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By H. W. McGerrigle

The St. Lawrence Valley of Quebec lies in an earthquake danger zone, the main centers of which appear to be Murray Bay, Montreal and Quebec City. This paper compares conditions in this "earthquake zone" with those of other, currently more active, earthquake zones of the Earth, deals briefly with the mechanics of earth movements and outlines the means by which earthquakes are re-corded. An account is then given of the earthquakes which have occurred in the Province of Quebec -from that of 1534-35 to the earthquake of 1925. A 50-year frequency cycle is implied. The general conclusion is that the 11robability of future earthquakes should be given more consideration in the construction of buildings in the St. Lawrence Valley region.

Jan 1, 1966

By D. W. Davis, P. Landais, L. J. Robb

Nodules of carbonaceous matter have been observed in several Archaean peraluminous granites adjoining the gold- and uranium-bearing Witwatersrand Basin. These nodules, which resemble solid bitumen, usually replace uraninite or uranothorite and are paragenetically late. They show many similarities to kerogen seams and nodules associated with the conglomerate horizons in the Witwatersrand Basin, including composition, maturation, intimate association with uraninite, and a late paragenesis. U-Pb isotope ratios for granite-hosted and Witwatersrand kerogens define a single, albeit somewhat diffuse, Ph-loss discordia suggesting an upper intercept age of circa 2300 Ma. Carbon isotope analysis indicates lighter Ol3C values for the granite-hosted nodules than for the sediment-hosted kerogen seams. Raman spectroscopy indicates that granite-hosted kerogen is more highly ordered than Witwatersrand seam kerogen. A model is presented which links the granitoidand sediment-hosted organic matter to an event of oil generation in the Witwatersrand Basin circa 2300 Ma ago. Light hydrocarbons migrated through the sediments and into underlying/adjacent granitic rocks, undergoing polymerization and accumulation in response to radiation accompanying the presence of high-uranium phases.

Jan 1, 1994

By Brennan Lang

Why Seal Tunnels? ?Restore groundwater levels to pre-mining conditions. ?Saturate rock mass to reduce acid rock drainage. ?Reduce discharge flow to the environment. ?Reduce or regulate flows to a water treatment plant.

May 1, 2002

By Q. Shi, Q. Feng

"The influence mechanism of chlorite on the flotation of pyrrhotite was investigated through flotation tests, Zeta potential measurements, optical microscope photos, solution ion measurements and absorption measurements. It is observed that, the addition of chlorite influences the flotation of pyrrhotite seriously. The IEP (isoelectric point) of pyrrhotite and chlorite is 7.6 and 4.3 respectively, in the condition of slight alkalinity, both pyrrhotite and chlorite are negatively charged, and the hetero-aggregation will not occur. With the existence of oxygen in solution, the iron ions are dissolved from pyrrhotite surface because of the oxidation of pyrrhotite during pulp conditioning. The dissolved iron ions adsorb onto the surface of chlorite in the forms of iron hydroxyl and iron hydroxide, and the adsorption results in the opposite charge of chlorite surface. As a result, the hetero-aggregation between pyrrhotite and chlorite occurs. The adsorption of chlorite changes the surface property of pyrrhotite and restricts the adsorption of amyl xanthate onto pyrrhotite surface, and then decreases the flotability of pyrrhotite.INTRODUCTIONIn sulfide ores processing, magnesium silicate bearing gangue minerals often report to the concentrate causing downstream processing problems as well as increased smelting costs. Talc,chlorite and serpentine are gangue minerals commonly encountered in complex sulfide ores (Engel, Middlebrook, & Jameson, 1997; Pietrobon et al., 1997; Witney & Yan, 1997; Chen, Grano, Sobieraj, & Ralston, 1999a, 1999b; Bremmell, Fornasiero & Ralston, 2005; Senior & Thomas, 2005; Wang, Somasundaran, & Nagaraj, 2005; WieseHarris & Bradshaw, 2007). All of them are characterized with uniform crystal structures consisting of silicon-oxygen tetrahedron and magnesium-oxygen octahedron. The magnesium-oxygen octahedron layers are broken during the liberation of serpentine to expose the magnesium-oxygen bonds with strong ionicity and hydration, which results in natural hydrophilicity of serpentine. In water, the hydroxyl on serpentine surface ruptures and dissolves into the aqueous phase while the Mg2+ remains, resulting in surface positive charge (Tao, 2012). As serpentine has a positive potential at alkaline condition, it is therefore likely that it will attach through electrostatic attraction to the negatively charged mineral surface and reduce both the flotation rate and recovery of target minerals (Lu et al., 2011). For talc, characterized as good natural hydrophobicity, the liberation comes from the interlaminar fracture to remain molecular bonds on surface with weak hydration. Being naturally hydrophobic, talc easily reports to flotation concentrates, thus reducing concentrate grade (Shortridge, Harris, Bradshaw, & Koopal, 2000; Beattie, Huynh, Kaggwa, & Ralston, 2006). Meanwhile, both molecular bonds and magnesium-oxygen bonds are exposed on chlorite surface when liberation occurs, resulting in moderate natural hydrophobicity between serpentine and talc (Tao, 2012). And thus the mechanisms of influence of chlorite on sulfide ores are more difficult to study."

