geological discontinuity, and the metric vector as the normal vector to this plane, whose magnitude is the spacing between two consecutive planes belonging to the same set of discontinuities. Furthermore, heterogeneities of rock masses can be also included in this description by substituting the plane of discontinuity by a zone of weakness. In this way, the magnitude of the normal vector results from the product of the spacing and the RQD or the RMR. In going further in this theory, at least three conclusions show up. First, in the evaluation of rock masses, it is necessary to determine the orientation, spacing, and RMR for each set of discontinuities. Second, the conventional method of finding the principal families from the contours of equal concentration of poles on the stereographic projection excluding the spacing does not have any physical meaning. Third, any problem of rock mechanics can be solved in a rigorous and reliable way by integrating the differential equation of the mean value."

Jan 1, 2015

By E. Onal

Mining operators are faced with complex information delivered by technology-centric rather than user-centric systems. The ability to achieve high situation awareness (SA) in the face of this data overload is a key challenge for effective decision making and information exploitation. Incidents that result in loss of revenue and even life, and compromise safety that are attributed to human error are often the result of system designs that overload human cognitive capabilities. This paper focuses on defining SA as it relates to mining operations and presents methods for improving SA through a scientific approach for electric shovel operator user interface development.

Aug 1, 2013

By Graham A. Davis, Michael Samis

The dramatic fluctuation in metal and energy prices in the last few years is causing mining companies to consider alternative analytical methods for mine design and valuation. The goal is to improve insight into unforeseen changes in the economic and project environment and to design projects that take advantage of management's ability to respond to these changes. Such flexibly managed projects will contain value that cannot be measured using traditional engineering economic approaches, further motivating the used of alternative, modern valuation methods. This paper considers the valuation of an uneconomic low-grade resource at an open pit gold mine using two of the most popular modern valuation methods, Dynamic Discounted Cash Flow and Real Options. Our analysis demonstrates that the conventional valuation method may incorrectly represent the value and risk attached to mining this type of resource because it is unable to recognize management's ability to respond to future events in a manner that minimizes losses and exploits advantageous business conditions. The results of this paper highlight the importance of recognizing flexibility and business uncertainty in both mine design and valuation.

Jan 1, 2003

By A. Ghahreman, M. Oliazadeh

"Heap leaching is an attractive process for low grade deposits where the metal content of ore does not justify the high cost of milling and agitated leaching. In general, the heap leaching capital cost is low, construction is rapid, and production costs are low. The primary disadvantage of heap leaching is low recoveries (typically 65% of agitated leach recovery); thus, recovery improvements could convert more lower-grade deposits into minable reserves. This will also allow to process primary sulfide ores and result in a dramatic expansion of the technology, as ~80% of total ore types are sulfidic in nature. The old gold operations tailings are another massive source of low-grade gold material that can be retreated by the heap leach process. The major parameters affecting the recovery rates are ore characteristics (mineralogy, particles size of metal and host minerals, ore permeability), and operational conditions (segregation of ore, compaction, leach cycle time, lift height, and solution application rates). The design of a percolation leach process is based on a limited number of coarse ore bottle roll and/or column tests under ideal conditions. The economics of the projects and engineering design of the process facilities are developed based on the bench scale test work results and the application of the conventional engineering scale up factors. This paper will deal with the parameters affecting the HL efficiencies and focus on the potential improvement of gold heap leach recoveries. This would include design of heap leach operation and evaluation of processes such as fine crushing and bioleaching.INTRODUCTION Gold heap leaching is a hydrometallurgical process designed to treat amenable low-grade gold ores that contain typically 0.5 to 1.5 g/t gold (Marsden and House, 2006; Wong Wai Leong and Mujumdar, 2010). Ore is stacked by various types of equipment on an impermeable barrier, and a sodium cyanide solution is applied to the surface of the stacked ore. Gold, silver, and other metals contained in the ore are dissolved to various degrees by the cyanide application, as it percolates downward through the stack of the ore. The pregnant leach solution (PLS) is then collected at the bottom of the heap. The PLS is then sent to the gold recovery plant where the metals are extracted. The barren leach solution (BLS) is then reapplied to the top of the pad. Ore is placed on a pad or previously leached lift by various means, including ramp conveyor belts, radial-arm and self-leveling stackers, and trucks. Once trucks are used to construct the pads, the pad can be compacted and hence the pad should be graded and ripped or scarified (roughened up) with wheeled or tracked equipment to produce shallow cuts to increase surface area exposure to the leachate."

