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By B. Mather

Discussion by B. Mather I was very glad to see an article on alkali-silica reactivity of concrete aggregate in Mining Engineering. It is a topic that should be reexamined every 45 or 50 years. The last paper (that I am aware of) that examined this topic in Mining Engineering was Rexford (1950), who, at the time, was chief of the Petrography Section, South Pacific Division Laboratory, Corps of Engineers, Los Angeles, CA. The paper was presented at the AIME meeting in Los Angeles in October 1948. Professor Leming's discussion generally reflects the current state-of-the-art; but it provides no references for more detailed information. I wrote a paper titled "How to avoid excessive expansion of concrete due to alkali-aggregate reaction," which was included in the Proceedings of the Second International Conference on Alkali-Aggregate Reactions in Hydroelectric Plants and Dams (Mather, 1995). In it, I emphasized that it was the hydroxide ion (not the alkali ion) that caused the pore fluid in the concrete to dissolve reactive silica, producing alkali-silica gel capable of taking up water, thereby, causing swelling and rupturing of the concrete. Excessive expansion will not occur unless there is a sufficient amount of potentially expansive alkali-silica reaction product and water. However, this will not occur unless there is a sufficient quantity of reactive aggregate and unless there is a mechanism by which the pH of the pore fluid can get well above the normal value of about 12.6. In many cases it is sufficient that the cementitious medium not only contains portland cement but also contains ground granulated blast-furnace slag (GGBFS) or a pozzolan such as fly ash, silica fume, metakaolin, volcanic glass or calcined shale - any of which may have been included for economic reasons. If, however, the person selecting the ingredients for concrete knows that the aggregate is reactive (with no alternate nonreactive aggregate available) and if the person knows that the only available portland cement has a high alkali content (i.e., more than 0.6% Na2Oe; Na2Oe = % Na2O + 0.658 x % K20), then it becomes irresponsible not to establish what reasonable precaution one should take. Professor Leming suggests using one of the following: 30% Class F fly ash, 50% GGBFS or 7% to 10% silica fume. My view is that one should prepare test specimens with the cement and aggregate to be used and then expose these both with and without at least two dosages of each of the various alternative materials (slag, pozzolan and lithium salts), so that the required dosage to control the expansion is established. Then, economic considerations can dictate the final selection. I would point out to the author that GGBFS is not a pozzolan and is not an admixture, but, rather, it is a latent hydraulic cement. An admixture is, by definition, a material other than hydraulic cement. I commend him for suggesting that aggregate producers should become more active in knowing about the reactivity of their products and the precautions needed to permit those products to be safely used in concrete. ? References Mather, B., 1995. "How to avoid excessive expansion of concrete due to alkali-aggregate reaction," Proceedings, Second International Conference on Alkali-Aggregate Reactions in Hydroelectric Plants and Dams, US Committee on Large Dams, Denver, CO, pp. 421-439. Rexford, E.P, 1950, "Some factors in the selection and testing of concrete aggregates for large structures," Mining Engineering, Trans. AIME, Vol. 187, pp. 395-402. Reply by M.L. Lerning Mr. Mather has once again provided a cogent and useful discussion. I am grateful to him for bringing these issues to my attention, as well as to the readers, and I am grateful for his discussion of several important factors in alkali-silica reactivity. Mr. Mather's discussion of the mechanism of alkali-silica reactivity (ASR) notes the critical role of the hydroxide ion. While the alkali hydroxides are highly soluble (reducing the solubility of the other primary hydroxide in pore solution, i.e., calcium hydroxide, and providing hydroxide ions that initiate alkali-silica reactivity), it is, as Mr. Mather quite correctly notes, the hydroxide ion that attacks the silica structure. This point was not clearly stated in my original article. As noted in the text, the requirements for mineral admixtures to mitigate or control ASR vary widely depending on the specific characteristics of the materials

