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By G. Macfarlane

"In 1985, a crushing plant, located in the feed preparation area of Kidd Creek's copper smelter, was modified. The plant, which originally had treated only copper-bearing materials for the smeller, at a rate of 10 tonnes per hour, was expanded to 40 tonnes per hour capacity and modified to treat silica ores to meet the smelter’s flux requirement and also gold ores from the company's Hoyle Pond property. In the storage area adjacent to the crushing plant a 300 tonne per day gold mill was constructed.This paper outlines the construction and first year of operation of the crushing plant and gold mill with emphasis placed on the capital and operating advantages to be gained by utilizing an existing metallurgical infrastructure.IntroductionThe Kidd Creek Mines Ltd. copper-zinc-silver orebody is located 24 km north of Timmins, in Northern Ontario . Mining of this massive sulphide de posit commenced in 1966 and, in the ensuing years, a concentrator and zinc plant were constructed at the metallurgical complex, approximately 17 km by rail from the mine site . At this same site, in 1981, a copper smelter and refinery were commissioned. The smelter which utilizes the Mitsubishi smelting process for the continuous production of blister copper from copper concentrate, has been comprehensively described in previous papers.At the time of the smelter construction, a small (10 tonnes per hour) crushing plant was installed in the smelter's feed preparation area. This plant was used to treat the copper-bearing revert material produced by the smelter, and, apart from periods when slags and matte were produced through furnace drainage at the end of each operating campaign, it was generally under-utilized. This fact, together with the requirement to devise a more cost effective method of providing silica flux for the smelting operation, was the basic rationale for modifying and expanding and crushing plan t. At the same time, development at Kidd's Hoyle Pond orebody had progressed to such a stage that a gold mill of 300 tonnes per day was warranted."

Jan 1, 1988

By R. B. Smith, D. C. Brady

"A review of the major renovation of the fifty-year old C-I headframe facilities at NERCQ's Con Mine, Yellowknife, N. W.T. is provided. The project was complete with only two 3-day shutdowns of the facility .Work consisted of four major phases: structural reinforcement of the adjoining mill building prior to dismantling; relocation of virtually all mine and mill utilities which transited the headframe collar en route from the adjacent powerhouse; dismantling of the entire headframe facility and portion of the mill building, including the removal of asbestos siding; and erection of a new headframe, idler rower, collarhouse and mill wall.This paper describes the work and methods used by NERCQ, J.S. Redpath Mining Consultants Limited, Acres International Limited, and others to bring the project to a successful completion. The history of the project is discussed from the planning phase through construction, including; design, procurement, and construction scheduling. The paper discusses the unique elements of the project, the problems encountered, and the methods used to achieve success.IntroductionThe Con Mine is a 800 tpd gold operation located in Yellowknife, Northwest Territories. Originally owned and operated by Cominco Limited , the Con Mine was sold in December 1986 to NERCO Minerals Company, and officially renamed NERCO Con Mine, Ltd. The Con Mine has been in operation for over 50 years. The first gold brick ever produced in the Northwest Territories was poured at the Con Mine in 1938. The original C-1 shaft was sunk in 1937 to a depth of 235 ft. In 1941 the shaft was deepened to a depth of 950 ft, and the original timber head frame was replaced with a steel structure. Following the discovery of the Campbell Shear in 1946, the C-l shaft was again deepened to its present depth of 2443 ft. Over the years, mine production gradually moved deeper and further from the C-l shaft."

Jan 1, 1991

By Ken Musgrave

"What Most Effective Management Does?It must satisfy the demands, of three groups of people: owners, employees and the public.OwnersOwners want profitable operations that will pay them a return on their investment, plus operations managed for long life. Many owners are also concerned with employee's well being and their company's public image.EmployeesJob security and a good income are probably still the most important demands, however, as time goes by, employees are also demanding more involvement in safety issues, working conditions, and even how work is done.The PublicThe public requires acceptable environmental standards, as well as enjoying the prosperity supplied by a successful operation.How are Mines Managed?Most mines are managed very traditionally. The basic credo is that the managers make the decisions, and the hourly, and junior salaried employees do what they are told. Clearly this is the extreme case, epitomized by the long service shift boss who might say ""It took twenty-years to become a shift boss, and no miner is going to tell me what to do"" .There are some exceptions, two notable ones being Campbell-Chibougamau, and T.G.S. (now Kidd Creek Mines) which, in very different circumstances, decided on a more participative style.Campbell-Chibougamau, having closed down due to unprofitable operations, reopened, using a team structure with working leaders, and a system for sharing gains against cost targets with employees. Although the system has some problems, and development miners are bonused conventionally, this has been a basically successful program.Texasgulf Sulphur, a company with a history of non-union operations, and a participative management style, made a corporate decision to operate their Kidd Creek Mine in the same way. The orebody was rich enough to allow them the luxury to do this, but they have never had a strike, so perhaps they have more than saved the cost of the benefits and salaries they needed to pay. They also reward employees with a stock distribution scheme based on service and earnings."

