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By R. Kumar, N. R. Mandre

"The beneficiation of iron ore slimes by selective flocculation was studied. Polyacrylamide (PAM) and guar gum were used as flocculant, along with sodium hexametaphosphate (SHMP) as a dispersant. The samples collected were initially subjected for chemical and mineralogical studies. The mineralogical studies indicated the presence of haematite and goethite as the principal iron-bearing minerals, with quartz and kaolinite as gangue minerals. The selective flocculation tests showed that the Fe grade can be enhanced from 58.24% to 64.60% Fe at a recovery of 66.33% using polyacrylamide as the flocculant. A Fe grade of 63.20% at a recovery of 68.04% was obtained using guar gum as flocculant. The efficacy of the studies was analysed using a separation index, which indicated that the higher separation index was obtained with the tests employing polyacrylamide as the flocculant. IntroductionIndia is endowed with huge reserves of highgrade iron ore and is one of the prominent producers in the world. As the high-grade iron ores have been exhaustively mined, it has become imperative to use low-grade iron ores, fines, and slimes to meet the growing demand. Also, generation of fines and slimes during mining and processing contributes to loss of minerals to the extent of 30% of the total value (Mandre and Panigrahi, 1997). These fines/slimes are often discarded as waste into tailing ponds, containing considerable amounts of iron. Therefore beneficiation of tailings/slimes has become necessary for optimal utilization of the resources.In recent years, interest in recovery of valuable minerals from low-grade ores, slimes, and tailings has increased worldwide. These activities are initiated not only to recover minerals but also to address various environmental issues associated with fine particle treatment. The literature reveals that several technologies have been employed in this regard, including flotation, advanced gravity separation, and high-intensity magnetic separation. However, these processes are found to be ineffective for fine particle treatment due to lower collision probability in flotation, decreased gravitational effect in gravity separation, and reduced drag force in magnetic separation, among other factors (Batisteli and Peres, 2008; Das et al., 2007; Dworzanowski, 2012; Rocha, Caneado, and Peres, 2010). Therefore, it is in this regard that selective flocculation has gained importance in treating slimes in recent years."

Jan 1, 2017

By C. J. Konya, A. Konya

"Presplitting is a form of blasting that when executed properly will create a smooth wall that will minimize secondary conditioning. In recent years, presplitting has become a popular technique to be used in both the mining and construction industry, however, presplitting fails on many occasions because of ‘geologic conditions’. This is typically because of improper explosive loads consisting of overloading and underloading blastholes. In this paper the authors introduce new factors such as a presplit factor which varies for different rock types, selection criteria for the spacing of presplits, the optimization of explosive load, as well as prediction methods for the environmental affects for the technique of Precision Presplitting. BACKGROUND Methods of controlled blasting in which a smooth, stable wall could be blasted have been an important consideration since blasting in mining and construction has begun. How can one use an explosive, with its massive energy and power, to break a non-homogenous rock without causing backbreak, poor walls, or deviations in the wall over large distances? The blasting industry has known for a long time that blasting ‘high benches’ would result in better walls and less backbreak. However, this was not a sufficiently controlled technique for critical projects, such as the Niagara Power Plant. This was one of the main challenges Hercules Powder faced when they began the project. In order to achieve this, they used a new tool in the explosive engineer’s toolkit, presplitting. The presplitting done on the Niagara Power Project utilized 1 ½” x 8” (38mm x 203mm) 40% gelatin dynamite in 2 ½” (63.5) diameter boreholes. This was done on multiple levels with benches between 12 feet to 52 feet (3.6m to 15.8m) with a total powder factor of 0.75 pounds/cubic yard (0.45 kg/m3). The spacing on these holes ranged between 18 inches to 24 inches (0.46m to 0.61m), center to center (Langfors & Kihlstrom, 1973). The strange aspect of this project was that this “lightly loaded” closely spaced row was fired before the production blast, yet this project produced some of the best walls seen at that time directly from blasting. In today’s blasting industry, presplitting is a more common way to produce smooth, stable high-walls for both the construction and the mining industry. Today, explosive loads for traditional presplitting is generally measured in pounds per foot (kilograms per meter) and typically ranges from 0.30 lb. /ft. to 0.5 lb. /ft. (0.45 kg/m to 0.75 kg/m). MECHANICS OF PRESPLITTING When analyzing a blasting process like the Niagara Project, many wonder how did that work when normally confining a charge would cause cratering and increased backbreak. Since then, presplitting mechanics have been studied extensively. It has been shown in many models that presplitting is not achieve through the collision of shockwave, but by gas pressure in the boreholes. In addition to models and finite element simulations, this has been proven full scale using Pyrodex, a propellant, to create a presplit plane with no shock wave present (Konya, Dean Barrett, & Smith Jr, Presplitting Granite Using Pyrodex, A Propellant, 1986)."