Jan 1, 1998

By Paul Nicholas, Raju Gottumukkala

"India’s economy is expanding with rapid urbanization. Improvement in infrastructure is a keystone to the Indian government’s economic plans, including expansion of transit systems. Most of these systems are designed with elevated and underground sections. New Delhi, a city of 23 million people, suffers from heavy traffic congestion and pollution. The transit (metro) system is seen to be of critical importance. Phase 3 of the Delhi transit system project is 140 kilometers (km), with 40 km of twin-bore tunnels requiring 35 tunnel boring machines (TBMs) for construction. Transit systems are also under design or construction in Chennai, Bangalore, Kolkata, Hyderabad, Jaipur and Mumbai. Most have twin-bore tunnels. At present, 55 TBMs are operating on underground metro construction, and future projects will increase this number. This paper highlights the unique challenges of construction of underground metro systems in highly congested urban environments, using Kolkata and Delhi systems as examples. INTRODUCTION The first transit system in India was constructed in Kolkata (India’s third-largest city by population.) This line is 28 kilometers (km) in length, has underground and at-grade sections, was built in phases, runs north/south, and has 24 stations. This project was sanctioned in 1972, started construction in 1984, and was completed in 1995. The tunnels were constructed primarily by cut and cover methods with some open-face, compressed air shield tunneling. In 2006, the urban transport policy recommended construction of transit rail systems in all urban cities with a population exceeding 2 million. Further, in 2014 the central government stated that it would assist in financing transit systems in all cities with a population of over 1 million. In 2015, the central government approved the development of transit systems in 50 additional cities across India with a budget of $75 billion (see Figure 1). Transit development is just part of a huge expansion in planned rapid transport systems, which include metro, bus, high-speed rail, and monorail systems. Metro systems will include elevated and underground sections with twin-bore tunnels and underground stations. In the Delhi/National Capital Region (NCR), now the largest city in India by population, Delhi Metro Rail Corp (DMRC) was formed in 1995 as a joint venture company between the central government and Delhi state. DMRC operates the Delhi metro system. (see the case studies described later in this paper). DMRC has developed an external function to carry out studies and to offer consultancy to other cities planning transits. Since its inception, DMRC has produced Detailed Project Reports (DPRs) for every metro and monorail project in India. In 2012, DMRC became a partner in the management consultancy for Jakarta Mass Rapid Transit System, its first project overseas. DMRC became a model organization for other cities to follow. These cities have included Chennai, Hyderabad, Kolkata, Bangalore, Jaipur, Mumbai, Ahmedabad, and Nagpur. Other cities (e.g., Lucknow and Patna) are in the metro planning phase. Metro systems are under construction or partially operating in Hyderabad, Bangalore, and Jaipur."

Jan 1, 2016

By Christian F. III Baiz

Introduction A new mining agenda is before us. Born of government policies and world economies, the years ahead will show reduced growth. Metals, mining, and most other natural resource industries are inflation dependant and capital intensive. For these, a long period of contraction of outmoded and costly capacity will evolve. In some cases, this will be supplanted by new, lower cost capacity. But this will occur only if the cost of production - including cash operating and debt service costs - comes in at or below current market conditions. The industrial outlook for metals mining can be characterized as bearish. A US bank economist said mining companies planning their five-year forecasts should take current market conditions and straight-line them to the end of the decade. With low growth and low inflation, the real price of metals will remain flat or decline slightly. As a result, capital funding will be problematic. The low-cost producer will succeed. That means only the producer with low cash costs and low principal and interest costs per unit of production, will succeed. Gaining a perspective One must look back 30 years to see the evolution of the capital markets, for metal mine financing. This period can be divided into two economic periods. The 1950s were a time of stable economics. Mining companies earned and reported good profits. They had good corporate creditworthiness. In essence, mining companies of the 1950s were highly self-reliant. When equity infusions were needed, they were easily obtained. Financial management kept long-term debt low. Fixed rate, long-term financing was readily available in the long-term bond markets. The commercial banks of the 1950s were rarely used by the mining industry. When necessary, mining companies went to the banks for short-term financing - the kind that was needed for working capital requirements. Changing rhythms In the old rhythm there was an orderly market. Even into the mid-1960s, 75% of world annual copper production was controlled by eight companies: two in the UK, one in Belgium, and five in the US. During times of poor market demand, these producers would control supplies to the market. They took their losses out of retained earnings. Capital was easily restored during the ensuing good times. The new rhythm called for copper to be an economic means to a political end. Third World copper producers brushed aside the market fundamentals of commodity economics. Yet, the political solutions of these countries - their lack of response to the market - have seldom achieved their original objectives. Each commodity has its own old and new rhythm. Aluminum may have a similar rhythm, coming 10 to 20 years later than copper. By the late 1960s, changes were occurring. Inflation was on the rise. Ore reserve quality was falling. As a result, capital expenditures were necessary to maintain existing output. These costs created an increase in corporate leverage, thus adding to inflation and rising interest rates. Indicators A comparison of the gross national product (GNP) deflator index, consumer price index, and the metal mining deflator index is instructive. (Fig. 1 and Table 1). With 1972 equal to 1, note that the metal mining deflator reached its lowest point in 1967. From that point, the metal mining deflator has accelerated severely for 16 years. Only two minor downturns occurred during the business recessions of the mid-1970s and the early 1980s. In essence, Fig. 1 shows that while GNP has grown modestly, a miner's dollar in 1972 would require $3.48 in 1980. This is an almost 17% compounded growth rate over that eight-year period. Beginning in the late 1960s, for the first time in mine financing, the risk of success shifted from the corporate balance sheet to third parties. Limited-resource financings were taken up by banking syndicates that accepted certain of these risks. The lender needed to be well acquainted with the types of risks to be evaluated. These include resource quality and quantity, construction and completion, capital and operations costs, market risks, force majeure, and political and exchange rate risks. Project financings became the basis whereby mining companies were able to maintain their borrowing capacities. In essence, the lender was willing to share some of the project risk while only receiving lenders' returns on these investments. Over the last two decades, mineral project financings have become highly complex transactions involving multicurrency, multi-