Jan 1, 1985

By W. L. Bancroft, Peter Price

"What is now known as the 'Waite section' of the property of Waite Amulet Mines, Limited, was originally the Waite-Ackerman-Montgomery mine. It includes a group of twenty-eight claims lying on either side of the Duprat-Dufresnoy township line, and about 8 miles north-northwest of Noranda, Quebec. ,The original discovery of ore on the property was made by Thomas Montgomery in Block Q, Duprat township, and was staked by him on behalf of himself, I. C. Waite, and C. H. Ackerman, in the summer of 1925. The property was optioned late in 1925 to N. A. Timmins who, after trenching and drilling the deposit, sold his 85 per cent interest to Noranda Mines, Limited. Early in 1927, a company known as Waite-Ackerman-Montgomery Mines, Limited, was formed and shaft sinking and development undertaken. Production commenced in August, 1928, ·and continued tJ May, 1930, when, owing to falling metal prices, shipment of direct smelting ore was stopped. Development of the mine by shaft sinking, drifting, and drilling continued until March, 1932, when owing to the prolonged low price of metals, the property was closed down. In September, 1933, an amalgamation with Amulet Mines, Limited, was effected, the new company being called Waite Amulet Mines, Limited. Early in 1937 operations were resumed, first by direct shipments to the smelter and later by aerial tramway to the Amulet mill. Operations have been continuous up to the present time, reaching an average of 720 tons a day milled for the year 1940. Production during the initial period (1928-1930) amounted to 75,811 tons, having an average grade of 7 .11 per cent copper, 4.47 per cent zinc, and 0.03 oz. gold and 0.65 oz. silver per ton. During the second period (1937 to end of 1946) , 1,060,524 tons, having an average grade of 4.53 per cent copper, 2.97 per cent zinc, and 0.03 oz. gold and 0.65 oz. silver per ton, were produced."

Jan 1, 1954

By Yves Gagnon, Nicole Gauthier, Daniel Kelly, Michel Rocheleau

"The Larder Lake-Cadillac Tectonic Zone in Rouyn-Beauchastel area (a 20 km long segment) contains distinct lithotectonic units bounded by artastomosing or sub-parallel shear zones. The gold deposits in the area are located within highly strained portions of E-W striking shear zones. The ore-zones are structurally controlled disseminated and vein-type lodes. They are spatially related either to megascopic z-shaped folds which refold E-W shears, or to areas of intersection between E-W shear zones and late NE-striking faults. Several ore bodies are enclosed within ultramafic rock units or at their contacts with adjacent units. All the lenses are elongated down dip and distributed, both loterally and vertically, ""en echelon"" . The ore zones tend to decrease in size and number away from the hinge zone of the megascopic z-shaped folds.Two main styles of mineralization are recognized at the Astoria deposit and at for most of the deposits in the Rouyn-Beauchastel segments a discontinuous ""Carbonate Ore"" and a more continuous, albitic sulphide-bearing ""Flow Ore"" (terminology used at the Kerr Addison mine).The typical ""Carbonate Ore"" zones are found within strongly arbonatized blocks bounded by conjugate late NE or NNW-striking fralls. The carbonatized blocks display features characteristic of llrittle to brittle-ductile deformation developed in shear zone system. Vein systems in the ""Carbonate Ore"" occur within shear fractures(D R)tension fractures (T), or within brecciated rocks adjacent to NS llrilcing faults. The ""Carbonate Ore"" zones are displaced by numer IU lole E-W, NNE, NW and N-S faults.The ""Flow Ore"" zones are found within predominantly ductile portions of E-W shear zones. The sulphide mineralization is finely disseminated and regularly distributed within mylonitic foliation or associated with veins and veinlets sub-parallel or oblique to the mylonitic foliation, in which case the mineralization is coarser grained. The gold mineralization is a late Archean event postdating the formation of the penetrative fabric, and synchronous with the end of the dextral oblique-slip motion."

