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By William Gleason

"In 2008, the world quietly hit a milestone that did not receive much fanfare or publicity, but it hit one that marked a turning point in the history of mankind and the way the planet is and will be inhabited in the future. It was in 2008 that, for the first time, more than half of the world’s population, an estimated 3.3 billion people, resided in urban areas. The world’s population had moved from being a majority of agricultural and rural society to an urban society. Andrew Whalley, deputy chairman for Grimshaw Architects, said this change in where and how people live will continue to increase urban populations and, by 2050, it is estimated that more than 6.4 billion people will live in urban areas. To meet these needs, Whalley, speaking during the keynote session of the 2105 Cutting Edge conference in Denver, CO, said future cities will have to be reimagined and redesigned to be more efficient. A large part of that design will be in the infrastructure of those cities.At the Fourth Annual Cutting Edge Conference, Sept. 21-23, much of the discussion was around how the tunneling and underground construction industries improve cities in a number of ways, including societal, environmental and in overall efficiency. There was talk, not just about the improvements that society enjoys at the completion of such projects, but also about the challenges faced by the owner and contractor in the construction of such projects.The conference, titled Urban Tunneling, brought together some of the top minds in the industry to not only discuss the issues faced today, but also those that will shape the industry in the future.For his part, Whalley spoke about a number of projects around the world and how they have impacted society in ways beyond simply providing a better means for transportation of people or water. Notably, he spoke about the Crossrails project in London.London is a city of 8.6 million people and a history that dates back nearly 2,000 years. It is a great example of the challenges the industry faces in bringing modern efficiencies to an established city.Running more than 100 km (62 miles) with more than 42 km (26 miles) of tunnels passing through 40 stations, including 10 new stations, the project is the largest in Europe and will transform the way people move about the city."

Jan 1, 2015

By Cyril W. Draffin

Introduction Americans have long been proud of their ability to recognize a need, define it, and then develop workable solutions. The U.S. underground coal conversion (UCC) program is a good example. In 1974, the need to utilize coal resources that could not be mined was recognized, and Congress gave the Bureau of Mines $1.7 million to gain a better understanding of the UCC technology and its potential applications. Since 1977 when the U.S. Dept. of Energy (DOE) was formed, workable technologies have been tested, the market for underground coal conversion has been recognized, and extensive industrial interest has developed. Perspective The U.S. is blessed with extensive coal resources. Coal supplies about 19% of our national energy needs, yet recently promulgated environmental standards and increasing mining costs have limited direct coal use. UCC potentially can use coal that is economically or technically infeasible to mine because the coal is too thick, deep, dirty, high-ash, and wet; has an excessive angle of dip; or has unpredictable and poor over-burden characteristics that make mining unsafe. Coal can be gasified underground or perhaps liquefied below the earth's surface. Underground coal gasification (UCG) is the prime element of the underground coal conversion program. It enables coal to be converted to gas underground. UCC Coal Resource Recent data show a total U.S. coal resource base of about 6.4 trillion tons. Of this, 1.8 trillion tons might be accessed by UCC, including some minable coal. A commercial UCC process could quadruple roughly the 450 billion tons of U.S. coal reserves, which presently are recoverable by mining. The U.S. coal resources are divided into five sectors. The two most eastern sectors contain relatively thin, swelling bituminous coals that are difficult to gasify. Of the two, the bituminous coal in the Eastern Province is the most difficult because it is deeper than in the Interior Province and the terrain is more rugged. The sector along the Rocky Mountains contains huge amounts of thick low-rank coals that are the most promising for early development by UCC. The coal has good continuity and the terrain is suitable. Low-rank coals such as subbituminous and lignite are highly reactive and tend to shrink and disaggregate when heated. Most of the coal in the Pacific

Jan 9, 1980

By S. J. Czehura

The Butte quartz monzonite is the host rock for early porphyry copper style mineralization and subsequent mesothermal-epithermal, enargite-gold style veining that transects the district. The east-west Anaconda system veins: including the Emma, Anaconda-Original, Syndicate, Badger-State, and Alice-Rainbow veins, each with strike lengths of over 12,000 feet and vertical continuity in excess of 4,500 feet, along with mining widths from 5 to 50 feet, are unique to Butte. The incredible dimensions and tenor of these veins have yet to be surpassed and Butte is still touted as being the “Richest Hill on Earth.” The Anaconda Company was formed in 1895 to develop underground copper reserves in the district and eventually consolidated all operations under one company. Production from underground mining continued through the mid-1970’s. Open pit mining in the Berkeley pit began in 1955 with the extraction of supergene ores, along with the remnants of deeper horsetail zones inaccessible to the underground mines. Operations in the Berkeley pit were suspended in mid-1982 because of the rejection of high arsenic, copper concentrates at major smelting facilities throughout the world. Bulk mining of porphyry style mineralization in the Continental pit began in January 1980 to supplement copper production while stripping “C” pushback in the Berkeley pit. The Anaconda Smelter closed in September 1980. A molybdenum circuit was added to the Butte Concentrator in 1981 and mining continued in the Continental pit to produce copper and molybdenum concentrates. A blend of Berkeley and Continental ores were processed to dilute penalty metals in the final concentrates marketed to toll smelters. All mining operations ceased mid-1983 when Anaconda closed the Continental pit and offered the property for sale. Montana Resources reopened the Continental pit in June 1986 to produce copper and molybdenum concentrates. Operations were suspended mid-2000, because of exorbitantly high power prices, but resumed in November 2003 with rising metal prices. Bulk open pit mining of the early porphyry style mineralization has sustained production over the past twenty years and mining is expected to continue in the Continental pit with favorable prices. Future mining will eventually develop a 500 million-ton geologic resource incorporated in a supergene enrichment blanket situated between the Berkeley pit and the Continental pit. Grades in this enrichment blanket range in the neighborhood of 0.5% copper.

