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By E. Ben-Awuah, H. Askari-Nasab, N. Al Habib

Open pit mines are getting deeper with time and transportation expenses are increasing because of the increasing haulage distance. In-pit crushing and conveying (IPCC) is getting popular in deeper open pit mines as a suitable alternative to truck haulage because it offers a lower operating cost due to shorter haulage distance and less truck requirement. Semi-mobile in-pit crusher, currently the most popular IPCC system, is relocated every two to five years and the short-term plan needs to be updated accordingly. Short-term planning with IPCC is an area of research that has not been explored extensively yet and hardly any model can generate short-term extraction sequence considering an IPCC in place. This research work proposes a mixed integer programming model to generate short-term production plan within a time horizon of 12 months. The objective of the model is to optimally allocate shovels to minimize cost of material handling and maximize revenue subject to meeting plant requirement, maximum allowable tonnage variation and IPCC location constraints to achieve production and NPV targets set by strategic plan. The proposed model has been implemented in a hypothetical case study for verification. The comparison of results between scenarios with and without IPCC is expected to justify the use of IPCC in large open pit mines from a short to medium term perspective. A semi-mobile IPCC system with one relocation is considered in the case study.

Feb 1, 2023

By Carl N. Anderson

This paper, which is one of a group presented by the United States Bureau of Mines covering mining practice in different districts, deals in particular with the methods employed in mining the horizontal type of zinc-lead orebodies of the West Baxter area of the Tri-State zinc and lead district. These orebodies, locally termed "sheet-ground" bodies, with their relatively low working faces and low tenor of zinc and lead, require changes in the mining practice which is used in the balance of the district. In addition to the general mining methods, this paper also describes the system of hand sorting employed prior to the direct milling of the ore. The Hartley mine is situated mile west of Baxter Springs, Kans., as indicated in Figure 1. Its general outline and the location of the shafts with respect to the milling plant are shown in Figure 2.

Jan 1, 1932

By Carl Fisher

A microseismic roof fall warning system was field tested in Western and Eastern coal mines in the United States to better define the capabilities and limitations of the microseismic method in predicting roof fall Microseismic event and energy type data were obtained and documented for twenty-three coal mine roof falls. Field test results indicate that both the microseismic event and energy methods, as applied in this development effort, have a high probability for successful application in a commercially practical roof fall warning system. Based on the results and experience obtained during field testing, a commercial prototype microseismic roof fall warning system was designed and constructed. Although it was the intent of the program to field test the commercial prototype system, delays in equipment fabrication and contract time limitations did not permit field testing.

Jan 1, 1980

By T. Veneroso

Comminution is the process by which rock or ore is reduced to a size where the liberation of the minerals is maximized without changing the chemical or physical properties of the rock. Comminution accounts for about 50% of all energy that mines use (Klein et al., 2018). Growing demands from stakeholders and shareholders for sustainable development and CO2 reduction will increase the need for energy-efficient comminution. Ores with different mineralogical characteristics, textures, strengths, apparent stiffness, and mass-specific fracture energies need to be processed with a specific, new, and efficient comminution technology to increase the production and profitability of mining companies. The adoption of crusher wear parts built with new materials, such as non-Hadfield manganese steel or synthetic materials or traditional manganese with wear-resistant components, could improve the efficiency of the crushers, reduce downtime, increase production, improve the desired P80 (the aperture through which 80% of the product will pass) product size distribution, and enhance mine safety.