Jan 1, 1990

By Russell Spry

THE cyanide plant of the Cariboo Gold Quartz Mining Company, Limited, at Wells, B.C:, in the Barkerville section of the Cariboo mining division, was erected during the latter part of 1932 and placed in operation in January, 1933. The initial capacity of the plant was about 50 tons daily. This was increased to 100 tons at the end of 1934 and stepped-up to 150 tons in September, 1935. It is expected that further plant extensions, for milling a total of 200 tons daily, will be required during the early part of 1936. Recent additions to the mill have been so planned as to permit ready expansion to 275 tons daily when underground development warrants the milling of this tonnage. It is the purpose of this paper to outline briefly the steps taken to bring the plant to its present capacity and to comment on possibilities for further plant expansion. The gold in Cariboo Gold Quartz ore occurs as free gold in the quartz and on pyrite surfaces. Some of it is very finely divided, specks l/10,000 inch having been observed microscopically, according to Dr. G. Hanson. In addition to pyrite, which comprises probably 15 to 20 per cent of the mill feed, galena and sphalerite are present, with small amounts of arseno-pyrite and chalcopyrite. Sulphobismuthites-chiviatite, galenobismutite, cosalite, etc.-occur more or less frequently in certain sections of the mine. Following tests on Cariboo ore by the Mines Branch, Department of Mines, at Ottawa, treatment of the ore by straigh t cyanidation was recommended and a 50-ton plant employing continuous counter-current decantation was erected. This plant was built as a pilot mill, the intention being to erect a larger mill after metallurgical treatment had proved successful over an extended period. This original installation, however, has been enlarged and retained as an operating unit, additional equipment being added to it as capacity increases have been required.

Jan 1, 1936

By R. R. Potter

"IntroductionAlthough iron formations were recognized in western New Brunswick about 1820, commercial production was not achieved until 1848. The Jacksonville deposits near Wood stock were developed first, and ore was treated at a blast furnace constructed 5 km to the southeast on the Saint John River. Several charcoal burners were also constructed. Pig iron was produced sporadically until 1884, when operations were finally suspended. There was little further interest in the area until 1953, when Stratmat Ltd. initiated an ambitious exploration program. At the same time, the company carried out research on an electric-reduction process to produce metallic iron and ferromanganese. The technical feasibility of this process (Udy process) was confirmed, however the high energy costs were prohibitive and work ceased in 1957. Some metallurgical research was carried out by Mandate Refining Ltd. about 1970 and more recently by the New Brunswick Research and Productivity Council.The Years Prior to 1848New Brunswick became a separate province in 1784, shortly after the influx of a large number of Loyalists from the United States. Man y of these people sett led along the Saint John River, including the area of excellent farmland near the present town of Woodstock.There were indication s that some of these earl y settlers recognized iron formations in the vicinity of Woodstock about 1820, however the nature of the deposits was not known until 836 when Dr. C.T. Jackson, a geologist commissioned by the Stat e of Maine, investigated the mineral potential of the area. Jackson (1836) described the ""ore"" as compact red hematite yielding 44 per cent pure metallic iron, and suggested that sizable tonnages may be present. It was about this time that an extensive limonite deposit was found in north-central Maine, near Mt. Katahdin. Other iron deposits were found in the Nictaux area of Nova Scotia, and a few years later at Londonderry."

Jan 1, 1983

By W. L. Single

"LET ME FIRST OF ALL thank you for sending this invitation to me to participate in this Annual Meeting of The Mining Society of Nova Scotia. As an engineer, although I don't get an opportunity to practice very much engineering these days, I am, of course, always pleased to be able to meet with my fellow engineers to exchange views, but I must admit that today I feel a little bit out of place in the midst of all the experts that are here today. Let me make one thing clear - I'm a generalist! Your program states that I am to make a keynote address. I have always thought that a keynote speech is one that says something that has not been said before - announces some new program or policy, or says something startling. I am sure you will understand when I tell you that keynote addresses are made by politicians and that civil servants speaking publicly are usually expected to give an account of what is, or what is likely, to happen as a result of policies that have been announced in keynote speeches by politicians. I am reminded of a statement by a distinguished parliamentarian who said, ""However exalted a civil servant may be, he is not above the elected officials."" So today I will confine myself to giving you an overview with particular emphasis on the development side of oil and gas. The theme of ""Oil and Gas"" is a topical one these days. I suppose it's because of all the copy that is being generated by the media on the U.S. energy crisis, the Mid-East situation, the fact that most of the world's proven oil reserves are in Arab hands and, of course, the accelerated exploration activity that is taking place in the North Sea, in the Arctic, and off the coast of Newfoundland and Nova Scotia."

