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By Paul F. Kerr

The uranium-producing areas near Marysvale, Utah provide an unusual group of veins and replacement deposits associated with a Pliocene-Oligocene intrusive and extrusive igneous complex. Aside from several scattered occurrences in older volcanics, uranium-bearing veins occur in intrusive quartz monzonite or granite and in nearby rhyolite. The veins in general trend within a few degrees of N55-65°E, are more or less vertical, and are found along the western margins of a quartz monzonite intrusive. The replacement deposits are found in rhyolite intrusives and also in the lower portions of rhyolite flows close to places where veins forming solutions rising through quartz monzonite came into contact with the rhyolite. At least 14 veins have been mined since uranium mining operations began at Marysvale some 16 years ago, while ore has been extracted through about 10 miles of underground workings. The active mmmg area is approximately 3000 feet long and in the central portion 1500 feet wide. Ores have been found over a vertical spread of almost 2000 feet, between the highest occurrence on Jungfrau Hill ( ca. 7200 feet) and the deepest discovery of ore in a drill hole near the Prospector Mine (ca. 5200 feet). Under existing mining and marketing conditions, mining along the veins in the main area appears to have reached an economic limit at about 500-600 feet below the surface. However, orebodies in the rhyolite and in the upper portions of veins close to the rhyolite are subject to active mining. Isotopic dates indicate that the uraninitebearing veins and replacements are late Miocene or Pliocene (ca. 10-13 m.y.), the rhyolite is Miocene (avg. 17.5 m.y.), and the quartz monzonite and granite intrusives are late Oligocene (23.6 to 26.0 m.y.). The older Bullion Canyon Volcanic Series which encircles the area is Oligocene (28 to 31.1 m.y.). Two generations of alteration occur: ( 1) Earlier alunitic veins and replacement bodies are found in the Bullion Canyon Volcanic Series, none of which have produced uranium; (2) Later argillic alteration associated with the introduction of uranium is found along veins, replacing glassy rhyolite dikes and replacing phases of the rhyolite.

Jan 1, 1968

By Unknown

We are indebted to the Consolidated Mining and Smelting Company of Canada Limited for the several pictures in this issue showing the progress made in new construction at Tadanac and at Bonnington Falls, B.C. Plate No. 1 shows the down-stream side of the No. 1 hydro-electric station of the West Kootenay Power and Light Co. Limited at Bonnington Falls, on the Kootenay river, at which 17,500 k.v.a. will be developed to meet the requirements of the enlarged zinc and lead refineries at Tadanac. The generators, exciters and switch apparatus were supplied by the Canadian General Electric Company and the step-up transformers by the Canadian Westinghouse Company Limited. The plant throughout was designed, manufactured and erected by Canadians. Plate No. 2 shows the casting room of the new lead-plant and plate No. 3 the new tank-room of the electrolytic lead refinery at Tadanac. These additions will increase the capacity of the lead refinery by some 150 tons daily. The extensions also include practically doubling the zinc plant, and the gold and silver refineries. With the exception of the new stack, all designing construction and installation of the various plants is being done by the Company's own employees at Tadanac. Plate No. 4 shows the progress made in the construction of the new reinforced concrete stack at Tadanac. This stack, of 21 feet inside diameter at the top, will be brick lined and when completed will rise 400 feet above its foundations: the tallest of its kind in the world. The construction is in charge of the Rust Engineering Company of Pittsburgh, Pa. Plate No. 5 is a panorama of the Tadanac smelters and refineries. The photograph was taken from the roof of the lead smelter and gives an excellent idea of the size of the plant and its surroundings. Tadanac produces electrolytic lead, zinc and copper and fine gold and silver. During 1924 the company treated 1,333,523 tons of ore and produced: 161,400,- 769 pounds of lead; 92,197,824 pounds of zinc; l,382,750pounds of copper; 4,074,044 ounces of silver, and 26,976 ounces of gold.