Jan 1, 2006

By William Gleason

"On April 4, 2017, the largest earth-pressure-balance tunnel boring machine (TBM) ever built finally completed its 2.7-km (1.7-mile) journey beneath Seattle, WA, nearly four years after it began boring the SR-99 tunnel that will eventually replace Seattle’s Alaskan Way Viaduct.By now, readers of this publication are familiar with the ordeal that began on Dec. 6, 2016, when the machine overheated and shut down. Until that point, the TBM, named Bertha, had grabbed headlines for being the largest in the world with a 17-m (57.6-ft) diameter face. But after the incident that stopped the progress, the headlines suddenly stopped focusing on the engineering marvel of the tunnel project and began asking what went wrong. Just 300 m (1,000 ft) into the tunneling journey the machine was stuck, unable to move forward and impossible to retract.The project’s contractors, Seattle Tunnel Partners,(STP JV) – a design-build consortium comprised of Dragados - USA, Tutor Perini, say a 20-cm (8-in.) steel well-casing buried in the path of the machine caused the problems that halted the TBM. The Washington State Department of Transportation (WSDOT) disagrees, and the issue over who ultimately pays for the repairs is now with the courts.Eventually, engineers determined the best way to get Bertha back to work was to dig a 36-m (120-ft), deep concrete lined shaft pit in front of the machine and lift the 1.8-kt (2,000-st) cutterhead out for repairs. The repairs took two years but eventually the machine resumed mining in January 2016. In April 2017, the TBM finished building the tunnel, breaking through at the tunnel’s north end near Seattle’s Space Needle, with little to no ground surface movement measured on the entire alignment.While the cause of the TBM breakdown and who should pay for repairs was debated above ground, the work below ground continued. David Sowers, deputy program administrator, engineering and program management with WSDOT, told Mining Engineering during the 2017 Cutting Edge Conference in Seattle that in addition to the technical marvel that the tunnel became, the fact the parties involved were able to work through the unprecedented challenge is a great example of how to overcome challenges on massive projects."

Jan 12, 2017

By Stephen Leslie

A multimedia tutorial illustrating the geology and mineral resources of Gulf Coast salt domes has been developed as a pilot project for a course on Texas geology and resources at the University of Texas at Austin. The principal objective of the tutorial is to educate students with limited background in geological sciences about a major local geologic feature and its geologic and economic significance. The content of the tutorial represents materials generally covered in 4 hours of lecture and laboratory instruction. These materials are presented using a flattened hierarchy of topics, each accessible from a central menu, including salt-dome history, salt tectonics, anhydrite and calcite cap rocks, mineral resources, and geophysical character. Information concerning salt-dome geology is communicated using interactive software; sound and animation are used throughout to reinforce key concepts. Visual aspects include field views, hand specimens, and photomicrographs that were digitized from slides. Illustrations were created using a graphics program or scanned from available archives. Each visual element has an accompanying explanation, which in many cases is linked to other parts of the tutorial. Navigation is accomplished using hyper-links or a set of unobtrusive buttons. The program includes a glossary of the major minerals associated with salt domes available as a pull-down menu and digitized video footage of the Gulf of Mexico sea floor to illustrate the active environment of the formation of the salt-dome cap rock. In addition, a group of interactive questions is presented to assess the level of mastery. Technical aspects were accomplished using a Power Macintosh ? 7100AV and a variety of commercial software. Actual linking of information was done with Authorware? 3.0 Educational, an icon-based, multimedia authoring program. Graphic illustration was done with Adobe Illustrator ?; image processing was accomplished using Adobe Photoshop?; video was cap¬tured, edited, and compressed using Adobe Premiere ?. The interactive nature of the project was designed with regard to future means of dissemination, especially on the World Wide Web. This tutorial serves as a template to build additional modules covering the entire course. A modular approach to the coding of the project and extensive self-documentation should ensure the viabil¬ity of the project by allowing other workers with moder¬ate computer skills to complete other lessons within the curriculum framework.