Jan 1, 2016

By J. R. MacDonald

MINING literature contains very few descriptions of operations producing as little as 25 to 30 tons of ore per day. The purpose of this paper is to discuss the gradual development of a mining method for such a small-scale operation at the No. 2 mine of Central Patricia Gold Mines, Limited, and in doing so to review several systems of stoping applicable to narrow ve1ns. MAIN ECONOMIC FACTORS TO BE CONSIDERED (1) Overhead.-The outstanding feature of most mines operating on a single narrow vein is chat of small daily tonnage. The reason for this low output can be readily understood. The strictly limited ore-reserves present a condition unfavourable for a policy of large initial expenditures, with the result that the capacity of the plant provided must be governed by the? amount of ore in sight at the time when mill construction is planned. This is immediately reflected in the cost per ton, there being relatively few tons over which to distributed the more or less static charges, such as hoisting, heating, supervision, administration, pre-production expenses, and other write-offs. (2) Development.-In mining this type of orebody, the main factor contributing to high costs is unquestionably the small tonnage of ore made available by a given amount of development. Thus, for example, from the data tabulated below it can be seen chat, assuming a 12-in. vein and a drifting cost of $9.00 per foot, the development charge for this item alone would be 71 cents per ton. Add to this the charges for raising, cross-cutting, shaft-sinking, and pumping, besides charges arising from development in waste, and the figure will be very considerably higher than 71 cents.

Jan 1, 1941

By A. B. Irwin

Abstract The techniques used in recognizing, recording, and interpreting structures in the Slocan sedimentary ceries in the West Kootenay district of B.C. are presented here in the hope that they will be of use to others working in complexly folded areas. The general pattern of folding in the Central Slocan area is obscured by lack of outcrop over a large part of the country, lack of marker horizons, local complexities of folding, and several systems of faulting. 1t was found necessary to observe and record in great detail all primary and secondary structural features and to analyse those observations on a statistical basis. The determination of bedding tops as well as bedding attitudes proved to be of fundamental importance in determining the pattern of folding. The interpretation of secondary linear and planar structures confirm:; this pattern. Accordingly, the thickness of the Slocan sedimentary series is computed to be several times that formerly postulated. Introduction The data upon which this paper is based were collected largely during the field seasons of 1947 and 1948 when the writer was employed by the British Columbia Department of Mines on detailed mapping in the Slocan mining camp. Additional observations were made by the writer during 1950 when engaged in property examinations for the Geological Survey of Canada. This presentation is largely a condensation of part of a more comprehensive thesis (5) and is restricted to observations and conclusions pertaining to the mapping of complex folds, stripped of any hypothetical discussion. Except for some unique occurrences of primary structures, there are no features or techniques described here which are not presented in standard textbooks on structural geology. It is felt that the major contribution of this paper is the method used in applying these standard techniques to a specific field problem and in interpreting the results.

