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By H. A. Pearse

The history of the flotation process, as applied to the practical field of ore dressing, dates back somewhat less than twenty years; yet this fact may be overlooked in view of the present universal dependence on this process in the concentration of almost all classes of ore. It is therefore interesting to recall that Britannia was the scene of some of the earliest experiments with flotation on a commercial scale. In 1911, sixteen years ago, an Elmore vacuum machine was operated for several months with encouraging results. In 1913, some time before flotation was very generally adopted elsewhere, a Minerals Separation plant, capable of treating 600 tons daily, was operating successfully at Britannia. Considering how little was known of the theory and practice of flotation at that time, the recovery was remarkably good, but it was not until the present mill was completed in 1923 that any attempt was made to raise the concentrate grade by differential flotation, or, in other words, to float the chalcopyrite while rejecting the then worthless pyrite with the tailings. The first ore passed through the present concentrator in March, 1923. As originally designed, the rated capacity was for 2,000-2,500 tons, but this has been increased year by year, until at the present time 4,000 tons daily can be treated economically. During the first few months of operation the concentrate grade averaged only 10-11 per cent Cu, but by the end of 1923 this had been raised to over 20 per cent, due to the elimination of the pyrite by differential flotation. In 1923 the tailings averaged approximately 0.25 per cent Cu, but also in the matter of recovery a considerable improvement has been effected, for during the year 1926 an average tailing of 0.15 per cent Cu was maintained.

Jan 1, 1927

By Bunting S. Crocker

Bunting S. Crocker (Executive Vice President, Kilborn Engineering Ltd., Toronto, Ont., Canada)-This is an interesting report on laboratory-scale testing with pebble charges in 12-in.-diam mills, and the author correctly concludes that useful laboratory testing can be done. In the opening paragraphs he states, "laboratory pebble-milling tests have been considered too difficult to perform and not relevant enough to the actual grinding situation"!! This is definitely not so! The writer published detailed tests done in a 12-in.-diam mill comparing pebble and steel balls in 1952 6 and again in 1954 7. We note that our review paper in 1963 was the only one mentioned in the References. The author states that "a smooth lining is more frequently used in commercial-scale pebble grinding." The writer has designed some 22 plants treating 134,000 tpd in North America and Australia. He has also visited pebble installations in Norway, Sweden, Finland, and South Africa. All these plants correctly use heavier lifters than would be considered standard for ball mills of a similar size. Smooth liners are not recommended in wet ore treating plants. In the 1952 paper6 it was clearly demonstrated in the test work that the individual ore pebbles had to be the same weight-not the same diameter or size-as the steel balls if comparable grinding results were to be obtained. In the tests performed in this study most of the pebbles "passed 50-mm round hole screen openings and remained on a 40-mm round hole sieve. Correspondingly most of the ball mill experiments were performed with 50 or 40-mm steel balls". Thus, the balls and pebbles were essentially the same size or diameter. This means we are comparing apples and oranges. When the feed being ground is fine enough the pebbles will be superior. When the feed is coarse the steel balls will be the better. This is similar to grinding tests in which two sizes of steel balls are compared. In the middle of the article, "Crocker5 has reported that equal weights of pebbles and steel balls achieve the same grinding capacities under similar conditions …", then he proceeds to question this postulate. In the eight reports that I have published, I have never made this statement. What I have said is "for comparable grinding, the individual pebbles in the mill charge should be the same weight as the individual steel balls." This is not the first time that I have been misquoted on this point! It is most unfortunate that this test work has been done with a ball and pebble of the same diameter and not the same weight. It is impossible to properly evaluate the results. I also regret the fact that the screen analyses of the tests were not shown. Many of us who have been 40 years or less in the grinding game can easily read screen analyses at a glance. We can tell whether the top meshes are being ground by the ball charge, etc., etc. The graphs are fine for the summary or to illustrate conclusions but are often misleading in the report itself. However, we do agree that useful pebble grinding tests can be done in a 12-in. mill.

