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By A. Bauer

The correct application of large crater blasts in unfrozen tar sand is an effective means of resisting frost penetration during winter. This results in much higher productivity and considerably less down time with excavators during winter operations. The resulting flatter angle of repose of the highwall has readily been accommodated with slight modification to bucket-wheel excavator operating procedure. Other mining schemes call for large draglines to be located on top of the highwall, which can be as high as J80 feet. The question now arises as to the effect that blasting might have on the stability of such a situation in which the highwall includes potential planes of weakness. This paper applies blast vibration data from large-scale blasting in tar sand to the stability analysis for typical production situations likely to be encountered and indicates safe scaled distances to be employed between dragline and blast.

Jan 1, 1979

By A. N. Kumar

Two fabrication stages in the manufacture of pipelines are identified as causing significant deterioration of the steel plates, namely cold bending and welding. The objectives of this work include quantitative assessment of the degradation of the quality of the steel plate by the loss of fracture toughness, and an estimation of the safe working pressure. The study was conducted on three API grade steels: X42, X46 and X60. After cold bending by the UO technique, various welding technique were employed for joining of the bend plates along the pipe length. A crack-tip opening displacement parameter (CTOD) was used at the onset of initiation and at the first load maxima as the criterion for assessing the toughness degradation under various fabrication stages and regime (such as WZ, HAZ and parent body). In addition, strength, hardness and impact energy values were determined to check the degree of degradation. Microstructural and microfractographic investigation of the fractured samples were also carried out in the scanning electron microscope (SEM). X60 grade steel plate suffers from CTOD toughness loss by around 38 and 55% after welding and cold bending, respectively, for the conditions studied. A combined effect of the two stages results in a loss in toughness by around 70% for the same steel. In X42 grade, the loss in impact energy ranges from 78 to 35 J due to electrical-resistance welding study (ERW). Crack resistance curves are also obtained for the steels to study the crack growth resistance behaviour and to supplement the information on crack initiation behaviour. The orientation of the pearlite banding, as observed on the fracture plane with respect to notch front and applied loading direction appears to have a strong influence on the microvoid-coalescence mechanism of fracture as well as the toughness values. An integrated analysis of the performance of pipelines, considering all these key factors, is discussed.

Jan 1, 2005

By AN. Kumar

Two fabrication stages in the manufacture of pipelines are identified as causing significant deterioration of the steel plates, namely cold bending and welding. The objectives of this work include quantitative assessment of the degradation of the quality of the steel plate by the loss of fracture toughness, and an estimation of the safe working pressure. The study was conducted on three API grade steels: X42, X46 and X60. After cold bending by the UO technique, various welding technique were employed for joining of the bend plates along the pipe length. A crack-tip opening displacement parameter (CTOD) was used at the onset of initiation and at the first load maxima as the criterion for assessing the toughness degradation under various fabrication stages and regime (such as WZ, HAZ and parent body). In addition, strength, hardness and impact energy values were determined to check the degree of degradation. Microstructural and microfractographic investigation of the fractured samples were also carried out in the scanning electron microscope (SEM). X60 grade steel plate suffers from CTOD toughness loss by around 38 and 55% after welding and cold bending, respectively, for the conditions studied. A combined effect of the two stages results in a loss in toughness by around 70% for the same steel. In X42 grade, the loss in impact energy ranges from 78 to 35 J due to electrical-resistance welding study (ERW). Crack resistance curves are also obtained for the steels to study the crack growth resistance behaviour and to supplement the information on crack initiation behaviour. The orientation of the pearlite banding, as observed on the fracture plane with respect to notch front and applied loading direction appears to have a strong influence on the microvoid-coalescence mechanism of fracture as well as the toughness values. An integrated analysis of the performance of pipelines, considering all these key factors, is discussed.

Jan 1, 2005

By T. Maruyama

Tamano Refinery started its operation in 1972 with the cathode production capacity of 84,000 tons per year. After several expansions, the production capacity increased to 228,000 tons per year in 2001. In this paper, we are introducing our next expansion projects starting October 2006. The project expands the production capacity of cathode from 228,000 tons to 260,000 tons per year. The main modification of this project is the conversion to the ISA process and the installation of additional seventy-two cells in the No.2 Tankhouse. The No.1 Tankhouse remains the same, using the conventional process. Half of our starting sheet cells are converted to commercial cells. The other expansions are to enlarge our purification plant. And the incidental facilities are enlarged, for example, circulation tanks, heat exchangers, filtration system, etc. Finally, this paper describes those improvements to our refinery performance and preparing for future expansions.