Jan 1, 1999

By Kwadwo O. Adutwum, George M. Potter, Todd J. Harvey

Ashanti Goldfields Company (AGC) is the largest gold mining company in Western Africa. Their jewel operation is a highly inte­grated set of mines and plants situated in Obuasi, in the Ashanti region of Ghana. This operation is responsible for mining 5.5 mil­lion tonnes of ore at 4.6 git annually from a series of open pits and underground mines. The sulphide ore is highly refractory with untreated gold recoveries in the range of 30-40%. The milling facilities are quite varied and include an oxide treatment plant (OTP), a heap leach, a sulphide roaster circuit (PTP), a tailings treahnent plant (TIP) and the most recent, a sulphide treatment plant (STP). The sulphide treatment plant, the world's largest stirred tank biooxidation circuit, was commissioned in early 1994. This plant treats approximately 800 tonnes of refractory gold con­centrate per day with overall recoveries of75-85o/o(ll, Biooxidation was selected for its ease of operation and less exotic maintenance requirements as well as its ability to achieve comparable gold recoveries. Unfortunately, despite stirred tank biooxidation being one of the lower cost methods of treating refractory ores, it is still too costly for low grade refractory materials. GeoBiotics, Inc. has developed a proprietary technology that allows the economic treatment of lower grade refractory gold ores through heap biooxidation. AGC has a significant resource of low grade refractory sulphide ore. This ore is too low grade to warrant BIOX™ sulphide treatment and its oxide plant recov­ery is low at 35-45%. GeoBiotics and AGC have completed pre­liminary testing and economic analysis of heap biooxidation for this material. Additionally, the feasibility of producing a gravity concentrate from the tailings of the sulphide treatment plant has been explored. This concentrate would be coated onto the low grade sulphide and biooxidized concurrently. Gold recoveries for the biooxidized low grade sulphide have increased to 75% and the gravity recovery from the STP tailings has achieved 80% uti­lizing a Falcon concentrator. The preliminary economics of the proposed full scale plant are very favourable with a pre-tax IRR of 233% and a payback period of 6 months. This paper details the biooxidation testing, flow sheet devel­opment and preliminary economics of employing heap biooxi­dation for low grade refractory ore treatment at Ashanti Goldfields' Obuasi operation.

Jan 1, 1998

By Doug Williams, Kent J. McGrew, Dr. Jack Murphy

For some time, selenate has eluded water treatment by conven­tional means such as precipitation and clarification. This has been due to the nearly universal solubility of all oxidized sele­nium species, the low levels required for surface discharge, and almost identical chemical behavior to all forms of oxidized sul­fur. The last characteristic makes selenate particularly difficult to deal with in acid mine waters where sulfate can easily be on the order of one gram per liter. Approaches such as ion exchange or reverse osmosis often tum into total demineraliza­tion schemes for the sake of removing a "trace" selenium species that may be present on the order of 108 sulfate/selenate. Reduction of selenate to selenite or elemental selenium by chemical means has been relatively unresearched. Some limit­ed success has been reported (Murphy, 1988) using ferrous hydroxide as a reducing agent; however, the reaction time as reported is too slow to be useful in a practical plant design. One significant difference between selenate and sulfate is that selenate has been discovered to be reducible in acid solution by elemental iron when catalyzed by soluble copper. Since the +4 or zero valence states of selenium are rather easy to deal with by coprecipitation with iron salts or alum, the elemental iron dis­solved in the process contributes to the overall process by pro­viding sludge upon raising the pH that can be separated by con­ventional clarification. Commercial production of selenium using copper catalyzed reduction of selenite to selenium metal via sulfur dioxide is well established, however, at the < 5 ppb requirements for surface discharge this reduction process equili­brates at levels far too high to be useful to low level selenate removal criteria. The reaction reported herein uses elemental iron for reduction of selenate with copper acting either as a cata­lyst or as a coprecipitation agent. This method, employed with conventional sludging, clarification and filtration can provide an economic method to achieve non detectable selenium (< 0.005 mg/I). Iron consumptions for the conditions tested have been on the order of 3 lbs/1,000 gallons treated. This represents as little as $0.06 or $2.00 per 1,000 gallons based on using a scrap source of iron or the purchase of electrolytic iron for the process.

Jan 1, 1996

By Cyril W. Draffin

Introduction Americans have long been proud of their ability to recognize a need, define it, and then develop workable solutions. The U.S. underground coal conversion (UCC) program is a good example. In 1974, the need to utilize coal resources that could not be mined was recognized, and Congress gave the Bureau of Mines $1.7 million to gain a better understanding of the UCC technology and its potential applications. Since 1977 when the U.S. Dept. of Energy (DOE) was formed, workable technologies have been tested, the market for underground coal conversion has been recognized, and extensive industrial interest has developed. Perspective The U.S. is blessed with extensive coal resources. Coal supplies about 19% of our national energy needs, yet recently promulgated environmental standards and increasing mining costs have limited direct coal use. UCC potentially can use coal that is economically or technically infeasible to mine because the coal is too thick, deep, dirty, high-ash, and wet; has an excessive angle of dip; or has unpredictable and poor overburden characteristics that make mining unsafe. Coal can be gasified underground or perhaps liquefied below the earth’s surface. Underground coal gasification (UCG) is the prime element of the underground coal conversion program. It enables coal to be converted to gas underground. UCC Coal Resource Recent data show a total U.S. coal resource base of about 6.4 trillion tons. Of this, 1.8 trillion tons might be accessed by UCC, including some minable coal. A commercial UCC process could quadruple roughly the 450 billion tons of U.S. coal reserves, which presently are recoverable by mining. The U.S. coal resources are divided into five sectors. The two most eastern sectors contain relatively thin, swelling bituminous coals that are difficult to gasify. Of the two, the bituminous coal in the Eastern Province is the most difficult because it is deeper than , in the Interior Province and the terrain is more rugged. The sector along the Rocky Mountains contains huge amounts of thick low-rank coals that are the most promising for early development by UCC. The coal has good continuity and the terrain is suitable. Low-rank coals such as subbituminous and lignite are highly reactive and tend to shrink and disaggregate when heated. Most of the coal in the Pacific