Jan 1, 1973

By R. V. Kumar, C. Schwandt, S. Subramanian, S. Sampath, S. Vimalnath

Waste waters and industrial effluents containing toxic heavy metals such as Pb, Hg, As and Cr, beyond permissible limits (0.1 ppm, 0.01 ppm, 0.02 ppm and 0.05 ppm, respectively) (WHO, 2011), pose serious threats to the ecosystem, when discharged into the environment. It thus becomes imperative to monitor the concentration of the toxic heavy metals and treat them before discharging. In this study with the above objective in mind, a microbial sensor has been developed using P. aeruginosa cells drop coated onto a carbon paste electrode (CPE) for the detection of Pb(II) ions in aqueous solutions. Detailed biosorption experiments confirmed that complete bioremoval of lead could be achieved using P. aeruginosa cells under optimum conditions of pH and biomass loading. For the electro-analytical determination, the Pb(II) ions were initially preconcentrated on the P. aeruginosa modified CPE, reduced and then anodically detected with oxidation of Pb0 using differential pulse anodic stripping voltammetry (DPASV) technique. All the experiments were carried out in the presence of 0.1 M sodium acetate buffer and KNO3 as background electrolyte. The parameters namely pH, scan rate, biomass loading, deposition potential, deposition time and concentration of Pb(II) ions were optimized for the biomass modified CPE, to obtain maximum voltammetric response. Under the optimized conditions of pH 5.6, scan rate of 0.008 V/s, biomass content of 1.7–1.8×109 cells/mL, deposition potential of –1.2 V, and deposition time of 500 s, the developed microbial sensor showed a linear response for the peak current between 0.5×10-5 M and 1×10-8 M of Pb(II) ions with the R2 value of 0.9979. The lowest limit of detection (LLOD) was determined as 2.4×10-7M of Pb(II) ions (0.05 ppm). The redox mechanisms of Pb(II) ions on the P. aeruginosa modified CPE surface are discussed.

Jan 1, 2016

By L. Semushkina

Next generation flotation agents of KSK series for the concentration of complex copper-lead ores and the new ferromagnetic nanosized depressor of galena are introduced in this work. The feature of the KSK flotation agents is that they exhibit both frother and collector properties. KSK-6 can successfully substitute the widely used in Kazakhstani industry reagent ?-80 at the reduced consumption of butyl xanthate by 30 %. The new ferromagnetic nanosized depressor of galena, besides a low consumption rate of 100 g/t, also makes possible to exclude sodium sulfite (3-4 kg/t) and ferrous sulfate (5-6 kg/t) from the separation process, and receive high metal grade concentrates. Keywords: complex copper-lead ore, flotation agent, concentration method, ferromagnetic nanosized depressor

Sep 1, 2012

By J. C. Bokros, A. S. Schwartz

The irradiation-induced densification of pyrolytic carbon has been shown to take place by a simple first-order rate process. The values of the densification constant, kd, for isotropic carbons are 1.0 x 10-22, 1.4 X 10-22 and 2.8 X 10-22 (nut)-1 at irradiation temperatures of 520o, 670°, and 1000oC, respectively. The apparent activation energy obtained from the temperature dependence of kd for isotropic carbons is about 0.2 ev. Estimates of the maximum stresses that develop in isotropic carbons owing to shrinkage onto an unyielding support show that stresses in excess of 15,000 psi may be avoided during irradiatian at 600" and 1000°C if the initial coating density is kept above 1.5 and 1.9 g per cu cw, respectively. If a carbon coating on an unyielding support is not completely isotropic, stresses may continue to build up at high neutron exposures after densification is complete because of continued anisotropic distortion. Stresses in excess of 15,000 psi from this source may be avoided during irradiation at 1000°C if' the Bacon anisotropy factor is kept below 1.3. A variety of applications of carbon and graphite in nuclear technology has stimulated numerous studies of the effect of neutron irradiation on these materials One of the most interesting problems associated with the use of carbons and graphites in the high fast neutron fluxes of nuclear reactors arises as a consequence of the neutron-induced dimensional changes that occur during irradiation. Energetic neutrons displace car- bon atoms from their lattice positions; and, depending on the temperature, the interstitials and vacancies produced migrate to form clusters, recombine, or are annihilated at dislocations or crystallite boundaries. The competition of various sinks for the point defects results in structure-sensitive bulk-dimensional changes that vary with temperature in a complicated way. This paper deals with the neutron-induced dimensional changes in turbostratic carbons that are being used as coatings on nuclear fuel particles. Details of the structure of these carbons and the variations of their structure with deposition conditions have been reported.'-3 It has been shown in previous work that the neutron-