Jan 1, 1925

By A G. Tomkins

"Approximately 32 per cent of the worldÆs total gold production has come from greenstone belts (Frimmel, 2008). In these regions numerous small to large, quartz vein-associated gold deposits are hosted in metamorphic rocks ranging from lower greenschist to granulite facies. Historical exploration has focused its efforts on rocks of greenschist to lower amphibolite facies because these seem to contain the most deposits. Consequently, these regions are now reaching mature brownfield status, particularly in developed nations, and finding new deposits is becoming progressively more difficult. In comparison, regions of similar tectonic setting, but instead containing high-grade metamorphic rocks of mid-amphibolite to granulite facies, have received little attention because the first explorers found fewer deposits there.However, these regions certainly do contain some very significant gold deposits such as Hemlo in Canada (~684 t Au), Renco in Zimbabwe (~31 t Au mined since 1980), the very profitable Challenger deposit in South Australia (~36 t Au; recovered + reserve + resource), and the most recently discovered Tropicana deposit in Western Australia (~182 t Au; reserve + resource). High-grade metamorphic terrains therefore offer explorationists an excellent opportunity to move onto relatively green fields within developed nations, which carry less risk associated with development. However, explorationists have comparatively little knowledge of the geological controls on gold distribution in high-grade terranes because: industry experience lies in traditional lode-style gold deposits these gold deposits are geologically more complex there has simply been minimal research conducted on these deposits as a holistic group.Furthermore, research conducted in the last five years has shown us that the previous widely accepted model for gold genesis in these terranes, the crustal continuum model, is incorrect. Here, I discuss features from several gold deposits hosted in high-grade metamorphic rocks and use these to define two categories for explorationists to consider: deposits that formed prior to peak metamorphism (metamorphosed gold deposits) deposits that formed well after peak metamorphism (post-metamorphic gold deposits). There are none that have been convincingly shown to have formed during peak metamorphism.CITATION:Tomkins, A G, 2013. Gold deposits in highly metamorphosed terranes, in Proceedings World Gold 2013, pp 23-26 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Sep 26, 2013

By L Snyman, S Webster and J Samosir

Northparkes is located 27 km north of the township of Parkes in central New South Wales, Australia. The operations consist of underground block cave mines and an ore processing plant, which produces high-grade copper and gold concentrate. Production is currently sourced from the E48 Lift 1 block cave mine and open cut stockpiles. Mining commenced at Northparkes in 1994 as an open cut operation. Underground block cave mining began in 1997. Northparkes was the first mine in Australia to use the highly efficient block cave mining method, which relies on gravity and natural rock stresses to fragment and recover the ore.The E48 extraction level is located approximately 581 m below surface and consists of ten extraction drives and 214 drawpoints. Several stress measurements carried out at various levels in both E26 and E48 indicated a low underground regional stress environment and a low to moderate stress state in the E48 orebody. The E48 block used a post undercut strategy that was initiated in 2009, with production commencing in September 2010 and cave-through to the surface occurring approximately four months later in January 2011. Changes to the cut-off grade and fixed prices meant that Northparkes was able to develop an additional two drives to the north of the existing cave and one additional drive to the south. Modelling of the cavability of the additional drives indicated that the cave would not readily propagate to sufficient height of draw to mobilise the ore and tended to arch over to the existing cave shape. Based on modelling and a desire to reduce the risk of the project, preconditioning was adopted. Along with the requirement of safely hydrofracturing an active cave, a measure of the success of the preconditioning was desired. Few block cave operations have developed extension drives adjacent to existing caves.This paper summarises the lessons learnt from the development of this extension project, with a specific focus on ground control and monitoring.CITATION:Snyman, L, Webster, S and Samosir, J, 2016. E48 cave extension at Northparkes, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 111–118 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