Feb 1, 2023

By Elaine R. Goetz

The eastern and southeastern regions of Ohio are part of Appalachian coal country, with a documented 1,300 miles of streams impacted by acid mine drainage. Since 1995, partnerships of local watershed groups, government agencies, private industries and universities have focused on the remediation of the affected waterways. The partnerships have spent $2.4 million annually in constructing and operating treatment systems such as wetlands, anoxic limestone drains, successive alkalinity production systems, steel slag leach beds, and lime dosers. These systems have been partially successful, and water quality has improved in the region, but no treatment has yet met the ultimate objectives of low cost, continual remediation. Our research group has investigated an alternative system for acid mine drainage treatment: electrolysis combined with pH adjustment to selectively precipitate recoverable, saleable metal products to offset the cost of treatment. We utilized a $0.2 million pilot facility constructed by the Wayne National Forest, in Nelsonville, Ohio, to process approximately 40 L/min of mine drainage from surrounding seeps. The pilot facility has multipurpose tanks that can be adapted to process acid mine drainage with different characteristics, such as high or low aluminum content or high or low carbonate concentrations. Sodium hydroxide was used for pH adjustment. Results indicate that saleable product can be generated with this treatment technology. In an initial, not optimized trial, one seep yielded 4.7 kg of dried 69 % iron oxide product from 10 hours of pilot facility operation, treating 23,000 L with electricity costs of $4.68. Dissolved iron concentrations in the water were reduced from 242 ± 18 mg/L to 2.8 ± 2.0 mg/L (99.0 % removal) while the initial acidity of 443 mg/L as CaCO3 was neutralized to an alkalinity of 160 mg/L as CaCO3. Iron oxide can be sold as pigment for an average of $1.47 per kg in the United States. Currently, approximately 80 % of all pigment used in the United States is imported, and there is incentive to purchase domestically recovered pigment. If the iron oxide from the pilot facility could be sold as pigment for $1.47 per kg, 43 % of the operating cost (raw materials, electricity and construction costs) of treatment could be offset by the sale of the product. Considering the cost offset, the electrolysis process shows potential as a sustainable low cost alternative for treating acid mine drainage.

Aug 1, 2013

By G Mendoza, L Taddei, C Robertson

It is a well-known industry fact that many mine sites around the globe experience blast holes that have elevated temperatures due to reactive ground, geochemical reactions, geomorphic reactions, sub level open fires, etc. It has been reported in many previous industry publications (White, Saputra et al) that elevated blast hole temperatures are a norm at many mine sites and are dealt with on an ongoing basis while the drill and blast process continues under elevated states of alert. The choice of explosives products including the type of detonator, primer, accessory, and bulk product all must be taken into consideration when loading blast holes where there is a known potential for elevated temperature conditions. It should also be noted that these elevated temperatures inside blast holes can be consistent throughout the blast hole length or vary in position within the blast hole depending on the mine conditions mentioned in the opening statement above. The risks associated with the loading of primers and bulk explosives in holes where elevated temperatures are present include: the ejection of the bulk explosives product, deflagration of the explosives during initiation of the blast or premature and uncontrolled initiation of a blast hole(s). This paper describes a unique and novel approach to the active monitoring of the temperature within blast holes with the application of electronic detonators with integrated temperature measuring (ITM) capabilities, deployed and demonstrated at a mine site in Mexico during January 2023 to March 2023. The deployment of ITM capable electronic detonators allows a user to monitor and record the real time temperature at multiple locations (e.g., toe, middle and collar) at each electronic detonator’s position within the blast hole, with the ability to react instantaneously to elevations in temperature reported. The measurement of the temperature can be for a single ITM electronic detonator, or simultaneous for an array of ITM electronic detonators, allowing for the potential for the data to be represented by a time/temperature/hole depth three-dimensional heat map. Changes in temperature at different borehole positions can vary, however, often the collar position is expected to be the most reactive due the combination of bulk explosive product and oxygen exposure, which can be confirmed with the application of the ITM electronic detonators. A mine site’s reaction to a measured elevated temperature within a blast hole can vary depending on the temperature measured and the change in temperature from the last measurement made. The study was found to produce real-time in-hole temperature information which was previously unobtainable without the introduction of another device. New and adaptive elevated temperature hole policies can be created and adapted proactively as real time information is received, leading to a safer mining operation.