Jan 8, 1985

By S. Budirsky

J.B. Gwiazda's article deals with an interesting problem that has not been studied thoroughly up to now. Gwiazda has proposed a technical solution that eliminates horizontal load imposed by the supports on the roof. That load is due to a discrepancy between the trajectories of the canopy and the overlying strata. It is true that the improved lemniscate guidance system proposed by Gwiazda could eliminate the horizontal load but, on the other hand, it would represent a higher cost of the supports, and it would complicate the already complicated mechanism of powered supports. That is why we must be convinced that the proposed improvement is actually needed. From that point of view, the following questions should be elucidated. As shown in Fig. 1, the point O travels along the curve a, which produces horizontal load on the roof in the direction of the coal face. As stated by Gwiazda, "such a high load is capable of destroying the roof above the support, causing rock debris to be scattered around the face." The results of our measurements carried out in coal mines of the Czechoslovak part of the Upper Silesian Coal Basin show that the convergence in the modern type of powered supports does not exceed several millimeters (not more than 10) per hour. Similar conclusions were drawn from field measurements performed in the Polish part of the Upper Silesian Coal Basin (see S. Romanowicz and H. Szopka in Proceedings, Scientific and Technical Symposium Simmex '85, Katowice, Poland, pp. 73 -83). The small vertical closure of the supports also induces a small horizontal displacement of the canopy. In addition to this, the transmission of forces between the roof and the canopy is influenced by the layer of debris on it, i.e., the canopy slides without imposing a greater horizontal load on the roof. This was deduced from field measurements of Voest-Alpine F 4/4600 powered supports equipped with rams instead of the common front lemniscate links (denoted as 1 in Fig. 5). In these supports, the pressurizing of the space of piston rod (la in Fig. 5) causes the rams to close, which induces horizontal load on the roof in a direction toward the coal face. After the setting of a support unit, the pressure in the piston rod space was found to drop, frequently to zero, because the small horizontal sliding between the canopy and the debris or inside the debris caused the horizontal reaction from the roof to fall to zero, although the debris were pressurized by vertical setting load. As a consequence, there was no horizontal load imposed by the canopy on the roof. A rise of the horizontal load was recorded only after the passage of the shearer and during lowering of the adjacent units. Thorough information on the results of the afore- mentioned field measurement is given in the paper, "Analysis of the performance of shield powered supports installed in a thick seam," published in Mechanizacja i automatyzacja gornictwa, 1982, No. 12, pp. 38-46. An English translation of the paper is available from the author. We have deduced from our observations that horizontal load imposed by the canopy on the roof in the direction toward the coal face is useful for roof control because it limits the displacement of the immediate roof toward the gob particularly during the rise working. We came to the conclusion that horizontal load on the roof toward the coal face should be induced on purpose. Could Mr. Gwiazda prove with the results of field measurements his contrary opinion? Has the small horizontal displacement of the canopy toward the face actually had an adverse effect on the roof? As shown in Fig. 2, during the setting of a support unit, the canopy imposes a horizontal load on the roof in the direction of the gob. In this case, I share Gwiazda's opinion that the horizontal load could have an adverse effect on the roof. Nevertheless, two important circumstances were neglected - the horizontal compliance of the supports and the interaction of adjacent units. The question of the compliance was analyzed precisely by I. Krumnacker in Gluckauf -Forschungshefle, 1984, NO. 5, pp. 219-223. He found that due to the clearance between the hinge pins and the eyes of the sheild and due to the elasticity of the steel structure,