Jan 1, 1951

By P. G. Keohane, J. S. Drake

"Cominco's Polaris Mine is a lead-zinc underground operation that began commercial production in early 1982. The mine site is located on Little Cornwallis Island in the Canadian Arctic at a latitude of 75°23'N and at a longitude of 96°55'W. The nearest settlement, Resolute Bay, some 100 kilometres southeasterly from the mine, is accessed from the mine by light aircraft. The mine site can be accessed by sea during the six-week ice-free period each year. During this time all major operating supplies are brought to the mine and the lead and zinc concentrate products are shipped out to market.GeologyThe Polaris orebody lies near the top of the Thumb Mountain Formation of Ordovician age. This formation is a relatively homogeneous section of thin to medium bedded limestones which are brecciated and dolomitized in the area around the orebody. The deposit is described as a Mississippi Valley-type orebody. Ore minerals are sphalerite and galena with the waste being predominantly dolomite with calcite and marcasite. The ore is quite porous with voids being generally filled with ice as the mineralized zone lies entirely within permafrost. Reserves consist of about twenty million tonnes of ore grading 3.8% Pb and 14.3% Zn with a further three million tonnes of inferred ore at a slightly lower grade.The orebody strikes in a north-south direction with the hangingwall dipping at about twenty degrees to the east. It is generally conformable with the stratigraphy. The footwall in the western portion of the orebody, the Panhandle, lies about 10 to 30 metres below the hangingwall. To the east the footwall drops sharply, creating an orebody height of up to 110 metres in what is referred to as the Keel Zone. The Panhandle Zone has typically 80 - 110 metres of cover above the hangingwall , About 25% of the reserves are in the Panhandle Zone, the remainder being in the Keel Zone. Figure I illustrates a plan view of the orebody relative to other surface facilities on the property and Figure 2 illustrates a simplified section through the Panhandle and Keel Zones."

Jan 1, 1985

By J. D. Botsford

Westmin Resources Limited's H-W orebody on Vancouver Island is a copper/lead/zinc massive sulphide deposit that has been in production since 1985. The sulphur content of the ore averages 28% and in some stopes is as high as 53%. Several minor secondary sulphide dust explosions have occurred in development headings and cut and fill stopes, but as yet no known explosions have occurred in blasthole stopes. This pa:per reviews the history of sulphide dust explosions in the H-W Mine, factors contributing to the explosions and the preventative measures taken to control them.

May 11, 1988

By Donald C. Craighead

"Overland conveyor systems to transport bulk materials from mines to processing plants or load-out stations have to be seriously considered as the viable alternative to truck and train haulage systems. With the ever-increasing price of diesel fuel, conveyors with their electrical drives generally prove to be the least costly when operating and maintenance expenses are analyzed for the life of the installation.IntroductionOverland conveyor systems are now considered as one of the viable alternatives which should be closely investigated when a mining company and its engineers initiate studies on methods of transporting raw materials from a mine to a processing plant or rail load-out station.This paper will generalize the overland conveyor and compare it with alternative handling systems. Each installation must be carefully studied to determine the best route, type of foundations river crossings; amount of excavation work for cut and fill, maintenance roads, etc. No two locations are the same nor are the principles of operation of an overland conveyor the same as those of a normal plant or mine belt conveyor. The proper time and effort must be put into the study of the system to evaluate the costs for comparison with other systems.The length of the overland conveyor depends on the layout of the mine and plant and the topography of the terrain between the terminals. Individual lengths may be up to 20 km, and a series of conveyors may extend 100 km or more.Advantages of Overland ConveyorsThe main advantage of overland conveyors is the low operating and maintenance costs which will be realized over its years of operation. Although truck haulage probably will permit lower capital investment during the initial phases of the mine development when the annual capacities have not developed to the design level, their operating and maintenance costs will be considerably higher. When evaluating the total costs per ton of material to be handled during the expected life of the operation and the projected increases for diesel fuel versus electricity, overland conveyor systems will generally prove to be economically more viable."