Jan 21, 2025

By Alfred Bellows

The objective was to develop and design a remotely operated temporary roof support system that provides a safe working environment between a freshly cut coal face and the last line of permanently supported roof in which to perform roof bolting activities. This system was to take the form of a retrofit package which could be adapted to a majority of the bolters currently in service. The result is an integrated piece of hardware which will retrofit onto virtually all single boom bolters by removing the drill boom and relocating and modifying the hydraulic controls. The system is composed of three basic components: a pair of cantilevered roof supports which reach forward to the desired line of support, a mast type drill unit for drilling holes and torquing bolts, and a transverse slide assembly to carry the drill from hole to hole across the entry without relocating either the bolter or the roof support. A converted bolter not only provides a much safer operation but is projected to increase the production speed as well as reduce the manpower required to operate a section.. This improvement in productivity is shown to make the job of retrofitting a bolter cost effective in less than two months and provide a potential profit thereafter. Based on the findings presented in this report it is recommended that a proto-type be built and tested in an underground coal mine to demonstrate its effectiveness.

Jan 1, 1976

By Ivette E. Torres

The final 26 tables of this chapter, tables 24-49, extend and expand the statistical series on production that was started in the 1963 edition of the Area Reports: International volume of the Minerals Yearbook and that was subsequently updated and expanded in the 1965 and 1967-90 editions. Mainly because of the separation of the countries that formed the U.S.S.R., the column that provided a 5-year average, was deleted. The listing order of major producers, however, continued to be, as in past years, in descending order of rank in the year of review, 1992. This group of tables now includes each of the top 19 crude minerals in 1983 in terms of value of their output plus 3 others of lesser rank. These 22 crude mineral commodities accounted for more than 98 % of the estimated value of world crude mineral production in 1983. Coal production, included in the 1983 value of production table as two separated entries, is combined in one table in this chapter. World output of five key downstream products-aluminum, steel, cement, nitrogen in ammonia, and refined oil-is included because of their significance as products.

Jan 1, 1994

By J. P. Smith

DURING the winter of 1946 to 1947 potash operations in Germany, France and Spain were visited by the author. The U. S. Department of Commerce, through its Field Intelligence Agency Technical, sponsored the inspection of German properties by representatives of the American Potash Industry. Vast deposits of potash salts occur in the above-mentioned countries, and the production and marketing of these salts are important in their respective economies. Germany An area of some ten thousand square miles of north central Germany is underlain by lenses of potash salts interbedded with greater thicknesses of sodium chloride or common salt. These salt measures are relict deposits from the evaporation of an arm of the Zechstein (Permian) Sea which formerly covered northern Europe. During the peak year of 1940, Germany produced salts containing 1,860,000 tons of K,O and, next to coal, potash is its most important mineral raw material. The German deposits are found in the synclinal areas flanking the Flechtingen, Hartz Mountain and Thuringian uplifts, and, to the north they lie beneath the plains of Hannover. The regional dip of the salt measures is to the north, hence, lenses of potash found at depths of 500 to 1000 m* in the Two general depositional provinces of potash concentrations are recognized: (1) the main basin, and (2) the subsidiary basin. The main basin embraces the producing districts of Stassfurt, Sud Hartz, Halle, and Hannover-Braunschweig. From J. P. SMITH, Member AIME, is Clzief Geologist, U. S. Potash Co., Carlsbad, New Mexico. New York Meeting, February 1948. TP 2621 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received May 26, 1948. 40 to 45 pct of the German production comes from this basin. The Werra-Fulda district, producer of 50 pct or more of the German potash, lies in the subsidiary basin. Salt Stratigraphy: Five commercial potash zones or beds are recognized. The three beds present in the main basin are: The Stassfurt bed that occurs near the top of the older salt, and the Ronnenberg and Riedel beds in the younger salt occurring 200 to 300 m, respectively, above the Stassfurt bed. The Stassfurt bed is found over the entire main basin, while the Ronnenberg and Riedel are restricted to the Hannover area. Two potash beds occur in the Werra basin. These are the Thuringian, or lower bed and the Hessen or upper potash bed that occurs 60 m above. These beds lie 210 and 150 m, respectively, below the position of the Stassfurt bed, and are restricted to the Werra basin. The upper bed is 2.5 to 11.5 m in thickness, and the principal ores are hartsalz with varying thicknesses of carnallite. The lower bed averages 3 m in thickness, and contains hartsalz. The upper bed is presently being mined, and the