Jan 1, 1973

By Terry J. Briggs, Ian J. Kelso

"Abstract- The Century mine in northwest Queensland,Australia, is a major sediment-hosted zinclead- silver orebody, located within the western Mount Isa Basin. Stratiform, carbonaceous shalehosted mineralization is developed within Proterozoic shale and siltstone sequences of the Lawn Hill Formation. CRA Exploration Pty Ltd (CRAE) discovered the Century deposit in 1990 and open pit mine development commenced in 1998 under the ownership of Pasminco Century Mine Limited (PCML). The deposit contains an identified mineral resource of 105 Mt @ 12.1% Zn + 1.7% Pb + 46 g/t Ag (Broadbent and Waltho, 1998) which is currently being mined at a rate of 5 Mt/y.In December 1998, an exothermic reaction was observed between explosives spilt onto shale drill cuttings from a blasthole. This reaction is referred to as ‘reactive ground’ and has the potential to cause premature detonation of explosives in a charged blasthole with catastrophic consequences. The incident led to a subsequent geological and geochemical assessment of the different rock types at the Century deposit and their potential reactivity with ammonium nitrate-based (AN) explosive products. This assessment was conducted in an attempt to characterize zones safe for blasting with uninhibited AN-based explosive products. A direct result of this study is a set of procedures outlining the use of different AN-based explosives in areas defined as potentially reactive or non-reactive for the different lithologies present. Black shales contained in both the waste and mineralized sequences contain the sulfide mineral pyrite (FeS2), which is thermodynamically unstable with AN contained in commonly used explosives. The potential reactivity of partially weathered pyrite can ultimately result in the decomposition of an AN-based explosive. It was determined that only pyritic black shales showed any potential to react with AN, irrespective of their position in the stratigraphic sequence. Other lithologies present at Century, including Proterozoic sandstone and Cambrian limestone, are classified as non-reactive with AN. This is a geological risk management situation present at the Century mine."

Jan 1, 2001

By A. Fyson, M. P. Smith, Margarete Kalin

"Drums containing 0.07 m3 of pyritic waste rock from a zinc/copper mine, with and without Natural Phosphate Rock (NPR) additions (40 kg/t waste rock) were set up outdoors. The effluent chemistry has been monitored over a period of 709 days. In the presence of NPR, the cumulative acidity discharged from 7 out of 8 drums was 80 % less than that discharged from drums without NPR. Mixing NPR throughout the drums was more effective in reducing acidity than placement of the NPR in a layer in the drum. NPR treatment was more effective for 'fresh' waste rock than for material which had weathered for more than 4 years. It is concluded that application of NPR to pyritic waste rock could greatly reduce acidity generated.IntroductionWaste rock dumps are significant sources of acid, generated by the oxidation of sulphides. Oxygen and water supply are oxidation rate-controlling factors, and oxygen supply is the reaction-limiting component. If secondary minerals can be formed on the surface of the rocks, pyrite oxidation and therefore acid generation may be reduced or inhibited. Neutralizing agents such as limestone do not prevent the oxidation of pyrite but treat the resultant acid mine drainage (AMD) or acid rock drainage (ARD).Meek et al. (1991) determined in pilot scale experiments that apatite mixed with tailings was more cost effective than liming or use of PVC liners to reduce oxygen and water penetration into the tailings. Phosphate is not acting as a neutralising agent but it creates a chemical condition on pyritic surfaces which prevents reactions of the oxidising agents. Precipitation of principally iron phosphates physically isolates the pyrite surfaces from further oxidation (Ziemkiewicz 1990; Evangelou 1994). The chemistry of the phosphate effects has been described by Spotts and Dollhopf (1990). Secondary mineral formation was evident in the leach pile at Gibraltar when leaching was impaired (Scott, 1991) indicating a link between oxidation and the mineral-formation process. This paper reports on effluent characteristics when waste rock is generating acid in the presence of NPR."

