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By C. Liu

Zeta potentials and settling rates of an oil sands tailings sample taken from Syncrude Canada Ltd's Aurora Stacker Pilot Operation were studied. Two clay minerals, kaolinite and illite, were also studied since these are the major constituent minerals in the fine oil sands tailings. Process water, decanted water as well as distilled water were used. It was observed that the zeta potentials of the tailings solids showed maximum negative values in process water at pH 8-10, which coincides with the pH of commercial hot water bitumen extraction process. The settling rates of the tailings sample were very low, at 0.008 m/hr in decanted water and 0.04 m/hr in process water. The rates were increased to 1.0 m/hr in decanted water and 0.13 m/hr in process water at an addition of 50 g/t Ciba AF1235. Although the settling rate was higher the supernatant was cloudier in decanted water than in process water. Further increase in the dosages of Ciba AF1235 did not increase the settling rates in decanted water but gradually increased it in process water, reaching 0.46 m/hr at 100 g/t. The results indicated that lower concentrations of Ca2+ and Mg2 (in decanted water) benefited settling rates but did not benefit supernatant clarity. Test results also indicated that Ciba AF1235 caused selective flocculation of the tailings sample, since the top sediments had higher illite and lower quartz contents than the bottom sediments.

Jan 1, 2006

By B. S. A. Tatone

The surface roughness of unfilled rock discontinuities has a major influence on the deformational and hydraulic behaviour of discontinuous rock masses. Although it is widely recognized that surface roughness is comprised of a large-scale (waviness) and small-scale (unevenness) components, most investigations of surface roughness have been restricted to small fracture surfaces (< 1m2). Hence, the influences of the large-scale components of roughness are often neglected. Furthermore, these investigations typically focus on analyzing roughness in terms of two-dimensional profiles rather than the complete three-dimensional geometry, which can lead to potentially biased estimates of roughness. This contribution demonstrates the use of a new optical digital measurement system, based on the principle of triangulation, to digitize a large-scale natural rock discontinuity surface at both outcrop-scale (~6m x 2m) and lab-scale (~100mm x 100mm). Subsequently, the digitized surfaces are systematically analyzed using the three-dimensional roughness methodology proposed by Grasselli to investigate the dependency of roughness on sample window size and measurement resolution.

May 1, 2009

1.Assessment of methane content in coal. Given a naturally low methane content in 11 Coal Seam and the fact that the areas of the 11 Seam adjacent to the tunnelway were degassing for a very, very long time due to previous development and mining, it is conservatively estimated that the remaining in -situ methane is less than 0.5m3/tonne. 2. Continuous miner cutting rate based on the past practice. The best loading rate recorded at the past operations of Smoky River Coal was 45 seconds for a 8 tonnes shuttle car. Using these records, the maximum peak loading rate potential at the face is calculated at 0.178 tonnes per one second or 10.67 tonnes per one minute. 3. Methane release at the face. ?The actual face methane liberation, conservatively assumed at 40% of the total in situ content: 0.20 m3/tonne ?The peak cutting rate: 0.178 t/sec ?The maximum (peak) face methane release: 0.20*0.178 = 0.036 m3/sec

May 1, 2003

By M. Labrie, H. L. Gibson, G. Doiron, J. J. Watkins, D. H. Watkinson

The Corbet deposit is one of 17 proximal Cu-Zn massive sulfide deposits located. within the Noranda Cauldron which occurs within the core of the Noranda Volcanic Complex. The deposit is located at the top of the Havrian Formation, the lowermost formation of the 3000 m thick Mine Sequence; the deposit has Yielded 2.75 M tonnes of ore grading 3% Cu, 0.96% Zn, 2.06 g/t Ag and 1.0 g/t Au. The multi-lens Corbet deposit formed within a 250 m wide crater that occurs within a localized breccia pile. The Corbet breccia pile was constructed by explosive magmatic and hydrovolcanic eruptions localized along the south flank of a larger andesitic shield volcano that formed during the waning stages of Havrian volcanism. The breccia footwall of the deposit allowed unfocused and widespread hydrothermal discharge that resulted in the development of a large area of chlorite and fringing sericite alteration that exceeds, by six times, the size of the sulfide lenses and represents a significant exploration target.