Jan 1, 1982

By R. D. Lindsay

Originally the focus of Deak Resources Corporation Business Plan was to establish a method for treating small gold and possibly base metal mineral deposits located in the Kirkland Lake - Noranda area of Northeastern Ontario, Northwestern Quebec at a central treatment facility. Since the establishment of this original business plan in 1988 other processing options have availed themselves for handling the base metal ores off site at a central facility which may preclude their processing at Deak.The site selected for the Deak processing plant was Kerr Addison's Virginiatown operation. At the time this property was acquired from the Receiver following Golden Shield Resources' Bankruptcy in 1989 the mill had been processing in the order of 1, 100 SDT per calendar day through one processing module; a module originally designed to process 1,500 SDT/day of a free milling gold ore. This tonnage level however had not been achieved since 1966 when an average throughput of 1481 SDT per calendar day was recorded.In order to realize Deak's overall Business Plan it would be necessary to expand the mills processing capabilities to provide:(1) A 2200 SOT per day circuit capable of processing a free milling, medium work index gold ore;(2) a 1 ,000 SDT per day flotation and/or cyanidation circuit capable of processing refractory gold bearing materials containing refractory minerals (i.e. copper) that should be removed prior to cyanidation; and possibly(3) a 1200 to 1500 SDT/day base metal circuit capable of processing the broad range of Deaks' Base Metal Property ores. If this circuit were introduced it would be completed in two tonnage phases.Further, it was recognized when operations resumed under Deak in 1990 that the existing tailings dam was suitable for, and had sufficient capacity for additional gold tailings with necessary physical modifications and changes in disposal techniques (including the construction of containment storage cells) that would allow storage for an additional 3.4 million tons of gold ore tailings approximately 1.0 million of which has been deposited since 1990. The focus of Deak's long term tailings disposal strategy is to acquire the Barber Lake Basin which is currently adjoined to our main tailings system which will provide an additional 10 million tons of storage capacity.

Jan 1, 1993

By David Minson

On many occasions, operators have seen that high density and paste thickener performance does not meet the initial design expectations, and the installed equipment produces lower than expected thickener underflow solids content. This lower-density tailings slurry leads to more water and lower beaching angles being seen at the tailings storage facility (TSF). This not only reduces storage capacity, but also increases the risks of non-compliance with environmental permit constraints. Reasons for the shortfall in performance can be many, but perhaps the main problem is inadequate understanding of the critical components of thickener design and operation, with too much reliance placed on standardised design terminology; such as tph/m2. This paper discusses additional factors that influence high-density thickener operation and presents test work that defines “Dynamic Slurry Fluidity”, “Bed Compaction Rates”, and “Natural Slump Angles” for these high yield stress/viscosity slurries. Finally, the paper uses this data to define thickener geometry, drive torque, floor slope(s), rake blade configuration and how this can be used to provide engineers with more tools for in depth evaluation of a High-Density thickener design. INTRODUCTION Generally copper mines are high tonnage operations, with copper flotation concentrators producing high volumes of tailings slurries of milled solids and water. As a result, Tailings Storage Facilities (TSF) become very large and this, along with potential scarcity of water, favours a design that maximises storage space. In theory, the use of filtration equipment that achieves immediate water recovery, high TSF in-situ densities and high stacking angles are preferred for this duty. However, large tailings filter units are uncommon, and have high operating and capital costs that favour the use of thickeners. The use of high-density thickeners to produce a thick or near-paste like slurry with high un- sheared yield stress values above 100 Pascals, is becoming more common. The higher thickener underflow is required to not only maximise water recirculation within the mill, but to also increase beaching angles, maximise tailings storage capacity and minimise evaporation losses arising from large areas of superficial water. However, the reliability of this equipment is also under scrutiny and there is perhaps a need to better define these designs.