Jan 1, 1988

By J. O. Marsden

Part I in this series of papers considered the chemistry of medium-temperature pressure leaching of copper sulfide concentrates and reviewed the metallurgical testing and initial process development for this technology. This initial work resulted in the development of a process flowsheet utilizing superfine grinding (to a P98 of 12 to 15 µm and P80 of 6 to 7 µm), pressure leaching at 160° to 170ºC, followed by conventional solution extraction (SX) and electrowinning (EW) for the recovery of copper. However, this process required a significant amount of concentrated acid to be fed to the pressure leach vessel to meet the reaction requirements, representing a cost of between $0.03 and 0.05/lb copper (2006 basis) produced by the process. It was immediately recognized that the development of a process that did not require such a large acid addition would be advantageous. This paper, which is Part II in the series, describes the development of a process that incorporates medium-temperature pressure leaching and direct electrowinning (DEW) in a novel and innovative process flowsheet. Phelps Dodge (a subsidiary of Freeport-McMoRan Copper & Gold Inc.) refers to this as the ?MT-DEW-SX? process. Part III in this series covers the large-scale demonstration of the MT-DEW-SX process at Bagdad, Arizona; and Part IV addresses the development and design of a full-scale commercial application at Morenci, Arizona.

Jan 1, 2007

This paper discusses the type of scaling operations presently in use in underground mines. It defines a scaling factor that is expressed as the area necessary to scale to extract 1 t (1.1 st) of ore. The problems associated with conventional scaling methods are discussed as well as recent developments in mechanized hand-held scaling de- vices and methods of loose detection. The elements of the scaling process are further analyzed. The conception and design of a mechanized scaling device, developed at Universite' Laval is also described. The results of field tests conducted in Quebec hard rock mines are presented. The experimental findings show that the mechanized hand-held scaling device shows potential in improving the productivity and safety of underground mining operations.

Jan 1, 1996

By J. J. Davis, D. G. Osborne

The Australian coal industry is characterised by efficient fines recovery, and the climate does not dictate the use of thermal drying. With the increasing trend to underground mining and hence finer ROM coal, and market pressures for reduced product moistures, dewatering issues are assuming increasing importance in the Australian coal industry. The greatest potential gains in dewatering performance undoubtedly lie in the treatment of finely sized material. This paper examines the dewatering issues which have been recognised by the Australian coal industry, and describes current Australian research in this field.

Jan 1, 1995

By Randy Fulton, Ignacy Puszkiewicz, George H. K. Schenck

Stuart Olson Construction, Inc. excavated 385m (1263 ft.) long twin tunnels with 6.25m (20.5 ft.) O.D., a portal and a cross-over for the Edmonton, "Metro". The work for Contract No. 76458, SLRT South Tunnel and Portal, in Edmonton, Alberta, Canada was performed in 1989 and 1990. NATM was specified because of the instability of glacial deposits. The contractor decided to use "two firsts" on this project: Fully mechanized NATM and a novel excavator¬mounted tunneling machine. An excavator was equipped with a cutter boom, and a loading and conveying system. This novel machine cost only about one-third of the price of a special NATM roadheader. This new combination of NATM and an excavator-mounted cutter boom offered a cost-effective alternative to shields, slurry TBMs and roadheaders for short tunnels in unstable ground. For NATM tunneling in hard rock or in mixed faces, the roadheader-type cutter boom can be replaced with an impact hammer (hydraulic breaker).