Jan 1, 1993

By Stephen Hardcastle

Presentation Outline ? Introduction ? Mine Ventilation Facts ? Ventilation-on-demand (VOD) ? Sensors/Controls ? Continuing Work ? Conclusions ? Mine ventilation engineering practices were published in 1556 in Agricola?s De Re Metallica (wind sails, bellows, etc.) That was then? This is today? Large fan installations handling >1M cfm

Nov 1, 2011

By A. C. Hitchins, C. N. Orssich

The Eagle zone is a gold-bearing, sheeted, quartz vein porphyry deposit hosted by the Cretaceous Dublin Gulch biotite granodiorite stock, located in central Yukon, 50 km north of Mayo. The deposit occurs at the western end of the stock and has been drill tested over a 1000 m by 500 m area, to a depth of 150m. Mineralization consists of 1 cm to 2 Cm wide, parallel, grey and white quartz veins with accessory gold, arsenopyrite, pyrite, pyrrhotite, chalcopyrite, native bismuth, bismuthinite, and scheelite. Vein densities range from two to fifteen veins per metre and average approximately five per metre. The majority of the veins have potassic and or phyllic alteration envelopes. Alteration of the granodiorite becomes pervasive as vein densities increase. Mineralization outside of the Eagle zone, but spatially related to the Dublin Gulch stock, includes: (i) scheelite bearing sheeted quartz vein zones within the east-central portion of the stock; (ii) a swarm of gold-bearing quartz-sulphide veins up to a metre wide in a zone along the northern contact of the stock and extending 7 km further west; (iii) the Mar scheelite skarn located on the south side of the stock; and (iv) cassiterite-bearing breccias located 2 km north of the stock. A geological potential resource of 64.5 Mt tonnes at a grade of 1.03 g/t Au has been estimated for the Eagle zone from widely spaced core and reverse circulation drill holes. Mineralization is open at depth, on strike to the west and beneath the hornfels to the north.

Jan 1, 1995

By Michael T. G. Brown

Syncrude Canada Ltd. has recently completed construction of its ambitious tar-sand mining and processing facilities at Mildred Lake, Alberta. The Syncrude site is approximately 48 km (30 miles) north of Fort McMurray near the west shore of the Athabasca River. The Syncrude operation involves the open-pit mining of approximately 270,000 tonnes (300,000 tons) of tar sand daily using draglines and a conveyor system. Bitumen is extracted from the tar sand, upgraded to oil-gas and transported to refineries in Edmonton and other locations for final processing . The Beaver Creek Dam forms part of the mining facilities for the Syncrude tar-sand mining operation. Water from Beaver Creek is diverted from the open-pit mine site south to the Athabasca River via Ruth Lake and the Poplar Creek Dam. General site layout is shown on Figure I . Design of the dam was completed by R. M. Hardy and Associates Ltd. of Edmonton, Alberta (now Hardy Associates (1978) Ltd) . Construction was undertaken during the summers of 1974 and 1975 under the direction of Canadian Bechtel Limited, managing contractors for Syncrude Canada Ltd. Total length of the dam is approximately 2.8 km (1.7 miles), with a maximum height in the order of 16 m (50 ft) The Beaver Creek Dam is located in a broad, flat, poorly drained valley. About 50 per cent of the site is covered with treeless peat deposits, which ranged in thickness from in excess of 5 m (15 ft) at the east end of the dam to less than 1 m (3 ft) at the central portion. The peat deposits overlie Pleistocene meltwater sediments on top of clay/clay shales of the Clearwater Formation and the tar sands of the McMurray Formation. The peat deposits and high groundwater levels at the site combine to make the excavation and construction of a conventional rolled clay cutoff impractical. For these reasons, the slurry cutoff trench technique was utilized to produce an efficient and effective seepage cutoff under various sections of the Beaver Creek Dam .