Jan 1, 2017

By Terrence M. Carroll, Chu E. Ho, Dimitrios Niarchos

"Construction of the new extension to Terminal 4 at John F Kennedy International Airport in Queens, New York involved the application of tapered piles for foundation support. Initial design based on past load testing experience had indicated that a pile depth of 65 feet was necessary to achieve a capacity of 394 tons. However, piles with a single length of 62 feet were ordered and installed, which led to marginal performance of the piles. This paper presents the technical considerations for mitigating the situation based on a systematic review of three preliminary pile load tests and dynamic capacity measurements on the shorter piles. Consideration was given to the long term gain in capacity of tapered piles due to set up in the Glacial Sand. The observational approach to the evaluation of piling performance adopted on this project resulted in salvaging many piles which failed to comply with the specified driving criteria. It was acknowledged that a more prudent course of action was for the Contractor to have ordered more appropriate pile lengths that allowed for potential variation in soil conditions across the site. The 4.6% saving in pile material did not warrant the effort that was required to mitigate the 39% of installed piles that did not meet the specified driving criterion.INTRODUCTIONAn expansion project at John F. Kennedy International Airport in Queens, New York involved the Phase 2 extension of Concourse B at Terminal 4. The site consists of 4 to 5 feet of pavement over about 10 feet of hydraulically placed sandy fill. Below this is a layer of marsh deposit, comprising of soft organic clay and peat, up to about 11 feet thick. Beneath the marsh deposits is a deep layer of glacial sand which extends to well over 100 feet below ground. Figure 1 shows the typical subsurface profile at the site. The new structures were supported on 223 steel tapered piles. The piles consisted of a straight steel pipe 18 inches in diameter and 0.375 inches thick, with a 25 feet long tapered bottom section (diameter varying from 18 inches down to 8 inches at the pile toe). The piles were first driven to the founding depth and subsequently in-filled with concrete. The concrete strength (fc) specified was 4,500 psi and the yield strength of the steel piles (fy) was 50 ksi. Initial pile design, based on earlier experience with static pile load testing for Phase 1 development at the same terminal, indicated that a pile length of 65 feet was necessary. However, piles with a single length of 62 feet were ordered and installed by the Contractor, leading to marginal performance of the piles. This paper presents the technical considerations for the adoption of alternative pile driving criteria, based on a systematic review of the preliminary static load test results for the present site and PDA measurements obtained during pile installation, in order to accommodate the capacity of the shorter piles."

Jan 1, 2017

By F. Mohamm

"Recent changes in regulations and scientific advances concerning selenium bioaccumulation have placed greater importance on non-biological treatment systems for selenium control. Mining wastewater contains hexavalent selenium (selenate) as the predominant aqueous species and possibly small amounts of tetravalent selenium (selenite). Non-biological removal of selenite is accomplished readily and cost effectively by ferric co-precipitation and adsorption using ferric oxyhydroxides sourced from ferric salts or iron anodes in electrocoagulation cells. Selenate on the other hand is difficult to remove. It is often incorrectly assumed that electrocoagulation is suitable for removing both selenite and selenate. Selenate removal efficiency by electrocoagulation is low and the amount of waste residue produced high. In contrast, BioteQ’s Selen-IX™ process removes selenate very efficiently and produces a much smaller amount of stable residue. The process combines selective ion exchange with electro-reduction of selenate in electrocells that physically resemble electrocoagulation cells but are designed and operated differently. This paper summarizes four years of pilot scale demonstration on a wide range of mine waters and discusses the key benefits and limitations of Selen-IX™. The pros and cons of combining ion exchange with electro-reduction in comparison to direct electro-reduction of selenate are also reviewed.INTRODUCTION Selenium in Mining and Mineral Processing Effluents Similar to sulphate and metals, selenium presence in mining and mineral processing wastewater originates from minerals contained in the mined deposits. Therefore, no organoselenium species are typically present in these kinds of waters and inorganic species including selenite (SeO32-), selenate (SeO42) and elemental selenium account for the majority, if not all, selenium present in these streams. Elemental selenium is not always present in mine water but is often contained in smelter wastewater. To comply with new environmental regulations, a successful water treatment plant must remove selenium from all these forms effectively and economically without generating by-products that are difficult to dispose of or manage. Elemental selenium can be dealt with using filtration or gravity separation in clarifiers often assisted by flocculants and/or coagulants. Mine impacted waters typically contain other forms of suspended solids which are removed from water in this way. Once the solids are removed, the selenium species that remain untreated are the dissolved forms. In mining waters this is mostly selenate with small fraction of selenite. In smelter wastewater, the majority could be selenite. It is well established that tetravalent selenium is readily removed from water by using ferric iron co-precipitation at pH > 4 (NAMC SWG, 2010). Ferric hydroxide/ferrihydrite solids act as the adsorbent for tetravalent selenium when it is in one of the ionic forms of biselenite or selenite. This treatment for tetravalent selenium is highly efficient. Therefore, tetravalent selenium removal is generally not an issue and is often easily incorporated into lime neutralization or unit operations such as electrocoagulation used for removing other metals present in mine impacted waters. Another option of dealing with tetravalent selenium is sulphide precipitation in acidic pH which produces SeS2 and can separate selenite from other metals and/or sludge produced at neutral to alkali pH."