"The Ohio River Bridge East End Crossing Tunnel requires a 75-year service life of its cast-in-place reinforced concrete lining. This twin-tube 3-lane (each tube) highway tunnel is about 55-ft wide and 30-ft high. Since unified tunnel service life prediction methodology does not exist in the US, challenges occurred during the design-build phase of this project. These challenges include the methodologies in determining the reinforced concrete’s degradation mode, the lining’s surface chloride concentration level, the reinforcing steel’s critical chloride threshold value for corrosion, how to verify the concrete’s diffusion coefficient, and the durability impact from concrete crack width. This paper presents a methodology in verifying the proposed concrete mix can achieve the service life requirements. This methodology can be adopted for service life prediction for future tunnels to be built in the US. INTRODUCTION The Ohio River Bridge East End Crossing (ORBEE) Tunnel is part of a Public Private Partnership project, which connects transportation network between Louisville, Kentucky and southern Indiana. Located in Louisville, Kentucky, the twin-tube highway tunnel provides northbound and southbound lanes of the Kentucky approach to a new bridge across the Ohio River. Tunnel lengths are approximately 1,675-ft and 1,685 for the northbound tube and the southbound tube, respectively. Connected by two cross-passages, both tubes have a roadway width of 40-ft, for 3-lane traffic, and a vertical clearance of 26.5-ft.The tunnel was constructed by drill and blast method with shotcrete and rock dowels as initial ground supports. The final tunnel lining system includes: •,Nominal 16” cast-in-place reinforced concrete lining, with 2” and 3” concrete covers over reinforcing steel for its extrados and intrados, respectively. •,20” cast-in-place reinforced concrete invert slab, with 4” and 3” concrete covers over reinforcing steel on its top face and bottom face, respectively. •,PVC waterproofing system around tunnel perimeter, excluding invert slab. •,Perforated PVC groundwater collection pipes on both sides of the tubes and one below the invert slab of each tube."

Jan 1, 2016

By S. Banisi, A. R. Hasankhoei, A. R. Ghasemi, E. Razi, G. A. Parsapour

At the pelletizing plant of the Gol-E-Gohar mining and industrial company two parallel dry closed grinding circuits, each with the average throughput of 300 t/h, are used. Each grinding circuit consists of a burner, a dry ball mill (6.2 m×13 m) and an air separator. The ball mill consists of a 2 m-long drying and an 11 m-long grinding chambers. The iron ore concentrate is fed to the drying chamber by a feed chute. The first 1.35 m of the drying chamber consists of 56 flights in two rows to decrease the feed moisture from 5% to below 0.2% and the rest of dryer chamber consists of deflectors to pass the material to the grinding chamber. It was found that when the feed moisture content increased from 1.3% to 3.5%, the throughput decreased by 12% (35 t/h) indicating a low performance of the dryer. Monitoring the wear rate of flights for a period of 12 months showed that the first 0.8 m length of the flights length did not experience any wear. This means that about 59% of flights do not contribute in material distribution and drying. A DEM simulation of feed chute and the drying chamber demonstrated that due to the inappropriate design of feed chute most of materials do not get in touch with the first 0.8 m length of the drying chamber. To overcome this problem, various feed chute designs with different geometries were simulated by the KMPCDEM© software. With the aim of arriving at a proper material trajectory, where the total length of the flights are used, a new feed chute was selected. The simulation results indicated that if the height of the flights is increased from 1.60 to 2.26 m and the slope is increased from 45 to 48 degrees the material arrive at the first 0.48 m of drying chamber. In this manner, the unused part of the drying chamber decreases from 59% to 36% of the length. After installation of the new feed chute, during a period of three months the throughput increased by 36 t/h.