Jan 1, 2007

By Ken Major, Anita Marks, Craig Sams

In 2007 Copper Mountain Mining determined that there were sufficient resources to consider reopening the Similco Mine near Princeton, BC. The mill facility that had been built by Newmont Mining in 1972 was closed in 1996 and all of the process equipment removed. The mill had been built on the west side of the Similkameen River to facilitate mining of the Ingerbelle Deposit. In 1980 construction was completed on a primary crusher and cable belt to deliver ore across the Similkameen River from the Copper Mountain deposit to the Ingerbelle mill. With the decision to restart mining of the Copper Mountain deposit, the decision was made to design and build a 35,000 tpd grinding and flotation mill near the mining operations. This paper discusses the application of historical operating data for the selection of the mills for the new grinding circuit and provides an overview of the grinding circuit design.

Jan 1, 2012

The Millennium deposit, located in northern Saskatchewan between the McArthur River mine and the Key Lake mill, has considerable potential for becoming an operating mine. Located in the Athabasca Basin, Millennium is blessed with both the riches and the challenges that this basin offers. The potential for ground water inflows during shaft sinking, and the unconventional challenges associated with underground uranium mining, render the successful development of a new mine in this area especially rewarding. The preliminary design and engineering of the Millennium Project has tapped into the experience of other operations in the area, and has utilized new applications of geophysics technology to enhance the understanding of the local geological structure. This discussion will focus on the innovations, and the challenges specific to uranium mining, for a deposit that is poised to be one of the next mines to supplement Saskatchewan's uranium resources.

May 1, 2008

By Peter R. Drummond

THE paper consists of four sections: (1) a discussion of early work; (2) presentation of the results of a chemical investigation of Falcon-bridge nickel-copper blast-furnace matte; (3) application of the results of this investigation to mattes in general; and ( 4) a summary of conclusions. All general discussion will be confined to the fundamental system of ferrous-cuprous sulphides (FeS-Cu2S). Nickel-, lead-, and zinc-bearing mattes will be treated as minor varieties of the parent system. DISCUSSION OF EARLY INVESTIGATIONS Matte investigations may be divided into three classes: first, those treating the sulphides as alloys-theoretical studies of the system FeS-Cu2S, such as those of Hofman (ll)t or Carpenter and Hayward (2); second, chemical research on mattes made under smelter conditions, such as that carried out by Hawley (3) or Avetision ( 4); third, combinations of the . above, such as the investigations of Gibb and Philp (5) or Fulton and Goodner (6). Much research in the first of these classes, including determinations of . freezing points and of data for structure-composition diagrams of the system FeS-Cu2S, was carried out in the period 1890-1910. LeChatelier and Ziegler (1, 13) proved that FeS, when heated alone above its melting point, dissociated in part, and Friedrich ( quoted by Fulton, 6) showed the same for Cu2S. The reactions may be expressed by the following equations: [xFeS yFe + yS (volatilized)+ (x -y)FeS xCu2S )'Cu+ yS ( vohtilized) + (x -)')Cu,S ] Sulphur is lost by volatilization, leaving a melt of metallic iron plus ferrous sulphide in the one case, and of metallic copper plus cuprous sulphide in the other. If the mixed sulphides are heated far above their melting points, there results either an iron-rich matte containing Fe-FeS-Cu2S or a copper-rich matte containing FeS-Cu2S3-Cu, depending on the relative amounts of each sulphide in the mixture.

Jan 1, 1940

By S. M. Grillo, M. Fiori, R. Bornioli, C. Marini, S. Fadda

The Na-feldspar orebodies in Sardinia occur along structures belonging to a regional ENE-WSW system that developed during the late stages of the Hercynian orogeny, and that was reactived during the Alpine orogeny. The main petrographic facies include granitoids, albitized granitoids, albitites (more or less rich in quartz), and albitites containing Mg minerals. Field observations and petrological and geochemical data point toward a metasomatic origin for the albitite orebodies. Albitization involved depletion of K from K-feldspar and biotite, dissolution of free quartz, and a lowering of Sr/Na2O and CaO/Na2O ratios in the protolith, under conditions of a positive thermal gradient. The absence of later K-metasomatism implies that, during a subsequent decline in temperature, the composition of the albitizing fuids was modified by an influx of meteoric or other waters. Copyright © 1996 Canadian Institute of Mining, Metallurgy and Petroleum.