Jan 1, 1979

By S. Koski

The Cerro Verde copper-molybdenum porphyry mine, majority owned and operated by Freeport- McMoRan Copper & Gold, is located 30 km southwest of Arequipa Peru. The concentrator, which produces copper and molybdenum concentrates, started operating in late 2006 with a 108,000 tpd design capacity, processing primary sulfide ore (mainly chalcopyrite and molybdenite). Plant capacity was recently increased to 120,000 tpd. The comminution circuit consists of one 60? x 113" primary crusher, four MP-1000 secondary crushers in closed circuit with dry screens, four 2.4m x 1.6m High Pressure Grinding Rolls (HPGRs) in closed circuit with wet screens, and four 7.3m x 11m gearless ball mills in closed circuit with cyclones. The HPGR circuit is fed by the Secondary Crushing circuit product at a T80 of 35 millimeters, and the HPGR circuit product feeds the Ball Mill circuit at a T80 of 3000 microns. This paper describes operating experiences with HPGRs during these first four years highlighting improvements in operating practices and presentation of data regarding operational efficiencies, also discussed are the primary benefits provided by HPGRs such as production of micro fractured particles, reduced ore Bond Work Index, and reduced power consumption. Cerro Verde represents the first installation of HPGRs in both the copper industry and large scale hard rock mining.

Jan 1, 2012

By S. Koski, J. Vanderbeek, J. Enriquez

"The Cerro Verde copper-molybdenum porphyry mine, majority owned and operated by Freeport- McMoRan Copper & Gold, is located 30 km southwest of Arequipa Peru. The concentrator, which produces copper and molybdenum concentrates, started operating in late 2006 with a 108,000 tpd design capacity, processing primary sulfide ore (mainly chalcopyrite and molybdenite). Plant capacity was recently increased to 120,000 tpd. The comminution circuit consists of one 60” x 113"" primary crusher, four MP-1000 secondary crushers in closed circuit with dry screens, four 2.4m x 1.6m High Pressure Grinding Rolls (HPGRs) in closed circuit with wet screens, and four 7.3m x 11m gearless ball mills in closed circuit with cyclones. The HPGR circuit is fed by the Secondary Crushing circuit product at a T80 of 35 millimeters, and the HPGR circuit product feeds the Ball Mill circuit at a T80 of 3000 microns.This paper describes operating experiences with HPGRs during these first four years highlighting improvements in operating practices and presentation of data regarding operational efficiencies, also discussed are the primary benefits provided by HPGRs such as production of micro fractured particles, reduced ore Bond Work Index, and reduced power consumption. Cerro Verde represents the first installation of HPGRs in both the copper industry and large scale hard rock mining."

Jan 1, 2012

By Frank Otten, Jochen Greinacher

"In July 2017, Deilmann-Haniel signed a contract with IOO Slavkaliy to sink two 750 metre (m) deep freeze shafts for the Nezhinsky Potash mine in Belarus. This is the second shaft project where this technology is employed after two over 1,000 m deep shafts for a potash project in Saskatchewan, Canada, have been sunk by the application of the Herrenknecht Shaft Boring Roadheader (SBR) technology. Developed to sink shafts for green field projects in soft to medium hard rock, in frozen or unfrozen geology, the machine technology utilizes a roadheader-type cutterhead on a telescopic boom to mechanically excavate the rock, and a pneumatic mucking system to convey the muck from the face into the buckets. This leads to a manless face during the sinking cycle and thus a significant improvement in terms of workplace safety. For the Nezhinsky project, the SBR design was significantly modified in close cooperation between Deilmann-Haniel and Herrenknecht, based on the available experience and the special requirements of the project. Such modifications have been for example the adaptation of the Pneumatic Mucking System to cope with wet material as well as dry as originally planned or the shaft lining moving from shotcrete application to jump for and tubbing installation for which a step-change hands-off tubbing erector for the placement of segments was designed. Further modifications to optimize the machine technology for the requirements during shaft sinking were applied on the two machines for the Slavkaliy project as well as several revisions to match the special project requirements. The SBR is suitable for soft and medium hard rock up to 100 MPa UCS. The main design limits are maximum weight of the machine with 400 tonnes and length or height of 50 m due to the depth of the pre-sink. The Pneumatic Mucking System was changed in all components, sheaves and rope arrangement were redesigned and the concept of the hoist system was also re-engineered. The SBR in Belarus is configured for a daily sinking rate of three metres and all systems behind the cutting boom are not the bottle neck of the whole system. The works on site started in November 2017 including freeze hole drilling, construction works, site setup, installation of hoists winches, permanent headgears and other components required for sinking. The SBRs arrived on site in June/July 2018 and were preassembled on surface and then lowered into the 50 m pre-sink. SBR #1 started full face sinking in mid-December 2018, and SBR #2 in late January 2019. By the end of February 2019, the machines were at 150 m and 100 m depth respectively, and achieved up to 3.6 m/day concrete lined shaft. The SBR and the sinking system, the experiences, problems, technical solutions and performances of this challenging project will be presented in this paper."