Jan 1, 2018

"A long-term mine production schedule optimization based on stochastic integer programming is presented that manages the risk due to geological variability and uncertainty while jointly considering mine profitability, truck fleet requirements and waste management. The evolution of the pit is controlled so that space within the mined-out pit is continuously provided to allow for tailings and waste rock to be replaced, thus minimizing the project’s environmental footprint. The method is applied at a taconite iron ore deposit and has a project net present value that is 16.9 percent higher than that of a conventional schedule and has a higher chance of meeting production forecasts. The stochastic schedule reduces equipment requirements by delaying waste mining and by mining selectively in order to meet plant ore feed requirements without stockpiling material, resulting in 23.7 and 26.2 percent lower mining capital and operating costs, respectively. IntroductionThe LabMag iron ore deposit is part of the Millennium Iron Range, a 210-km (130-mile) belt of taconite in northern Quebec and Labrador, Canada. LabMag has significant economic potential. It contains more than 3.3 Gt (3.6 billion st) of proven reserves at an average total iron content of 29.8 percent and is very low in impurities. However, on a project basis of 23 Mt/a (25 million stpy) of product, the required capital expenditure is estimated at more than $5 billion (SNC-Lavalin, 2014), which necessitates careful evaluation of all sources of risk. Resource estimation is one of the main sources of risk in a mining project as knowledge of the orebody is primarily based on drilling, which is often sparse because it is expensive. If the expected ore tonnages and qualities are not obtained when mining, the project cash flows are directly affected. The expected quantities and qualities of ore and waste are defined by the mine production schedule, which specifies the sequence of extraction and is dependent on the resource estimation. The goal of mine production scheduling is thus to maximize the expected profit, while also meeting all production targets and constraints, by creating an extraction schedule that is robust in the face of geological risk and has the highest chance possible of actually being realized."

Jan 10, 2017

By D. Harrington

We Americans fondly believe that ours is the greatest , most progressive and most civilized country in the world , but the cold statistical facts reveal that we apparently have little appreciation for the lives of those whose efforts of various kinds give us whatever greatness we have . The latest available statistics indicate that the annual death rate per 100,000 persons , from accidents is nearly 80 , or far higher than that of any other country in the world ; it is more than double the rate in Denmark , Belgium, Netherlands , France , Sweden , Italy , Ireland , Germany , Norway, and Austria; it is practically double the rate in England and Wales and in Japan , and is from 25 to 60 per cent higher than the corresponding rate in Scotland , New Zealand , Australia, Switzerland , and Canada . And added to this is the fact that of the main industries of the United States mining has by far the worst accident record , viewed from either the frequency or the severity basis . The National Safety Council's latest figures for 27 of the country's main industries give mining the highest or worst accident frequency rate, 74.43 in 1929 ; meat packing comes next with 55.94 , and construction third with 50.41 . In accident severity in 1929 mining also was the worst offender , with a rate of 9.99 ; quarrying is second with 6.11 , and construction third with 4.62 . Metal mining's frequency rate in 1929 was 52.16 , as against 69.25 for bituminous coal mining ,99.68 for anthracite mining, and 74.43 for the entire mining industry ; accident severity for metal mining in 1929 was 5.99 , as against 11.69 for bituminous coal mining , 10.87 for anthracite mining , and 9.99 for the mining industry as a whole . Although this figure indicated that metal mining is on a much safer plane than either bituminous or anthracite coal mining, the frequency rate of 52.16 for metal mining was higher in 1929 than the frequency rate of all but one of the other 26 industries about which data are available ; the metal mining severity rate of 5.99 , while much better

Mar 1, 1932

By Michael Gauna

A high seam geologic condition exists at Jim Walter Resources No.4 mine which requires immediate rib support after a continuous miner cuts a place, High seam mining heights are typically 2.4 to 2.6 m (8 to 9 it) and a thick rock layer is present in the upper portion of the coal rib side. The 610 m (2000 ft) mining depth at No. 4 mine induces horizontal rib line movement after excavating an entry. Horizontal rib movement greater than 30.5 cm (12 in) have been noted. When lateral movement occurs in the high seam areas, the thick rock layer, and the coal above the rock layer, become fractured and tend to roll into the entry. If not adequately supported, large rib rolls develop in the place being bolted as well as in previously mined areas which endanger men and equipment. To control this situation rib support must be installed in the roof bolting operation as the bolting crew works towards the face. A rib control system has been developed for this high seam mining condition utilizing a mechanical anchor bolt placed at an angle in the rib and installed with a modified Fletcher DDFI- 15 bolting machine. Over 120 km (75 mi) of entry have been supported with angle rib bolts. These angle bolts constitute 43 to 60% of the total roof and rib bolt footage in the high seam port- ions of No.4 mine. The angle rib bolting system has proven safer, more effective, and more efficient than the earlier support method employing bolts installed horizontally into the rib. No rib roll accidents have occurred since the angle drill bolting machines were placed in service in 1981. By securing the broken rib material with an angle bolt anchored above the coal pillar, a more effective means of rib control is achieved. Angle rib bolting does not interfere with continuous miner productivity. In addition, bolting crews are now reduced from three to two people and the total drilling supply costs are approximately 60% the previous cost, when compared to the previous support method using horizontal rib bolts.