Jan 1, 1993

By Ted D. Maki, Joseph B. Taylor

INTRODUCTION FMC's Paradise Peak mine, located 13 Ian (8 miles) south of Gabbs, Nevada, became opera¬tional in April of 1986 (Figure 1). It was designed and built by Davy McKee, who was instrumental in equipment design and selection. It is the 7th largest gold deposit in the United States with defined reserves of 10.9 mt (12 million st) containing 34 million gms (1.1 million tr oz) of recoverable gold and 933 million gms (30 million tr oz) of recoverable silver. The mine extracts ore by the open-pit method, taking advantage of a 1.5:1 stripping ratio. The mill operates at a 3600 mt/day (4000 st/d) capacity. Crushing is done in three stages to achieve an ore size of minus 0.635 cm (0.25 inches). Grinding further reduces the ore to 85% minus 100 mesh. The ground ore is treated with cyanide in agitated leach tanks and then washed in the counter-current decantation (OCD) thickeners. Zinc is added to the clarified, deaerated pregnant solution to precipitate the precious metals. The precipitate is acid digested to eliminate excess zinc, filtered and retorted to drive off contained mercury. The retorted precipitate is then fluxed, melted and poured through cascade molds. Dore bars are cleaned for shipment and the slag is sent to an in-house slag treatment system. ELECTROSTATIC SEPARATION The dry electrostatic slag treatment system at Paradise Peak is the first installation of its kind. Electrostatic separation has been widely used in mineral processing since the early 1950's. A brief discussion of the theory behind the process is helpful to those not familiar to electrostatic separation. Charging and sepa¬rating slags, dry minerals or other materials by ion bombardment is the most common form of electrostatic separation. Millions of tonnes of minerals are processed each year by this method. In an ion bombardment separation, granular material is fed onto a grounded metal cylinder (or roll) and charged by a corona-producing electrode placed above the roll's surface (Figure 2). While both conductors and non¬conductors become charged, only the conductor is able to lose its charge. The charged non¬conductor, as it rests on the roll's surface, "sees" an oppositely charged image of itself in the metal surface. It is attracted to the image charge, becomes electrostatically pinned to, and moves with the roll's surface. The conductive particle also sees an image and is attracted to it. But upon touching the roll's surface, it discharges rapidly to the grounded surface and is thrown free from the roll's surface with a projectile motion. In the case of precious metal slags, the con¬ductive particles would be metallic prills of dore metal; and the nonconductive particles would be the slag, free of metal. Middling grains, generally, are in the form of a metallic prill encased in or incompletely liberated from slag. LABORATORY TESTING A number of precious metal slag samples have been tested in the laboratory. It has been found that the composition of slags vary widely from one refinery operation to another. Typical laboratory procedure is to crush and size the slag followed by two-stage lab-scale electro¬static separation. The conductor fractions from the first and second pass are then combined as a prill concentrate; middlings fraction and clean slag tailings from the second pass are held separate. Selected results from laboratory testing are shown in Table 1. It is important to note in Table 1 that the assays represent overall silver and gold, not metallic values. From experience, the laboratory results are typically lower in grade and recovery than the industrial installations. This is largely due to the hydroscopic nature of precious metal slags coupled with the high local humidity in Carpco's Jacksonville, Florida, location. Relative humidity in Jacksonville can range form 60-95%, while most mining locations in the

Jan 1, 1987

By John Ralston, Rossen Sedev, Nicola Richards, Quinn Anthony, Geoffrey Reynolds

Derivatised 2,4-diketopyrimidine groups have been anchored to planar gold surfaces as self¬-assembled monolayers (SAMs) for chemical and photochemical studies. Three derivatives, 5-methyluracil (thymine), 5-trifluoromethyluracil and 5-nitrouracil, were chosen for the specific moiety present at position C-5, which can exert strong or weak inductive effects on the adjacent aromatic ring. The measured acidity constants decrease from 5-methyluracil to 5-nitrouracil as the inductive properties of the group at C-5 position increase. When grafted to the surface the acidity constants increase, albeit still adhering to the expected inductive trend. The functionality also exerts an influence on the surface wetting properties. Surface wettabilities indicate that the orientation of the pyrimidine groups at the surface occurs so that the C-5 position has a direct effect upon the surface chemistry. Upon dimerisation the surface contact angles increase signifi¬cantly. At pH 11 the contact angle changes reflect that the dimer has a higher pKa than the monomer. The monomer surface is deprotonated at this pH while the dimer surface is partially deprotonated, leading to changes in contact angle of up to 22 degrees. Large changes in contact angle were observed for 5-nitrouracil which were independent of surface charge and were appar¬ently due to conformational changes alone. Electrowetting of three fluoropolymer surfaces by electrolyte solutions was studied with the sessile drop method. The electrowetting curve (contact angle/potential) is analogous to the elec¬trocapillary curve (surface tension/potential) and may be described by a combination of the Young and Lippmann equations. We demonstrate that for some specific surfaces a systematic deviation occurs at positive potentials. This departure is attributed to double layer effects, namely the adsorption of hydroxide and halide anions. pH, ionic strength and polymer composi¬tion can all influence electrowetting behaviour.

Jan 1, 2003

By L. W. Laage

Throughout recorded history, underground fires have plagued mining operations. Compared with other hazardous situations in underground mining, a fire can become a global problem by swiftly spreading deadly carbon monoxide and other products of combustion (POC's) throughout the whole mine, often without warning. Experience has shown that when fires are detected and located in their early stages, they are much easier to control and proper escape routes can be more intelligently selected. Recent advances in sensor and data communication technology have made reliable mine fire .detection system installations possible. Unlike building construction. in a mine, it is impractical to install detectors at every desired location. Abandoned workings and unsafe locations preclude sensor installations from both safety and economic standpoints. The net effect is that some fires are detected and located early while the location of others, even if detected early, remains unknown too long for effective evacuation and fire fighting.