Jan 1, 2019

By Shinji Shirakawa

The gold wire bonding technology is widely used for electrically connecting aluminum pads on IC devices to external terminal of semiconductor. In 1960?s, the tensile strength of gold bonding wire was about 100 MPa only. However, recently the wire with 300 MPa in strength has been developed. Since the gold bonding wire was started to use for IC package in early 1960?, the many types of gold bonding wire has been developing continuously in order to meet market?s needs, such as high density, small size and thin type of IC packages. Moreover, electronic equipments were used in automobile applications and also demanded good reliability under severe temperature environment. Thus Au-Pd alloy wire has been introduced instead of the wire with 4N(99.99wt%) purity gold to provide excellent reliability in Au-Al bond interface. The phenomena at Au-Al interface in the high temperature storage have been already explained by several studies as follows. There are five phases of inter metallic compound (IMC) between gold and aluminum. Just after bond gold ball on aluminum pad, Au5Al2 phase was formed and occupied almost all of IMC area. During high temperature storage test, Au5Al2 phase transformed to Au4Al phase, then this phase was corroded by Bromide (Br) ion that was isolated from molding resin. Thus the reliability between Au and Al degraded. In case of Au-Pd alloy wire, the concentrated palladium phase (Au-Pd-Al) was formed at the interface between Au alloy ball and Al pad. The Au-Pd-Al phase prevented from transforming between Au and Al, therefore Au5Al2 phase was hard to transform to Au4Al phase. Consequently the reliability at the interface between Au alloy ball and Al pad was less degradation. However, in recent study using several advanced analysis technique, Au4Al phase was found at the interface between Au alloy ball and Al pad same as the case of conventional gold wire (4N purity gold bonding wire). It was suggested that the formation of concentrated palladium phase (Au-Pd-Al) related the excellent reliability at ball bonding process.

Oct 1, 2003

By R. B. Humphrey

Cut-and-fill mining methods at the Thompson Mine have undergone significant changes since production began in 1961. Changes have been introduced in almost all aspects of the mining operation, from the original timbered chutes and air-operated slushers to the present-day bored raises, footwall ore passes and load-haul-dump units. This paper describes the evolution of mechanized cut-and fill mining methods, highlighting the changes in design for stopes and service openings, ground support methods; hydraulic filling and ventilation systems. A mechanized stoping area is described from the initial mining concept through design, development, construction and mining cycles.

Jan 1, 1979

By J. Ipinza, J. P. Ibáñez, Casas J., S. Cortés

Copper powder generated during the treatment of electro-metallurgical bleeds by reactive electro- dialysis (RED) had totally reduced the content of ferric ions in acidic synthetic electrolytes (Cu (II)- Fe (III) – H2SO4). In batch and columnar experiments 100% of the ferric ions were reduced to ferrous ions with an extremely high kinetics. A preliminary conceptual design for reducing ferric ions based on an integrate RED- Column for treatment of electro-winning bleed system is discussed. This system is able to recover simultaneously sulfuric acid and copper powder, which will reduce ferric ions. INTRODUCTION It is well known the negative effect of ferric ions in the copper electro-winning process. The increase of ferric content in the electrolyte reduces the current efficiency, i.e., there is an increase in the energy consumption, and changes the surface morphology of the copper deposit (Das, Gopala & Krishna, 1996; Vargas & Parra, 2013). The increase of the specific energy consumption due to the loss of current efficiency occurs by the reduction reaction of ferric to ferrous ions in the cathode, at the anode the ferrous ions are slowly oxidized to ferric ions (Shaw, Dreisinger, Lancaster, Richmond & Tomlinson, 2004). To control the level of iron in the electrolyte, bleeds are needed; this bleed electrolyte containing copper, sulfuric acid, cobalt, iron and other impurities normally is waster, which increased the production cost of the plant (Milbourne, 2002; Mckevitt & Dreisinger, 2009). Then, permanent efforts oriented to treat those bleeds to avoid losses of valuables components are made; among them are use ion exchange resins (Shaw et al., 2004; Mackevitt et al., 2009), chemical precipitation (Izadi, Mohebbi, Amiri & Izadi, 2017) and electrodialysis (Cortés, Ipinza & Ibáñez, 2013). The last one allows to the simultaneous recovery of sulfuric acid and copper powder. If one of the valuables obtained by treating the bleed can be used to reduce ferric ion concentration in the electrolyte could be an important asset. This issue at laboratory scale is reported in this work, the use of copper powder, generated by reactive electro-dialysis when treating synthetic acidic electro-metallurgical bleeds, for reducing ferric ions in acidic electrolytes.