Jan 1, 2019

By G. Goodlad

This paper describes some of the main aspects involved in implementing the European “Equipment and Protective Systems for Use in Potentially Explosive Atmospheres” (ATEX) Directive (No.94/9/EC) in the British mining industry by the year 2003. It provides a brief background to the Directive and considers its main objective of removing all barriers to trade in Europe. It outlines the need for the European Standards making bodies to write new editions of the ’explosion protection concept standards’ and for manufacturers to adopt new quality management procedures and develop new versions of equipment in conformity with the ATEX categories of “M1” or “M2.” Finally, it assesses the possible future relationship between the European enforcing authorities, manufacturers and users of explosion protected equipment.

Jan 1, 1999

By F. L. Casey

"Part 1 – Published Papers Baria, G., and Boshkov, S., ""Insights into Undercuttingin Block Caving"", Proc. 5th Rock Mech. Symp.,Mines Br. - Information Canada, 41-58, 1970.Barron, K., Coates, D. F., and Gyenge, M., ""ArtificialSupport of Rock Slopes'', Mines Branch ResearchReport R228, July, 1970.Barron, K., ""Criteria for Brittle Fracture Initiation inand Ultimate Failure of Rocks and Their Applicationto Fracture Zone Prediction"", Proc. 2ndCongr. Int. Soc. Rock Mech., Belgrade, 1970, 3-37.Bauer, A., and Calder, P. N., ""Projectile Penetration inRock"", Proc. 5th Rocle Mech. Symp., Mines Br. -Information Canada, 157-170, 1970.Bauer, A. M., ""Site Development and Rehabilitation ofPits and Quarries"", Canadian Mining Journal,Vol. 91, No. 7, 77, July, 1970."

Jan 1, 1971

By C. Cutts, K. Wheatley

The Dragon Lake Zone (uranium prospect) comprises a N160º trending vein that is approximately 110 m long, up to 5 m wide and from 1 to 40 m high. The zone is mostly within altered sandstone of the Fair Point Formation, the basal unit of the Paleoproterozoic Athabasca Group, but it also extends downward (up to 3.5 m) into basement rocks that occupy the northerly trending Maybelle River Shear Zone (MRSZ). This subvertical shear zone separates graphitic paragneiss and weakly deformed granitoid rocks to the east from non-graphitic paragneiss and foliated granitoid rocks to the west. The Dragon Lake Zone cuts across both the MRSZ and N060º-N070º-trending cross-faults that offset the shear zone but not the mineralized vein, which contains values up to 54.5% U over 50 cm core length. Strong chloritization occurs in graphitic paragneiss and sandstone adjacent to mineralization. Silica was depleted from the basement rocks and redeposited at the borders of the 5-15-m-wide alteration zone, as well as within overlying sandstone of the Manitou Falls Formation. There are minimal alteration effects in the overlying Lazenby Lake Formation. Geochemistry shows that the alteration zone in the Fair Point Formation is highly anomalous in metals, most notably in Ni, As, U, Pb, Co, and B. Clay within this zone is illitic and Fe-rich chlorite is also present. Brittle deformation is not as widespread or as well developed in the Fair Point Formation as it is in the overlying Manitou Falls Formation, which is extensively brecciated and contains some anomalous metal values.