Jan 1, 1983

By F. Qunfei, W. Laijun, X. Lyu, A. Cosmos, L Huijuan, Z. Jinggang, G. Yao

In this paper, four cyanide-utilising bacteria strains were isolated from a gold mine tailings pond, in the Hexi gold mine (Shandong), China. Based on the analysis of 16S rRNA gene sequence, four strains were identified as Aeromonas hydrophila strain C1, Bacillus mojavensis strain C3, Shewanella oneidensis strain C2, and Aeromonas media strain C4. Based on the cyanide detoxification efficiencies results of the four strains, Bacillus mojavensis C3 had the highest detoxification rate and was selected to assess its efficiency to detoxify cyanide in gold mine tailings at different pH (9.5, 10.5, and 11.6) using cyanide as the sole carbon and nitrogen sources. Strain C3 was capable of detoxifying 46.5mg/L of cyanide-containing gold mine tailings at all tested pH ranges at 30oC, 196 rpm after 15 d. The amounts of cyanide detoxified were 87.5%, 84.1%, and 73.5% at initial pH 9.5, 10.5, and 11.6, respectively. The highest cyanide detoxification was observed at pH 9.5. The viable cells growths in the samples were 5.7E+09CFU/mL, 4.9E+09CFU/mL, and 2.7E+09CFU/mL at pH 9.5, 10.5, and 11.6, respectively. Free ammonia concentrations were observed as a byproduct of the biodetoxification process. The residual ammonia concentration in the sample was observed at 0.35 mg/L, 0.33 mg/L, and 0.62 mg/L at pH 9.5, 10.5, and 11.6, respectively. The highest residual ammonia concentration was noticed at pH 11.6. The amount of cyanide removed by mixed culture strains C3 and C4 was 94.6% and showed a residual free ammonia concentration of 2.6mg/L, as biodetoxification byproduct after 120 h. The pure culture Bacillus mojavensis strain C3 and its mixed culture strains exhibited higher potential to detoxify cyanide-containing gold mine tailings at alkaline pH values and could be cultivated to detoxify gold mine effluents at large-scale. Keywords: Bacteria strain, cyanide, detoxification efficiency, gold mine tailings

Jan 1, 2020

By M Z. Abzalov

Drilling mineral sand deposits has specific challenges, which are different from drilling hard rock deposits: Drilled sequences are represented by non-consolidated free-flowing sands to semi-consolidated clayey sands and indurated layers. Valuable heavy minerals (ilmenite, rutile, zircon, leucoxene) have an average density of 4.6 g/cm3, which is significantly heavier than the hosting sand matrix; mainly represented by quartz (density = 2.6 g/cm3). The large difference in the mineral densities can result in sample segregation in the drilling rods and capturing devices causing biased composition of the recovered samples. The drilled strata can contain beds of consolidated sediments, hard pan lenses and concretions. Therefore, chosen drilling equipment should be flexible and permit a switch from drilling the free-flowing sands to hard rock drilling. Thickness of the sand hosted mineralisation varies from several metres to more than 170 m. Mineralisation is distributed above and below the water table. Evaluation of the mineral sands deposits is based on hundreds of drill holes and many thousands of samples, therefore drilling methods should be fast and cost effective. The drilling methods commonly used for mineral sand deposit evaluation, include aircore, triple tube diamond and sonic; including its variants, such as vibrocore drilling technique. The ability of the studied drilling techniques to properly address the above listed challenges in mineral sand deposits, whilst maximising the quality of the obtained samples has been compared in three deposits: Corridor Sands (Mozambique), Richards Bay (Republic of South Africa) and Fort-Dauphin (Madagascar).This study has shown that aircore samples tend to be negatively grade biased, underestimating the heavy minerals content, in particular when drilled intervals are located below the water table. Triple tube diamond core method can produce non-biased samples. However, results can be marred by incorrectly chosen drilling mud. Sonic drilling is a relatively new technique based on a high frequency vibratory technology. It is the only production drilling technique that allows a continuous relatively undisturbed sample of unlithified sands to be obtained. The main limitations of this technology preventing its wider use for developing mineral sand resources are the relatively high drilling costs, in particular when the depth of drilling exceeds 100 m, and speed of drilling advance.

