Search Documents

By T. J. Haddon

For the mining industry in general and gold mining in particular the past few years have been difficult. Economic conditions coupled with growing regulatory restrictions have led to increasingly important reasons for managing assets wisely. From a public company perspective this makes sense because it ultimately provides for an increase in shareholder wealth and that's what we're in business to do. Regardless of whether a company is public or not, assets must be managed effectively to remain a player in today's competitive environment. So, as the CEO of a gold mining company, how do we manage our assets to ensure that we are satisfying our obligations to our customer --the Amax: Gold shareholder? Let me tell you. As I said earlier, the gold industry has been under siege. We have just completed our third year of average prices below $400 per ounce. When you add finding and/ or acquisition costs and development costs to operating costs, there isn't much of a profit margin left. What this environment provides, however, is opportunities for growth. In the gold business, you either grow or you die. Today, most North American gold producers have at most a 10 year mine life. We are all scrambling to add reserves so we don't become extinct. As prices remain low, opportunities for increasing diminishing reserves abound. This is the time for the opportunistic company to take advantages of market conditions and that's precisely what we, at Amax: Gold, have done!

Jan 1, 1992

By Yuting Xue, Todd Minoski, Daniel Su, Mark Van Dyke, Mark Mazzella, Steve Hicks, Zoheir Khademian, Nicole Evanek

Design of chain pillars in a longwall mine setting is one of the main engineering controls on the global and local stability of the mine. Challenges with roof instabilities, floor heaves, and pillar failures can be alleviated, to some extent, by careful design of gateroad configurations and chain pillars. This work investigates three different alternative designs in a three-entry-system gateroad configuration with an inter-panel barrier in a longwall mine in Virginia. Starting with the current design, which has faced challenges such as excessive floor heave, a geomechanical model developed in 3DEC software is validated by monitoring in-situ pillar stress in the yield, abutment, and barrier pillars in the mine. Next, three different chain pillar designs are evaluated in the model with a focus on roof sagging, floor heave, and pillar average stress. Results show the relationships among the alternative designs and instabilities associated with roof, pillar, and floor in the study mine, providing insights for optimizing longwall mining designs in deep-cover settings.

Feb 1, 2025

By Yoginder P. Chugh

Physical mine modeling and computer simulation techniques were used to conduct technical feasibility and economic analyses of selected pit configurations for surface mining a single coal seam dipping at about 15 percent to a maximum depth of 150 feet. The pit orientations along the dip (dipline), strike (strikeline) and along an apparent dip were studied. Physical models of each pit configuration were prepared first to delineate potential operational problem areas, and also to furnish a data base-for the computer simulation models. A dragline micro- simulation model (Fluor), a shovel/truck simulation model and a cost analysis model were used to conduct the study. Sensitivity analyses were conducted on variables such as the pit width, block length, and angle of dip. The productivity (Cu.yds/hour) of the mining equipment (dragline or stripping shovel) under each set of conditions was used to optimize the pit configuration. Both block length and panel width were found to significantly affect the dragline productivity larger size trucks were preferable to smaller size trucks in the pit configuration along an apparent dip. The cost analysis results showed that the pit configuration along the seam dip was more economical than the pit along the strike or along an apparent dip.

Jan 1, 1984

By Y. Numata

2-Mercaptobenzothiazole (MBT) is a heterocyclic-mercaptan and belongs to a group of molecules containing nitrogen and sulfur as electron donor atoms. MBT has been employed as a collector in flotation under names of MBT, Captax, and Flotagen. It is a major constituent of reagents in the 400 series of the American Cyanamid Co. Current applications of the reagent include flotation of copper minerals, tarnished or oxidized lead, copper activated sphalerite, and nickel ores. It has been used successfully in the flotation of pyrite in acid circuits in the range of 4 to 5.

Jan 1, 1990

By D. C. Lynch

While silica ores are generally considered to have little value, those with exceptionally low B and P content have, potentially, significant worth if processed for use in photovoltaics. The sales value of a cubic meter of ?run of the mill? silica sand is $42 per cubic meter (retail and delivered). If the silica sand can be used to produce solar silicon, its value after processing is $34,000 per cubic meter. Factors that impact the suitability of silica for direct processing in production of photovoltaics are examined.