Jan 1, 2013

By S. R. McCutcheon, D. R. Lentz

"Abstract - The 12.1 Mt Brunswick No. 6 Zn-Pb-Cu-Ag massive sulfide deposit (New Brunswick, Canada) lies between two subaqueous felsic volcanic formations near the base of the Middle Ordovician Tetagouche Group. The footwall comprises rhyodacitic pyroclastic rocks (Nepisiguit Falls Formation), whereas the hanging wall comprises a sequence of rhyolitic flows, breccias, and hyaloclastites (Flat Landing Brook Formation), with an age difference of less than 3 m.y. The Brunswick No. 6 deposit is a proximal autochthonous deposit with a well-zoned massive sulfide body and a basal Cu zone that, at depth, develops into a stockwork stringer sulfide (feeder) system. The massive sulfides are capped sharply by a layered magnetite-chert unit that extends regionally beyond the deposit. The deposit has a keel shape (sheath) formed by the F1F2 interference pattern. The heterogeneous ductile deformation and upper greenschist-grade regional metamorphism has transposed many of the epigenetic stockwork structures and alteration within the host sequence. Hydrothermal alteration is much more extensive, both vertically and laterally, in the footwall (a few 100 m) than in the hanging wall (<100 m). Anomalies farther up in the hanging wall sequence seem to be associated with independent alteration systems related to the rhyolite domes. The least-altered footwall units exhibit keratophyric alteration (albite or adularia) and weak Mg enrichment (chlorite). More intense micaceous (sericite and Mg-rich chlorite) alteration occurs around the footwall sequence. The transposed stockwork stringer sulfides are typically composed of Fe-rich chlorite (± sericite, ± silica) with pyrite, pyrrhotite, chalcopyrite, arsenopyrite, and sphalerite. The (Fe2O3T+MgO)/(Na2O+K2O) and base metal alteration indices are the best practical lithogeochemical vectoring tools at this deposit. Sommaire - Le gîte de 12.1 Mt de sulfures massifs de Zn-Pb-Cu-Ag Brunswick No.6 est localisé entre deux formations volcaniques felsiques sous-marines situées près de la base du Groupe de Tetagouche d’âge Ordovicien Moyen. Le mur consiste en roches pyroclastiques rhyodacitiques (Formation de Nepisiguit Falls) tandis que le toit consiste en une séquence de coulées, de brèches et d’hyaloclastites de composition rhyolitique (Formation de Flat Landing Brook); la différence d’âge entre les deux est de moins de 3 m.y. Le gisement Brunswick No. 6 est un gîte autochtone proximal comprenant un corps de sulfures massifs clairement zoné ainsi qu’une zone cuprifère basale qui passe progressivement à un stockwork de filonnets de sulfure en profondeur (feeder). Les sulfures massifs sont abruptement surmontés d’une unité laminaire de chert et de magnétite dont l’extension par delà le gisement est régionale. Le gisement affecte la forme d’une quille (fourreau) résultant d’un patron d’interférence F1F2. Plusieurs des structures épigénétiques du stockwork et des patrons d’altération de l’encaissant ont été transposés par la déformation ductile hétérogène ainsi que par le métamorphisme régional au faciès des schistes verts. L’altération hydrothermale est beaucoup plus importante, tant verticalement qu’horizontalement, dans le mur (quelques 100 m) que dans le toit (<100 m). Les anomalies situées plus haut dans le mur semblent être reliées à des systèmes d’altération indépendants associés aux dômes de rhyolite. Les unités les moins altérées du mur une altération kératophyrique (en albite ou en adulaire) ainsi qu’un faible enrichissement en Mg (chlorite). Une altération en mica (séricite et chlorite magnésienne) plus intense est présente autour de la séquence du mur. Les sulfures du stockwork de filonnets transposés consistent typiquement en chlorite ferrifère (± séricite, ± silice) accompagnée de pyrite, de pyrrhotite, de chalcopyrite, d’arsénopyrite et de sphalérite. L’indice d’altération (Fe2O3T+MgO)/(Na2O+K2O) et l’indice des métaux usuels sont les outils d’exploration les plus appropriés dans ce gisement."