Jan 1, 1968

By Edward (Ted) Dowey

City Water Tunnel No. 3 is being built in order to provide a redundant system for delivering 4,542,500 m3 (1.2 billion gallons) of water per day throughout the five boroughs of New York City. The present tunnels, City Tunnel No. 1 and 2 (Figure 1), were put in service in 1917 and 1936 respectively, and both have been in continuous service. Tunnels No. 1 and 2 can-not be shut down for inspection or repair until City Tunnel No. 3 has been completed. City Tunnel No. 3 has two stages (Figure 1).Stage One is 20.1 km (12.5 mi) long and is presently delivering drinking water to the boroughs of the Bronx and Manhattan. Stage Two consists of the Brooklyn-Queens Tunnel 17.7 km (11 mi) and the Manhattan Tunnel 14 km (9 mi). The Brooklyn-Queens leg is complete and is set to be put in service soon. The Manhattan Tunnel is being constructed and is the subject of this paper. The Manhattan Tunnel (Figure 1) is comprised of three legs: the north, south and east legs. The con-tract includes the excavation of 8,976 m (29,500 ft) of tunnel and the lining of 14,561 m (47,773 ft) of tunnel. The south tunnel, 5,705 m (18,720 ft), was excavated but not lined, under a previous contract. The tunnel ranges from 181 m (600 ft) to 152 m (500 ft) in depth in the bedrock under Manhattan. The tunnel contract also includes the construction of nine new shafts, 161 m (530 ft) average depth, and the completion of a tenth shaft (the previously built construction shaft). The Manhattan Tunnel contract was awarded to the sole bidder Schiavone/Shea/Frontier Kemper Joint Venture (SSF) for $668,532,680.00, on January 31, 2005. The Manhattan Tunnel was designed ?in house? by the New York City Department of Environmental Protection (DEP) and the construction management was provided by a combined staff of the DEP and a joint venture of Jenny Engineering and URS.

Jan 1, 2008

By J Loeb, D Thornton

Blast movement measurement is a key component of grade control, enabling mines to carry all of the precision involved in locating ore blocks, through a chaotic blast, to production. A monitoring system was developed about ten years ago so that mine operation personnel can measure three dimensional movement in every production blast. Research has shown that blast movement is always highly variable with a significant non-deterministic component. The consequence is that modelling will be inaccurate and therefore optimal ore recovery can only be achieved by direct measurement of blast movement.Three dimensional movement vectors are measured at discrete locations in the blast. If blast movement is only measured at one location in a blast pattern, data is severely limited and has all of the inherent inaccuracies of modelling a variable population. It is intuitive that the accuracy of defining the post-blast ore block mark-ups are proportional to the number of movement measurement locations, however increasing the number of monitoring locations causes increases to the direct operating costs of the mine.A cost-benefit analysis is used to explore the optimal number of monitoring locations. Using the method described in this paper, an automated selection of the number and location of blast movement monitors are analysed against the value of recovered ore compared to the control, for each trial.CITATION:Loeb, J and Thornton, D, 2014. A cost benefit analysis to explore the optimal number of blast movement monitoring locations, in Proceedings Ninth International Mining Geology Conference 2014 , pp 441–450 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By J. R. Foster

Investigations into the effect of a variety of acid digestions on common rock-forming minerals demonstrated a considerable variation in the proportion of the total metal extracted. A similar situation applies to the analysis of a number of common rock types. In addition, grain-size variation has an effect on the amount extracted by certain digestions. This work indicates that considerable care needs to be given to selecting an appropriate acid digestion capable of determining metal held in particular components of a sample. By using partial extraction techniques, the identification of certain patterns of metal distribution, undetectable by normal methods involving total metal analysis, may be achieved.