Jan 1, 2010

By Thomas F. Evans

Niagara Mohawk, like other northern coal users, has long experienced increased costs during winter months as a consequence of handling frozen coal. These costs include operating labor, maintenance labor and materials, equipment rental, fuel for thawing, demurrage, carrying costs for excess coal inventory and substitution of higher cost electrical generation for lost production capacity. In the winter of 1978-79, limited trials of several coal freeze conditioning agents were made to determine whether these would provide a cost-effective means of reducing the problems of handling frozen coal. Since these trials were generally effective, the Coal Purchasing Department arranged that all rail coal received during the 1979-80 winter from early December to mid March would be treated at the mines during loading. Inasmuch as the treatment costs are themselves appreciable and our various coal suppliers made a variety of choices of chemical supplier and equipment for application, an attempt was made to evaluate the unloading performance of each train received at the two generating stations which use coal as an aid in the future optimization of the use of freeze conditioning agents. These stations also receive coal by truck but this coal is not treated because of the short hauling time. The C.R. Huntley Station in Tonawanda, NY has a capacity of 800MW and burns about 5000 tons of coal per day during the winter months. The coal is primarily from central and western Pennsylvania but one source in southern Kentucky is also used regularly. Most of the Pennsylvania coal is washed and the fines are thermally dried. The coal is delivered in unit trains made up of mixed cars of 60, 75, and 90 ton capacity. Time between loading and arrival at the station averaged five days and ranged from three to eleven during the 1979-80 winter season.

Jan 1, 1980

By P. Semberg, B. Bjorkman, C. Andersson

"In this study, the interaction between magnetite and the additive olivine was studied after oxidation as well as after isothermal reduction at temperatures in the 500º-1,300º C range. In the olivine sample, the forsteritic olivine particles react partly during the oxidation pretreatment to form magnesioferrite and vitreous silica along the particle corona. This breakdown of the olivine particles during oxidation liberates magnesium from the particles, which do not react until temperatures of above 1,150° C in reducing atmosphere. When the hematite in the sample is reduced, and when the temperature is high enough to allow solid-state diffusion at ~800º C, the magnesium of the magnesioferrite redistributes, so that the magnesium concentration approaches the same level throughout the iron oxide structure. For magnetite, this did not occur until 800° C. At 1,000° C, this magnesium reacts further with the silica in the glassy slag phase, which crystallizes into fayalitic olivine. At this temperature, the magnesium has diffused over distances of more than 600 µm from large olivine particles after 2 hrs reduction. From this point, the primary slag phase in the pellet, until melting, is solid fayalite. Upon reduction to metal, the metallization front concentrates the MgO in the remaining wustite, which can lead to MgO levels of up to 10 mole% locally. The melting point of the fayalite is raised from 1,145º C to a melting range of 1,238-1,264º C due to the MgO increase, as estimated based on phase diagrams, which were adapted to the pellets tested. Much of the olivine that remained unaltered in the oxidation process will be encapsulated by iron before the particles begin to dissolve in reducing conditions and, therefore, plays no role in the reduction before final melting of the particles occurs.IntroductionEnvironmental concerns have placed increasing pressure on the CO2-intensive iron and steel industries to further optimize operation toward higher productivity and more efficient use of reducing agents. At the same time, high demand for iron ore has caused new iron ores with lower grades and more complex chemistries to be taken into production. In Scandinavia, the main iron-bearing material is magnetite-based pellets with olivine, which allow operation with very low slag volumes and, hence, lower fuel rate ( Slagnes and Sundqvist Ökvist, 2008; Hallin et al., 1994). Both the MgO and SiO2 of the olivine, as well as the elements of the gangue of the ore, i.e., Si, Mg, Ca, Al and alkali, interact with the iron oxides to give the finished pellet product its final properties in the blast furnace. Tailoring the additive mix to be used for a certain iron ore for optimal product properties, therefore, requires detailed knowledge of the mechanisms by which these elements dissolve and react in the pellet."