Jan 1, 2006

By E. Ghali, W. Zhang, M. Robichaud, G. Houlachi

"The performances of six mesh and one plate dimensionally stable anodes have been investigated in the zinc electrolyte at 38oC and compared to conventional silver-lead alloy anodes. The electrochemical behaviour of these anodes was studied by cyclic voltammetry, galvanostatic, potentiodynamic, open circuit potential and electrochemical noise measurements. It was observed that at the end of 24 h of galvanostatic polarization at a current density of 50 mA/cm2 in the zinc electrolyte at 38oC, the mesh V had the lowest potential followed by mesh III, mesh VI, plate-V, mesh IV, mesh I, and mesh II electrode. When the potentials for mesh I, mesh III, mesh IV, VI and plate-V, mesh V were compared with that of the Pb-0.7%Ag anode (1920 mV/SHE), they were found to be 6%, 16%, 7% 10%, 17% lower, respectively.After 16 h decay following the 24 h polarization, the mesh III had the lowest corrosion rate (51 mS/cm2) among the seven specimens, followed by the plate-V, mesh V, VI, IV, I and II (221 mS/cm2). For the fresh surface as received, after two hour open circuit potential measurements, mesh I had the most positive (noble) potential (1382 mV/SHE) among the seven electrodes followed by the plate-V electrode, mesh V, mesh IV, mesh VI, mesh III and mesh II (751 mV/SHE). After consecutive series of electrochemical testing, mesh V, mesh III, mesh VI had lower corrosion rates (~ 343± 6 µA/cm2) than that of mesh IV, mesh I, plate-V, and mesh II (~ 644± 143 µA/cm2).INTRODUCTIONThe continued use of lead based anodes in the electrowinning process has been mainly due to their relatively low cost as compared to other materials (Msindo, Sibanda, & Potgieter, 2010). They have good electrical conductivity, stability during electrolysis, very good corrosion and wear resistance (Ivanov, Stefanov, Noncheva, Petrova, Dobrev, Mirkova, Vermeersch, & Damaerel, 2010). However, lead-based anodes give their high energy consumption and low corrosion resistance for electrowinning. To overcome these problems, Dimensionally Stable Anodes (DSAs) has been chosen (Msindo et al., 2010). Comparing to the traditional lead anodes, the DSA anodes have many advantages: (1) no contamination; (2) no loss of anode material, pure electrolyte; (3) lower gas bubble effect through special anode designs; (4) use of higher current densities (Msindo et al., 2010). The success of DSA anodes has been attributed to their chemical stability that allows them to maintain nearly constant dimensions during the life of the anode even when operating in environment with low pH values. Due to their stability, the consumption rates of DSA are as low as 1 mg/amp.y (Alfantazi & Moskalyk, 2003)."

Jan 1, 2016

By J. C. O'Rourke

"This paper presents an overview of Alberta Oil Sands Technology Research Authority&apos;s in situ oil sands recovery project at its underground test facility near Fort McMurray. It covers the rationale for horizontal wells drilled from tunnels below the oil sands, the tunnel layout, the drilling program, and the recovery process and facilities. INTRODUCTIONThe Alberta Oil Sands Technology and Research Authority (AOSTRA) has constructed an Underground Test Facility (UTF) in the Athabasca oil sands deposit in northern Alberta, approximately 50 km northwest of Fort McMurray (Fig. 1). The facility initially is to serve as an underground laboratory from which to field test new approaches to in situ recovery of the bitumen (oil) in the Athabasca deposit.The UTF employs mine access to get at the effective exploitation of an oil reservoir. It does so, however, under conditions considerably more difficult than those faced by most other applications of the technology. Because of the unconsolidated nature of the McMurray formation, the development of drilling chambers directly in the reservoir would be prohibitively expensive. Further, mobilization of the extremely viscous bitumen requires intensive steam stimulation with the attendant concerns of large, complex well completions and potentially hazardous operations Over-all, the oil sands presents mining, drilling and operations challenges which push the technology to a new state-of-the-art. The UTF as it now exists consists of:• a 20 000 ha lease with all-weather road access;• 925 m of 5 m by 4 m tunnels in the limestone 15 m below the oil sands at roof depth of 178 m below the surface (Fig. 2);• two 3 m diameter steel lined shafts to a depth of 213 m;• administration and sur face mine support facilities from which to operate the UTF (Figs. 3 and 4);• underground piping systems to transport drill fluids, steam, produced fluids and services to and from the future well heads underground (Fig. 5);• surface steam generation and production handling facilities (Figs . 3 and 4).• a drilling rig and auxiliary systems specifically designed to drill long horizontal wells into the oil sands from tunnels in the underlying limestone (Fig. 6);• an experienced staff to engineer and operate the facility (Fig. 7)"