Jan 1, 2019

By C. D. Anderson, C. G. Anderson, P. R. Taylor

"Rare Earths have become critical for our energy and materials driven society. While primary production must ramp up to fill growing demands, the recycle of these strategic elements must also be initiated. This paper will outline the potentials of and metallurgical technologies employed for Rare Earths recycle from various secondary sources.INTRODUCTIONWhen investigating the recovery of rare earths from secondary sources it is necessary to understand the nature of the sources to be processed. As of 2009, the USGS reports the distribution of rare earths by end use, in decreasing order, was: chemical catalysts 22%, metallurgical applications and alloys 21%, petroleum refining catalysts 14%, automotive catalytic converters 13%, glass polishing and ceramics 9%, rare earth phosphors for computer monitors, lighting, televisions 8%, permanent magnets 7%, electronics 3%, and others 3% (USGS, 2011). Additionally, it is reported that the amount of recycling was quite small and consisted primarily of rare earth magnet scrap, although with the sharp increase in the price of rare earths, and the potential shortage in supply, focus in this area is continually growing. Thus, the opportunities, whether realistic or not, are extremely vast in the recycling of rare earth containing materials. The following subsections will present methods proposed for the recycling of rare earths based on the feed material type.RECYCLING OF RARE EARTH MAGNETSCurrently, investigations are ongoing in the area of recycling production scrap and end-of-life materials in the rare earth magnet industry. The following techniques illustrate potential methods for the recovery of rare earths from secondary RE magnet sources.Recycling of Rare Earth Magnet Sludge by Selective ChlorinationDuring the production of neodymium magnets, a by-product of production is the so called “magnetic sludge”. This by-product contains high amounts of neodymium, thus a process for the recovery of Nd has been proposed. In this process the Nd present is selectively chlorinated/distilled by the addition FeCl2 via the following reactions."

Jan 1, 2012

By Scott Hommema

? Bitumen in the oil sand averages Bitumen 12 wt%; water makes up 4 wt% ? Imperial Oil (operator) and ExxonMobil developing Kearl Leases #6, 36, 87 (~200 sq. km) ? Target resource > 4 Billion Barrels ? 3 Phase development, ~100 kbd per phase High-quality resource ? Surface mineable bitumen ? Average properties (full lease) ? TV:BIP < 8 ? Grade > 12 wt% bitumen ? Ore thickness of 40 m ? Overburden thickness of 20 m Status ? Phase 1 ? Front end engineering and design underway ? Detailed engineering to start 2008

May 1, 2008

By R. C. E. Bray, Peter Price

"The property of Pamour Porcupine Mines, Limited, is in the eastern Porcupine area, about 12 miles northeast of Timmins. It covers parts of lots 3-6 (inclusive), cons. 5 and 6, Whitney tp., Cochrane district, Ontario. It is an amalgamation of the old La Palme, Three Nations, and Porcupine Grande properties.The first discovery on the Pamour was made in 1911. During 1911 to 1914 a small shaft was sunk on the Three Nations vein to the 200-foot level and a 100-ton stamp mill was put in operation, but only a small production was achieved. The property remained idle from 1914 to 1923. In 1924 some diamond drilling was done along the Keewatin-Timiskaming contact with some encouraging results, but the gold mineralization encountered was too low in general at the price of gold then prevailing to warrant further · work. The increase in the price of gold revived interest in the area, and the above noted amalgamation was effected in 1934. Options were taken and successively dropped by two large mining companies. Following this, exploration both on the surface and underground was started by Quebec Gold Mining Corporation, Limited. Diamond drilling in the early part of 1933 showed considerable ore in conglomerate and greywacke to the east. In the middle of 1935 Noranda Mines, Limited, assumed control and management. A new shaft was sunk and the property was ·equipped with a mill having an initial capacity of750 tons a day. This was subsequently enlarged to 1,500 tons a day.From the commencement of milling (May 16, 1936) to the end of 1946, 5,025,545 tons of ore having an average grade of 0.1327 oz. a ton ($4.64 at $35 an oz.) have been mined and milled."