Jan 1, 2015

By J. V. Belke

There are many aspects of a grinding mill supply and installation that can result in the mill not operating as intended, resulting in unplanned shutdowns, poor reliability and headaches for all concerned. One of the largest aspects, both in criticality and physical mass, is the mill foundation; if not properly specified, designed and constructed the mill operation will be compromised. The mill foundation starts with the design specification and loading diagram supplied by the mill builder. The importance of understanding these documents and how to apply them is critical for the start of any mill project. When designing mills with a GMD the foundation must be properly modeled to ensure that all interactions of the mill and motor structures, foundation and subsurface strata behave in the intended fashion. All of the previous efforts can be laid to waste if the construction of the mill foundation is poorly handled. Mill foundations are typically the most massive concrete structure at a processing plant and can easily cause nightmares for inexperienced contactors. This paper explores some of the more important facets along the lifecycle of the grinding mill foundation.

Feb 23, 2014

By Stephen Meyer, Véronique Falmagne, Antti Pyy

Kittilä Mine is a gold mine located in Northern Finland, owned and operated by Agnico Eagle Finland. The mining operation started in 2008 with open pits which were completed in 2012 and production moved to being entirely underground. The mine consists of four separate orebodies, currently producing 2.0 Mt per annum, and requiring over 19 km of development on a yearly basis. The principal mining method is long hole stoping with delayed backfill. Microseismic monitoring was initially tested in the Suuri orebody, in the mining block between the 625 and 500 Levels. A small monitoring system was installed in 2018, after observations of stress-induced damage to the rock mass and in anticipation of future mining at greater depth. The initial goals were to evaluate if useful data could be recorded and test the signal quality of different sensor types. Prior to 2018, ground control related challenges in the Suuri and Roura mining blocks had been mostly related to sill pillar mining and block movements between weak graphitic schist planes. Seismic monitoring in the Suuri deposit proved successful for following the rock mass response to mining and helped support the rationale for early monitoring in the new, deeper mining block in Rimpi at 900 m depth. Preliminary geotechnical characterization and testing of the rock mass showed areas of silicification in the footwall rocks with potentially continuous structural features. Seismic monitoring of the Rimpi area has confirmed the brittle behavior of the rock mass, with strong seismic responses induced by stress changes following blasting being recorded. In addition to this expected behavior, a large magnitude event associated with significant damage in the footwall drive development 90 m above the stoping was experienced. Initial investigation of the rockburst showed two separate damage locations with over 300 tons of displaced rock. The seismic monitoring system recorded this rockburst as a single seismic event with magnitude significantly larger than previously recorded. Additional analyses indicated a possible thrust fault event, triggering two separate strainbursts. Following this event, the ground support system was upgraded in parts of the footwall drives to improve dynamic capacity. Systematic calculation of the source mechanisms for the small footwall events down to Mw-2.5 has revealed potential planar structures in the footwall. Early implementation of microseismic