Jan 1, 2017

By L. Vollert

"The application of pneumatically-assisted fluidized bed flotation, specifically the ERIEZ HydroFloatTM technology, has found widespread use in non-sulphide processing (particularly phosphate, potash and coal) for the recovery of coarse particles. In gold processing, centrifugal gravity circuits and flash flotation cells are typically employed for the recovery of coarse and near-size liberated gold which tends to accumulate in recycled grinding streams due to bypass to the classifying cyclone underflow streams owing to the higher specific gravity of these gold particles. Early removal of this gold from the grinding circuits contributes toward the reduction of downstream processing losses as a result of overgrinding and particle shape deformation in the case of flotation processing, longer leach residence times in the case of extractive processing, and surface occlusions in the case of both flotation and extractive processing. Furthermore, recovery to bullion attracts higher payment terms by way of comparison to concentrate sales in the case of copper-gold concentrators. This paper assesses the application of pneumatically assisted fluidized bed flotation for the recovery of coarse liberated gold through the analysis of piloting data generated from the operation of a continuously operated 16 inch ERIEZ HydroFloatTM unit treating a sized coarse grinding circuit stream. The paper presents the stage recovery by size performance of the ERIEZ HydroFloatTM unit together with quantitative gold deportment analysis to demonstrate the efficiency with which gold was recovered during this pilot campaign and then compared to more conventional processing units – centrifugal gravity and conventional flotation.INTRODUCTION Newcrest Mining Limited (NML) has identified that an opportunity exists for improving the recovery of coarse composite particles at its current operations, both as a means for improving overall recovery and as an opportunity for early gangue rejection. As such, Newcrest Mining Limited has taken an interest in the application of pneumatically-assisted, fluidized-bed flotation i.e. ERIEZ HydroFloatTM technology. This interest dates back to 2011 with sighter testwork carried out on their Australian-held ore bodies (Kohmuench et al, 2013). Newcrest continues to assess ERIEZ HydroFloatTM potential for coarse composite recovery and, in the case of the topic of this paper, coarse gold recovery based on examples in the literature which demonstrate the efficiency of this technology in this regard (Carmona Franco et al., (2015); Miller et al., (2016); and Kohmuench et al, 2013)."

Jan 1, 2017

By J. Bilant, J. Donnelly

"A spatially-discretized heap leach simulation model was developed within the GoldSimTM modeling framework for the Coeur Rochester Mine, Nevada. The model is being evaluated by Rochester as a means of improving forecasting of silver and gold production on an actively loaded valley fill heap leach facility. Optimization of future heap leach operations and crusher product size is also being evaluated with this software. Simulation modeling incorporates metallurgical data such as mine planning, particle size distribution of placed ore, stacking plans, pad geometry, and leaching dynamics. The heap leach model is represented as a discretized matrix with lifts providing vertical separation to form 19,840 spatially discrete cells. Each cell is quantified with regard to ore properties and metal content. The model tracks solution and cyanide-complexed silver and gold within each cell as a function of time based on pad operations, unsaturated flow properties, and kinetic rates. Practical heap leach management and daily operation will be discussed as it relates to the simulation and forecasting model. INTRODUCTION The Rochester mine, and associated valley fill heap leach facilities, is an open-pit silver and gold mine located in Pershing County, Nevada. In 2015 Rochester produced approximately 4.6M troy oz silver (Ag) and 52K troy oz gold (Au) through combined cyanide heap leaching of the Stage II, III and IV heap leach pads with recovery via a Merrill Crowe process plant. Rochester is actively stacking and leaching ore on the Stage III Heap Leach Pad. The Stage III pad contains approximately 70M tons of ore in a valley fill configuration and is forecasted to receive an additional 20M tons of ore by the life-of-pad in 2019. The ultimate height of the pad is permitted to be 400 ft above the liner and solution collection system. Approximately 80 percent of the ore loaded on the Stage III pad is produced from tertiary crushing split from two crushing circuits that produce a nominal 80 percent passing 3/8” product. The remaining material consists of run-of-mine (ROM) from off loads of re-mined waste rock stockpiles. Metallurgical testing of the ore indicates low variability in geologic properties and high variability of both Ag and Au extraction rates for the various particle size distributions. Historically the valley fill configuration provided metal forecasting challenges that could not be predicted as the leach pads grew. Forecasting metal production and leach optimization of the Stage III pad became challenging due to the extreme variation in heap leach pad depth, application of available barren leach solution and the variability of Ag and Au extraction rates for the various ore particle size distributions. To assist with understanding and optimizing metal production forecasting with these factors for the Stage III pad, Hatch developed a spatially-discretized heap leach simulation model using GoldsimTM dynamic simulation software. The model incorporates metallurgical data such as mine planning (tonnages and grade), particle size distribution, ore crushing, ore stacking plan, pad geometry, and both leaching dynamics and kinetics to forecast metal production from a consolidated mine plan."