Jan 1, 1997

By A. Dave Mackintosh

The Current demand for potash has pushed world prices to record levels and allowed Saskatchewan producers to operate at maximum capacity. Several years of strong demand has resulted in a permitting rush not seen since the 1960?s. Even though Saskatchewan has a huge potash resource, mining and processing constraints will limit the areas of economic development. Introduction In 1969 M. E. Holter prepared a report ?The Middle Devonian Prairie Evaporite of Saskatchewan?, published by the then Dept. of Mineral Resources. Holter?s report has remained the bible of geologists working in potash ever since. It is astounding that a report written before there was really much potash mining history in the province, is as valid today as when it was first published. This paper re-iterates much of Holter?s report and adds a little mining experience to essentially confirm the importance of his work.

May 1, 2008

By R. R. Song, Q. Li, H. Xing, X. C. Wei

Geomechanical modelling of fault stability is an important part of CO2 capture and storage (CCS) research program to ensure its viability and long-term safety. Storage of CO2 in deep subsurface reservoirs requires estimates of sustainable fluid pressures that will not induce fracturing or fault reactivation which could cause fault slip. More importantly, fault slip gives rise to seismic waves, which inflict severe damage to the cap rock of the reservoirs and may result in leakage pathways for the buoyant plume and thus compromise the integrity of seal formations. In this study, we generated a numerical model tuned towards conditions usually encountered in the Shenhua CCS demonstration project in Ordos basin. In particular, with the purpose of evaluating the sensitivity of the parameters as well as their relationships with the moment magnitude of the seismicity, a variance-based sensitivity analysis was adopted for importance ranking of the uncertain input parameters of the given model. However, such flow models could involve one or several spatial inputs, for instance, fault dip, distance between fault and the injection well, friction angle of fault, permeability of fault, Young's Modulus of fault, injecting pore pressure and the stress field. Studying the sensitivity to each of these spatial inputs motivated the present work. The results led to the following conclusions: fault depth and fault dip significantly affect the behavior of the fault and the fault slip becomes violent with the fault depth increasing and fault dip decreasing. Furthermore, the results indicate that slip time increases with the fault dip decreasing and decreases with permeability of fault increasing. In addition, the calculation of seismic moment and moment magnitude of an earthquake demonstrates that it rises significantly with the decrease of fault dip and the increase of fault depth, and it decreases with the rising stress field. The results show that variations of the only controllable parameter injection pressure do not show great influence to the fault behavior compared with other geological factors, which imply site selection is extremely important for the CCS projects.

Jan 1, 2015

By M. Jackson, R. McDonald, T. Osken, D. Robinson

"The addition of sulphidic materials in the leaching of nickel laterites has several potential benefits that include improvements in the rheological behaviour of the blends, reductions in the sulphuric acid and steam requirements, sweetening of the nickel grade of the combined feed material, and the option of using a range of sulphidic materials. The current study examines the oxidative processing of 30% w/w blends of nickel concentrate and nickel laterite ore at 250oC. In particular, the viability of using low oxygen overpressures was examined for significant amounts of nickel flotation concentrate, up to 12% w/w. Further to this the impact of process water quality, in particular the presence of chloride, upon the kinetics of extraction and the mineralogy of the residue were also investigated. This paper will present the results of preliminary work in this area.INTRODUCTIONThe use of sulphidic ore as a source of sulphuric acid during oxidative hydrometallurgical processing is not a novel idea and can be traced back over forty years (Seidel & Fitzhugh, 1968). A series of patents pertaining to application of this approach for laterite ores can be subsequently found (O’Neill, 1973; Opratko et al., 1974; Rodriguez, 2008a, 2008b; O’Callaghan, 2010; Liu et al., 2011) and a few of articles containing published data (Ferron & Fleming, 2004; Quinn et al., 2009). The use of pyrrhotite tailings as a source of sulphuric acid has also been employed for the leaching of converter slags (Perederiy et al., 2011).It is known that hydrometallurgical operations such as the HPAL plant at Murrin Murrin have added small amounts of sulphides as either transitional (mixed sulphide/oxide) ore or non-smeltable concentrate. This can improve the rheology (Rodriguez, 2009) and provide heat credits. However these plants are generally limited in their ability to add such materials due to their reducing nature and given under highly acidic conditions that the breakdown potential of the titanium used for the autoclave lining is significantly reduced. A suitable oxidant must therefore be present to ensure the potential remains sufficiently high."