Jan 1, 1973

By Thomas E. McNider

On July 20, 1990, a Spontaneous Combustion Control Plan was approved by MSHA for Jim Walter Resources, Inc. No. 5 Mine. The mine operated under the plan for twenty-nine shifts before MSHA revoked the plan. A total of fifty-five seals were constructed during the process of implementing the plan. The major difficulty in applying the control concept resulted from problems controlling leakage at the numerous seals. The number of seals was a direct consequence of attempting to merge the traditional bleeder system, used in all the Blue Creek Mines, with the bleederless system Jim Walter Resources proposed as a means of controlling the spontaneous combustion hazard. This paper discusses Jim Walter Resources experience with spontaneous combustion control and the new safety standard which applies to underground coal mines that demonstrate a history of spontaneous combustion and how this standard would apply to our No. 5 Mine. The new Ventilation Standards were published in the Federal Register on May 15, 1992, and scheduled to go into effect August 16, 1992.

Jan 1, 1993

The changes in layout and design over 48 years, from the first crushing and shiploading plant installed at Whyalla in 1915, up to the present plant being built at Koolan Island, Yampi Sound may be summarized as follows.1. Smaller motors and lower power consumption have resulted froma. oil lubrication, where practicable, in place of grease,b. a high pressure "oil-lift" device to reduce starting torque in crushers with white metal bearings, andc. roller bearings in place of white metal.2. Short centre drives with pivot-based motors permit controlled loads on bearings and a reduction in size of crusher houses and motor rooms.3. Dust proofing and pressurising of crusher houses and motor rooms reduces working temperatures and improves cleanliness.4. Variable speed pan feeders give improved plant control, with the use of one man only, and remove the need for large surge bins between crushing stages.5. The use of diesel quarry trucks to dump ore through a "drive over"tip results in faster truck cycles.6. Elevated cylindrical silos with air operated chutes, replace long horizontal storage bins with multiple manually operated side discharge chutes for loading rail cars, and give faster and more accurate truck loading. The silos are charged by belt conveyors with automatic distributors.7. Increased shipping bin capacity (40,000 tons at Whyalla and 6O,OOO tons being built at Koolan Island) caters for the bigger vessels now in use and envisaged for the future. Automatic travelling distributors deliver ore to the bins in layers to ensure, maximum mixing and a more uniform grade for blast furnaces.INTRODUCTIONThe Broken Hill Proprietary Company Ltd has six major crushing plants in Australia at its principal sources of raw material. These are at Iron Knob bench G, Iron Knob scree ore plant, Iron Baron, and Cockatoo Island, all of which are ironstone operations. The other two plants are at Rapid Bay, a limestone quarry, and at Ardrossan, which is the Company's only source of dolomite.

Jan 1, 1963

By Xiao Caimei

The method of Substance Flow Analysis (SFA) provides a helpful tool for the study of the industrial metabolism of a certain metal within a regional level, such as copper, lead etc. In this paper the flows and inventory of cobalt in China, 2008, were traced with the STAF model, and the situations of production, fabrication and manufacturing, use and waste management in China were introduced. The result revealed that the resource of cobalt in China was so poor that a large amount of cobalt material or cobalt products needed to be imported, and there was lack of recycling of cobalt resources.0

Jan 1, 2010

By Dean E. Bluman

The Bureau of Mines investigated the convective heat transfer coefficients for recirculated suspensions of 30-micron glass beads in inert gas. Coefficients for the suspensions were not significantly higher than coefficients for the pure gas. Reynolds numbers of the suspensions were 20,000 to 35,000; solids loadings ranged from 0.0 to 0.7 pound per pound of gas. Comparison of the results with published data showed the direction of heat flow to have no effect at the test conditions.

Jan 1, 1967

By R. M. Morris, G. A. Matthesius, M. J. Desai

Slow pyrolysis experiments were caried out on samples of coal from north em Natal having a volatile mater content ranging between 10 and 25 per cent by mass and a constant particle-size distribution. The tests were conducted at atmospheric pressure in a helium atmosphere, and ata heating rate of 10 KImin and a final temperature of 1220 K. The pyrolysis products were taken as coke and volatile mater (tar and eight other species), and tMir respective yields were measured. The experimental results were used in the modification of an existing mathematical model and in the derivation of new corelations for some of the parameters of the model, so that the volatile products of a coal can be predicted from ultimate and proximate analyses of the coal. The pyrolysis tests showed good agreement between the model predictions and the experimental results.