Jan 1, 1954

By D. Van Tonder

ABSTRACT The comparative economics of ion exchange and direct precipitation have been analyzed by developing indicative CAPEX/OPEX models for typical alkaline leach uranium circuits. An agitated tank ore leach application is considered, comparing resin-in-pulp and direct precipitation as recovery options. Capital and operating cost estimates are derived from rates and factors used in recent studies and projects. The analysis showed the main economic drivers to be the ore grade and resin loading. The resin loading is impacted by the prevalence of competing ions, especially chloride. At resin-in-pulp solution chloride concentrations below 3 g/L and uranium grade below 640 g/t U3O8, resin-in-pulp is more economical, whereas at chloride concentrations above 3 g/L, direct precipitation is more economical.

Jan 1, 2012

By P. R. Geoffroy

Extensive and Intensive Prospecting As often pointed out, the search for new ore deposits may be carried on along two widely divergent lines. It may be either extensive or intensive. By extensive prospecting we mean a search for mineralized outcrops conducted over large and more or less virgin tracts of land, followed by a test of the practical value of any discovery made, the test being limited to the immediate vicinity of the outcrop. Extensive prospecting, was, in fact, the only policy known to the early prospectors. of the American continent and is still the rule in the remote sections of the Canadian Shield and in the African wilderness. It can be further characterized by the lack of emphasis placed on geological and other technical means as an aid to the search for ore. However, very little virgin ground is left, nowadays, for a man or a company to test their luck, and whatever is left is so remote from the necessary means of development that any new discovery will require considerable delay and expenditures to be amenable to profitable mining. On the other band, a great many of the favourable and still unexplored areas (and this is particularly true in Canada) are overburdened to such an extent that outcrops of bed-rock are extremely scarce and extensive prospecting is nearly impossible. Intensive prospecting, to the contrary, must be understood as a detailed search for ore carried on over limited areas in districts already known to be mineralized and covering regional structures whose relation to mineral deposits bas been estahlis.hed with a fair degree ot certainty. The search for deep-seated deposits in the vicinity of successful mines is a perfect example of intensive prospecting. As a matter of fact, the majority of the geologists employed by the mining industry are engaged in some kind of intensive prospecting.

Jan 1, 1950

By M. J. Nicol

The leaching of gold with thiosulfate solutions in the absence of ammonia and copper ions was investigated using electrochemical and kinetic methods. In sodium thiosulfate solutions, the electrochemical oxidation of gold was hindered and the leaching rate using pure oxygen as the oxidant was very slow. It was found that the addition of a small amount of thiourea or formamidine disulfide to the thiosulfate solution greatly enhanced the oxidation of gold. However, the enhancement to the overall dissolution of gold using dissolved oxygen as the oxidant was limited. This was attributed to the fact that the oxygen reduction process was unaffected by the additives and remained slow, which limited further improvement to the overall kinetics. An alternative oxidant, the ferric EDTA complex having a reduction potential of approximately 0.12 V, was found to be promising. The cathodic reduction of ferric EDTA was fast and reversible at neutral pH. While the ferric EDTA complex was effective in oxidizing gold, it did not appear to significantly oxidize thiosulfate. A thiosulfate solution containing thiourea and ferric EDTA has been shown to quickly dissolve gold in the form of a powder as well as from gold ores. In the leaching with the ores high gold recoveries of above 90% was achieved with the best performance obtained in neutral solutions (pH 6-7).