Jan 1, 1989

By H. A. Kortmann

Blast furnace operators have been steadily increasing their quality requirements on burden material. These requirements include a high reduction rate of the ore burden as well as good strength at the beginning and during the whole blast furnace operation. Special requirements are made on the softening and melting behavior which ensures an optimal formation of the cohesive zone in the blast furnace. On the whole, the self-fluxed and olivine-based pellets meet these higher requirements due to their chemical composition, especially in the pure pellet operation. In the mixed sinter/pellet operation partly-fluxed and acid iron ore pellets can be appropriately used. The paper describes the demands on, and characteristics of, different kinds of pellets used in the Federal Republic of Germany.

Jan 1, 1989

By Liuming Wu, Willis Forsling

The complexation of Alizarin Red S (ARS) at the surface of hydrous fluorapatite particles has been investigated at 25.0 C by means of potentiometric titrations, adsorption experiments in 0.1 M NaCI ionic medium, and Diffuse Reflectance Infrared Fourier Transform spectroscopy (DRIFT) studies. A chemical model describing the reaction at the aqueous mineral surface has been established under equilibrium conditions according to the reaction [== -CaOH2 + HA2 - « _ =PO CaA2 - z + H30&apos;; to (int.) = -5.61 ± 0.02 3 a =PO- logBs-1110(int.) ( ) within the range 8.40 <_ -log[H ] <_ 9.80.] Experimental data were evaluated using the computer program FITEQL on the basis of a constant capacitance model for the electric double layer. The surface complex structure, has been proposed to involve the quinoid and alpha phenolic oxygens of ARS coordinated to the calcium sites on the fluorapatite surface.

Jan 1, 1993

The global tunneling industry has many challenges, both technically and financially, with the latter likely to be a significant feature over the next few years. Many industries have ad-opted standard practices in order to overcome specific technical challenges, from which standard out-comes are anticipated. In tunneling, however, every project is unique so standardized practice is not easy. This requires original thinking and being open to adopting new techniques and technologies as they emerge. Sprayed waterproofing is a hot industry topic and a paper recently presented at the North American Tunneling conference not only discussed this methodology for achieving a dry tunnel but also how it can reduce build costs and construction time. Despite their being relatively few ?universal truths? throughout the tunneling industry, there is one opinion that appears to be shared by most in the sector, ?all tunnels leak.?

Jan 1, 2010

By Agus Sasmito, SRIVATSAN JAYARAMAN SRIDHARAN, Akash Adhikari, Purushotham Tukkaraja

The amount of toxic fumes produced during blasting operation depends on the types of explosives used, product formulation, degree of confinement, conditions of use, and reactivity of the explosives with the host rock. Ammonium nitrate and fuel oil (ANFO) is still the explosive of choice due to its lower price and ease of use/loading. If not properly monitored and controlled, toxic fumes produced from blasting operations can have serious health hazards [1]. This paper presents a novel technique developed to track trapped fumes during a development blasting.

Aug 1, 2022

By N. Piret, B. Shoukry, S. Buntenbach

Traditional or artisanal goldmining, also known as small scale goldmining, has a strong and probably a negative environmental impact. The processing methods applied are very frequently a source of severe pollution due to the emissions of mercury by the extraction of gold by means of amalgamation as well as the emissions of cyanide through cyanide leaching of gold bearing ores. The emissions find their way into the environment and contaminate soils, sediments, water and atmosphere. Abnormal concentrations of mercury and cyanides in waterways are known to occur year after year destroying irreplaceable regions of the world. Mercury and cyanide compounds are highly toxic and may directly create permanent damage to the whole ecosystem. Existing methods for recycling of mercury and for decontamination of mercury and cyanide contaminated tailings are not customary applied in small scale mining and are ineffective as well. Based on investigations of traditional and small size goldmining, this paper presents: -processing methods of gold and discarded tailings under consideration of environmental protection; -figures on actual situation; -recommendations for equipment; -some decontamination methods for mercury and residual cyanide. Mineral Processing methods in traditional gold mining Gold is usually existing in its ores as the metal alloyed with metallic silver and perhaps copper. The element may occur in the form of: -native gold -inclusions also of microns or submicroscopic size metal sulfides (auriferous) such as pyrite, pyrrhotite, stibnite, arsenopyrite and galena -combined as telluride or sulphotelluride. The separation process selected depends on whether the gold can be freed from its unfavorable associations (e.g. gangue) at a sufficiently coarse grain-size, or whether it is carried in a heavy sulfide which can be freed similarly. The usual practice is to concentrate the goldbearing mineral at a relatively coarse grain-size and to regrind the ore if necessary. The gold content is concentrated by secondary or tertiary gravital methods or is extracted by chemical methods (amalgamation, cyanidation etc.) Gold, even when of fine grain-size, settle readily due to its high specific gravity from pulps in which the main gangue mineral is quartz or silicates. Amalgamation is the process of separating gold and silver from their associated minerals by binding (entrapping) them into a mixture with mercury. The cyanide process is applied to separate gold or gold-bearing compounds by dissolution from the finely ground ore (CIP, CIL, RIP), or as heap leaching. The dissolved gold is separated from the solids and the metal-rich or pregnant solution is then treated to recover its gold. Gold is also recovered by flotation methods. This process is widely used in treating base metal ores and in separating various sulfide components of ores, as well as in removing the barren gangue. The gold usually associates with a specific product in a sequence of flotation operations and is recovered subsequently in the smelting of the sulfide concentrates and refining of the metallic products, or by cyanidation of the roasted concentrates. Froth-flotation can be applied to separate gold and sulfide minerals from a finely ground pulp. The Amalgamation Process Amalgamation is the main method for the recovery of gold in traditional mining and is applied for the extraction of gold from placers as well as primary ores. The mineral technology used depends on the nature of ore deposits. In winning gold from solid ore, the matrix of minerals and rocks must be crushed and ground to sufficient fineness to liberate the gold. The liberated gold could be treated similar as free gold from placers. Gold is mainly separated from the valueless gangue (barren rock) by utilizing the difference between the density of the impure native metal (density about 16-19) and the gangue (density about 2.5). In simple operations the material is carried by a stream of water down a sluice generally equipped with small transverse barriers (riffles) against which the gold collects. The riffled sluice is the principal device used by artisanal gold miners. Nowadays, spirals as well as centrifuges, such as Knelson separator or Falcon separator, are occasionally applied for gold recovery. Gold may also be recovered from the pulp, by passing it over corduroycovered tables that catch the heavier particles - a method maybe as ancient as gold mining itself. In history, sheep skins were used to catch gold particles in this manner. Furtheron, gravity separation of gold is practiced on jigs, hydraulic traps, shaking tables and