Apr 26, 2022

By Takuya Tani, Tomoyuki Aoki, Kazuo Saskai

"The authors have proposed a unique methodology in the evaluation of rock mass stiffness for the range of five to twenty meters ahead of the tunnel face, with the use of small diameter inclinometers. The inclinometers are installed at the tunnel crown at a pre-determined constant interval, about one meter behind the face in order to obtain the slight change in inclination angles during tunnel excavation. This method was developed as a daily control measurement such that the geological condition and its possible change ahead of the face could be detected early by monitoring the trend of inclination angles. This paper describes the results of the ground stiffness ahead of the tunnel face evaluated at a mountain tunnel construction site. This stiffness is evaluated automatically and displayed visually by the current system equipped with a result display screen, together with a warning message against a weak zone when detected. 1. INTRODUCTION In a conventional planning of a mountain tunnel, according to the Japanese Standard (JSCE, 2006) seismic refraction surveys from the ground surface are performed and tunnel support patterns are then designed based on the seismic wave velocity. The resolution of the velocity profile estimated by this technique, however, is poor as the tunnel overburden becomes large. The tunnel face is therefore often encountered by unforeseeable fault zones or weak rocks. Such an unpredictable weak zone occasionally causes the collapse of the tunnel face. Clarifying the forward geological information during excavation is hence significant for safe and well-controlled construction. According to the geological conditions ahead of a face, auxiliary methods typified by forepiling and face bolting could be preliminarily prepared for prompt execution. For the purpose of acquiring forward geological information, pilot borehole drilling and seismic reflection methods represented by the TSP (Tunnel Seismic Prediction) or HSP (Horizontal Seismic Profiling) are often applied. Those exploration are carried out near the tunnel face for a couple of days. This often leads to the interruption of excavation procedure and has no small effect on the construction schedule. On the other hand, some forward ground prediction methods in the daily control measurement have been developed e.g. by Schubert et al. (1995). In this method, tunnel displacements in the direction of the tunnel axis are used together with the crown settlements. We have tested their method at a few of our tunnel projects and found difficulty in the evaluation of the forward ground condition due to insufficient resolution of the axial displacement measured by a total station theodolite that we usually use for the tunnel measurement. We have therefore developed a novel methodology of monitoring slight inclination at the tunnel crown near the face, so that the forward ground conditions can be predicted and evaluated as part of a routine measurement procedure."

Jan 1, 2016

By S. S. Xue

A study has been conducted to remove iron from sulfate solutions using Eichrom's Dipex® extractant, P,P?-di(2-ethylhexyl) methanediphosponic acid (H2DEH[MDP]). It is found that this extractant can effectively and selectively extract ferric ion from highly acidic solutions. Ninety-nine percent of the iron was extracted from a simulated feed solution in a single stage. The separation factor, Fe/Zn, was 10 6 Over 90% of the iron in a loaded organic can be recovered in an aqueous solution by reductive stripping. This paper presents the experimental results. A new process for iron removal using Dipex extractant is described and discussed.

Jan 1, 1999

By Edward C. Dowling

Froth flotation has been used industrially for many years to separate valuable minerals from their associated gangue constituents. Notwithstanding the significance of this process to the mineral industry, many of the engineering tools needed for the evaluation of the flotation process are still in a developmental stage. Traditionally, the evaluation of a flotation test on both the laboratory and industrial scale has been based largely on a rougher recovery value achieved at a sufficiently long time to give a "final" or "equilibrium" recovery value. The assumption of a time equivalency for a given recovery value between various flotation tests or between laboratory tests and industrial operation underlies this method of evaluation. Such an assumption implies that flotation is an "equilibrium engineering process" and that the time-rate recovery of material from the flotation cell is not affected by changes in the flotation system. It has been shown by Bushell (1) and by Klimpel (2) that the assumption of time equivalency between various flotation tests is unjustified. Consequently, better tools are needed for evaluating flotation test results, Because mathematical flotation models which incorporate both a recovery and a rate function can completely describe flotation time-recovery profiles, they can provide an excellent tool to evaluate flotation tests.

Jan 1, 1984

By Zhe Wang, Tichang Sun, Yongqiang Zhao, Zhengyao Li

Coal-based reduction followed by magnetic separation is a promising technology for processing refractory iron ores. Analysis of the size distribution of metallic iron particles is an effective approach for exploring the mechanisms of coal-based reduction. As metallic iron particles exhibit very complex shapes compared with nature ores, approaches for measuring and characterising metallic iron particles still remain unclear. In this study, the size of metallic iron particles was measured and characterised using manual point-counting by optical microscopy technics and image analysis. The effects of CaF2 as an additive on behaviour of the particle size distribution (i.e. mean size and cumulative frequency distribution) were also determined. Results indicate that CaF2 had a significant effect on the growth of metallic iron particles. When the CaF2 dosage increased from 0 to 8%, the mean size according to the area method increased from 9.03 to 16.28 μm, while the promotion effect gradually weakened and generated many fine metallic iron particles when the CaF2 dosage exceeded 8%. An increase in CaF2 dosage also led to cumulative distribution curves gradually shifting to the right mainly because F− can break the crystal structure of each phase during the carbothermic reduction of titanomagnetite and reduce the stability of the lattice structure, thus reducing the melting point and accelerating the growth of metallic iron grains. Compared to different measuring methods, the area method is a more suitable approach for measuring and characterising the size of metallic iron particles.