Oct 5, 2011

By George A. Savanick

The objective of this paper is to review advances in the art of borehole (slurry) mining made by the Bureau of Mines. The introduction gives historical and general background information on borehole mining. The borehole mining concept is defined and justified in terms of economics, health and safety, the environment, and resource conservation. This is followed by a description of the design of a prototype borehole mining tool (BMT) developed by the Bureau of Mines. Included are design drawings and performance specifications. The next section describes the application of the BMT in the extraction of metallurgical coal from steeply pitching seams near Wilkeson, Wash., during 1975-76. The field test showed that the BMT produced coal at the rate of 8 tons per hour (tph) (7,272 kg/hr) from a depth of 88 feet (26.8 m) and was effective in producing coal as far as 15 feet (4.6 m) from the center of the borehole. The data are used as a basis for an economic analysis of borehole coal mining and for the suggestions for the design of a second-generation borehole coal mining tool. The following section describes Bureau efforts in applying borehole mining to the extraction of uraniferous sandstone at Rocky Mountain Energy Corp.'s Nine-Mile Lake site, Natrona County, Wyo. BMT design changes appropriate to uranium mining are described along with production, reliability, and reclamation data taken during 1977-73. The field test revealed that the BMT produced 900 tons (81.8x105 g) of uraniferous sand at an average race of 8 tph (7,272 kg/hr) from a depth of 75 (22.8 m) to 90 (27.4 m) feet and created a cavity 25 feet (7.6 m) in radius. These data are used as the basis for an economic analysis of borehole mining of shallow uraniferous sandstones. Reclamation data included will be the results of backfilling experiments conducted at Nine-Mile Lake during 1978. The most significant of these data were that backfill was emplaced at the rate of 16 tph (14.5x103 kg/hr) and that 90 percent of the material extracted from a cavity can be backfilled.

Jan 1, 1979

By H. J. Rose

Sixty-five million net tons of coal were carbonized in the by-product and beehive coke ovens1 of the United States during 1924. This tonnage represented 13.4 per cent. of the bituminous coal which was mined in that year. Table 1 shows production and disposition.2 Table 1.—Production and Disposition of Ocen Coke in 1924 By-product Beehive Total Coal charged into ovens, tons................ 49,061,339 15,914,310 64,975,649 Coke produced (exclusive of screenings and breeze), tons............................. 33,983,568 10,286,037 44,269,605 Average yield of coke, per cent ............. 69.3 6.46 68.9 Disposition of coke, per cent.: Blast furnace.......................... . . 82.6 Foundry............................... . 13.3 6.6 Domestic.............................. . 1.4 6.7 Water gas manufacture and all other uses.. 4.6 . 4.1 Cokes having somewhat different combinations of chemical and physical properties are desired for the various uses listed above, and these properties depend primarily upon the types of coal used. American coals have a wide range of purity and coking quality, and the chief subject of this paper will be the selection of coals for the manufacture of blast furnace, foundry, domestic, and water-gas coke, in by-product coke ovens. It will obviously be necessary to include a discussion of those physical and chemical characteristics that are generally considered essential or desirable for each type of coke. Suitable Coals The examination and development of new sources of coking coal are important because of the irregular and limited distribution of high-grade