Jan 1, 1980

By Z. Kowalski

In the defunct Gorka quarry, there are situated both a waste disposal site of an area of 6.7 ha and containing approximately 1 million ton of high aluminium wastes and a pond of an area of roughly 3 ha, up to 15 m deep, with about 400000 m3 of polluted water. The reservoir is fed with the filtrates arriving from the disposal site mentioned, at a rate ~130 m3/day. An alarming phenomenon is the subsidence of the pond bottom and the infiltration of solutions into the Triassic and Jurassic water resources, assessed at ~40 m3/ day. The basic problem related to the effluents in the Gorka pond is their high alkalinity (pH 12-13.4) and variable pollutants whose level increases appreciably with the pond?s depth. The solution offered provides pumping out and treatment of about 500000m3 of effluents retained in the Gorka reservoir in an about yearly time. Effluents would be treated in a reverse osmosis plant, according to a process so far verified on a quarter-commercial scale. The treatment process by-product would be discharged into the Ropa stream. The brine solution (containing about 25% NaCl), whose amount may reach up to100000m3, might be supplied to the Debiensko Mine which produces technological salt from salt-bearing mine waters. Similarly, there would be provided an intake of underflows of infiltrates (by means of deep wells, and possibly girdling ditches) arriving from the high aluminum site, and such effluents would be subject to ongoing dilution in water courses. It would be possible to test the system of underflows reaching the Gorka reservoir by means of geochemical identification, previously performed. The next stage following the pumping, would be the utilization of approx. 50 000m3 of bottom slurry. Highly alkaline slurries would be utilized in the production of self-solidifying mixtures. Slurries would be pumped out and temporarily dumped on drying beds, located nearby the Gorka reservoir, and later on used successively for making self-solidifying mixtures in a plant similar to a concrete making seat. Those mixtures would be applied for sealing the bottom of the Gorka reservoir and the necessary part of the edges of the defunct quarry. An alternative solution is the utilization of slurries in the facilities of the Trzebionka Mining Works and the application of self-solidifying mixtures, made in the Trzebionka Works, for sealing the reservoir bottom.

Jan 1, 2006

By Joseph Fernandez, Luminita Calin

INTRODUCTION The Tunnel and Reservoir Plan (TARP), commonly known as the Deep Tunnel Project, is the daring and unique concept developed by the Metropolitan Water Reclamation District of Greater Chicago (MWRDGC) as a two-phase solution to the problem of waterway pollution and basement flooding due to sewer system backups created by the combined sewer over-flows (CSOs). Phase I, primarily for pollution control, includes the tunnels, shafts, connecting sewers and structures. These Phase I facilities intercept, during wet weather events, the CSOs within the MWRDGC TARP service area. The tunnels and shafts provide temporary storage of the CSOs for subsequent treatment at Water Reclamation Plants (WRP) before discharging into waterways. Ultimately, upon completion of Phase II, the tunnels, located from 45 m(150 ft) to 100 m (330 ft) below ground, will convey the CSOs to reservoirs for storage until WRP capacity is available. Phase II, which includes the reservoirs, pumping facilities, additional tunnels and shafts, has been planned to serve a dual purpose. The reservoirs provide storage of excess flows during wet weather events, those beyond the WRP capacity. Following the wet weather events, the stored flows can be conveyed, at controlled rates, to the WRPs for full treatment. These excess flows previously would have resulted in both overflows into the waterways and eventually backups in the local sewer systems. The storage capacity of the reservoirs also will promote the full utilization of the conveyance capacity of the tunnels and shafts, which in turn results in optimizing the conveyance capacity of the existing sewer systems. Therefore pollution control is further enhanced by allowing for capture of a greater volume of CSOs and greatly minimizing backups in the local sewer systems. The TARP project consists of four systems: O’Hare, Mainstream, Des Plaines, and Calumet, covering a total area of 971 km2 (375 mi2) and serving the City of Chicago and 51neighboring communities as shown in Figure1. With only two contracts left before completion of the Phase I tunnels, the 12.7km (7.9mi) Torrence Avenue leg, to be awarded in 1997, and the 12.4km (7.7mi) Little Calumet leg unfunded as of yet, both part of the Calumet System, the pollution control phase of the project is now 86% complete (Table 1).

Jan 1, 1997

By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By Michael Drozd

Introduction Flotation is a recovery process that utilizes natural or chemically-introduced hydrophobicity of minerals to separate ore from gangue by attaching the ore particles to bubbles. Flotation has been known since 1906, when it was patented by E. L. Sulman, H. G. K. Pickard and John Ballot. Flotation&apos;s first documented plant use was in 1911, when it was used on gravity tailings from a zinc plant. Cyanide leaching was first patented in 1887. Most early precious metals flotation was linked to lead or copper concentrates. Most precious metal flotation is linked to sulfide recovery and utilizes such reagents as xanthates, dithiophosphates and thiocarbanilides. But the involvement of flotation in precious metals recovery does not have to be limited to active flotation of the gold and silver bearing minerals. Flotation can be used to remove interfering minerals, such as carbon, or cyanicides, such as copper, arsenic bearing minerals or other cyanide complexing minerals. Cyanide Leaching The use of cyanide to recover precious metals is the backbone of the industry. The various recovery processes that use cyanide are: 1. Leaching-Counter Current Decantation(CCD) Merrill-Crowe 2. Carbon-in-pulp (CIP) 3. Carbon-in-leach (CIL) 4. Vat leaching 5. Heap leaching These processes are selected depending on ore grade, ore leaching characteristics, reserve size and capital limitations. Ore grade is used as the first criteria to determine the type of cyanide plant required. However, the reserve size can determine whether a ground ore alternative is used. The ore leaching characteristics may indicate that the higher recovery, associated with a process that employs grinding, is the only economic recovery alternative. There are some extremely large leaching plants with grades that would be non-economic for some existing heap leach plants. Many factors, such as strip ratios, metal recovery and reserve size, determine what the ultimate plant design will be. The exact processing scheme that is used, is a multifaceted decision matrix, which gives the process metallurgist some job security. Precious Metal Recovery Using Flotation Flotation recovery can conceivably be used on any ore that has been ground. Many operations utilize gravity separation on the ore to remove native gold from the ore prior to sending it to cyanide leaching. On certain ores, where the gold is in visible nuggets or particles, this processing alternative is advantageous because the native gold cyanide recovery is