Jan 1, 2019

By Theo. Koulomzine

"The property of Consolidated Central Cadillac Mines, Limited, consists of 1,103 acres formerly held by Central Cadillac Mines, Limited, and Wood Cadillac Mines, Limited. Both mines have been partly developed to depths of 565 and 1,000 feet respectively, and from 1939 to 1943 they produced gold, with small amounts of silver and tungsten, having a total value of $2,208,528 from 377,489 tons of ore. The average grade of the ore was $5.85 a ton . Less than 10 per cent of the ore hoisted was discarded by sorting. New development work, carried out since the former Wood property has been purchased by Central Cadillac, has shown that the tonnage indicated by mine workings and diamond drilling exceeds 2,000 tons per vertical foot.The property lies between the claims of Kewagama Gold Mines, Limited, and Pandora Cadillac Gold Mines, Limited, and straddles the Cadillac 'break' for a length of nearly 1 ½ miles. The main workings are 1 3/4 miles east of the well-known O'Brien gold mine.GEOLOGYOutcrops are very scarce, with none at all of the most important horizons, so that the general geology shown in Fig. 1 is deduced mostly from underground workings and diamond drilling. Fig. 2 is a vertical cross-section and Fig. 3 a longitudinal section through the ore zone.The ore lies within a belt of Blake River volcanics which is approximately 600 feet wide and is flanked on the north and south, respectively, by Cadillac and Kewagama sediments. All the formations have a nearly vertical attitude. The Blake River belt consists essentially of well sheared volcanic tulfs, but associated with these in the northern part are a number of bands of iron formation and the central part, to a width of 100 feet, is occupied by talc schists, making the Cadillac fault proper. Intruding the volcanics south of the Cadillac shear zone are dykes of two types- finegrained aplite or quartz albitite, and porphyry; both consist essentially of albite and quartz and are distinguishable only by the grain size. They have the general east-west trend of the formations they intrude."

Jan 1, 1954

By Stanley Davidson

"The ore deposit mined by Falconbridge Nickel Mines, Limited, is one of the numerous nickel-copper sulphide bodies marginal to the large oystershaped mass of differentiated norite in the Sudbury district, famous as the world's most important nickel producing field. The unique character of the ore deposits of the district has resulted in its intensive geological study over a period of more than 50 years. The results of these studies have given rise to a voluminous geological literature and to a great deal of controversy as to the genesis of the ores. This description of the structural geology at the Falconbridge mine is presented in the hope that it may form a small part of the chain of evidence leading to a reasonably satisfactory explanation of the ore genesis.The mine, named after Falconbridge township in which the deposit occurs, lies along the southeast margin of the norite mass, which, however, is h::re concealed beneath a deep mantle of glacial sands and gravels averaging over 100 feet in thickness. As a consequence, the deposit was not discovered by the early prospectors, and it was not until 1916-17 that E. J. Longyear, of Minneapolis, located the ore by drilling(ll). The discovery was foreshadowed in 1899 by the work of Thomas A. Edison, who found indications of the deposit by what is believed to have been a combination of magnetic and electrical geophysical surveys. He sank a prospect shaft over what has to date proved to be the largest and richest ore section in the mine. A wet quicksand layer at the rock surface proved impenetrable, however, and the 'Edison' shaft never reached bed-rock.The Falconbridge property was acquired by the Lindsley interests in 1928, and ore output, since its inception in 1929, has gradually risen from a modest beginning to the present 2,000 tons and more daily. Total production in this period is 7,500,000 tons of ore averaging 1.72 per cent nickel and 0.89 per cent copper, together with a small but valuable content of precious metals, principally of the platinum group."