Jan 1, 2017

By Gilbert R. Tallard

"Pouring deep foundations by the tremie pipe method is nowadays such a common process that it has becomes an unspecified practice. Means and methods are almost entirely at the contactor’s discretion. A variety of practices are permitted and seldom associated with the discovery of abnormalities and areas of random defects in concrete. This writer wishes to call the attention of stakeholders to the initiation details of the tremie pour operation.KEYWORDSINTRODUCTIONThe advent of slurry as a means of shoring an excavation of relatively small cross section but of possibly great depth has become part of state of the art in foundations construction, with the substitution of said slurry by concrete to provide the structural end product element through the use of tremie pipes. In appearance, there is nothing more simple than pouring concrete into a tremie pipe through a hopper after the basic work has been done i.e. the excavation or drilling, the cleaning of the slurry, the placement of the reinforcing cage, then the tremie concrete pour is almost a reward for a successful achievement. Force of habit has removed the awareness on the process fundamentals for initiating the pour; a variety of practices used call for a review and a reassessment as a means of reducing construction defects that may result from flawed practices, and and especially we are pouring deeper holes. This writer intends to address this issue focusing both on the drilled shaft and the diaphragm wall industries.TREMIE PIPES AND HOPPERSA tremie pipe diameter needs to be sufficiently large to allow the rapid descent of concrete with a minimum of friction loss and avoiding possible bridging in relation to aggregate size. Impervious joints between pipe sections are essential to prevent slurry “raining” over the concrete which can cause concrete segregation pockets and possibly plug the pipe with damaging consequences. That tremie pipes be long enough to reach the bottom may be self-evident but poorly written specifications can be erratic on that subject and seldom specific in terms of inches. A standard pipe of 10 inches is well accepted throughout the foundations industry; double O rings and spline joints are common and function well if properly maintained. However the practice to cover the joint with the ineffable duct tape indicates that all is not always well with this detail. Also, spline cables may shear in case of very long tremie pipes causing the separation of the tremie pipe requiring a restart in midstream with another probable source of inter mixing of concrete and slurry with subsequent defects."

Jan 1, 2017

By John Shirokoff, Ladan Khaksar

"Cyclic polarization technique and also potentiodynamic measurements were applied to study corrosion behavior of 13% chromium steels in chloride containing environment. Potentiodynamic measurements were performed at 25 and 75 °C with 0, 10 and 20 g/l NaCl in 0.01 M hydrochloric acid solution. At the same time, the significant pitting parameters such as corrosion potential (Ecorr), pitting potential (Epit), protection potential (Epro) and passive current (Ipass) of cyclic polarization curves with various NaCl concentration were analyzed and compared to describe the corrosion behavior of 13% chromium stainless steel. Surface measurement techniques such as SEM and EDS were used for morphology and analysis of corrosion products. The results showed that both temperature and Cl- concentration had significant effects on corrosion behavior of 13% chromium stainless steel in acidic solution.INTRODUCTION Of all the corrosion resistant alloys (CRA) used in oil and gas developments, martensitic stainless steels (MSS) are some of the most common ones these days. These steels not only have the advantages of normal steels: high strength, cost-effective, easy to manufacture, available in a wide variety of product forms, but they also have good CO2 corrosion resistance at higher temperatures as a result of their 11% minimum chromium content. One of the most common types of MSS used in the oil and gas industry is 13% chromium steel. NACE and API standards both reference several different varieties of 13% chromium steels for use in oil and gas projects(“Corrosion Resistant High Cr Steel for Oil and Gas Wells,” 2013). All types of 13% chromium steels have about 12 to 14% chromium and other minor alloying additions, and can be placed into two general groups: conventional and modified. Conventional 13% Cr steels (13Cr) are those most common in oil and gas, some of which are CA6NM (UNS J91150), AISI 410(UNS S41000), and AISI 420 (UNS S42000). These alloys have excellent corrosion resistance when compared to carbon steels up to about 300°F (150°C) and are used for a variety of wellhead, tubular, and downhole components however they have very limited resistance to cracking in H2S (Vallourec Group, 2015). Modified 13% Cr (M13Cr) steels came from the need for increased corrosion resistance at higher temperatures, increased H2S cracking resistance, and increased yield strength above 85 ksi (586 Mpa) primarily for tubular goods and other downhole equipment. These features were developed by taking a conventional 13Cr steel and adding nickel (Ni), molybdenum (Mo), copper (Cu), and other alloying elements. M13Cr provides better corrosion resistance than 13Cr up to about 350°F (177°C), yield strengths up to 110 ksi (758 Mpa), and also improved H2S cracking resistance(“Corrosion Resistant High Cr Steel for Oil and Gas Wells,” 2013)."