Jan 1, 1995

By Stephen Mcgibbon, Tim Twomey

"Abstract - The High-grade Zone (HGZ) was discovered 5000 ft (1520 m) below surface in the hangingwall of the existing orebody, 46 years after the Red Lake mine first went into production. The very rich ore of the HGZ is characterized by a remarkably abundant distribution of visible native gold. The location of this discovery challenged conventional wisdom at the time and its genesis has yet to be fully understood. Textural observations of ore within carbonate veins as well as crosscutting relationships suggest peak-metamorphic emplacement for at least some of the gold. Post-lamprophyre dike remobilization of gold is locally observed in the HGZ. Structure and dilatancy were the key elements for localizing ore at the Red Lake mine. Ore occurs where a fault trend intersected folded ultramafic volcanic rock that created a semi-permeable cap-rock to ore fluids ascending the "feeder" structure. This was enhanced by strong competency contrasts between the eastern ultramafic rocks, felsic volcanic rocks west of the HGZ, and the mafic volcanic host to ore - an exceptional environment for the development of protracted dilatant fluid pathways for gold deposition.Development in the HGZ began in February 2000 and reached commercial production on January 1, 2001 at a capital cost of US $53 million. December 2001 reserves are 1.9 million short tons (1.73 Mt) at an average cut and diluted grade of 2.05 ounces per ton gold (70.5 g/t) containing 3.8 million ounces of gold. The extremely rich ore presented a unique challenge in grade estimation for the Geology Department at the mine. The historic cutting factor was replaced by a statistical method as more data were collected from mining and mill reconciliation. The mine produced 503 000 ounces of gold in 2001. Exploration at the mine is targeting HGZ-type ore formation within basalts associated with the intersections of fluid pathways and folded ultramafic rocks, as well as other targets."

Jan 1, 2001

By D. Ibarra, J. Jacob, A. G. Guezennec, P. d’Hugues, A. Archane

The lack of adequate gas mass transfer is a rate limiting step in many bacterial leaching processes. Oxygen often becomes a limiting factor because of its low solubility compared to the high demand induced by sulfide oxidation. One way of increasing the solubility of oxygen in water or in media solution is by increasing the driving force, i.e. raising the oxygen partial pressure in the gas stream supplied to the leach pulp. The use of oxygen is a well-known practice in high-temperature bioleaching reactors (above 70°C) whereas air is usually preferred in medium and low-temperature operations in order to avoid too high dissolved oxygen concentrations which could impact negatively the bacterial activity. The purpose of this study was to investigate the use of oxygen-enriched gas in bioleaching reactors at 40°C in order to improve the gas transfer in the system when operating at high solid load (20% and more). Bioleaching experiments were performed on a sulfide-rich tailing waste (pyrite 60%, copper 0.8%, cobalt 0.06 %, gold 1 g/t) using the “BRGM-KCC” bacterial consortia. The reactor used for the tests was designed on the basis of a new bioleaching reactor concept developed by Air Liquide, Milton Roy Mixing and the BRGM which uses floating agitators to inject gases and to mix and to suspend solids in the bioleaching solution. Two types of tests were carried out: with air injection and with oxygen enriched gas mix injection. The aim of this work was to confirm the capacity of bacteria to grow and to dissolve pyrite in this type of bioreactors in oxygenrich atmosphere, and to compare sulfides dissolution rates during both types of bioleaching tests. The results obtained show that the use of oxygen enriched gas mix does not decrease bioleaching performances compared to air injection. High metal extraction yields were achieved (above 80%). No deleterious effect due to oxygen use was observed on the bacteria despite the high level of dissolved oxygen reached in oxygen tests (up to 14 mg.L-1).

Jan 1, 2016

By Haibin Wang, Zhiwei Huang, Zhixiang Cui, Wenzhao Cui, Zhi Wang

Fangyuan Two-Step copper smelting process was commercialized in October, 2015 and has been operating for 3 years and 5 months. As the unique continuous copper smelting process with simplified flow in the world, production parameters have been significantly improved after three years. In the first half of 2018, the feed rate was above 200t/h, availability reached 97.5%, operation cost was about 20% lower than the conventional three-step processes showing obvious economic advantages. The Two-Step process in Fangyuan is composed of 1 Submerged Lance Smelting furnace (SLS) and 2 Submerged Lance Converting& Refining furnaces (SLCR). Two SLCRs work alternatively to ensure continuous production. High grade and liquid matte from the SLS is introduced to the SLCR by sealed launder to complete converting and refining. The SLCR is an original and independent innovation of Fangyuan, for which Fangyuan completed lots of development and improvement. This paper will introduce the process and experience. INTRODUCTION China is the largest producer and consumer of refined copper. Most of the processes adopted in China are conventional "Three-Step" copper smelting processes, which are comprised of smelting, converting and refining. There exists some problems in the conventional processes, such as high energy consumption, large heat loss, low altitude emission of SO2, etc. To deal with these problems, Fangyuan developed the new copper smelting process with simplified flow, low cost, clean and high efficiency. The SLCR that combines converting and the refining process was also developed, which shortens the process flow from conventional three steps to two steps. The SLCR is of great significance in promoting the development of the copper industry. Development and industrialization of SLCR At present, copper concentrates are treated in three steps: smelting, converting and refining. Anodes are also produced by continuous converting processes such as the Mitsubishi process and the Double Flash process. Problems like matte transportation by ladle, capacity limitation, and low altitude pollution exist.