Jan 1, 1954

By C. Larsen, R. A. Walker, A. Voth, C. Daoust

"The 3000 t/d mill at Hernia Gold Mines Inc. was commissioned in March, 1985. By Canadian standards the concentrator is highly automated, however, unlike the retrofit systems in many other plants, the Hernia process control system was installed during plant construction. This paper describes the Hernia control system beginning with system development (justification, architecture, start up, etc.), proceeding to current applications (crushing, grinding, leaching, etc.) and finally discusses some of the short term objectives for system improvement.INTRODUCTIONThe Hernia Gold Mines Inc. operation is located in the Marathon area of Northwestern Ontario. Controlled by Noranda Minerals Inc. (50.1 %), the cost of bringing the mine into production was nearly $300 million. A unique feature of the mine development was the extraordinarily short elapsed time of three years from Noranda's initial involvement to the start of milling operations. Additional background information on this project can be found in the trade literature (e.g. Tinter (1986). Scales (1987).)The Hernia concentrator is designed to process 3000 tpd of ore and commenced operation in March, 1985. Figure 1 provides a general plant flowsheet indicating the major circuits. Typical annual production statistics are included on the figure and these have been extrapolated from the available 1987 year-to-date data. Larsen and Tessier (1986) have previously described the operational aspects of the mill.In a metaphorical sense, the Hernia concentrator is like the traditional bride, ""wearing something old (e.g. the cyanide leach process), something new (e.g. a plant wide digital control system installed during construction), something borrowed (e.g. crushers from Gaspe) and something blue (e.g. the cladding on the surface buildings). While the prospect of fully undressing this bride is of considerable interest, here we choose to focus on something new. In particular, we choose to examine the process control system in some detail since Hernia is one of Canada's more highly autcmated gold mills. The control system is of interest because it has received suostantial credit for maximizing metal production and minimizing operating costs. With reference to the latter, it is worthy of note that Hem lo reports recent operating costs of less than $100 U.S./oz."

Jan 1, 1988

By S. Mamonov, V. ?shikhin

"The purpose of work was the significant increase of copper recovery from slag at a concentrating plant. Analysis of study of slag phase composition showed the perspective of a slow slag cooling method due to its recrystallization and transition of primary copper minerals to secondary.Copper content in base slag varied between 0.71 and 0.96%. The study of a degree of copper minerals release of the slowly cooled slag in comparison with a yielding slag showed that in a flotation size grade –0.071 mm of yielding slag the weight ratio of released copper minerals grains is 58.95% against 86.73% in samples of slowly cooled slag. The process flotation sheet consists of three main, one control and two cleaner flotation operations with aftergrinding of middling products. A reagent mode was differed from a factory one by decreasing the slurry pH from 11–12 to 8–9. This paper shows research results of the slurry alkalinity impact on technological indexes of beneficiation (flotation).Industrial testing of a resulted batch of slowly cooled slag at “SUMZ” concentrating plant showed the efficiency of such processing: copper recovery to a copper concentrate increased by 15–19%.INTRODUCTIONIntensive development of mining and metallurgy industries within a 250-year period in the Ural region led to forming of approximately 315 man-made mineral formations at the territory of the Sverdlovsk region. The region contain over 32.5 billion tons of barren rocks and out of balance ores, over 1.5 billion tons of tailings, over 115 million tons of slimes, and over 100 million tons of metallurgical slag (Demin et al., 2000). Latters, at average, contain 407 thousand tons of copper, 2519 thousand tons of zinc, 7 tons of gold, and 150 tons of silver (Rudoy et al., 2011). Wastes generated in mining processing and metallurgical sector of economy contain not only thousands tons of copper and zinc, but cause significant damage to the environment where they are located. The reduction of negative impact of anthropogenic wastes on the environment as well as the storage volume reduction of valuable metals from dump slags is possible by improving the efficiency of the melting and further flotation of slags formed during this process."