Jan 1, 2012

By John Jiang, Brian R. Hart, Stamen S. Dimov, Aparup Chattopadhyay, Barun K. Gorain

"This paper presents a systematic approach to metallurgy diagnostics using an integrated gold deportment methodology developed for complex refractory gold and copper-gold ore bodies. Quantification of gold deportment for these ore bodies is a challenging task due to reasons such as lower grades, high variability of visible and invisible gold content in different sulphides and iron oxide phases, significant presence of disseminated and composite carbonaceous matters along with associations of various copper sulphide phases. Some case studies are presented in this paper to demonstrate the value of a high confidence gold deportment methodology in developing insights for problem diagnostics and to identify opportunities for metallurgical improvements. A holistic approach involving different fields of expertise is critical for successful development and delivery of solutions for complex ore bodies.INTRODUCTION Studies were initiated on various complex ore deposits and also on samples representing various process streams. These complex copper-gold single refractory, double and triple refractory gold ores are mainly composed of silicates and/or carbonates with moderate amounts of clay and mica, along with minor amounts of sulphides and organic carbonaceous matter. Variable proportions of visible gold and “invisible gold” (sub-microscopic gold) were noted in different mineral phases for various ore types and process stream samples. Invisible gold commonly occurs as solid solution or as colloidal-sized particles in different sulphides and other phases for different refractory ores. The dominance of invisible gold and presence of highly preg-robbing along with finely disseminated carbonaceous matters adversely affects gold recovery in flotation and leach circuits for most of these double and triple refractory ores. This paper discusses applications of an integrated gold deportment technique developed for different types of gold and copper-gold ores, namely, complex copper-gold, double and triple refractory gold ores (Chattopadhyay & Gorain, 2014; Chattopadhyay et al., 2016). Various process streams from flotation and gold leaching circuits were investigated. Different concentrates and tailings samples from flotation circuits whereas feed and discharges from pressure oxidation (POX) and carbon-in-leach (CIL) residues were used from gold processing circuits."

Jan 1, 2017

By The Mill Staff

Introduction The Hollinger cyanidation mill commenced operating in June, 1912, and 49,779,978 tons of solids had been discarded by the end of 1950. The present tailing disposal plant has been in operation since 1924 and holds 39,762,250 tons. The tailing disposal plant is on the property of Kayorum Gold Mines, Limited, about a mile and a quarter south of the mill. The area available for tailing covers 640 acres and is roughly l ¼ miles long by ¾ mile wide. Tailing was first deposited on the McKay veteran claims, l mile long by 1/2 mile wide, which were later included in the Kayorum property. This site sloped from northeast to southwest, the slope amounting to 106 feet between the northeast corner and the low land to the southwest. The low land was wet and a considerable area was covered with muskeg up to 17 feet deep. It was drained by a brook running across the property in a southwesterly direction. Later, when more ground was needed, the low-lying mining claims adjacent to the south and west boundaries were acquired to allow for the extension of the plant. This entailed some extra difficulties and expense that would not have been encountered had the whole of the present site been acquired with the original purchase. At the present time, the top of the tailing deposit stands at two elevations, but eventually the tailing deposited on the extension to the original plant will be raised to the same height as the older deposit. Previous to 1924, all of the tailing was deposited on Low land and in lakes near the mil1. Between 1924 and 1926 an increasing tonnage was sent to the McKay claims until, in 1926, all of it was being sent there. In 1936, the area of the deposit was extended and shortly afterwards all of the tailing was deposited on the new site.