Jan 1, 2005

By R. O. Burt

"IntroductionThe introduction to this section on gravity concentration indicated the wide range of equipment types that exist for the concentration of minerals, in what is the broadest size range of any concentration process.The equipment described in this chapter includes jigs, tables and sHme plant equipment. In contrast to the equipment described in the two succeeding chapters, this range of equipment, in nearly every instance, superimposes a mechanical movement on the fluid motion to effect separation of the valuable mineral from the waste mineral. The motion may be vertical Gigs), reversing (tables), or rotary (slime plants): however, in common they ""dilate"" the bed of particles to allow preferential settlement of the heavier particles.Description of Process UnitsJiggingJigging is the sorting of different minerals by stratification, based on the movement of a bed of particles relative to a fluid in a vertical plane.The basic construction of a jig is shown in Figure 1. Essentially it consists of an open tank filled with a fluid, normally water, with a horizontal jig screen at the top and with a spigot in the ""hutch"", or bottom compartment, for removal of heavies. The jig beri consists of a layer of coarse heavy particles (ragging) on the jig screen. This ragging is of a specific gravity between that of the minerals it is intended to separate.In operation the bed is made fluid by a pulsating current of fluid to provide stratification. On the upward or pulsion stroke the bed of ragging and slurry is normally Hfted as a mass, then as velocity of the upcurrent decreases it dilates, loosening the whole mass. On the downward or suction stroke it closes slowly again. The ideal is to control this dilation of the bed of material so that the heavier and smaller particles penetrate the interstices of the bed, while the larger heavies fall under a condition akin to ""hindered"" settling."

Jan 1, 1989

By D. Sharma, A. K. Singh, R. Singh, P. K. Mandal, M. Sarkar

"This paper deals with the stability investigations carried out at the XIVth seam of Jitpur colliery. The study was conducted to investigate the possible causes of roof cavity formation and face spalling at the middle slice face of block-3 longwall panel and to suggest remedial measures. The instrumentation and monitoring work was carried out in the panel over a period of nine months. On the basis of observations, it was found that if the maximum face span was reduced from the existing 4.8 m to 3.6 m, and if the face time schedule cycle was reduced from a maximum of 37 days to about 18 days, this problem could be eliminated. The recommendations of CMRI have been implemented in the middle slice face of block-4 panel with success. IntroductionModerate to steeply dipping, undisturbed seams of uniform thickness are worked by way of inclined slicing parallel to walls. In this method, the seams are treated as a combination of several moderately thick seams dipping parallel to roof and floor. Numerous variations of multislice mining have been used worldwide to get high coal recovery under local conditions of geology, subsidence control, proness to spontaneous combustion, statutes and economics. Multislice mining can proceed either downward descending from the roof or upward ascending from the floor. Ascending methods require backfill behind the working face to provide a base for the next slice. It is also adopted for surface protection against subsidence or mining under a massive roof. This paper presents a scientific study conducted in Jitpur colliery, IISCO where a 9 m thick seam is mined in three slices in ascending order with stowing."