Sep 18, 2020

By R. R. Bhappu

The metallurgical processing industry is in the midst of a hydrometallurgical revolution with mining companies looking for low cost, energy efficient and environmentally friendly technologies for processing sulfide ores and concentrates. The Intec Copper Process is one of the most advanced of these technologies and has the potential to greatly change the way copper concentrates are treated in the future. Specifically, the Intec process has the ability to be built on a small scale and at mine sites while being much more economic than conventional smelting. Intec Copper has constructed a I tonne of copper per day demonstration plant in Sydney, Australia designed to verify metallurgical and operational performance. This paper will describe the process in more detail and provide preliminary results of the process based on the plant's performance. It will also compare the Intec Copper Process with a few of the competing technologies and evaluate the relative economic merits of Intec.

Jan 1, 1999

By R. Flores V.

The Refugio Project is located in the Andes mountains of northern Chile, in the Maricunga Hineral Belt. The Verde, Pancho and Guanaco Au ore bodies have been recognized and are Early Miocene in age. Mineralization consists of fine-grained disseminated gold hosted by dacitic to dioritic subvolcanic porphyries. Alteration comprises massive to veinlet-controlled silicification overprinted on earlier potassic and propylitic alteration and surrounded by later argillic alteration. Geologic relationships suggest that the Refugio ore bodies occur within an Early Tertiary volcano overlying Paleozoic and Mesozoic basement rocks. Geologic reserves at Refugio are approximately 300 million grams (9.7 million ounces) of gold. The Verde deposit contains 185 million tonnes of mineable and leachable ore with an average grade of 0.90 glt Au and is currently in the pre-construction stage after a positive feasibility study. Proven and probable mining reserves at Verde deposit are 101 million tonnes at an average grade of 1.02 glt gold. Additional potential exists at to other deposits within the Refugio property (Pancho and Guanaco) but no reserves have yet been defined due to the limited drilling to date.

Jan 1, 1994

In some 35 years in the explosive and construction industries, we have probably seen the extremes in drilling and blasting practices. All of you know that it is difficult if not impossible, to separate those two operations, particularly when we are searching for an optimum. Early in my career while doing engineering survey work for a major cross country pipeline construction company, we witnessed some of the wildest blasts imaginable, and they were wild for two reasons. Both the drill hole diameter and the drill pattern were such that when loaded with explosives, the factor in terms of pounds/cu. yd. was extremely high. On the other hand, in 1982 we were in an underground coal mine where just one hole was used to prepare a 20-ft. (6m) wide face. In this coal mine, they brought down 7 yd3 (5.35m3) of coal with just 1 lb. (.454 kg) of explosive. On a pipeline job, the factor may only be .1 yd3 per lb. of explosives. Open pit operations in the same geographical area may have almost the same type of rock, but similarities could end there. The several operators over the years will have installed different size-crushing equipment or variances will exist in other facets such as drilling, excavating, and haulage units. That tells us that if we have 6 operations in that common area, we are faced with a fact that there is an optimum drilling and blasting pattern for each of those six. We start at that point. We will note from Figure No. 2 the simple equations for determining either cu. yds. or tons. For a rock heavier than limestone, it may require only 10 to 11 cu. ft. (.31m3) for 1 ton (.91 metric ton) of rock in place.