Jan 1, 1951

By Rick Coleman, Chelvan (Chevy) Veloo

"P.T. Freeport Indonesia's lrian Jaya Operation has been producing copper concentrate continuously since 1972. Numerous ore discoveries and five mill expansion programs have seen the initial milling rate of 7,200 MTPD escalate to the present rate of 125,000 MTPD, producing 2.5 million pounds of copper and 3,500 troy ounces of gold per day. The latest expansion, commissioned in February 1995, incorporates a 34 foot Semi- Autogenous Grinding (SAG) Mill, two 20'x30' Ball Mills, conventional roughing, cleaner column flotation, and attendant upgrades to the existing concentrate pumping and dewatering facilities. This paper will describe the SAG Plant flowsheet, and focus on the key training issues which contributed to a successful plant start-up.IntroductionP.T. Freeport Indonesia's operations in the Ertsberg mining district are located in the Jayawijaya Mountains of Irian Jaya, the easternmost province of the Republic of Indonesia. P.T. Freeport Indonesia (PTFI), the operator, is owned by Freeport MacMoRan Copper and Gold (81.3%), the Government of Indonesia (9.4%), and private Indonesian investors (9.4%). Geographical location, and sailing distances are provided on the attached location map.The operation stretches over 7 4 miles ( 114 km) from the dewatering and concentrate shipping facility at the Port of Amamapare on the Arafura Sea to the Grasberg open pit located at 3400m. The regional geology is covered in detail by Van Nort et al (1991 ), with current proven and probable reserves estimated at 1.13 billion metric tonnes grading 1.30% Copper, 1.42 ppm Gold, and 4.06 ppm Silver. The Grasberg open pit mining operation provides 92% of the mill feed, and is described in detail by Hickson (1991). The remainder of the mill feed is derived from several underground block caving operations. Current production rates provide a mine life in excess of 25 years. However, very favorable geology in the area leads to expectations that additional reserves will extend the life of the operation considerably."

Jan 1, 1996

By Z. S. Shao, X. Z. Li

"Sub-critical crack growth plays an important role in long-term creep failure in brittle rock. Based on micromechanics established by using relationship between micromechanical wing crack growth model and macroscopic strain, and comparing theoretical model with testing results, a method, which applies the theory to real rock, is suggested in this paper. Using this method, properties of creep, damage and progressive failure in brittle rock are investigated. Effects of stress state on creep failure time and steady-state creep rate are analyzed and graphically presented. The results are helpful to support structure design and service life evaluation of underground structures with high geostatic stresses. INTRODUCTIONDuring the long-term operation in underground space with brittle surrounding rock, due to chemical action between pore water and rock, the failure of rock often occurs at stresses below its short-term failure strength, which brings serious effect on the long-term stability of underground structure. Thus, the long-term strength of brittle rock attracted wide interests of scholars around the world. The static fatigue limit was predicted by exponential function that is derived from abundant experiment data of stress level and failure time(Schmidtke & Lajtai, 1985), and similar result by relationship between stress state and failure time was also suggested (Lau & Chandler, 2004; Aubertin et al., 2000; Sangha & Dhir, 1972). According to relationship between steady creep rate and stress state during creep, a critical stress that transforms steady-state creep to accelerated creep was studied and the long-term strength was suggested (Tsai et al., 2008; Ma & Daemen, 2006; Lajtai et al., 1991). Based on the results of experiment in compression, the long-term strength of rock was defined by stress at volume strain reversal. When applied stresses are larger than the long-term strength, the microcracks in the rock will grow constantly, damage of rock will increase and failure of rock will occur finally (Chandler, 2013; Bieniawski, 1967; Brace et al., 1966)."