Jan 1, 1987

By R. C. M. Minnitt, W. Dr. Assibey-Bonsu, F. A. Camisani-Calzolari

Born in Bothaville (OFS) on the 26th August 1919, Daniel G. Krige grew up in Krugersdorp on the West Rand. Hematriculated from Monument High School in Krugersdorpin 1934 at the age of 15 and graduated with a B.Sc. (Eng.)degree in mining engineering from the University of the Witwatersrand at the end of 1938 (age 19). In 1938 he joined Anglo Transvaal where he worked on a number of gold mines until 1943, gaining a wide range of valuable practical experience in surveying, sampling and ore valuation. He then joined the Government Mining Engineer’s Department where he worked for 8 years during which time he participated in the uranium negotiations with the British and American authorities and designed the uranium pricing formula for the contracts which led to the establishment of South Africa's profitable uranium industry in the early 1950s. During the same period he began his pioneering work in the application of mathematical statistics to the valuation of new gold mines using a limited number of boreholes and of ore reserves for existing mines. His early papers in the application of mathematical statistics, some republished in French, created world-wide interest that led to the development of the science and concepts that surround the spatial evaluation of mineral resources and reserves known as geostatistics. This technique has contributed to improved ore evaluation techniques as well as the reduction of the financial risks associated with investment in mining enterprises. Furthermore, it led to the recognition of his contributions to the evaluations of mineral deposits through the coining of the term ‘kriging’ that is used to describe a spatial mineral evaluation process known and practised in international mining circles. He continued his work in geostatistics throughout his career as Group Financial Engineer of the Anglovaal Group until his retirement in1981 and afterwards for ten years as Professor of Mineral Economics at the Witwatersrand University. He is still active as a consultant and is a registered Professional Engineer. He has published some 90 technical papers both locally and overseas, some in Russia, and has lectured and participated in international congresses in many countries. His contributions were recognized by the Witwatersrand University through the award of the D.Sc. (Eng.) degree in1963, by three honorary degrees (from the universities of Pretoria, UNISA and Moscow State Mining University), by many awards from the South African Institute of Mining and Metallurgy, the International Association for Mathematical Geology, Die Suid Afrikaanse Akademie vir Wetenskap en Kuns, the SME of the American Institute of Mining Engineers, the International APCOM Council in1999, the University of Antofagasta in Chile, and by the South African President (Order for Meritorious Service, Class 1, Gold). He has served on various Government committees, notably that for State Aid for gold mines (1967/8) and designed the State Aid formula which enabled many mines to survive a period of low gold prices; also various taxation committees, and the mining mission to Iran in 1974. He is a fellow and/or honorary life member of various technical societies locally and overseas, including the Royal Society of South Africa. During the period in the Government service he also handled several of the post-war lease applications in the Free State and Klerksdorp goldfields. The fact that decisions on new gold mines of critical importance to the State and the economy as a whole, were being taken on a limited number of drillholes without any scientific analysis of the risks of failure, stimulated him to start basic research into ore evaluation. His approach was based on the application of Mathematical Statistics to these problems, an approach of which very little was known world-wide at that stage but which had already been initiated in South Africa by Sichel via the lognormal frequency distribution model. In Krige's 1951 paper published in the Journal of the Chemical, Metallurgical and Mining Society of South Africa he covered the statistical explanation of the conditional biases in ore block valuations and stimulated the use by several gold mines of regression corrections for routine ore reserve valuations, a technique which, in effect, was the first use on an elementary basis of what is now known as kriging. His 1952 paper introduced, inter alia the basic geostatistical concepts of ‘support’, ‘spatial structure’,’ selective mining units’ and ‘grade-tonnage curves’. His 1951 paper was based on his M.Sc. (Eng.) thesis submitted in the Department of Mining Engineering to the University of the Witwatersrand, and expounded his pioneering work in geostatistics in more detail. His early research papers that had stimulated interest in several mining circles overseas, were republished in French in1955, resulting in a major research effort by French mining engineers in this field. This, in turn, led to the establishment of the now world-renowned French school of ore evaluation in Fontainebleau, Le Centre de Geostatistique de l’École des Mines de Paris. As the Group’s Financial Engineer at the Anglovaal Head Office until his retirement in 1981, he was responsible for