Jan 1, 1995

By Kenneth P. Katen, Robert H. Trent

GENERAL DESCRIPTION Shortwall mining is one of the newest mining meth¬ods to be introduced into the world coal industry. The technique of shortwalling seems to have a number of origins, but it is generally agreed that it is an outgrowth of pillar-and-stall mining where timbering was employed as the primary means of roof support. In early opera¬tions the face was undercut by a shortwall cutter, hand¬bored, and blasted. The coal was loaded by hand or gathering arm loader. Australia, having experienced success with this method, advanced the technique through the use of mechanized roof supports in 1968. Although there were several attempts at shortwalling in the US, it was not until 1973 that Eastern Associated Coal Corp.&apos;s Federal No. 1 mine successfully extracted a shortwall panel. Table 1 is a list of shortwall opera¬tions in the US as of April 1976. Shortwall mining is a method of mining in which a continuous miner cuts and loads from the short end of a rectangular pillar of coal while hydraulically powered self-advancing roof supports provide protective cover. Characteristics of a shortwall system include panel di¬mensions (Fig. 1) of 610 to 1219 m (2000 to 4000 ft) in length, face widths of 30 to 61 m (100 to 200 ft), straight-line ventilation, and production machinery, with the exception of roof supports, commonly used on con¬tinuous miner units. GENERAL REQUIREMENTS AND LIMITATIONS The justification for shortwalling is primarily eco¬nomic, i.e., increased production and recovery concur¬rent with savings in material and labor. The method also has a number of secondary benefits including the much needed progress it has stimulated in deep mine technology, the alternatives it presents to some of the problems resulting from the 1969 Coal Mine Health and Safety Act, and is perhaps a key to the solution of the enigmatic labor problems which plague the industry. The advantages of shortwall mining over continuous miner units include the following: a highly effective method of ventilation and dust control; a greater per¬centage of extraction (approximately 85%) ; fewer equipment moves, hence more efficient operating time; cost savings on materials such as roof bolts, rock dust, ventilation curtains, timbering materials, and the labor needed to supply, load, and transport these materials; a cyclical pattern of mining, each phase of which, i.e., development and retreat, lasts for approximately three months and appears to be psychologically motivating as well as highly productive; and improved safety. In addition to these positive aspects, a shortwall sys¬tem also has advantages over a longwall system, and these are listed as follows: the capital outlay for a shortwall system is one-third to one-fourth that for a longwall (1976 prices); the shortwall system utilizes equipment that is used on a continuous miner section, with the ex¬ception of the roof support system; there is a great deal of flexibility in areas with numerous gas wells and in areas where surface subsidence cannot be tolerated; the shortwall has reduced tailgate problems, since the tail¬gate entry serves as a return airway only; and the cyclical pattern of mining insures that the mining machinery pro¬duces coal even while the roof support system is being moved to another panel. There are several parameters which govern the se¬lection and use of a shortwall system. They are all of equal importance, and hence are not listed in any par¬ticular order. These are deposit size, seam height, seam inclination, roof and floor conditions, depth of over¬burden, subsidence and chain pillar extraction, man¬power, ventilation, and materials handling. Deposit Size Reserves available for shortwalling should be suffi¬cient for at least an eight to ten-year period, since this seems to be the effective life, without rebuilding, of a hydraulic roof support under average conditions. Of