Jan 1, 2019

By Andrea R. Gerson, Roger St. C. Smart, Russell Schumann, Joan E. Thomas, Warwick Stewart, Stuart Miller

Sequential Net Acid Generation (NAG) accelerated oxidation testing using peroxide was carried out on single sulfide mineral samples. The calculation of the nett acid production potential (NAPP) assumes that all the sulfide is present as pyrite and that dissolution results in a 2:1 ratio of H+:SO42-. NAPP as kg H2SO4/t is calculated as 30.6 x wt.% S present. To test whether this calculation is valid a NAPP* value was defined in which instead of using the wt.% S present for the calculation the wt.% S dissolved is used. Stoichiometric analysis of the dissolution products of the sulfide minerals as they would occur at pH 7 indicates, that in all cases, the ratio of H+:SO42- of 2:1 is viable. The NAG to pH 7 and NAPP* closely coincided for pyrite, bornite, chalcocite and covellite. For the three Cu-containing minerals, the NAG to pH 4.5 value was 0 as is predicted on the basis of the speciation likely to be present at pH 4.5. Acid generation is due to hydroxylation of the dissolved Cu between pH 4.5 and 7.0. The NAG to pH 7 and NAPP* values for chalcopyrite, arsenopyrite and pyrrhotite were in reasonable agreement but were not for either sphalerite or galena. Zn is not completely hydroxylated at pH 7 as is assumed by the calculation. For galena, surface analysis identified polysulfide formation on of the leach residue polysulfide. The formation of polysulfide is an acid consuming reaction. Therefore under conditions of accelerated oxidation the NAPP value is a reasonable estimate of the possible acid production in all cases except for sphalerite and galena. However, under conditions of less intense oxidation, for instance in columns, waste rock and tailings dumps, the formation of polysulfides is likely to be more widespread thus reducing the correlation between NAPP and the maximum possible acid production.

Jan 1, 2004

By C. Barnett

"The mill at Noranda Mines Limited, Geco Division in Manitouwadge, Ontario went into production in 1957 treating 3,500 tons per day of copper, zinc, silver, lead ore. It was not until 1964, however, that the zinc concentrate drying plant was constructed. In 1970, milling capacity was increased from 4,000 to 5,000 tons per day. The zinc content of the ore is quite variable so that the tonnage of zinc concentrate dried varies from 200 to 650 tons per day.The drying plant was designed to dry 25 tons of zinc concentrate, dry basis, per hour from an initial moisture of 8.5% to a final moisture of 2%. Major components were selected so that capacity could be increased to 35 tons per hour with minor modifications.The rotary dryer is a Hardinge type ""XH-12"" with a diameter of 80"" and a length of 50 feet. The all welded construction shell is 1/2 inch thick mild steel plate with 3/4"" thick reinforcing bands under the tires arranged for operation at a slope of 1/2 inch to the foot. The 3/8"" x 12"" radial flights are welded to the shell. There are no flights in the last ten feet of the shell in order to reduce dusting. This unit is driven by a 50 HP motor through a Hamilton type CD250 double reduction helical gear speed reducer. The shell speed is 5.8 RPM.The standard Peabody horizontal air heater has a rated capacity of 10,000,000 B.T.U. 's/Hr. The combustion chamber is lined with high temperature refractory brick and is air cooled. The conical outlet section of the air heater is lined with insulating fire-brick. The burner is a type AT wide range mechanical atomizer fired by a type ""CR"" gas-electric ignitor.The Bunker ""C"" fuel is supplied by a Compower duplex pumping, heating and straining unit, including tank hotwell, Powertronic oil line heating system with a 550/24 volt transformer."