Jan 1, 1993

By Peter P. Harvey

The Nonel Primadet is a non-electric blasting initiation device. The system incorporates the safety advantages of a cap and fuse system with the precision blasting afforded by electrical blasting caps. Each unit is made up of a delay blasting cap factory-crimped to a Nonel® shock tube lead line. The delay elements in the blasting caps are manufactured in both millisecond and long period series. There are 16 periods available in both series - period 0 through period 15. The units are manufactured in 12&apos;, 16&apos;, and 20&apos; lead lengths. The hollow shock tube is made of surlyn plastic and it is coated on the inside with a reactive material. The powder train has a density of only 1/10 grain per linear foot of tube. A typical blast round is timed precisely the same way it would be with electrical caps. After the round has been timed and loaded, the tube leads are gathered in convenient bunches and tied into a detonating cord trunkline. The Nonel® tube bunches are initiated by the output of an 18 grain per foot detonating cord trunkline. Once initiated, the Nonel® shock tube will reliably propagate at 1800 meters per second. The energy from this shock front initiates the delay elements in each delay cap. The shock tube is inert to electrical energies, including; stray current from AC/DC power sources, as well as, electrostatic discharges. The tube is also inert to initiation by impact. The system has proven to be particularly advantageous in both tunnel and shaft blasting applications. The major advantages in using the Nonel® system in this type of work are: 1) faster and easier tie-in techniques through the use of bunch-blasting. 2) electrical power can be left on during loading and tying-in. a. electrical lights can be turned on at the working face - better lighting facilitates loading and tying-in. b. electrical pumps can continue to function throughout loading and priming procedures (this is particularly advantageous in shaft work with water problems). 3) down-time due to approaching electrical storms is minimized. 4) increased overall efficiency in the blasting cycle: a. it is easier to train a man in the proper use of the system. b. fewer missed holes per round. c. less time required for tie-in. 5) safety-advantage--The system is not susceptible to involuntary initiation from extraneous electrical energies.

Jan 1, 1979

By Richard L. Bullock, Bruce Kennedy

Vince Lombardi once said, "Practice doesn&apos;t make perfect, perfect practice makes perfect." When it comes to building a mine that will operate at the optimum level for the set of geologic conditions from which it was developed, Lombardi&apos;s remark might be paraphrased to describe the problem: planning won&apos;t guarantee the best possible mine operation unless it is the best possible mine planning. Any sacrifice in the best possible mine planning introduces the risk that the end results may not reach the optimum mine operation desired. This section addresses many of the factors to be considered in the initial phase of mine planning. These factors have the determining influence on the mining method, the size of the operation, the size of the mine openings, the mine productivity, the mine cost, and, eventually, the economic parameters used to determine whether or not the mineral reserve even should be developed. A little-known fact, even within the metal-mining community, is that room-and-pillar mining accounts for most of the underground mining in the united States. According to a 1973 study on noncoal mining (Anon., 1974), more than 76% of the producing mines [of over 1089 t/d (1200 stpd) capacity] produced approximately 70 000 000 t (77,000,000 st) or 60% of the nation&apos;s underground tonnage of material by room-and-pillar mining. That same year, 96.8% of the nation&apos;s under- ground coal mines produced 262 950 000 t (289,911,000 st) of coal extracted from room-and-pillar mines (Anon., 1976). Thus, nearly 333 000 000 t (367,000,000 st) of the United States&apos; raw material is produced from mines using some form of the room-and-pillar mining system. Because approximately 90% of all mining in the United States is done by some variation of room-and- pillar mining, it is appropriate to give special emphasis to the effects of the various elements of mine planning on room-and-pillar mining. The relationship of these elements to other mining methods will become apparent as the elements are described in later sections herein. TECHNICAL INFORMATION NEEDED FOR PRELIMINARY MINE PLANNING Assuming that the reserve to be mined has been delineated with diamond-drill holes, the items listed in the following paragraphs need to be established with respect to mine planning for the mineralized material. Geologic and Mineralogic Information The geologic and mineralogic information needed includes the following: 1) The size (length, width, and thickness) of the areas to be mined within the overall area to be considered, including multiple areas, zones, or seams. 2) The dip or plunge of each mineralized zone, area, or seam, noting the maximum depth to be mined. 3) The continuity or discontinuity within each of the mineralized zones. 4) Any swelling or narrowing of each mineralized zone. 5) The sharpness between the grades of mineralized zones within the material considered economically minable. 6) The sharpness between the ore and waste cutoff, including whether this cutoff can be determined by observation or must be determined by assay or some special tool; whether this cutoff also serves as a natural parting resulting in little or no dilution, or whether the break between ore and waste must be induced entirely by the mining method; and whether or not the mineralized zone beyond (above or below) the existing cutoff represents submarginal economic value that may be- come economical at a later time. *7) The distribution of various valuable minerals making up each of the minable areas. 8) The distribution of the various deleterious minerals that may be harmful in processing the valuable mineral. 9) Whether or not the identified valuable minerals are interlocked with other fine-grained mineral or waste material. 10) The presence of alteration zones in both the mineralized and the waste zones. Structural Information (Physical and Chemical) The needed structural information includes the following: * 1 ) The depth of cover. 2) A detailed description of the cover including: the type of cover; * the structural features in relation to the mineralized zone; * the structural features in relation to the proposed mine development; and * the presence of and information about water, gas, or oil that may be encountered. 3) The structure of the host rock (back, floor, hanging wall, footwall, etc.), including: * the type of rock; * the approximate strength or range of strengths; * any noted weakening structures; * any noted zones of inherent high stress; noted zones of alteration; the porosity and permeability; * the presence of any swelling- clay or shale interbedding; the rock quality designation (RQD) throughout the various zones in and around all of the mineralized area to be mined out; the temperature of the zones proposed for mining; and the acid generating nature of the host rock. 4) The structure of the mineralized material, including all of the factors in item 3 plus: * the tendency of the mineral to change character after being broken, i.e., oxidizing, degenerating to all fines, recompacting into a solid mass, becoming fluid, etc.; * the siliceous content of the ore; the fibrous content of the ore; and the acid generating nature of the ore. Economic Information The needed economic information includes: *1) The tons of the mineral reserve at various