Jan 1, 2003

Elongates are commonly used as stope support in South African gold and platinum mines. To evaluate the effect of the inherent strength, stiffness and yield variability associated with elongates, a statistical method is presented which addresses and quantifies the support performance at various levels of certainty. The analysis, which is based on a normal distribution, makes use of force-deformation curves determined from multiple laboratory compression tests on the same support type. The applicability of the method is illustrated making use of actual performance capabilities of various types of elongates. In order to ensure a high probability that support units installed underground exceed the standard support performance curve established by means of laboratory tests, the statistical mean of a suite of tests based on the same support type is downgraded. The ?correction factor? is a function of the mean and standard deviation of the test data, as well as the sample size of units deemed to control the local rock mass stability. It is shown that downgraded performance curves based on a high probability (90 to 95percent) of support units exceeding this performance level are suitable to ensure that the design of support systems will in variably meet the design criteria. In the highly discontinuous hangingwall rock mass typically associated with intermediate- and deep-level mining operations, the interaction between adjacent support units is comparatively limited (at the support spacings commonly used). Therefore, the appropriate sample size (n) of the support units controlling the local rock mass stability should be n= 1. In shallow mining conditions, sample sizes of up to n=5 may be warranted, depending on the spacing of joints and other discontinuities in the hangingwall.

Jan 1, 2001

By Kazutoshi Haga, Ariunbolor Narankhuu, Ariuntuya Battsengel, Yasushi Watanabe, Altansukh Batnasan, Akita University

"Rare earth minerals are usually concentrated by flotation, magnetic or gravity separation methods to produce concentrates that are subsequently leached with hydrochloric (HCl), sulfuric acid (H2SO4) or nitric acid (HNO3). About 75 % REEs are lost in gypsum by-product when apatite is dissolved in H2SO4, whereas more than 80% of REE can be extracted from the apatite using HCl and HNO3 leaching processes, respectively. Therefore, the main aim of this study was to find an efficient process of dissolution of rare earth elements (REEs) from apatite ore by sulfuric acid solution due to its lower cost. In order to optimize the dissolution process, effects of different parameters, namely, concentration of H2SO4 (0.05 – 6 M), solid/liquid phase ratio (1:50 – 1:100), temperature (20 – 80 ºC), agitation speed (300 – 900 rpm), and leaching time (0.5 - 6 h) on the dissolution of REEs from apatite ore were examined in various intervals. The contents of main REEs in the apatite ore used in this study were (as % wt): 3 % La, 6 % Ce, 0.6 % Pr, 2.1 % Nd, 0.2 % Sm and 0.3 % Y. X-ray diffraction (XRD) analysis identified fluorapatite and phlogopite as main minerals in the ore. The results obtained from leaching experiments showed that more than 85% of REEs were dissolved from the apatite ore under the optimized conditions, which were: 1 M H2SO4, particle size of -150 + 100 µm, S:L phase ratio of 1:50, agitation speed of 500 rpm at 20 ºC for 1h. The XRD analysis revealed that main mineral compositions of the residues from 1M and 6 M H2SO4 leaching are gypsum (CaSO42H2O) and hemigypsum (CaSO40.5H2O), respectively.INTRODUCTIONThe term rare earth elements (REEs) denote the group of 17 chemically similar metallic elements consisting of the 15 lanthanides, plus yttrium (Y) and scandium (Sc) in the periodic table. These metallic elements are generally divided into two sub-groups; (1) the cerium sub-group, which is called “light rare earth elements” (LREEs), includes the elements from lanthanum (La) to europium (Eu), and (2) the yttrium (Y) sub-group elements, which are known as “heavy rare earth elements” (HREEs), are the remaining lanthanides from gadolinium (Gd) to lutetium (Lu) plus yttrium (Gupta and Krishnamurthy, 2005). The REEs occur together in the minerals due to their chemical and physical similarities, which cause problems of separating them from one another (Abreu and Morais, 2014). Although REEs are relatively abundant in the earth’s crust, they are typically dispersed and rarely occur in concentrated forms, making them economically challenging to obtain. There are a wide variety of rare earth minerals, however economic deposits of REEs in mining are bastnasite (La, Ce) FCO3, monazite (Ce, La, Y, Th) PO4, and xenotime YPO4 (Binnemans, 2013; Xie, et al, 2014). Due to the unique physical, chemical, catalytic, electrical, magnetic and optical properties of REEs, demand of these elements has increased dramatically with the advent of new technologies in electronics, communication, construction, energy and medicine (Jordens, et al, 2013). Since 1990, China has provided about 90 % of REEs worldwide. Japan and United States are the largest consumers of China’s REEs. In 2010, China announced to reduce REEs exports to Japan, as it tended to increase prices of REEs and to reduce a stable supply of REEs in the world market (Humphries, 2013). Therefore, there has been an increase in recent exploration activities intended to reveal new commercial sources of REEs and to bring them into production (Antico, et al, 1996). Apatite is an important host mineral for REEs in igneous and metamorphic rocks and can be used as an ore for those metals (Kanazawa and Kamitani, 2006). Apatite is also the main source of phosphate fertilizers and phosphoric acid (Habashi, 2013; Emsbo, 2015)."