Jan 1, 2015

By M. Tanaka, H. Narita, K. Morisaku

"The selective separation of palladium (Pd), platinum (Pt), and rhodium (Rh) from a hydrochloric acid solution by solvent extraction was investigated using amide-type extractants, i.e., N,N,N’,N’-tetra-n-octyl-thiodiglycolamide (TDGA) and N-n-hexyl-bis(N-methyl-N-n-octyl-ethylamide) amine (BisAA). The separation experiments employed two separation schemes: (1) extraction of Pd with TDGA and that of Rh and Pt with BisAA; and (2) extraction of Rh, Pd, and Pt with BisAA. Separation scheme 1 achieved the highly selective separation of each metal. For separation scheme 2, Rh can be recovered prior to Pd and Pt, despite having a lower selectivity for Pd/Pt than separation scheme 1. INTRODUCTIONThe recent increase in the use of platinum group metals (PGMs) in industry (especially auto catalysts containing palladium (Pd), platinum (Pt), and rhodium (Rh)) has lead to their increasing waste; therefore, improvements in the recovery techniques for PGM have received significant attention. In metal refineries, PGMs are mainly separated and purified by solvent extraction (Cox, 2004). The extraction process used in INCO’s Acton precious metal refinery is one of the most famous commercial processes and is utilized worldwide. In this process, silver is first removed as a residue (AgCl), and then gold, Pd and Pt are extracted from hydrochloric acid (HCl) solutions using dibutylcarbitol (DBC), dialkylsulfide (DAS), and tri-n-butylphosphate (TBP), respectively. Finally, Rh remains in the raffinate (Barns & Edwards, 1982).The development of a successful solvent extraction process depends on the choice of the appropriate extractants. We have studied the extraction of PGM using N,N-disubstituted amide compounds (Narita, Tanaka, Morisaku, & Abe, 2004; Narita, Tanaka, Morisaku, & Tamura, 2006; Narita, Tanaka, & Morisaku, 2008a, 2008b). They are relatively stable against acids, and their metal extractability can be improved by changing the side chain length and/or introducing additional donor atoms like sulfur and nitrogen. We have recently found that thiodiglycolamide and an amide-containing tertiary amine have good extractabilities for Pd and Rh, respectively."

Jan 1, 2012

By Wynne Potter

"On the west coast of Cape Breton Island four coal basins have been identified, varying in age and extending from Port Hood northeasterly to Chimney Corner. Those of Chimney Corner - St. Rose in the north and Port Hood in the south belong to the Pennsylvanian Riversdale Group while lying between them are the coal basins of Mabou and Inverness belonging to the younger Pictou Group. Almost all of the coal basins are submarine and coal mining was carried out on all basins at various times. The seams in all basins dip steeply seaward at angles from 10 to 20 degrees but were found to flatten seaward. The coal is ranked as a high volatile C bituminous coal. Ash and sulphur contents are high compared with Sydney coal and the coal has a tendency to spontaneous combustion when stored in large quantities.Of the four basins, the most significant is that of Mabou. At Mabou recent seismic and drilling investigations have indicated that although a sizable reserve of coal may be found offshore the structural problems, particularly faults, and the poor quality of the coal encountered will make it a resource of the future, if at all. Inferred resources have been calculated at 150 million tonnes in four seams over a 10 km2 area and extending 3 km from shore. The quality of the Mabou coals varies with ash, ranging from 5 to 25 per cent and sulphur up to 6 per cent. Calorific value ranges from 25.3 Mj I kg to 31. 7 Mj I kg. The Port Hood and St. Rose -Chimney Corner fields (Fig. 1) contain 16 million and 3 million tonnes respectively, but the Inverness basin is considered to have been depleted. The active mine in this area is the Evans Mine at St. Rose that is currently producing between 30 -50,000 tonnes a year from proven reserves of approximately 3 million tonnes."

Jan 1, 1985

By L. P. Srivastava

Rock bolts have been used since long to enhance the shear strength response of jointed rocks in civil and mining engineering structures. However, the assessment of strength response of a rock mass reinforced with rock bolts remains an unresolved problem till date. In the present article, an attempt has been made to study, through physical model tests, the shear strength enhancement of a blocky mass due to provision of passive rock bolts. Direct shear tests were conducted on large specimens of size 750 mm x 750 mm x 900 mm of rock mass under a normal stress varying from 0 to 2.0 MPa. Concrete was used as a model material to simulate intact rock and, elemental blocks of size 150 mm x 150 mm x 150 mm were used to form the mass specimens. Steel bolts of 6 mm diameter with 2 mm thick grout were used to reinforce the mass. Three different configurations of bolts were adopted. Observations were made on shear stress, horizontal shear displacement, and dilation during the tests. Effect of normal stress and reinforcement on peak shear stress and shear stiffness was studied. It is observed that the rock bolts enhance the shear strength of the mass by preventing sliding along the shearing plane, and by inducing additional normal stress due to tension mobilized in the bolts. The rock mass, therefore, becomes stronger due to the high degree of interlocking produced by the bolts. Shear strength of the mass increases with increasing numbers of bolts. In the case of reinforced mass, the enhancement in shear strength due to provision of bolts decreases with an increase in normal stress. The provision of bolts improves the cohesion of joints substantially. The effect on friction angle is low. The shear strength enhancement depends on the area of the bolts, spacing of the joints, spacing between the bolts and the cohesion of the intact material. A correlation has been established using which the enhancement in shear strength due to the provision of bolts can be assessed.