Jan 1, 2017

By K. K. Murray, D. G. Gelderland

"The pressure oxidation autoclave at the Goldcorp’s Red Lake Gold Mines (RLGM) was commissioned in 1991. Since commissioning, the circuit has been well maintained and its capacity increased to mechanical and gaseous oxygen limitations. With advances in available technology, opportunities to produce more sulphide concentrate and a changing business environment, an optimization project has been initiated at RLGM. A multi-phase approach has been developed to optimize both the operations and economics of the circuit. This paper describes the history and future of the autoclave circuit at RLGM.AcknowledmentsWe would like to thank the operations and maintenance staff at Red Lake Gold Mines for their support and openness of mind to enact changes to enhance the operation. We would also like to acknowledge the members of the integrated team from Red Lake maintence, metallurgy, operations and corporate technical services that did the hard work to push past the status quo in order to advance productivity.OpportunityThe Red Lake Gold Mines (RLGM) pressure oxidation (POx) circuit has been in operation for over 25 years processing the high sulphide content gold bearing minerals. Increased sulphide concentrate production from improved flotation recovery and increasing gold sulphide association at depth has led to the POx circuit becoming a throughput bottleneck. The opportunity exists to process increased tonnages of sulphide concentrates and increase gold production. In order to realize this value, it will be necessary to increase the throughput capacity of the POx circuit.BackgroundThe POx autoclave at the RLGM, commissioned in 1991, was Canada’s first gold autoclave. It was installed to replace the two-stage fluid bed roaster that was built when the Campbell mine and mill were constructed in the late Forties. The circuit was originally designed to process 12.5 t/d sulphide sulphur in about 50 t/d of Campbell flotation concentrate. Following the merger of the Campbell and Red Lake Mines in 2006, the POx circuit started processing flotation concentrate from both Campbell and Red Lake Mills. In recent years, the circuit has been processing about 18 t/d sulphide sulphur in 75-85 t/d flotation concentrate."

Jan 1, 2017

By R. Agius, A. Nacu

"Cyanide destruction technologies are implemented at gold and silver operations to comply with the International Cyanide Management Code and meet ever more stringent discharge limits. Currently, the most common chemical cyanide destruction processes used in industry are sulfur dioxide/oxygen, hydrogen peroxide, and Caro’s acid. The cyanide destruction process used at an operation must reliably generate streams of acceptable quality at all times, and depends on numerous factors, including permit discharge requirements, tailings chemistry, location, site-specific climatic conditions and size of the operation. Selecting the most appropriate cyanide destruction process and determining the process design criteria requires accurate laboratory testing. This paper presents and discusses the results of a number of cyanide destruction evaluations that have been conducted to reduce cyanide concentration in process solutions or cyanidation tails. Laboratory testing provides important insights into reagent consumption and other factors required to achieve the target residual weak acid dissociable cyanide (CNWAD) levels.INTRODUCTION Process solutions associated with the extraction of gold and silver contain several forms of cyanide. The most toxic form of cyanide is free cyanide, which can be present as a dissociated cyanide anion (CN-) or gaseous or aqueous HCN. Weak or moderately stable cyanide complexes such as those of zinc, copper and nickel are classified as weak acid dissociable (CNWAD). Although less toxic than free cyanide, weak acid dissociable cyanide complexes are considered “ecologically sensitive” forms of cyanide and are regulated in most jurisdictions. In practice, reported WAD cyanide values comprises both free cyanide and weak acid dissociable cyanide complexes. Cyanide complexes with gold, mercury, cobalt and iron are referred to as strong acid dissociable complexes (CNSAD) and are much more stable, even under mildly acidic conditions. Another term commonly used in cyanide chemistry is total cyanide, which represents all cyanide compounds present in solution. Total cyanide levels in solutions are always equal to or higher than WAD cyanide values and the difference between the two usually indicates the presence of iron-cyanide compounds. Typically in Canada, CNTOTAL levels cannot exceed 1 mg/L in solutions discharged to the environment."

Jan 1, 2017

By E. Eccleston, E. Gardner, F. Letarte, R. Schonewille

"In 2005, after nearly a decade of process development and feasibility work, the commercialization of the Falconbridge Nickel Smelting Technology (NST) began in earnest with the design and construction of a grassroots ferronickel plant at Koniambo Nickel in New Caledonia. The NST process was developed as an alternative to the conventional Rotary Kiln-Electric Furnace (RKEF) process for production of ferronickel from saprolitic nickel laterite ores. The NST process promised environmental and energy efficiency advantages over the conventional route. A key step of the NST process involves fluidized bed pre-reduction of calcined laterite ore using direct injection of pulverized coal as the reductant. Two parallel fluidized bed reduction roasters (FBRs) were installed at Koniambo Nickel, each designed for a maximum capacity of 4,200 MTPD of calcined nickel laterite ore. This paper describes the process development and scale-up of the fluidized bed pre-reduction process, the design challenges associated with the prototype commercial units, and the current status of the FBR operation at Konaimbo.INTRODUCTION In the mid-1990s, Falconbridge began to develop a new flowsheet for production of ferronickel from nickel laterite ores. The process that developed came to be known as the Falconbridge Nickel Smelting Technology (NST). The flowsheet, which is particularly appropriate for saprolitic ores, involves several major unit operations, including: drying in a hammer mill-flash dryer, calcination and preheating in a flash calciner, pre-reduction in a fluidized bed reactor, and smelting in a DC arc furnace to produce the ferronickel product. This paper focuses on the fluidized bed pre-reduction step of the process and provides an overview of the process chemistry, laboratory and pilot plant development work, scale-up and commercialization of the pre-reduction step, and the results obtained in the prototype commercial units."