Jan 1, 1972

By R. H. Coloquhoun

"INTRODUCTIONDome Mines Limited is located on 3,149 acres in Tisdale, Whitney and Shaw Townships in the City of Timmins in Northern Ontario. The mine and mill facilities are two miles from the town of South Porcupine, now part of greater Timmins.In April of 1909, J.S. Wilson of Massey, Ontario headed a party of prospectors into the Porcupine region. Disembarking from the Temiskaming and Northern Ontario Railway at mileage 222, they established an initial camp on the shore of Frederick House Lake. Following the lake ""break up"" they canoed to their destination, Porcupine Lake, and established a base camp on the west side of the lake by May 23, 1909. At a point approximately two miles from Porcupine Lake the moss on an outcrop was rolled back and gold was discovered.In the spring of 1911, the first mill structure was in the process of being built and consisted of 40 stamp mills, five S'x22' pebble mills and the ancillary equipment for a 300-400 T.P.O. operation. This structure was destroyed along with much of the Porcupine Camp by the famous Porcupine Fire of July 1911. By mid-1912, a steel ""fire proof"" structure was built and utilizing the salvaged mills from the fire, supplemented by the installation of three, 8'x30"" cylindro-conical Hardinge ball mills purchased third hand from Vipond Mines, processed 75,088 tons by year's end.The mill was again levelled by fire in October of 1929, this time caused by an overheated boiler stack. This 11 fire proof"" structure was immediately replaced by a wood structure and again the three Hardinge conical mills and the five 5'x22' pebble mills were restored and utilized. The new mill operated close to its daily capacity of 1500 tons during the last two months of 1930. During the rebuilding phase, the crushing plant was remodelled and a new crusher installed. These plants were protected from fire by a 60,000 gallon water tank and an automatic sprinkler system."

Jan 1, 1984

Forward-Looking Statements Information set forth in this presentation may involve forward-looking statements under applicable securities laws. Forward looking statements are statements that relate to future, not past, events. In this context, forward-looking statements often address expected future business and financial performance, and often contain words such as "anticipate", "believe", "plan", "estimate", "expect", and "intend", statements that an action or event "may", "might", "could", "should", or "will" be taken or occur, or other similar expressions. Forward-looking statements or information relate to, among other things expected iron ore demand, timing for a production decision, expected share price drivers, potential railway options, timing, grade and quantity of potential resource estimates, employment and economic benefits, taxation estimates and timing of key project milestones. By their nature, forward-looking statements involve known and unknown risks, uncertainties and other factors which may cause our actual results, performance or achievements, or other future events, to be materially different from any future results, performance or achievements expressed or implied by such forward-looking statements. Such factors include, among others, the following risks: the risks associated with the need for additional financing; operational risks associated with exploration; fluctuations in commodity prices; title matters; environmental liability claims and insurance; reliance on key personnel and other risks and uncertainties discussed in the management discussion and analysis section of our interim and most recent annual financial statement or other reports and filings with the TSX Venture Exchange and applicable Canadian securities regulations. Forward-looking statements are made based on management's beliefs, estimates and opinions on the date that statements are made and the Company undertakes no obligation to update forward-looking statements if these beliefs, estimates and opinions or other circumstances should change, except as required by law. Investors are cautioned against attributing undue certainty to forward-looking statements. NI 43-101 Qualified Person Alderon?s exploration work on the Kami Project is supervised by Edward Lyons, P.Geo., the Chief Geologist for Alderon and a Qualified Person as defined byNational Instrument 43-101. Mr. Lyons has reviewed and is responsible for the technical information contained in this presentation. Mr. Lyons is a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC) and the Professional Engineers & Geoscientists Newfoundland & Labrador (PEGNL) as a registered Professional Geoscientist (P.Geo.).

May 1, 2011

By J. Holtzhausen

"An analysis of operating costs for underground mining operations in the current global economic climate considers the financial burden for a ventilation system to be in vicinity of 30–50%. These operating costs are often overlooked by management teams in the greater scale of things when looking to save money in the procurement phase. This oversight to operational costs can be as little as the effective delivery of ventilating air to the work zone. One of the biggest opportunities for substantial savings and improvements can be realized in the selection of a quality vent duct that is suitable and fit for purpose. The key component to the selection process for quality products is a solid selection criterion that is followed during the tender process. This allows the management team to filter out unsuitable suppliers and products. The case study identifies some of the key components that were considered in the procurement and selection process for lay-flat ducting products. These included the characteristics and strength of the materials. The selection criteria also considered a ducted performance comparison that offers the best aerodynamic performance with some of the following components. • Airflow at Duct Face (m³/s or cfm), Fan Pressure (Pa), Air Power (kW or hp), Assumed Fan Efficiency, Annual Running Costs, K – Factor or Lambda, Leakage Factor, Joining Method and available Duct Lengths with weights. • Critical components considered were the current Testing standards available for: o Bursting Strength AS 4878.5, Mass AS 2001.2.13, Tensile Strength AS 4878.6, Tensile Strength AS 2001.2.3.1, Tear Strength AS 2001.2.10, Tear Strength AS 4878.7 • Other key considerations were: o Warranty Claim, Damage claims, Local storage / operations, community engagement and relationships, current testing data, previous issues, after sales service with technical assistance, etc. Hidden costs associated with product failures and operational stoppages are often not quantified during the tender process. Buyer beware. INTRODUCTION The largest constraint underground mining operations currently face are their operating efficiencies and electrical costs. The overall cost of an operations ventilation system and its components requires more focus to identify what opportunities exist to maximise efficiency and effectiveness. This tender process may not always be clearly understood and it should be broken down into fewer components to achieve quantifiable results."