Jan 1, 2002

By Q. G. Reynolds, R. T. Jones

"Residence time distributions are well defined for batch and continuously stirred reactors. However, there exist many smelting furnaces that do not fit either of these classifications. A mathematical description is presented for the case of a furnace that is fed continuously, but where the liquid slag and metallic products are tapped intermittently. The effects of slag and metal tapping intervals and the fraction of liquid retained in the furnace, are shown to have a significant effect on the residence time distribution.INTRODUCTIONThere are some processes in which the time spent by the feed material in a smelting furnace plays a significant role in determining the extent of the reactions between metal and slag and influences the degree of settling of metal droplets from a molten slag.Residence time distributions are well defined for batch and continuously stirred reactors. However, there exist many smelting furnaces that do not fit either of these classifications. Here we will focus on producing a mathematical description for the case of a furnace that is fed continuously, but where the liquid slag and metallic products are tapped intermittently. The effects of slag and metal tapping intervals and the fraction of liquid retained in the furnace, are studied.Previous work (Jones, la Grange and Assis, 1997) has studied the effect of retention time on the degree of cobalt recovery achieved in a slag reduction process. In that work, a method was proposed for the calculation of the mean residence time of the contents of a pilot-plant furnace. That approach is to be extended here for the purposes of providing a more general design or scale-up tool.The maximum allowable depth of molten material in a given furnace is affected by such factors as the desired batch size (or the size of the ladles) and the pressure of the molten material that the tap-hole and surrounding system can withstand. Within these constraints, there might be some scope for choice of taphole position. The effects of this choice need to be understood mathematically."

Jan 1, 2012

By S. Parent, J. -F. Lepage, S. Glover

"Goldcorp – Eleonore’s 7 000 tonnes per day gold mill was started up in October 2014 and achieved commercial production in April 2015. During the first year, operation has faced challenges from grinding, flotation, sulfide cyanidation and tailing filtration to achieve targeted throughput, availability and recovery.The optimization of non-sulfide tails filter cloth lifetime and filter press availability had the most significant impact on ramp-up of both mill throughput and availability. A change in steel ball addition size to the grinding circuit was made to meet the targeted p80 of 65 µm. Various optimization efforts were conducted on the flotation circuit, intended to recover all sulfide-bearing minerals. The goal is to attain the environmental target of 0.12% sulfur in flotation tails (to eliminate acid mine drainage or AMD), while an average of 0.31% sulfur was obtained over the first months of 2015. Collectors, frothers, activators and hydrodynamic parameters were tested to improve the circuit desulfurization efficiency. High sulfide reactivity in pre-oxidation and cyanidation severely impacted dissolved oxygen levels and cyanide consumption. Studies were undertaken to understand the effects of mineralogy, grind size and leach conditions on gold dissolution kinetics and cyanide consumption. Efforts were focused on pre-oxidation to restore dissolved oxygen levels as well as optimizing both lead nitrate addition and pre-oxidation time.INTRODUCTIONGoldcorp Eleonore’s mill commissioning started in September 2014. It was considered to operate the mill at 3 500 tpd during design, as the mine was planning to operate at this throughput for the first year. However, with the stockpile of ore and a higher throughput from the mine, the mill was able to run at higher tonnage than initially anticipated.While increasing tonnage, a few bottlenecks were discovered and overcome in areas such as filtration and cyanide destruction. There were also recovery issues which were addressed in areas such as grind and sulfide leach. This paper focusses on the first complete year of operation ending September 30th 2015. A process description and the flowsheet are presented below and it will then go into detail about operational, metallurgical, mechanical and design issues experienced during the year."

Jan 1, 2016

By Marc Piard

In a cool mine, tapping high-pressure lines through reducing valves lowers the temperature beyond remedy. It is the writer&apos;s suggestion that, for low-pressure work, law-pressure air ,should be generated in a compressor installed underground adjacent to the work, and that the jacket-water of the compressor should be utilised for re-heating purposes. Such a system of compressed air distribution, designed especially for service in cool mines, is described in the following notes. The suggested system includes a three-wire pole line from the sub-station to the opening nearest to the working face, with a cable connection extending to the latter from the openings; an electrical centrifugal unit; a reservoir excavated in the ground; and a distribution line, re-heated by jacket-water. As regards the reservoir, it may be pointed out that in many mines there are blind or abandoned tunnels that could be used for the purpose. The centrifugal type of compressed air machine is a comparatively new departure, and it was formerly considered that such a machine, operating at less than 1,000 kilowatts per unit, would be commercially impracticable. At the present time, however, these machines are being built in Europe for as little as 400 kilowatts per unit, and as much as 110 lb. per sq. inch pressure. One advantage which these centrifugal machines have over other types is that they may be driven by plain induction motors, without the help of special starters-a consideration of the first importance where the electrical regulations are severe.