Jan 1, 1989

By X. Cao, P. Wanjara, J. L. Fihey, X-T. Pham, F. Mirakhorli

Single-pass hybrid laser-arc welding of 10-mm thick cast plates of CA6NM, a low carbon martensitic stainless steel, was carried out in butt joint configuration using a continuous wave fiber laser at its maximum power of 5.2 kW over welding speeds ranging from 0.75 m/min to 1.2 m/min. The microstructures from the fusion zone to the heat affected zone and base metal were characterized after post-weld tempering at 600°C for 1 hour. From optical and scanning electron microscopy examination, the fusion zone was observed to mainly consist of tempered lath martensite and some residual delta-ferrite. The mechanical properties including microindentation hardness, ultimate tensile strength and Charpy impact energy were evaluated in the post-weld tempered condition and correlated to the microstructures and defects. The ultimate tensile strength and Charpy impact energy values of all the fully penetrated welds in the tempered condition were acceptable according to ASTM, ASME and industrial specifications. The failure mechanisms during tensile deformation and Charpy impact loading were examined by scanning electron microscopy, fractography and optical metallographic analysis near the fracture origin.

Jan 1, 2015

By Jason Clarke, André R. Laplante

"The gravity-recoverable gold (GRG) ore characterization test was developed 15 years ago and has been applied to hundreds of ore samples. The test accurately determines the size distribution of GRG, at which grind it is recoverable, and the extent of its liberation above 150 µm. This contribution shows how the test can be simplified, what are the likely problems to occur, and what information is then lost. Two simplified versions of the test are presented, of which the simplest, grinding a 20-kg sample to a P80 of 75 µm, is retained as the best one. A database of 18 comparative tests is presented. The choice of the full GRG test vs. the simple test is discussed, as is the deportment of the GRG recovered in the standard test but not in the simplest test.INTRODUCTIONThe GRG test was developed in 1990 to characterize the progressive liberation of gold in representative samples of ores, with the primary objective of generating data for a gold gravity recovery model (Woodcock and Laplante, 1993; Laplante et al., 2000). Characterization includes the size distribution of GRG as well as the grind size at which it becomes recoverable. This is achieved in a 3-stage protocol of liberation and recovery using a laboratory centrifuge unit, a Laboratory Knelson Concentrator 3 inches in diameter (KC MD3). This protocol is shown in Figure 1. The first stage takes place at a F100 of 850 ?m (F80 of ~550 ?m), the second at F80 of 180 ?m, and the third at a P80 of 75 ?m. The rationale for the three cycles is the desire to preserve the original size distribution of the gold grains in the ore and to obtain a recovery grind size curve. The feed is a representative sample of ore that can be extracted from ore samples (e.g. drill core reject) or stream products of the grinding circuit other than circulating loads or product having undergone hydraulic classification (e.g. rod mill discharge or SAG screen undersize). Ideally the samples should weigh between 60 and 100 kg, depending on feed grade and gold top size. It is beyond the scope of this paper to fully discuss measures of representativity in the ore sample, for example Gy’s minimum sample mass model (Gy, 1979). However, further work in this regard may be an opportunity. The large initial weight facilitates representativity and minimizes weight and sulphide recovery to the concentrate (since concentrate weight is relatively constant at ~100 g), which limits the recovery of gold in nonsulphide gangue or gold carriers (i.e. sulphides). Thus, the 100 g concentrate carries principally discrete gold species, such as electrum, and native gold. The concentrate and a 600-g representative sample of gangue of each stage are screened down to 20 ?m and each size class is fire assayed (the concentrates to extinction, the tails up to 30 g per size class). The five coarsest size fractions of stage-1 concentrate are individually upgraded using a hydrosizer and the concentrates examined with an optical microscope to assess the degree of liberation of the gold phase. Scanning electron microscope work has established that the GRG has an excellent degree of liberation below 150 ?m in stage 1 and generally a good degree of liberation in stages 2 and 3 (Guerney et al., 2003); as a result the size distribution of the GRG has been found to approach closely that of gold grains (Nesset et al. 2005)."