Jan 1, 1951

By J. P. Smith

DURING the winter of 1946 to 1947 potash operations in Germany, France and Spain were visited by the author. The U. S. Department of Commerce, through its Field Intelligence Agency Technical, sponsored the inspection of German properties by representatives of the American Potash Industry. Vast deposits of potash salts occur in the above-mentioned countries, and the production and marketing of these salts are important in their respective economies. Germany An area of some ten thousand square miles of north central Germany is underlain by lenses of potash salts interbedded with greater thicknesses of sodium chloride or common salt. These salt measures are relict deposits from the evaporation of an arm of the Zechstein (Permian) Sea which formerly covered northern Europe. During the peak year of 1940, Germany produced salts containing 1,860,000 tons of K,O and, next to coal, potash is its most important mineral raw material. The German deposits are found in the synclinal areas flanking the Flechtingen, Hartz Mountain and Thuringian uplifts, and, to the north they lie beneath the plains of Hannover. The regional dip of the salt measures is to the north, hence, lenses of potash found at depths of 500 to 1000 m* in the Two general depositional provinces of potash concentrations are recognized: (1) the main basin, and (2) the subsidiary basin. The main basin embraces the producing districts of Stassfurt, Sud Hartz, Halle, and Hannover-Braunschweig. From J. P. SMITH, Member AIME, is Clzief Geologist, U. S. Potash Co., Carlsbad, New Mexico. New York Meeting, February 1948. TP 2621 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received May 26, 1948. 40 to 45 pct of the German production comes from this basin. The Werra-Fulda district, producer of 50 pct or more of the German potash, lies in the subsidiary basin. Salt Stratigraphy: Five commercial potash zones or beds are recognized. The three beds present in the main basin are: The Stassfurt bed that occurs near the top of the older salt, and the Ronnenberg and Riedel beds in the younger salt occurring 200 to 300 m, respectively, above the Stassfurt bed. The Stassfurt bed is found over the entire main basin, while the Ronnenberg and Riedel are restricted to the Hannover area. Two potash beds occur in the Werra basin. These are the Thuringian, or lower bed and the Hessen or upper potash bed that occurs 60 m above. These beds lie 210 and 150 m, respectively, below the position of the Stassfurt bed, and are restricted to the Werra basin. The upper bed is 2.5 to 11.5 m in thickness, and the principal ores are hartsalz with varying thicknesses of carnallite. The lower bed averages 3 m in thickness, and contains hartsalz. The upper bed is presently being mined, and the