Jan 1, 1926

By P. L. Schwartz

The present method of shipping potash to the myriad of consumers is inefficient, not only because of the extreme and unpredictable nature of seasonal demand, but also because railway car turnaround times are very lengthy. The Floating Train system employs unit trains of potash operating from Saskatchewan mines to distribution warehouses located throughout the six corn-belt states of Minnesota, Iowa, Wisconsin , Illinois, Indiana and Ohio, transshipment points at Thunder Bay and the U.S. waterway network, and, where applicable , large consumers. These unit trains would "float", meaning that they would not originate at the same mine nor terminate at the same warehouse on each trip . They would " float " freely among the origins, destinations and participating railways. The potash would be shipped uniformly throughout the year and usually stored in warehouses located within the six states. It would be subsequently delivered to consumers within a 75-mile radius. Distribution warehouses and rolling stock would be provided by a new company, Potash Distribution Limited, consisting of major potash producers and the Canadian railways. Potash Distribution Limited would be responsible for the distribution of potash from the origin through the warehouses to the final consumer. The additional costs of providing warehouse space and outbound delivery will be offset by the reduction in both rail rates and "peaking" expenses.

Jan 1, 1978

By G. H. Montgomery

THE PROPERTY of Aunor Gold Mines, Limited, consists of eleven claims, located in Deloro township, approximately three miles southeast of the town of Timmins. Production began in 1940 and currently the mine is operating at a rate of 485 tons per day. GEOLOGY AND MINE DEVELOPMENT The ore zone at Aunor occurs in altered lavas bounded on either side by soft talc-chlorite schist, locally known as &apos;soapstone&apos;. The zone has a strike of N.80°E, dips 60° North, and varies in width from 0 to 250 feet. Within this zone the main orebody has a westerly plunge, away from the No. 1 shaft area (Figure 1). The mine is developed by a vertical three-compartment shaft from surface to the 3,082-foot horizon. Working levels are spaced at 125- foot intervals from the 625-foot to the 1,625-foot horizon, from which to tl1e 2,125-foot level the interval is 250 feet.

Jan 1, 1957

By A. R. Farkoosh, K. Korgiopoulos, McGill University, O. Rodop, L. A. Villegas Armenta

"New generation of Al-Mg-Si alloys with thermally stable dispersoids that prevent recrystallization and grain growth during hot deformation are currently used in automotive components requiring good formability. Alloy chemistry controls the phase formation especially age-hardening precipitates, dispersoids, and primary intermetallic which have pronounced effect on grain structure, strength and formability. In this work, thermodynamic calculations have been used to predict the effects of varying levels of Si (0.5-1.5wt%), Mg (0.5-1.5wt%) and Mn (0.1-2.5wt%) on phase selection. Thermodynamic calculations have been performed in equilibrium and non-equilibrium conditions via the FactSage™ software using the FTlite database. Selected compositions have been cast, heat treated, extruded and characterized through scanning electron microscopy with energy dispersive spectroscopy and electron channeling contrast imaging techniques. The calculations showed that at the hot deformation temperature of ~540 oC, exceeding certain Mg and Si levels would lead to retained *-Mg2Si intermetallics and Si particles at the grain boundaries in addition to a-Al-(Mn,Fe)-Si. It was found that retained *-Mg2Si phase which adversely affects formability can be avoided for Si+2Mg levels lower than 2.3wt%. FactSage calculations also showed that increasing Mn from 0.5wt% to 1wt% leads to an increase in the amount of dispersoids (from 1wt% to 2wt%) that can potentially slow down recrystallization and grain growth during hot deformation and improve formability. Calculations at room temperature also show that for an alloy with 0.7wt% Mg, excess Si beyond 0.7 wt% promotes the precipitation of Si clusters in the Al matrix that accelerate natural age-hardening. Results were confirmed by experimental investigation.INTRODUCTIONAluminum alloy 6082 (Al-Mg-Si) is used in the automotive industry because of its high strength, and good weldability (I. J. Polmear, 2006). The alloy is age-hardenable through the precipitation of ß-Mg2Si and its precursors, and the recrystallization is usually inhibited by Mn- and Cr-containing dispersoids formed through high temperature heat treatments (I. Polmear & Couper, 1988). However, it is known that minor changes in alloy chemistry (Mg/Si ratio) can have significant influences on the ductility and formability of these alloys by altering phase selection and the resultant microstructure (Gupta, Lloyd, & Court, 2001; McQueen & Celliers, 1997; Zajac, Hutchinson, Johansson, & Gullman, 1994). For automotive applications where wrought alloys (usually extrusions) are cold-formed to produce suspension components, the chemistry of the alloys and the extruded structure are of crucial importance as they determine the formability at room temperature (Asensio-Lozano, Suárez-Peña, & Vander Voort, 2014). In this study, thermodynamic calculations along with experimental investigations were carried out on the Al-Mg-Si system to understand their phase formation and precipitation hardening as a function of alloy composition."