Jan 1, 1982

By Deborah McCombe

-Canadian Regulators -Qualified Person -Professional Association -Resources & Reserves -Foreign Codes -Disclosure of Historical Resources -Issues in Application of Resources & Reserves -Industry Best Practices developed by CIM -Technical Report and Reserves and Resources -Summary

Jan 1, 2003

By Joyce A. Ober

Minerals for electronic and optical uses divide easily into two sections: 1) quartz and 2) minerals other than quartz. QUARTZ The properties of quartz crystal that make it useful for radio communications were discovered in 1918. Since that time, an in¬dustry for the mining and processing of natural quartz crystal has grown, matured, and been almost entirely replaced by new tech¬nology. The new technology still involves quartz crystal, but ma¬terial that is grown rather than mined. An economic summary of the commercial growing of quartz crystals has a place in a handbook directed to the mineral engi¬neering industry because quartz crystals have long been an impor¬tant commercial mineral, and the raw material for cultured quartz - ¬that is to say, quartz crystals grown through the ingenuity of man - is still natural quartz. Nearly all the natural crystals that have been used for elec¬tronics and optics came from Brazil. The larger pieces which met rigorous standards of quality were used for electronic and, to a lesser extent, optical components. Smaller pieces and fragments were used for vitreous silica. The need for high quality material in quantity led to US government sponsored research and exploration programs in the 1940s. No deposits meeting the very rigid requirements for electronic-grade quartz were found, but other projects resulted in the development of a process for the factory growth of beautiful crystals of prescribed shape, size, and quality. Domestic deposits of appropriate quality were identified to use as raw materials for the quartz culturing process. The development of the cultured quartz crystal illustrates the success that technology can have in adapting a product of the mine to increasingly sophisticated uses. A remarkable achievement per¬haps, but foreshadowed by experiments by Giorgio Spezia (1908), an Italian geologist studying the relative effects of temperature and alkaline environment on the solubility of quartz. Modem radio equipment is most often controlled as to fre¬quency by the presence in the circuit of a separately added crystal¬ - the 1918 discovery responsible for the existence and growth of the quartz industry. The crystal is quartz, but this component is a carefully oriented and prepared slice from a crystal, but not a crystal as recognized by a rock hound or seen in a museum. How quartz operates to control frequencies is not a proper subject for a handbook on industrial minerals, and references should be consulted (Cady, 1964, Mason, 1964). Quartz belongs to a class of materials called dielectrics: those that do not conduct an electric current but permit electric fields to exist and act across them. Quartz shows the piezoelectric effect, which means that when a quartz plate is mechanically deformed against its natural stiffness, one of its surfaces becomes negatively charged, the other positively charged. When the plate is released quickly from the stress, the charges disappear as the plate regains its original shape, but because of mechanical momentum the plate deforms in the opposite direction (to a lesser amount) and the surfaces correspondingly become charged in the opposite direction. By thinly coating the two surfaces with metal and attaching flexible wires, these charges can be brought into an electronic circuit. If the surfaces are suddenly electrically charged by movement of current through the wires, the converse piezoelectric effect occurs and the plate deforms. Carry the thought further and it is realized that an alternating current flowing through the wires responds to the mechanical oscillation. By controlling the thickness of the plate, its mechanical vibration frequency can be varied through a wide range. One type of quartz plate, the AT-cut, has a precisely defined orientation with respect to the crystallographic axes of the crystal and vibrates on a microscopic scale much as a book would deform when placed flat on a table and the top cover moved parallel back and forth with the hand. At least 17 other orientations have been studied, some of which have preferred uses in various applications (Cady, 1964). The quartz crystal industry is composed of three main segments (excluding fused quartz and quartz used for optical purposes): 1. Natural electronic-grade quartz crystals. Mined quartz suitable for fabrication into piezoelectric units. Zlobik (1981a) esti¬mated the waste to ore ratio at 1:1000 to 1000 000, depending upon the deposit. 2. Lasca. Mined quartz usable as feedstock in the production of cultured quartz. Approximately 0.63 kg of lasca are required to produce 0.45 kg of cultured quartz. 3. Cultured quartz. Cultured quartz is produced from lasca feed¬stock in a process of crystal growth in an autoclave under conditions of heat, pressure, and time. It is estimated that 0.45 kg of cultured quartz is equivalent to 1.4 to 4.5 kg of natural quartz crystal in yield of commercial quartz suitable for slicing into piezoelectric units. The chronology of the development of the quartz crystal industry both natural and cultured follows: Date Comment 1918 Discovery of the piezoelectric effects of quartz crystal 1921 Application of the piezoelectric effects of quartz crystal in the circuitry of radios 1948 Establishment of a quartz crystal commodity stockpile by the US Government 1952 US consumption of natural quartz crystal at an all time high of 228 t 1958 First commercial production of cultured quartz crystal 1970 Cultured quartz crystal production exceeds imports of nat¬ural quartz crystal 1971 Cultured quartz crystal consumption surpasses natural quartz crystal consumption