Jan 1, 2017

By C. Liu, G. Zhu, Y. Liao

"The effect of particle size on the flotation kinetics of a low-rank coal was compared using air bubbles and oily bubbles. Five kinetic models were applied to the data from the tests to derive the relationships between the flotation rate constant, the ultimate recovery, and the particle size fraction. The results show that oily bubble flotation (OBF) resulted in a lower ash content and higher combustible recovery than air bubble flotation (ABF), especially for the -125 ??m size fraction. It was found that the flotation of a low-rank coal in individual size fractions using air bubbles and oily bubbles can be best described by a first-order model with rectangular distribution of tendency to float and the classical first-order model respectively. OBF yielded higher ultimate combustible and ultimate ash recoveries. The flotation rate constants for OBF were higher in all the models except for the fully mixed reactor model. The ultimate recovery values and modified rate constants first increased and then decreased as the size fraction decreased. The trend of the flotation rate constants as a function of the size fraction was similar in all the models except for the fully mixed reactor model. IntroductionFroth flotation is a physical-chemical separation method that is widely used in coal beneficiation for the separation of fine coal particles from associated minerals in water slurries. Since the cumulative recovery of a component in the concentrate is proportional to flotation time, the flotation process can be considered as a time-rate recovery process (Emad, Mohammad, and Bahram, 2010). Therefore, the general mathematical flotation models that incorporate both a recovery and a rate function can describe flotation time recovery profiles.The kinetics of coal flotation, as a measure of coal recovery as a function of time, have been studied in detail by many researchers. Different kinetic models have been proposed and published in the literature. The order and the rate constant are strongly dependent on the flotation conditions, such as the particle size and size distribution, the bubble concentration, the reagent concentrations, and the operating parameters (Zhang et al., 2013). The particle size has been reported to affect the flotation kinetics to a great extent (Yoon, Luttrell, and Asmatulu, 2002; Ntengwe and Witika, 2011). Numerous researchers have studied the aspects of flotation kinetics while paying special attention to particle size (Meftuni and Ibrahim, 1997; Rubinstein and Samygin, 1998; Graeme, 2012). However, the effect of particle size on the flotation kinetics of a low-rank coal has been relatively little discussed."

Jan 1, 2017

By Alan B. Richards, Aleks Todoroski

The undesirable effects of blasting fumes can be controlled by minimising the amount of fume created, and by avoiding the movement of the fume plume to sensitive locations. Substantial progress has been made by the blasting industry to reduce the amount of fume typically created by blasting, but at this time no specific blast can be guaranteed to be ‘fume free’.

Jan 1, 2017

By Takefumi Takuma, Ali Fayad, Shigeru Kambe

"This paper discusses the unique design and construction circumstances of the cantilevered dual sheet pile walls and the Accordion Sheet Pile Wall for the repair and upgrade of the East Garden Grove-Wintersburg Channel in Huntington Beach, California that was concluded in 2014. It also describes the 2008 Emergency Sheet Pile Project for the same channel. Both projects utilized the Press-in Piling Method for its low noise and almost no vibration advantages.BackgroundThe East Garden Grove-Wintersburg Channel is part of the Orange County Flood Control District (OCFCD) and is operated and maintained by OC (Orange County) Public Works (see Figure 1 for its location). The channel, referred to as OCFCD facility # C05 (C05), had suffered significant toe undercutting and slope erosion of its levee due to a series of storms. These began after the facility was constructed in the 1960’s and culminated in 2005. Consequently, the levee was deemed significantly deficient and was eventually declared as a local emergency in 2007. C05 was an incised channel and its levee was constructed primarily from native materials. Operation and Maintenance (O&M) personnel notified the design staff at OC Public Works to be aware of major consequences associated with the dilapidation problem in the most downstream reach of the channel in the vicinity of the Bolsa Chica Wetlands Restoration Project. In 2007, the Orange County Board of Supervisors declared a local emergency to fix the levees with sheet piles to be driven by the Press-in Piling Method, hydraulic pushing piles into ground by holding onto previously installed piles. The project that ensued represented the first in-tandem use on the West Coast of the United States of the Press-in pile driver with its GRB System (Press-in pile driver, a support crane and pile transporter travel on top of already-driven sheet piles). This is further discussed in this paper in the sections titled “Press-in Piling Method” and “GRB System”, respectively. The project repaired approximately 3,800 feet of erosion on the north levee of C05 with 45-foot long cantilevered sheet piles. The sheet piles design constraints are discussed in this paper in the section titled “C05 2008 Emergency Sheet Pile Project (Phase 1)”. Between 2012 and 2014, another mile’s worth of levee reach was improved with the Soil-Mix-Sandwich (SMS) Method. This concept consisted of deep soil cement mix (DSCM) columns sandwiched within the dual rows of sheet piles. Additional details on the SMS, its DSCM construction, and its hybrid geo-structure design concepts are well beyond the scope of this discussion, which is primarily focused on sheet piles. Instead, the interested reader is referred to two papers on the SMS and its DSCM metrics (Fayad, 2015; and Arora, et al., 2015). These were presented and published in the Proceedings of the Deep Foundations Institute (DFI) 2015 Conference on Deep Mixing. Hence, in the section of this paper titled “C05 SMS Dual Sheet Piles Project (Phase 2)” only the sheet pile design constraints of the SMS are covered. Finally, in the area where the project traversed the Alquist-Priolo zone of the Newport-Inglewood tectonic fault, an expandable and deflectable sheet pile wall was used to isolate the flood channel from the Bolsa Chica wetlands. Its details are addressed in the section titled “C05 Accordion Sheet Pile Wall (Phase 2)” of this paper."