Jan 1, 2016

By M. E. Wilson, A. C. Lee

"The property owned by Senator-Rouyn Mines, Limited, is in Rouyn township about 17'2 miles southwest of the town of Rouyn. It adjoins the Chadbourne claims of Noranda Mines, Limited, on the south and the Bagamac property on the ·west. It includes seven claims-block 44, blocks 80 to 83 and blocks 96 and 97; the mine is wholly in blocks 44 and 82.Late in 1936 gold was discovered in an outcrop of rusty weathering carbonate rock near the north end of block 44, and the following year the present Company was organized to proceed with the development of the property. During 1937-38, 17,000 feet of diamond drill core were bored to determine the extent of the ore deposit at depth. From October 15th, 1940, to April 15th, 1941, the ore was hauled to the mill at the Arntfield mine for treatment. The mill on the Senator-Rouyn property started on April 25th, 1941, and has been in practically continuous operation since than time. To the end of 1946, 574,874 tons of ore, from which gold having a value of $3,836,450 was recovered, were treated in the mill.GEOLOGYThe rocks exposed at the Senator-Rouyn mine, except in the southern part of the property, are almost wholly quartz diorite, part of an irregular mass that extends west from the town of Rouyn along the north shore of Noranda lake and thence southwest across the Senator-Rouyn claims. In the northern part of the propertyt inclusions of rhyolite occur here and there in the diorite. In the southern part the predominant rock is dacite and in the east-central part of block 44 claim andesite. Dykes of andesite, quartz syenite, and later diabase also occur on the property.The andesite of the Abitibi volcanics is not exposed on the property, but has been intersected in three diamond drill holes, one on the eastern boundary of block 44 and two farther west within the block. This andesite is probably the western end of a band of pillowed andesite well exposed in the Larose claim to the east."

Jan 1, 1954

By Lawrence Devon Smith

The selection of the production rate is one of the most crucial decisions to be made in the development of a mineral property. However, it is not uncommon for the production rate to be decided arbitrarily, before any technical evaluations begin, and it is often not reconsidered when better data becomes available. A number of methods are used in the mineral industry to select the production rate for a project. There is general agreement that the size of the ore reserve has a strong influence on this decision, but what other factors are involved? Is there a way to identify an optimum rate? This paper examines the factors that influence the selection of a production rate from the perspective of project economics. The NPV maximum value should be viewed as a “failure” point. The “two times capital” maximum is proposed as a lower limit. A practical check is suggested to confirm that a selected rate is in an optimum range. Production rate assumptions are much more likely to be reviewed for marginal projects, but they should also be reviewed for viable projects.

Jan 1, 1997