Jan 1, 1982

By R. A. Metz, Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (1), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profitability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus minimum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12). Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-lowgrade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1992

By Rolf Hindstroom

Outokumpu Mintec started development of the pressure filter in 1991, and the first unit was delivered to Tara Mines Ltd. Ireland in 1992. The filter replaced a conventional disc filter and a thermal dryer. The replacement of the thermal dryer has improved both environmental and working conditions. The filter is currently dewatering 80,000 tons per year of lead concentrate, producing a cake with less than 6 wt % residual moisture.

Jan 1, 1995

By Charles T. Steuart, Richard B. Berg

Tripoli and the related mineral commodities such as micro- crystalline silica have been mined for more than 100 years for their abrasive properties. Although abrasive and buffing compound markets are still very important, within the last 15 years the filler and extender markets in paint, plastics, rubber, adhesives, and sealants have increased substantially. Tripoli or microcrystalline silica consists almost entirely of very small quartz crystals, many less than one micrometer in length. Material mined from different districts differs in crystal shape, grain size, and texture of the rock, all of which influence markets. Most US deposits are now mined by surface methods, and both air floated and micronized products are marketed. Deposits of tripoli now mined in the United States occur in chert-bearing Paleozoic limestone in the central part of the country with producing districts in southern Illinois, central Arkansas, and eastern Oklahoma. Although deposits within each district are confined to specific formations and extend laterally within those formations, individual bodies form minable deposits that are typically several hectares in areal extent. Tripoli is white to cream to rose and characterized by high porosity and ease of disaggregation. DEFINITIONS Tripoli In the United States, tripoli was first used to describe the fine-grained, easily disaggregated material from Seneca, MO, be- cause of its similarity to a rock from the Tripoli region of North Africa (Hovey, 1894). The North African rock is actually diatomaceous earth, a material that is similar in appearance to the rock from Seneca, but is of entirely different origin having formed by the accumulation of siliceous remains of microscopic marine or fresh- water animals. Tripoli is best defined as a very fine-grained, generally porous rock that consists of microcrystalline quartz, typically formed by the alteration of a chert-bearing limestone. Tripolite A term used to describe a rock from the vicinity of Tripoli in North Africa which is diatomaceous earth (Quirk and Bates, 1978) Microcrystalline Silica Microcrystalline silica is the same material as tripoli, but the distinction between the use of these two names is dictated largely by convention and markets. Material produced from southern I1linois deposits and used in white pigment and filler applications is generally referred to as microcrystalline silica, whereas that used in abrasive applications, both from the Illinois district and from other states, is commonly called tripoli. Amorphous Silica Amorphous silica, a term formerly used to describe the material produced from the deposits in southern Illinois, is now replaced by the term microcrystalline silica. Amorphous silica came into use when even optical methods for the identification of very fine- grained quartz were not widely available and the Illinois product, composed of quartz grains too small to be seen with the unaided eye, was thought to consist of amorphous silica. The Illinois material is clearly crystalline quartz, as shown by X-ray diffraction analysis and scanning electron microscopy (Fig. 1). Novaculite Although originally used to describe a rock suitable for the manufacture of whetstones, novaculite is now defined more generally " - as a homogeneous, mostly white or light colored rock, translucent on thin edges, with a waxy or dull luster, and almost entirely composed of microcrystalline quartz" (Steuart et al., 1983). The more compact rock mined in central Arkansas from the Arkansas Novaculite is referred to as novaculite, whereas the more porous rock is referred to as tripoli. Rottenstone The commodity rottenstone is sometimes included within the general mineral commodity category of tripoli. Rottenstone is mined in Northumberland County in eastern Pennsvlvania and formed by the weathering of a siliceous shale of Devonian age (Faill, 1979, Berkheiser, private communication, 1991). This material is used as a filler and extender, but is apparently unlike tripoli both in origin and physical properties. Spiculite Spiculite is a rock consisting of siliceous sponge spicules having formed by the removal by solution of the carbonate matrix of a spicule-bearing limestone. Spiculite has been mined in Texas and, because it resembles tripoli in several aspects, is included in the discussion of deposits.

Jan 1, 1994