Jan 1, 2017

By C E. Johnson, S Hosein

Previous researchers have put forward two different theories as to the origin of air overpressure from quarry blasting. In 1980, Siskind et al postulated that the initial face movement gave rise to the displacement of air and that this resulted in the air overpressure pulse. However in 1990 McKenzie et al carried out research that indicated that the air overpressure pulse came into being as a consequence of the shock wave created by the detention of the explosive charge in a blast hole. In an attempt to evaluate these hypotheses, an investigation was undertaken over a 4-year period with the aim of determining the precise origin of air overpressure. The research was carried out at a chalk quarry in the North of England.

Jan 1, 2017

By Jean Wehbe, Daniel C. Brahana

"This paper presents a case history of drilled pier construction for a new stadium in Atlanta, Georgia. The stadium structure is supported on several different foundation types, with the heaviest ""mega columns"" typically bearing on drilled pier supported mats. The project includes 141 drilled piers with design diameters of 5.5, 6 and 6.5 feet. The piers bear on bedrock at depths of 20 to 110 feet below the ground surface. Many of the piers included rock sockets with design lengths of 5 to 18 feet. A pre-production drilled pier load test program was completed including one Statnamic load test, two Osterberg Cell tests and two lateral load tests. Non-destructive testing (crosshole sonic logging and sonic integrity testing) were also performed during construction. Production pier installation with over 3800 linear feet of earth excavation and over 1350 linear feet of rock excavation was completed in a little over four months. The piers were installed using both the cased hole method and the slurry method. All cased holed required downhole inspection and slurry holes were initially inspected with a mini-SID device and subsequently by sounding. Advanced drilling equipment and tooling were used to meet the production requirements for the project. This paper reviews the highly variable soil and rock conditions encountered at the site, briefly presents the results of the load tests, and discusses pier installation procedures and the equipment used for installation. The paper also presents conclusions related to the installation and lessons learned.INTRODUCTIONWith the opening of AT&T Stadium in Dallas, a new era of stadium construction began in the U.S. The new age stadiums are larger and have more amenities than their predecessors. The new Atlanta Falcons stadium is no exception. Many of the new features, including Atlanta’s innovative retractable roof, result in extraordinary structural loads. The roof structure is supported by a series of mega columns that bear on eight large caps, six of which are drilled pier supported. Each of the pier caps contains between 8 and 16 drilled piers that bear in the underlying bedrock. The design pier diameters are 6 and 6.5 feet. Augercast piles were used to support two of the mega column caps. Drilled piers, augercast piles, aggregate piers and spread footings are used to support the more lightly loaded portions of the structure."

Jan 1, 2017

By Arash Khosravifar, Ali N. Ibrahim

"Large areas of Iraq are covered with collapsible soils namely gypsiferous soil which is often used as the foundation soil for structures. In particular, the Mosul Dam foundation is comprised of high contents of gypsiferous soils. Intrusion of water through the foundation soil dissolves gypsum which acts as the binding material between soil particles. This phenomenon results in the formation of collapsible soils which can cause serious damages to the dam structure. The failure of Mosul Dam, with a reservoir as large as 11 billion cubic meter of water, can be catastrophic. In the past, Mosul Dam Authorities have been remediating the soil foundation by injecting cement mortar. This method has been proved not economic; therefore, new, more efficient, methods are desperately needed.This research study focuses on finding practical methods to remediate collapsibility of gypsiferous soils by improving their geotechnical properties. This study presents remediation of collapsible gypsum-rich soil by grouting. Three types of stabilizer materials were used to make the grout; (1) sodium silicate (Na2 O3 Si) which is a well-known chemical soil stabilizer, (2) bentonite slurry, and (3) silicone oil. The purpose for using these stabilizer materials is to provide an impermeable layer (waterproofing coat) around the gypsiferous soil particles and to fill the voids that may occur in the soil as a result of gypsum dissolution when subjected to water.A series of double oedometer (collapse) tests and direct shear tests were carried out on gypsiferous soils before and after remediation by grouting. The results show that the soil compressibility decreases in the grout-treated gypsiferous soil by reducing volumetric strain. Consequently, the potential for collapsibility decreases too. It was also shown that grouting, increases the shear strength of the soil at confining stresses smaller than 250 kPa by increasing the cohesion combined with a slight reduction in the internal friction angle.INTRODUCTIONGrouting is one of the ground improvement techniques which is often used to improve the loose foundation soils. Bruce (2003) provides step-by-step procedures on using grouting techniques for sealing dams and other water reservoirs. Grouting can be classified as permeation grouting, compaction grouting or hydraulic fracturing. Permeation grouting is an effective way to send the grout into the ground without disturbing the soil structure, i.e. free flow of grout into the soil voids where grouting is accomplished by injection of a fluid-like material that is capable of forming a gel and binding the soil particles together. Grout materials may be solution grouts or suspension grouts. The permeation grouting technique is generally used to minimize ground permeability and to control ground water flow, but it can also be used to strengthen and harden the ground (Raman et al., 2012)."