Jan 1, 1986

By Tsunenori Shibata, Kazunori Jimbo, Kaoru Toya, Masafumi Matsumoto, Jinya Mitsui, Toshinori Tsuji

"The Toyama Line is under construction as a natural gas pipeline with a length of 102 km from Itoigawa City to Toyama City. A 2 km-long tunnel section of the pipeline that we are reporting here is being excavated in a mountain area away from an urban area. The geology under which the tunnel is constructed is collapsible fractured breccia slate and Holocene sandy soil. Other tunnels of the Hokuriku Expressway and Hokuriku Shinkansen are in service above the tunnel route. From the starting shaft on Toyama side, conventional tunneling was adopted to facilitate excavating the base rock layer and maintain enough distance from the Hokuriku Expressway tunnel. From the starting shaft on the Itoigawa side, TBM tunneling was adopted to cope with the unconsolidated soft ground for a 200 m-long section from the shaft. We will discuss the issues and countermeasures concerning the selection of a tunnel route and excavation methods, performance of both conventional and TBM tunneling, and underground connecting of these tunnels excavated by conventional (NATM) and TBM methods under a high groundwater pressure up to 1.2 MPa. INTRODUCTION AND BACKGROUND The International Petroleum Exploration Teiseki Co., Ltd. (INPEX) is constructing a high-pressure main pipeline network in the Toyama Prefecture area to promote the use and reliable supply of natural gas as shown in Figure 1. As part of the Toyama line project that has a total length of 102 km, A-1 construction, the nearest section to the tunneling starting point in Itoigawa City, where various tunnel excavation methods have been adopted to install the high-pressure pipeline 20 inches in diameter, a 2 km tunnel in the A-1 construction section discussed here is to be excavated by conventional and TBM methods in a mountain area. Based on environmental studies such as traffic jams due to frequent traffic of heavy vehicles disposing of soil, and the effects of dust, noise and vibration during construction, we selected the route to install the gas pipeline in a mountain area as shown in Figure 2, where the impact on the surrounding environment could be minimized."

Jan 1, 2016

By T. McCallum, C. Cooper, C. Lewis

"Following a successful solvent extraction (SX) pilot plant demonstration in 2016, CYANEX® 600 has been used commercially to recover molybdenum from concentrate leach liquor since March 2017. Consistent with the piloting results, the commercial plant has achieved expected molybdenum recovery, a significant reduction in base consumption, and a step-change improvement in physical performance. This paper highlights the optimization activities and plant performance from startup through normal operation. INTRODUCTION Historically, tertiary amines have been the most commonly used extractant for molybdenum and rhenium recovery. Tertiary amines utilize a protonation of the ligand combined with an anion exchange mechanism as proposed in equations (1-2). R3N + H+ ???R3NH + (1) 4R3NH+ + Mo8O264- ???(R3NH+)4Mo8O264- (2) In addition to the transfer of Mo and Re, the amine transfers acid and other anions to the strip liquor. The undesirable transfer of other species increases base consumption during stripping, impacting operating costs. Additionally, lower solubility of molybdenum-amine complexes create operational challenges, resulting in diminished physical performance. The presence of colloidal Si in concentrate leach liquors and the formation of Pickering Emulsions adds to the physical plant performance challenges with the amine extractant. CYANEX® 600, a phosphinic acid-based extractant, uses a cation exchange mechanism proposed by equation (3). 2(H2R2) + MoO22+ ??(HR2)2MoO2 + 2H+ (3) The cationic exchange mechanism eliminates chemical transfer of acid and other anionic impurities and avoids some of the physical challenges of the amine extractant (emulsion and 3rd phase formation). In addition, the reagent offers high selectivity for Mo over Re, which may be desirable in certain moly pressure oxidation (POX) flowsheets. Alkali base (ammonium hydroxide, sodium hydroxide, sodium carbonate, etc.) is used for molybdenum stripping. Therefore, reduction or elimination of chemical acid transfer from extract can significantly reduce base consumption in strip. Extensive laboratory and pilot plant testing was completed to evaluate the use of CYANEX® 600 for molybdenum recovery at Freeport-McMoRan’s Bagdad operation. The test work and pilot demonstration confirmed a reduction in base consumption along with improved selectivity. In addition, the evaluation highlighted a step-change improvement in physical performance over the incumbent amine extractant. The decision was made to convert the commercial operation to CYANEX® 600 in March 2017."