Jan 1, 2016

By D. Harrington

An explosive may be defined as a substance (usually a mixture) that has the property of yielding, readily to combustion or oxidation upon application of flame or shock, the combustion or oxidation process being rapid and accompanied by formation of such relatively large quantities of gases that much violence and pressure are produced; these, in turn, react against their surroundings and cause what is termed an explosion. By their very nature their action is destructive and disruptive of adjunct materials; hence, all explosives axe potentially hazardous and every type of explosive, irrespective of its relative safety, should be handled, .stored, and used with utmost care at all times and under all conditions. It should always be remembered that any explosive can be rude safe only in certain of its features and under certain conditions and circumstance. If these facts (and they certainly are facts) are always kept in mind, most of the modern types of explosives can be utilized advantageously in our industrial life and with a considerable degree of safety. That commercial explosives can be handled. safely is borne out in part by the fact that in the past 15 years at least 3,000,000,000 pounds of the commercial explosive dynamite have been transported by railroads in the United State, without injury to any person, insofar as available records indicate.

Jan 1, 1939

By S. Gopalakrisbn, J. D. Miller, M. Misra

Pilot-scale tests indicate that effective concentration of fine gold from Colorado River sands can be achieved by flotation with the air-sparged hydrocyclone. The fine sands (55% -400 mesh and 0.01 opt*) from gravity concentration were processed with a 5-cm air-sparged hydrocyclone (capacity ~1.0 t/h dry solids). Concentrate grades as high as 5.0 opt were produced in a single stage at a recovery of 75%. The separation efficiency was significantly better than that achieved in conventional batch flotation experiments.

Jan 1, 1987

By Cathy Aimone-Martin, Robert Brentar, Brent Meins, James Lauer

Vibration measurements of multi-story structures in an urban environment began 3 years ago in New York City during close-in blasting to adopt correct measurement procedures and recommend revised ground vibration limits that best characterize cracking potential in mid- to high-rise structures. Two broad categories of structures included non-engineered 4- to 6-story multi-family unreinforced brick masonry buildings, built from the late 1800s to around 1930, and post-WWII, high-rise structures of reinforced concrete and steel rising 12 to 40 stories. In-ground vibration limits similar to those recommended by the U.S. Bureau of Mines safe criteria, developed for 1- and 2-story residential structures, are imposed by the New York City Fire Code. However, these limits are often lowered by project engineers or New York Department of Buildings when blasting is planned near buildings designated Landmark or historic and classified as fragile or vibration-sensitive. Seismograph measurements to determine compliance with in-ground velocity limits currently employ geophones mounted to basement walls and on concrete slabs overlying rock. These measurements cannot be correlated with upper structure cracking potential and are often exceeded, resulting in excessive and costly downtime to allow post-blast damage investigations.

Jan 1, 2014

By P. Ashtari

Influence of Sr and Mn single and combined additions on the modification of the Fe-containing intermetallic compounds and mechanical properties of Fe-rich AA319 aluminum alloys have been investigated, The microstructural features e,g, morphology of the Fe compounds and eutectic Si has been correlated to the mechanical properties, It was demonstrated that modification of the Fe compounds from the platelet ß-Al5FeSi to the Chinese script a-Al8Fe2Si compound improves the UTS and elongation of the alloys by 10.3% and 83.9%, respectively. Moreover, the simultaneous modification of the Fe compounds and eutectic Si increases the UTS and elongation of the alloys by 15,3% and 116,1%, respectively.

Jan 1, 2005