Jan 1, 1994

By E. A. Martell, K. S. Sweder

INTRODUCTION Lung cancer in uranium miners has been attributed to alpha irradiation of basal cells of the bronchial epithelium by radon daughters, primarily by 7.7 MeV alphas from polonium-214 (Altshuler et al., 1964). It also has been was observed that for a given cumulative radon progeny exposure, uranium miners who smoke cigarettes have an incidence of lung cancer about 10 times higher than nonsmoking miners (Lundin et al., 1969). It has been pointed out that the large excess of lung cancer deaths among smoking uranium miners is a multiplicative effect (Doll, 1971), which suggested possible synergistic interactions between airborne radon progeny and cigarette smoking. Experimental studies of the complex pattern of interactions between radon progeny, cigarette smoke particles, and the cigarette smoking process are in progress in our laboratory. Preliminary results, reported elsewhere (Martell, 1981), implicate alpha radiation from indoor radon progeny in the etiology of lung cancer in all cigarette smokers. Cigarette smoking produces high concentrations of smoke particles of low mobility and respirable size--particles between 0.5 and 4.0 µm in aerodynamic diameter (see below). The attached fraction of indoor radon progeny is highly dependent on the air concentration of small particles from cigarette smoking and from other combustion sources (Martell, 1981). The size distribution and other properties of radon progeny associated with cigarette smoke particles enhances their effectiveness in the induction of bronchial cancer in man. In this paper we discuss the properties of radon progeny associated with cigarette smoke, the fractionation of radon progeny and modification of their aerosol properties in burning cigarettes, the role of 218Po in these processes, the production of insoluble 214Pb and 212 Pb enriched particles in burning cigarettes, and the consequent differences in the patterns of polonium isotope alpha irradiation in the bronchial epithelium of smokers. EXPERIMENTAL PROCEDURES Experimental methods used in these studies involve the use of small experimental chambers of known radon and radon progeny concentrations in combination with aerosol collection and sizing techniques and sensitive radioactivity detection methods. The use of low-level [ß-] counting for radon progeny determination, providing a measure of 214 Pb plus 214Bi activity, makes it possible to carry out chamber experiments with small radon emanation sources and relatively low air concentrations of radon and radon progeny concentrations in the range from 100 to 1,000 pCi per liter. Thus, for example, in a typical experiment we use a 10 nanocurie 226Ra solution standard in a 10 liter chamber, providing an equilibrium concentration of 1,000 pCi of radon per liter. In small sealed chambers, radon progeny plate out rapidly on the chamber walls, with steady-state concentrations of airborne progeny less than 2 percent of equilibrium levels. This is experimentally convenient because, upon introduction of high concentrations of cigarette smoke particles or small particles from other sources, there is a systematic ingrowth of attached radon progeny, providing a tagged aerosol source of known age and radon progeny composition. In some chamber experiments a 226Ra solution standard of small volume, acidified to O.1N HNO3, was used as the radon emanation source. When used with a bubbler the holdup of radon in an 8 ml volume of 226Ra solution standard at 0.1N HN03 was only 2% of the total radon in the chamber at equilibrium. For experiments with 212Pb-tagged aerosols, we used a dry Ba(228Th) stearate emanation source prepared by the method of Hursh and Lovaas (1967). 226Ra and 222Rn determinations were made by radon gas counting. The 222Rn in a sealed air or water sample is transferred, using helium gas as a carrier, successively through a dry ice cooled trap at -80°C to remove water, through ascarite to remove C02, and through a small activated charcoal trap at -80°C to collect the 222Rn. Subsequently, by heating the charcoal to 400°C, the 222Rn is transferred next to an LN2-cooled capillary trap, and finally into an alphascintillation counting cell of the type described by Lucas (1957). As already stated, radon progeny activities were determined by low-level [ß-] counting, which provides a measure of 214Pb plus 214Bi. The radon progeny samples, collected on efficient Delbag polystyrene micro-fiber filters or on impactor foils, are placed in close, sandwich geometry between two thin-walled flow counters inside shielding anticoincidence counters and a 15 cm thickness of steel shielding. This configuration provides nearly 4II geometry and a low background of only 0.25 to 0.30 cpm. Aluminum absorber was added to provide a combined thickness of absorber and counter wall exceeding 7.0 mg/cm2 to eliminate the variable contribution of 7.7 MeV alphas from 214Po. 212 Pb determinations also were carried out by low-level [ß-] counting, in this case using a combined absorber and counter wall thickness of 9.0 mg/cm2 to eliminate contribution of 8.8 MeV alphas of 21 Po. In each experiment the [ß- ]activity data were corrected for decay to an appropriate common reference time for assessment of activity distributions.