Jan 1, 2018

By G. Wang, L. R. Moore, Y. Xiong, A. Gorken

"Feldspar may represent 60% of the earth’s crust, but the mineral does not exist in a pure state. It is generally associated with quartz, iron containing minerals such as mica, and titanium containing minerals such as rutile. Feldspar itself can even exist in different lattices from those containing sodium to various divalent ions. As such, its beneficiation has become a process that is quite complicated, when compared to the beneficiation of other minerals. It is this complexity that has made feldspar beneficiation dependent on tailor made reagents to successfully yield the industry specified feldspar products. This paper reviews the mineralogical variations found from two feldspar mines and the development of collectors to specifically target a Fe+Ti concentration of < 0.05%, while maintaining a feldspar recovery greater than 90%.INTRODUCTION Feldspar is a class of aluminum silicates that represents the most abundant mineral on earth (Vidyadhar, 2002). This class of mineral can further be defined by the types or concentrations of cations present in the crystal lattice, MAlSi3O8 (M = K, orthoclase and microcline; M = Na, Albite; M = Ca, Anorthite) (Heyes, 2012). Feldspar is most commonly mined and processed for either the alumina used in glass making or for the alkalis used in ceramics. The alumina is known to improve product hardness, durability, and resistance to chemical corrosion. The alkalis act as flux, which means they not only lower the melting temperature of the mineral mixture, but also assist in connecting the other components present in ceramics. It was stated in 2009 that the world production of feldspar was in excess of 18.3 million tons produced, with the majority coming from China, Italy, and Turkey. Raw feldspar is typically found with quartz, iron containing minerals such as mica, titanium containing minerals such as rutile, and other feldspar minerals (Celik, 2001). Quartz is stated as having a negative effect on the melting point of the feldspar, whilst the iron and titanium based mineral contaminants are known to have negative effects on color and quality of the product produced; these mineral contaminants must therefore be separated prior to application (Barns, 1997 & Celik, 2001). Mineral flotation and magnetic separation are the most common means of feldspar beneficiation. However, conditions vary greatly from mineral contaminant to mineral contaminant (Table 1). Some literature suggests feldspar can be floated with an amine, but this generally yields lower than acceptable recoveries and often still requires the magnetic separation of iron based minerals to achieve the desired grade (Amarante, 1997 & Heyes, 2012). Feldspar flotation can be further improved with the addition of hydrofluoric acid (HF) as a silica depressant and feldspar surface modifier, but this still requires the removal of some of the other minerals first. Rutile can be floated with an anionic fatty acid collector at a pH range of 7-8 or with a petroleum sulfonate at a pH < 4 (Celik, 2001), while mica can be floated with a petroleum sulfonate or an amine collector at a pH of 3.5."

Jan 1, 2016

By Bruce E. Kester

The concurrent observation of the phenomena of alkali-aggregate reaction and development of efficient dust collection systems on rotary kilns made mandatory the process modifications to produce low alkali cements. Such low alkali cements as were available prior to these recent changes were either the rare result of low alkali raw materials, the loss of alkalies with dust loss from the kiln stack, or in some instances by the use of CaC12 in the kiln feed to increase volitalization of the alkalies and their subsequent loss or discard with the normal dust emission from the kiln. The relative slowly soluble nature of natural occurring alkali minerals in cement raw materials as compared to the very soluble nature of the alkali salts as present in cement kiln reaction products, either in clinker or in dust emission led to the development of the water leaching system. Two distinctly different problems of leaching are created by the two major systems of portland cement production, the wet and dry processes. The wet process is characterized by moderate dust loads and concentration of soluble forms of alkali in the dust. Further the return of the leached dust slurry does not materially change the characteristics of kiln feed on reintroduction and subsequently does not materially influence kiln operation. The dry process differs in high levels of dust emission and the presence in the collected dust of large amounts of unreacted kiln feed with its sparingly soluble alkali minerals. In the dry process reintroduction of leached dust slurry changes the dry process to what may be classed as a damp process. This change results in some decrease in kiln capacity and creates conditions of operation which are uncommon to either the wet or dry process.

Jan 1, 1963