Jan 1, 1951

By P. Chryssanthakis

"The aim of this study is to give a better understanding for the behaviour of the composite material RRS and the synergy effect of its components while an excavation face advances during tunneling. The lower part of the Q-system scale (i.e. Q-values of 0.4 or less for a tunnel span of 10 m) is commonly used for tunnel applications in softer rock formations. This is implied in the updated Q-chart, where the dimensioning of rock support by using Reinforced Ribs of Shotcrete (RRS) comprises the use of fiber reinforced shotcrete, S(fr), rock bolts and shotcreted beams reinforced with steel rebars. The newest part of the lower part of the Q-chart gives more information on required support compared to older Q-system versions. The RRS can be described as a three part composite material with the 3rd component being the reinforced concrete beam (with no fibers) and with steel rebars that are placed in groups and sprayed with pneumatically applied shotcrete in layers. The RRS usually have a total thickness varying between 0.20 and 0.50 m and a width between 0.25 and 0.60 m respectively. They are placed at different distances along the tunnel axis depending on the rock mass quality.In this study, an effort has been made to simulate RRS in a numerical code and to study its behaviour in a real case study. A double tube, triple lane tunnel in weathered tuff has been chosen as reference material for testing RRS under changing load. This was done by writing a FISH subroutine which is implemented in the discontinuous numerical code Universal Distinct Element Code (UDEC) that takes into account the characteristics of the reinforced beam (height, width, number of steel rebars, and spacing) in the calculation process. A rather extensive parametric study has been run by using the same key input rock mass parameters and in situ stresses (?H/?v = 0.5), and changing only the RRS key input parameters in order to simulate the changing loads during the face excavation process in tunneling. The critical part of this numerical modelling work that is under investigation in this study, is the period just after the excavation: i.e 4 hours; 16 hours after excavation; and finally after the complete curing (28 days) of the cement based components. The RRS are installed in three consecutive steps as follows: a) the application of S(fr) on the rock surface in order to even it up for the RRS application; b) radial bolting; and c) installation of the reinforcing steel rebars, that form the reinforced beam.The results show that: a) the S(fr) is almost overloaded in the first 4 hours after application in both tunnels; b) on the S(fr) and steel rebars, the stresses, axial and shear loads, as well as moments are decreasing after 16 hours and 28 days; and c) the bolts loads are increasing with time. The observed reduction in stresses and loads in both the S(fr) and reinforced beam is surprising. A possible explanation is that tunnel arches are self-supported to a lesser extent as time progresses due to redistribution of stresses in the nearby rock mass through the radial bolts. The loads in radial bolts are increasing by almost 25%. This can be seen from the overall results as we observe an increase in maximum principal stresses, displacements, shear displacements along joints and hydraulic apertures compared to tunnel baseline that has no support of any kind."

Jan 1, 2015

By John R. Goode

The fist part of this paper comprises a general discussion on refractory gold ores and their mineralogy. Processing methods now in use are reviewed and test procedures for the important methods are discussed. Factors determining the selection of a process are outlined. The paper includes a survey of production plants using pressure oxidation, roasting and biological oxidation to process refractory ore. The status of processes under development is briefly reviewed.