Jan 1, 2019

By E. Landry, W. Naudts

"The amenability theory was developed by Alex Naudts and considers grouting operations as continuous permeability tests where Newtonian test fluid water, is replaced by a cementitious or solution grout. It utilizes hydraulic conductivity and grouting data to establish the amenability coefficient of a particular grout zone. The amenability coefficient reflects the quantity of apertures or voids accessible to the selected grout, divided by the number of apertures accessible to water, in the zone intersected by the grout hole. By calculating, in real time, the effectiveness of a particular grout at the start of a grouting operation, the most appropriate grout to match design criteria can be targeted. By incorporating the amenability theory in every grouting operation, schedule savings and an increase in productivity are typically achieved. Three short case studies illustrate different enhancements in productivity: At an underground mine during the construction of declines, a cement based suspension grouting program was highly ineffective and did not appear to affect the inflow into the mine workings. The amenability theory was utilized to evaluate the grouting program and based on this evaluation, changes to the grouting program were made which allowed the grouting to meet the project’s objectives. On another project, a pre-excavation grouting program implemented during shaft sinking operations utilized the amenability theory to select the most suitable grout for the program. Schedule constraints also influenced the selection of the grout. In the third case history presented, the amenability theory was used to select the grouting sequence of every individual grout hole for a grouting project in a highly heterogeneous soil formation. INTRODUCTION The amenability of a given soil, civil structure or rock formation for a particular grout formulation is a measure of the ability of that grout to permeate the pores, fissures and apertures water will flow through. The amenability coefficient of a particular grout for a specific zone equals the percentage of apertures, pores and fissures accessible by water that are also accessible by grout. Grouting should be viewed as a continuous permeability test, where the apparent in-situ hydraulic conductivity value for a particular grout zone is continuously evaluated, using grout instead of water as test fluid."

Jan 1, 2019

By Michael Arriagada, Ivan Arriagada

"As mines return to shafts to economically exploit deeper reserves, two methods mine owners use to minimize short term costs are: sinking shafts in phases and refurbishing old, existing shafts. While there may appear to be benefits to utilizing existing infrastructure, often this is not the case, resulting in missed deadlines and added costs. Similarly, if proper long term planning is not executed when sinking a shaft in phases, additional costs and downtime will be accrued. This paper will take an in depth look at both of these options with an eye towards optimizing costs, minimizing downtime, and facilitating construction. INTRODUCTION With the onset of mechanized mining and large capacity trucks, many mines turned away from shafts and began to use ramps to haul ore to surface. However, continual exploitation of mineral reserves located closer to surface, combined with the subsequent unearthing of additional resources at greater depths, forced mines to go deeper and deeper to extend the life and keep the operation running. In shallow mines, hauling ore by truck was an economical option. However, as mines get deeper, driving ramps takes additional time, and hauling over longer and longer distances becomes less economic, both in terms of cost and time. For this reason Mexico, in particular, has seen a jump in the number of shafts which have been put into production, as well as having a number of shafts in different stages of development and still more in the conceptual stages. Many of the shaft projects in Mexico are being designed in phases, with the hoist and equipment purchased capable of reaching the ultimate planned depth of the shaft, but with sinking operations stopping short at an intermediate depth. A loading pocket is then installed to start exploiting mineral at this depth sooner, with subsequent phases planned for a later date using either a rock pentice or bulkhead to separate the sinking operations from the active shaft above. Currently, we are directly involved in two such projects which are actively being deepened, as well as having seen various others already reach completion. If done properly with the long-term plan in mind, sinking in phases can be a very viable option, allowing for faster production while fully utilizing the equipment at various depths. However, there are many issues which can arise when sinking a shaft in phases that if not properly examined and planned for from the outset, have the potential to eliminate the gains hoped for when it was decided to sink in this manner."

Jan 1, 2019