Jan 1, 1981

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today&apos;s DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine&apos;s ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982

By T. Umadevi, A. Pratap Singh, R. Sah, K. Abhishek

"It was possible to beneficiate a low-grade iron ore fines sample from Bellary-Hospet, Karnataka, India assaying 56% Fe, 9.1% SiO2 and 5.2% Al2O3 and produce two concentrates, namely, a sinter-grade fines concentrate (43.7% by weight), assaying 63.6% Fe, 2.6% SiO2 and 1.7% Al2O3, and a pellet-grade concentrate (30.7% by weight) assaying 64.7% Fe, 2.9% SiO2 and 1.7%Al2O3. Rejects (25.6% by weight) assayed 32.5% Fe, 27.9% SiO2 and 15.4% Al2O3. The mineralogical characterization indicated that the sample consisted of 69.5% ironbearing minerals (56% hematite, 7.5% martite and 6% goethite-limonite), 12.5% quartz and 15.2% clay, and the rest were minor minerals. Liberation size was found to be in the size range of 25 to 100 µm. The process flow sheet consists of jigs, spirals, grinding and low- and high-intensity magnetic separation of the deslimed feed. IntroductionJSW Steel Ltd is a 10-Mt/a integrated steel plant. To produce 10 Mt/a hot metal requires nearly 16 to 20 Mt/a iron ore fines. The fast depletion of highgrade iron ore fines, increased industrial demand and an increase in iron ore prices have necessitated the exploitation and use of low-grade iron ore fines in the steel-making process after beneficiation. Thus, the beneficiation of low-grade iron ores by removing the associated gangue minerals and enhancing their grade is essential. The use of these iron ore fines in sinter and pellet making has prompted the undertaking of systematic studies on low-grade iron ores.Finely disseminated low-grade iron ores with siliceous impurities have been upgraded by gravity, magnetic and flotation-based processes (Burt, 1994; Svoboda, 1987; Houot, 1983). The processing of low-grade Indian hematite ore containing high levels of alumina-bearing minerals as impurities present separation challenges and pose special problems in martite and magnetite, in association with quartz and clay as the gangue-forming minerals. Alumina is mainly contributed by clay (kaolinite, montmorillonite, illite, alunite), lateritic material and gibbsite; some alumina occurs as solids in iron oxide minerals, namely, goethite and limonite. Silica is mainly contributed by quartz and the associated clay. achieving a high-grade concentrate (Pradip, 1994; Singh and Mehrotra, 2007). Hematite ore basically consists of varying amounts of hematite, goethite,"

Jan 1, 2015

"A cationic surfactant, tetradecyl tributyl phosphonium chloride (TTPC), was first used as a collector. Flotation characteristics of muscovite using TTPC have been investigated through flotation tests. And dodecyl trimethyl ammonium chloride (DTAC) was tested for comparison. Their adsorption mechanisms on muscovite surface were clarified using zeta potential measurement, DFT calculation and molecular dynamics (MD). The flotation results indicate that the recovery of muscovite decreases when the pH increases in the presence of surfactants. When a maximum of recovery was reached, the dosage of TTPC was 2×10-5mol/L which was half of that of DTAC. Zeta potential measurements confirm that the TTPC displays a stronger collecting power than DTAC and is a superior collector for the muscovite flotation. DFT calculation and MD simulations show that TTPC interacts with the muscovite mainly through electrostatic attraction and hydrogen-bonding. The conclusions drawn from purely theoretical computations matched remarkably well with the experimental results. Keywords: Quaternary phosphonium salts; Flotation collectors; Flotation INTRODUCTION Muscovite mica is widespread in igneous, sedimentary and metamorphic rocks. It has the general composition KAl2(AlSi3O10)(OH)2. Muscovite mica exhibits perfect cleavage along the (001) plane, which exposes interlayer potassium ions[1-3]. In aqueous suspensions, the potassium ions readily enter into solution, leading to the development of a net negative charge on the (001) plane[4-6]. The surface of the muscovite mica is negatively charged at all pH > 1[7]. Due to the negative charge surface over the whole pH range of 2–12, the muscovite mica would not be expected to respond to anionic collectors in the absence of an activator. Conversely, the mineral is readily recovered using cationic collectors. However, until now, various amines and their derivatives (with N atoms as the center) are still the only available cationic collectors which still had their shortcomings such as weak selectivity, low flotation speed. Since few types of cationic collectors could be used, in some sense, the limitations of collectors are increasing significantly and the cationic flotation is facing more and more challenges. Quaternary phosphonium salts, as novel cationic surfactants, are widely used in antimicrobial, catalysis and surface modification field due to their biological activities, hydrogen bonding capability, stability and superior group electrical property[8-10]. There are a number of processes where quaternary phosphonium salt surfactants have shown better properties than quaternary ammonium salts and amines. Although the adsorption of quaternary phosphonium salt surfactants at the bacterium–liquid interface has been studied thoroughly, the case of adsorption in mineral surface has not been investigated thus far."

Jan 1, 2016