Jan 1, 1993

By GEORGE S. POPE

The Government of the United States annually expends approximately $8,000,000 for coal. Prior to 1906, this coal was purchased mainly on its reputation, or trade name, rather than under contracts specifying the heating value of the coal, its contents of ash and moisture, and other characteristics. In recent years the Government and other large consumers have appreciated more and more the desirability of definitely determining by chemical analyses and tests the character and quality of coal deliveries. Studies of engineroom and boiler-room efficiencies and economies point forcibly to the necessity of devoting more thought to the improvement of the boiler-room end. This calls for an intimate knowledge of the character and quality of the fuel used. The analyses and tests of delivered coal furnish such data, whereby the power-plant results can be comprehensively studied and a continuous check maintained on the coal and on the conditions of plant operation. Under authority of acts of Congress making appropriations for analyzing and testing fuels belonging to or for the use of the United States Government, a laboratory is maintained at the headquarters of the Bureau of Mines, 'Washington, D. C., where samples representing deliveries of coal purchased for Government use are analyzed and tested. The Bureau of Mines has been active in promoting the purchase of coal by the Government under specifications,a that is, by contracts that make definite requirements regarding the heating value of the coal expressed in British thermal units, and the composition as shown by proximate analysis. Under many of these specification contracts the bidders guarantee the quality of the coal offered, and that guaranteed by the successful bidder becomes the standard of his contract. The deliveries are sampled and analyzed to determine whether the coal is of the quality guaranteed by the contractor, and if it is not, the price to be paid is decreased in proportion; but if the coal is of higher quality, the price is proportionately increased. The analyses of coals or the guaranties offered by the bidders may in some cases be used for determining the award of a contract, but if the coals offered are of widely different characteristics, it may be difficult to determine which would prove the cheapest fuel, without having their relative evaporative efficiencies. Hence, it may be necessary to conduct evaporative tests under service conditions in order to determine which coal is particularly adapted for successful and economical use in the plant. But once a particular coal is selected, regardless of the procedure of determining the award of a contract, then the greatest advantage of the specification method is realized, in that it guarantees to the purchaser coal of uniform character and quality, thereby insuring the best results in the plant. The cost of purchases, under such specifications, of fueL for the use of the Government now aggregates approximately $4,100,000 per annum, and the cost of additional fuel for which samples are taken and submitted to the Bureau of Mines for analysis in order to determine the character and quality of the coal furnished aggregates some $3,000,000 per annum more. In fact, information is now obtained regarding the quality of fuel purchased on almost all contracts large enough to warrant sampling and analysis and heatingvalue tests. It may be said that the purchase of coal under specification is still in the stage of evolution and development, and that experience and time will be required to develop fully standard specifications and methods of sampling and analysis that may be used universally. The success of the specification method for the purchase of coal and the reliability of analysis now rests almost entirely on sampling.a Improper sampling leads to controversies, with resulting condemnation of the specification method in general and the value of laboratory analyses and tests in particular. A number of bulletins b have been published showing the results of the purchase of coal by the Government under specifications. This bulletin is a compilation of the results of coal analyses by the Government, including those made in the fiscal years 1908-1910, which have been previously published, and later analyses representing coals analyzed during the fiscal years 1911-1915. The preceding bulletins gave individual analyses or average monthly analyses, whereas this bulletin gives the average analyses only for the entire contract. It is hoped that this compilation of a series of comparable analyses of the coals purchased by the Government will be of value alike to coal dealers, to Government engineers, to purchasing agents, and to all users of coal.

Jan 1, 1916

By Glan Dale Coe, Will H. Coghill

"INTRODUCTION Characteristics of the mineral products that are to be contrasted herein are net energy consumed in grinding to a definite top limiting size and size distribution of the ground products. Through contrasts of these grinding characteristics this paper to differences in related mineral products and in unlike products. For,example, in anticipation of the data to follow, although Joplin chert and Babeitt taconits are running mates in defying the best crushing machinery, taconite is one of the more docile and Joplin chert one of the most refractory mineral products ordinarily fed to a grinding Routine grinding tests on a great variety of ores and tabulation of the findings have been covered amply by Lennox 4/ and by Maxon 5/ and associates. In contrast to that work, the present paper is a record of random tests made over a long period, and each answered an urgent question at the moment. The work was done at the Southern Experiment Station of the Bureau of Mines, United States Department of the Interior, in cooperation with the University of Alabama. The pressure of war work prevented an organized investigation."

Apr 1, 1943

By William H. King

The Mining Division of the Bureau of Mines has been making development and power-requirement studios for the Interior Department Missouri Basin Field. Committee. The studies by the Bureau of Mines have as their purpose, the summarizing of potential future power requirements in the mineral industry within the Missouri Basin, special emphasis being placed on possible expansion of these requirements through construction of new mining enterprises by private capital. In preparing a proposal for the production of phosphoric acid from phosphate rock near Lander, Wyo., it was found that the basic information on the physical and chemical characteristics of the phosphate beds to determine costs of mining and treatment was lacking.