Jan 1, 2017

"This paper challenges the traditional notion that mine planners need to plan production so as to incur the lowest mining cost. For a given mine configuration, a mine that increases its mining rate will incur increased mining costs. In an environment in which operations are fixated on cost reduction, a proposal that increases costs will not be readily accepted. Such a proposal requires financial justification—the increase in costs might be recuperated by the additional production. This paper evaluates the net present value (NPV) across a range of copper prices for two underground orebodies located at different depths, using a production rate of 300 kt per quarter and a scenario that introduces additional equipment and costs for 450 kt per quarter. The evaluation was based on the changes of NPV for the orebody located at a shallow depth compared with the orebody at a greater depth. Discrete event simulation combined with mixed integer programming was used for analysis. Unlike traditional sensitivity analysis, this study re-optimizes the mine plan for each commodity price at each production rate. The results show that, for the low mining rate at the final copper price, an NPV of A$1530.64 million is achieved, whereas an NPV of A$1537.59 million is achieved at a higher mining rate. Even though pushing mining rates beyond traditional limits may increase mining costs, this option may be beneficial at certain commodity prices, particularly when prices are elevated. IntroductionA traditional notion among mine planners is that mining rates should be planned in such a way as to reduce production costs. This concept generally relates to the mining rate that results in the highest machine utilization for the selected equipment. Although an existing mine may consider increasing its mining rate by adding equipment to a given mine configuration, this typically increases mining costs. In an environment in which operations are fixated on cost reduction, a proposal to increase costs may be very difficult to argue for. The value of the mine plan is well recognized as varying with the mining rate for a given orebody (Smith, 1997)"

Jan 1, 2017

By T. BoBo

"In the 1990’s metal prices were trending in a long term decline and the usual cost cutting exercises were adopted. Cuts to research and development made innovation difficult and the outlook was grim. But necessity is a strong motivator, and these conditions spurred the Mine-to-Mill movement which optimised across organisational “silos” and utilising new technology tools, innovative software and increased computing power. Many applications of Mine-to-Mill exploited the fact that comminution is usually the site processing bottleneck, and that blasting is more efficient at breaking rock than grinding. This approach sought to: • Understand and characterise rock breakage from mining to the mill • Develop models and simulators for blast design, fragmentation, crusher and mill circuits • Develop tools to measure in real time the particle size distribution of rocks on conveyors and Run of Mine (ROM) muckpiles • Ensure effective communication across the silos between geologists, blast design engineers, mining engineers and metallurgists These methods and the technology tools available in the 1990s were widely adopted at numerous sites around the world with significant benefits. SAG mill throughput increases of 10-20% were common. This was the advance the industry desperately needed. It was low capital cost; it was obvious. It was here to stay. Except, in too many cases, it didn’t. In the minerals price boom of the 2000’s Mine-to-Mill was no longer necessary to “survive”, even though it was still good business. Some operations, including some of the early success stories, slipped back to old habits of optimising within organisation silos instead of across them. Fortunately, other operators still embraced Mine-to-Mill and developed new techniques. Mine to Mill was still successful, but it not as widely adopted as we had expected it to be.Now the boom has ended and operations are again under cost pressure. Since these are the same circumstances that created Mine-to-Mill, it seems time to achieve the wider adoption. Since the 1990s many new or advanced technology tools are available for Mine-to-Mill projects: blast-hole sensors, new explosives formulations, new blasting techniques and modelling, ore tracking devices, improved image analysis to determine size, texture and colour of coarse ore, grade sensors, whole-stream simulation tools, even more powerful computing hardware and data analysis software. If we could achieve so much in the 1990’s, how much more can we achieve today when we have the same imperative, the same potential and a larger number of high technology tools?"

Jan 1, 2017

By T. P. Letsebe, L. Magwaza, T. A. Sethu, S. M. Rupprecht

"The authors were tasked with introducing safe and efficient drilling and blasting practices to a small artisanal mine located in the Eastern Province of Rwanda. The current mining methods employed at the mine are inefficient as rockbreaking is conducted utilizing hammer and chisel, or by drilling and blasting with no understanding of the rockbreaking process. The paper discusses the drill and blast cycle, as well as the associated activities necessary to support the mining cycle. The challenges and critical aspects in implementing drill and blast activities are discussed, as well as some of the technical aspects, i.e.drilling pattern, initiation and timing of the blast, and handling of misfires. Ancillary issues such as ventilation, training, communication, effective planning, logistics, and human relations are also examined. IntroductionRockbreaking accounts for 60% of the direct mining costs in small-scale mining. When drilling and blasting practices are poorly implemented the process becomes inefficient and dangerous. The introduction of a proven drill and blast system to artisanal mining offers a number of challenges, with safety being considered the backbone of any sustainable operation and a standard that cannot be compromised.The paper is based on a project at an artisanal cassiterite (tin) mining operation in a small village in the Kayonza District in the Eastern Province of Rwanda. A drill and blast programme utilizing industry best practice was introduced at an underground development end; a raise developed at an inclination of 23 degrees. The underground working is accessed via a 17 degree incline shaft with an operating length of 105 m and serviced by a single drum winder. The winder operates at a speed of about 3.5 m/min hoisting two 1 t wagons from the bottom of the incline to surface for tipping (Figure 1)."

Jan 1, 2017