Jan 1, 2015

By E. Barnard, J. Tero, J. Rajala, N. Avenido, M. Fullam, A. Nashitati, S. Mbalamba

"The Otjikoto mining and milling complex is 90% owned by B2Gold Corporation, and began commercial operation on February 28, 2015. The milling circuit operates at a throughput 3.3 million tonnes per annum, with an average reserve gold grade of 1.28 g/t. The ore is high in Gravity Recoverable Gold (GRG), which varies between 73-91%, depending on ore type. Overall gold recovery since start-up is 98.2%, with about 63.5 % coming from the gravity circuit as doré. 2016 cash operating costs, and all in sustaining cost (“AISC”) were $361 and $604 per ounce, respectively.During design, three ore types were tested for GRG content, and based on the high GRG value of all three ore types, the decision was made to treat the full ball mill circulating load by gravity. Once operating, the gravity circuit performance was characterized by performing a detailed gravity audit, which included the ore, cyclone classification, Knelson concentrators, and the Consep Acacia intensive leach system. The deeper ores contain significant pyrrhotite, and this created some unusual challenges that had to be overcome. This paper presents the gravity circuit design process, the circuit commissioning, and the audit results and ongoing improvements.IntroductionHeadquartered in Vancouver, Canada, B2Gold Corp. was founded in 2007. Currently, B2Gold has five operating mines, and numerous exploration projects in various countries, including Nicaragua, the Philippines, Namibia, Mali and Burkina Faso. The Otjikoto Mine is located in the north-central part of the Republic of Namibia. It is situated approximately 300 kilometers north of Windhoek, the country’s capital, within the Province of Otjozondjupa. The Otjikoto mining and milling complex is 90% owned by B2Gold Corporation, and began commercial operation on February 28, 2015. The milling circuit operates at a throughput 3.3 million tonnes per annum. During design, a large emphasis was placed on gravity recovery, as the Otjikoto ores are unusually high in Gravity Recoverable Gold (GRG)."

Jan 1, 2018

By N. W. Xu, M. D. Wei, F. Dai, B. Li

"Chevron notched samples achieve general acceptance for measuring the Mode-I (opening mode) fracture toughness of rocks. The cracked chevron notched semi-circular bending (CCNSCB) sample, which is developed from the International Society of Rock Mechanics suggested the cracked chevron notched Brazilian disc method, has recently received much attention. However, these chevron notched methods have not been fully mastered, some ideal assumptions based by their theories and formulas, such as the straight-through crack assumption (STCA), have not been verified detailedly. Fractures of CCNSCB specimens were numerically simulated considering different heterogeneities, geometrical and mechanical parameters. The results show that, the real fractures of CCNSCB specimens are never satisfied with the STCA. In fact, micro fractures not only occur in the notch tip but also the two chevron edges. Consequently, the crack front features prominently in a shape of parabolic curve, far from being straightthrough. Moreover, curvature of the crack front is only affected by shape of the chevron ligament. Cracks usually start from two sides of the notch tips rather than center of the notch width, and propagate deviating from the postulated fracture routes. These fracture mechanisms induce errors with the measurements, and thus more attentions should be paid to the real fractures of the chevron notched samples for measuring fracture toughness more accurately. INTRODUCTIONWith the increasing requirement of people for resources and energies, frequent rock engineering activities have been conducting, such as underground excavation, oil exploration and deep burial of nuclear waste. Many problems appear in rock fracture. Consequently, there are more and more attentions paid in rock fracture mechanics due to its broad application prospects for solving these engineering problems. Since Mode-I (opening mode) fracture is the most common failure mode in rocks, Mode-I fracture toughness is a primary parameter to predict rock mechanical properties in actual engineering, and thus its accurate values are necessary to the researches of rockburst, rock mass slope stability, rock cutting, hydraulic fracturing, tunneling, et al. Therefore, it is thus significant to evaluate and study relevant methods to obtain the reliable value."

Jan 1, 2015