Jan 1, 1949

By James R. Finlay

SOMETIMES the American Institute of Mining and Metallurgical Engineers appears to be a strictly technical society, and if so my paper should deal with the technical operations of finding and producing gold. But I have not been engaged in practical gold mining since 1910. What I have been doing mainly for the last 25 years is to try to figure out what the income of a mine-gold, or any other-would be worth to an investor. So I have been in contact with gold mining, in one way or another, for some 42 years. I say that with a purpose. In that time the industry has been through one of its cycles. The value of gold mines is very dif¬ferent at different points in one of these cycles. For this reason the cycle of gold mining is a highly practical study. The cycle of gold is a more or less rhythmical change in its value or purchasing power measured in the grand average of other commodities, which is the same thing as its price. Taking the price of gold in 1897 as a stand¬ard, a dollar at that time declined to 55c. in 1913, to 26.8c. in 1920, and since then it has risen to 47.4c. in 1930 and to about 64c. in 1932. It is worth about 21/ times as much now as it was in 1920, but less than two¬thirds as much as in 1897. When this change is applied to the value of a gold mine still greater differences are evident. The cost of producing gold is of course measured in labor and other commodities, and as that cost goes down the margin of profit measured in the gold itself increases. Thus a mine which might make only 10 per cent profit in 1920 on its gross output may easily make 25 per cent now. The stockholder gets 2 ½ times more gold that is worth 2 1/2 times as much in purchasing power, therefore the real value of his holding has risen by the one factor multiplied by the other, giving him 61/4 times as much goods or services as he had in 1920.

Jan 1, 1933

By W. K. Katoto

The Tailings Leach Plant (TLP) of the Zambia Consolidated Copper Mines Ltd at Nchanga was commissioned in 1974 as the world's first largest successfully operated agitation leach SX-EW plant for copper production. The design, development and operation of the plant soon after commissioning termed as TLP stage II has been described by Holmes et al (1). In the initial phase the plant was designed to treat approximately 25,000 tonnes per day of current flotation tailings from the Nchanga Concentrator operations, and had a design capacity to produce 75,000 tonnes per annum of copper. Further expansion took place in 1986, and TLP stage III was successfully commissioned to treat an additional 20,000 -25,000 tonnes per day of large reserves (over 135m tonnes) of reclaimed tailings stockpiled since the early 1930's in four paddock type tailings dams (T01 to T04). The salient feature of TLP stage III expansion described by S K Mwenechanya et al (2 -3) was the installation of the world's largest Horizontal Belt Filter Plant with 26 filters each of 80 m2 filtration area.

Jan 1, 1996

By M. Omneya El-Hussaini, M. Galal Abdel Wahab, A. Mohamed Mahdy, M. Fathi El Shahat

"Um-Gheig Pb-Zn representative sample was collected from the main adite of Um-Gheig mine along Red Sea coast in Egypt. Its chemical analysis shows that, it contains 16% Zn, 2.3% Pb, 0.1% Cd and 0.1% B as elements of interest. After H2SO4 agitation leaching, the chemical composition of Um- Gheig Pb-Zn deposit ore residue assayed 2.26% Pb, 1.5% Zn, 0.085 % B and 0.01% Cd. This residue was leached with 20% HCl at S/L ratio 1/3 for 2 hours at 80o C. The chloride leach liquor was found to contain 18 g/l Pb, 13 g/l Zn, 0.1 g/l Cd and 0.75 g/l B with the corresponding leaching efficiencies 79%, 86%, 100% and 88% respectively. Lead was precipitated from the chloride leach liquor using concentrated H2SO4 to obtain PbSO4 with precipitation efficiency 96%. While boron was extracted by the selective resin Amberlite IRA743 with loading efficiency of 80%. On the other hand, Zn and Cd were precipitated using Na2S solution with 86.7% and 100% recovery efficiencies respectively.IntroductionLead-zinc ores occur in Egypt in a few localities distributed along the Red Sea coast between El Qusier and Ras Banas.. The reserve of the ore deposits has been estimated by Soviet and German Experts to be about 900,000 Ton, [1]. Onal et al., [2] recovered Pb from Pb–Zn sample containing 10.17 % Pb and 10.98 % Zn from Aladag in Turkey. El-Sherif, [3] used concentrated HCl to leach lead-zinc ore sample of Um-Sammuki area in Egypt. As had been published by Marchioretto et al. [4], the precipitation of heavy metals with lime, sulfide and caustic soda are the most commonly used economical methods for metal removal. Lead was completely removed from the anaerobically digested sludge by precipitation with NaOH and Na2S at waste water treatment plants. Mahdy et al., [5] had successfully extracted boron from the bittern solution of Qaron Lake by ion exchange resin (Amberlite IRA 743). The extraction of boron was mainly depending upon the pH value. They extracted 80% of boron at pH 5.5, while at pH 3.9 about 74% boron was extracted. Kabay et al. [6] used the column-mode to remove 1g/l of boron from geothermal waste water of the Kizildere-Geothermal Power Station by using Diaion CRB 02 Nglucamine- type chelating resins. In this regard, the removal of lead, cadmium and boron allow the residue to be reclassified as non hazardous material."