Jan 1, 2016

By S. Yadav, T. G. Murthy, C. Saldana

"An experimental study has been carried out in order to study the deformation of porous brittle materials under plane strain indentation. Gypsum, which is well known model system for understanding the behaviour of soft rocks, is used in this study. The effect of wedge angle is investigated in this research through this series of experiments, using wedges of 3 different apical angles. High-speed, high-resolution images are captured using a CMOS camera during the experiment. The images obtained are analyzed using a Particle Image Velocimetry (PIV) algorithm in order to obtain the evolving velocity, strain-rate and strain fields.INTRODUCTIONThe problem of indentation, which is usually referred to in the context of hardness testing is ubiquitous as a material characterization technique for a large class of solids. Mechanical properties of materials such as yield stress, flow stress, fracture toughness, elastic moduli etc. can be related to the hardness of a given solid. The indentation problem involves a given material or solid (whose properties are to be determined) indented by a relatively harder material (indenter) allowing the solid to undergo plastic deformation, which is then related to the yield stress of the material. The problem of indentation and its variants see many applications ranging from mining and petroleum industry, infrastructure industry, manufacturing, and biomedical sectors. Indentation has been extensively studied in metals, with accurate measurement of deformation parameters, stresses and comparison with analytical solutions. On the other hand, understanding of indentation in other class of solids such as cellular solids, porous solids, rocks, and soft solids is still nascent. In case of rocks, a comprehensive understanding of the indentation problem is very crucial because of the obvious applications in mining, petroleum and infrastructure sectors. The inherent porosity, the brittle nature of the rocks, adds to the complexity of indentation in these porous brittle solids. Various theoretical models with degrees of simplifying assumptions have been proposed to study the indentation in rocks. The most common model used in the study of indentation in rocks is a cavity expansion model, wherein a blunt indenter indenting into a rock is approximated with the expansion of cylindrical or spherical cavity in an elastic-plastic material. Complementary studies in understanding fracture initiation and propagation due to indentation has been largely based on the studies of traditional brittle materials such as glass."

Jan 1, 2015

By J. Chlan

"Dryer DescriptionProctor and Schwartz Continuous Conveyor DryerThe Proctor and Schwartz Continuous Conveyor Dryer has been in operation since July 1958, and comprises six unit• - an 8• wide single conveyor with roller extruder, insulated housing, type 304 stainless steel circulating and exhaust fall system, steam heating coils, temperature controls and discharge crusher rolls. Entire dryer dimensions are 52’ x 12'0""weighing 53,000 lbs,Extruder rolls - setting is 1/8"" (Drive 3 HP 1800 RPM) clearance off the extruder screen.Conveyor - consists of 151 interlocking pans, measuring 8"" x 8' (2 HP 1200 RPM) One Exhaust Fan (5HP - 900 RPM)Steam Coils -- l"" steel pipe - with steel fins -- each coil measures 16' 2’’- 64 coils in the No, 1 Compartment- 64 coils in the No, 2 CompartmentDryer OperationThe yellowcake slurry at 28% solids is pumped to an 81 x 8' Peterson Filter. The filter media presently being used is a Neotex 4372 blanket, being 100% staple fibre nylon, 12 oz per square yard with air porosity of 9 CFM's. The cake discharge at 45 - 50% moisture falls into the roller extruder. Extrusions are formed by forcing the material by rolls through a perforated plate moving back and forth and depositing the spaghetti-like extrusion on onto the dryer-conveyor.The conveyor now carries the extruded material through the heating chamber•, The feed end of the dryer will be referred to as No. 1 Compartment, with temperature control set at 280""F. The circulating fans take the hot air off the coils and circulate the heated air through the perforated openings in the dryer conveyor pane, After passing through the set of steam coils and circulating fan in No, 1 Compartment, the material then goes through No, 2 Compartment where the temperature is set at 305'F, Final material then discharges through a set of crusher rolls set at ~"". Moisture of final dried product runs at less than 2%.The dryer conveyor load is usually run to a 3"" - 4"" depth load. Conveyor speed at approximately 90 minutes for load to pass through. Load onto the conveyor is controlled by boot level and speed of the 8' x 8' filter. The dryer steam coils are water-washed twice a month."

Jan 1, 1972