Jan 1, 2006

By Glenn S. Dobby, Glenn A. Kosick

"Column flotation circuit engineering has matured to a good level of understanding regarding testing, scale-up and design. Some of the factors important to the process of implementing columns are addressed in this paper, highlighting the following issues• operating bias and the maximum achievable solids carrying rate for a column application.• internal circulating loads of full scale and pilot plant flotation columns• short-circuiting oflarge scale columns and the impact on separation selectivity• the number of stages in a column circuit and the impact on selectivity.The paper uses several case studies, from pilot plant to full scale implementation, to illustrate these design issues.1. IntroductionTen years ago, in 1984, there were very few column flotation installations world wide and the largest columns were about I . 5 to 2 m diameter or square. During the following decade there have been in the heighbourhood of 80 column installations that have been successful and profitable, as both retrofits and green-fields plants. Now, columns as large as 3 to 4 m are not uncommon. Examples of installations with many large columns include Escondida, with ten columns each at 4 m square, and P T Freeport Indonesia, with 21 columns: nine 3 m diameter columns, four 2.4 m diameter columns and, starting up in April I 995, eight 3.6 m diameter columns.During this growth period there have also been column installations that can be considered as failures. In the majority of these cases, the reason for their failure has been an insufficient installed capacity for applications where there is a high weight recovery of solids to the froth product These include zinc cleaning, copper cleaning and iron ore flotation (silica flotation from a feed that is 15 to 30 %Si02) An application with high weight recovery to the froth product makes column scale-up and circuit design more complicated. This is particularly so if a high degree of selectivity is also required while attaining a high weight recovery, which can happen with iron ore flotation in columns."

Jan 1, 1995

By P. Ramirez

"To achieve an optimal metallurgical performance in the flotation process the bubble diameter should be between 0.5 and 2.5 mm. To control bubble size, frothers are added, reducing bubble size and conferring bubble stability against coalescence. However, the bubble generation mechanism determines predominantly the bubble size distribution produced. In mechanical flotation cells, bubbles are produced by the shear stress generated by a rotating blade on a gas stream. This mechanism produces a log normal bubble size distribution. In contrast, in column flotation, jetting gas injectors are commonly used in Chile and the bubble size distribution is bimodal with small and large bubbles. Bubbles larger than 2.5 mm mean a loss in surface area, while bubbles smaller than 0.5 mm move with the pulp stream and therefore collide less frequently with particles. In this paper, a new device based on a jetting sparger and a deflector disc (impinging jetting) is evaluated as an alternative sparger for column flotation.The new device was tested in a rectangular acrylic cell 1 m × 0.5 m × 3 m. High-speed cinematography was used to compare bubble generation with jetting and impinging jetting at different gas velocity in the nozzle (5–50 m/s). From the image analysis, it was observed that the deflector induces a shear stress, in analogy to a rotating blade but in this case the shear results from a high velocity gas hitting a static surface. The device was evaluated in a 0.5 m diameter 3 m high column. The McGill bubble viewer technique was used to compare the bubble size generated by jetting and impinging jetting. It was found that the Sauter mean bubble diameter reduced from 7.4 mm to 2.8 mm, when the jetting sparger was modified with the deflector. In the case of the impinging jetting sparger, 70% of the bubbles were in the effective range for flotation, while for the jetting sparger, only 10% were in the effective range."

Jan 1, 2016