Jan 1, 2008

By J B. Pateiro-Fernßndez, M Fuchsschwanz, G Deissmann, S Ahmad

Quang Ninh province is the most significant coal producing region in the north of Vietnam. The Nui Beo Coal Company (NBCC) operates open pit mines in Quang Ninh province and the waste overburden produced is dumped as sidehill fill in the Chinh Bac waste dump site. In the past, no special consideration has been given to the environmental impacts during the dumping operations. The mine waste is transported in trucks from the mine site to the dumping site and the loading, hauling and dumping operations are the main sources of dust emission. Rain events, especially during the rainy season, cause the generation of contaminated surface water as well as seepage of water through the dump at a number of locations. The existence of cracks at the top of the dump can pose serious stability problems with heavy rain events leading to failure of slopes which can potentially disturb vast areas of land currently used for agricultural purposes and human settlement. For sustainable development at the Chinh Bac waste dump site in the future, apart from the social and economic considerations to be assessed, there is an absolute necessity for environmental protection measures to be in place both during and after the dumping of the mine waste overburden. Research Association Mining and Environment in Vietnam (RAME) is working to develop and implement concepts for the mitigation of environmental problems emerging from dust, contaminated water and lacking slope stability to ensure successful rehabilitation of the coal mine waste dumps in the north of Vietnam. Within the framework of RAME, the Institute of Mining Engineering I (RWTH Aachen University) is managing and working to develop the stabilisation concepts for the solid mine waste dump in collaboration with the Chair of Geotechnical Engineering (RWTH Aachen University), Brenk Systemplanung and the Vietnamese group of companies VINACOMIN (Vietnam National Coal-Mineral Industries Group). The most economical and environmentally sound operations, as well as locally adaptable environmental protection measures, are essential for the sustainable development of the dumping operations at Chinh Bac dump. This paper describes the dump site and the sustainable development possibilities with an emphasis on the environmental protection and stability problems at the Chinh Bac waste dump.

Jan 1, 2009

By T Horberry

This paper describes work conducted as part of an overall research project that aims to identify current and emerging trends in the development of automation and remote mining technologies, and to provide an operator-centred perspective to the design and deployment of such systems. More specifically, this paper outlines a research and development process being constructed to examine the issues of the human skills and capabilities required to operate or maintain automated/remote controlled mobile mining equipment. A human factors approach presented here, where the focus is upon designing equipment interfaces to fit operators, rather than purely designing and deploying the equipment based on a technology-centred approach. However, the importance of training to help develop the required skills and capabilities is also emphasised.The work built on a full review of the literature about the effects of automation in mining and elsewhere, extensive interviews with different stakeholders in mining automation and recognised best practice regarding designing for the human element in automated systems. The human element research and development process proposed is in four stages, ranging from understanding the tasks involved in both manual and automated equipment operation through to user trials to ensure that the automated systems are acceptable to operators and are fit for purpose.It is outside the scope of the current paper to present the full data collection described by this four stage process, but an overview of what will be collected is presented. As such, this paper focuses on the broad methods being developed, rather than the specific results to be obtained. The ultimate objective is to apply a broad ‘technology assessment’ approach to the design, deployment and evaluation of automated mining technologies, in particular by considering the user-centred design implications of these technologies.

Nov 22, 2011