Jan 1, 1927

By Roger Amato

The 1982 United Nations Convention on the Low of the Sea includes provisions under Article 76 for extending a coastal State's jurisdiction beyond the 200-nautical-mile (nm) Exclusive Economic Zone (EEZ) and for establishing the seaward limits of sovereign rights over natural resources. Article 76 prescribes a set of methodologies for determining these outer limits based on physiography and sediment thickness. The methodologies were applied to each of the five continental* margin areas of the United States - Atlantic, Pacific, Gulf of Mexico, Alaska, and Pacific Islands-using existing digital hydrographic and geologic data. Figures 1 and 2 show the results of applying the methodologies to the Atlantic, Pacific, Alaskan, and Gulf of Mexico continental shelves. Figure 1 depicts the preliminary limits of each of the Article 76 determinations for the Atlantic and Gulf of Mexico continental shelves. The Atlantic limits were described by Carpenter, Thormahlen, and Amato (1995) at the 26th Underwater Mining Institute.1 The limits for the Alaskan continental shelves are shown in Figure 2. Additional areas beyond the 200 nm EEZ are highlighted. These include a small area north of the U.S. - Canada boundary, containing 7,000 km2, a large area north of Point barrow (395,000 km2), and a third area north of the Aleutian Islands (178,000 km2). The shelf between the first two areas appears to have sediments thick enough to qualify for the 1 percent method (Irish Formula); however, year-round ice cover has prevented acquisition of seismic or drill data. The entire northern Alaska shelf is prospective for petroleum The Gulf of Mexico (Figure 1) has two small areas beyond the EEZ known as "doughnut holes." One is south of Louisiana and includes part of the Sigsbee Escarpment, an area believed to have potential for oil and gas. Oil was found on Sigsbee Knoll in cores from Deep Sea Drilling Project's Leg 1 cruise in 1968. The Sigsbee areas contains about 6,000 km2. The second area is south of the Florida panhandle and contains about 12,000 km2.

Jan 1, 1996

By Andres M. Baquerizo, Jeremiah J. Filjones, Nicholas Feldt

The Estates at Acqualina located in Sunny Isles Beach, FL, is comprised of two, ocean-front, 50-story, luxury towers with podiums and villas. With city regulations now requesting that above-grade floors be limited to commercial or residential units, parking facilities for the tower structures were situated within two levels of underground basement parking. However, ensuring dry excavation through south Florida’s highly pervious limestone, immediately adjacent to the beach where the groundwater table is very shallow, can be extremely challenging and costly for developers. In recent years, significant advances have been made to overcome this challenge with the innovative use of deep soil mixing to create a “dry hole”, assisting in the construction of foundations and underground usable space for parking, storage, etc. The “dry hole” system includes a soil mix bottom seal, with soil mix secant pile walls in combination with auger cast deep foundation elements to support the future structure. This approach enables groundwater control to greater depths than can be achieved by conventional sheetpiling and dewatering. For the Estates at Acqualina, a successful turnkey, design-build joint venture between two foundation and geotechnical contractors resulted in a dry excavation to 34 ft (10.4 m) below grade and to a record- breaking auger cast pile depth of 182.3ft (55.6 m), all within an aggressive 26-week schedule to deliver the “dry hole” excavation for the first tower. This presentation will discuss the logistics, quality control, design, and safety for the different techniques involved. INTRODUCTION AND BACKGROUND The Estates of Acqualina is a pair of 50-story luxury towers located on Sunny Isles Beach in Florida with a two-level underground basement for parking. The project’s location just steps away from the Atlantic Ocean, geology, and high-water table created a challenging project. The owner, Trump Group, contracted the joint venture of HJ Foundation and Hayward Baker (HJ/HB) to design-assist-build a system to control the groundwater, allowing the 2-story underground parking structure to be constructed. In order to accomplish this task, the JV utilized an installation technique known as deep soil mixing. This technique mixes the in situ material with a cement slurry to create a very impermeable material called “soilcrete”.

Jan 1, 2019

By Georgios N. Anastassakis, Nickolaos Georgiou

"Soil contamination by oil as well as by other organic pollutants poses a major problem to the environment. Various methods have been used to remove organic contaminants from soil, such as: biological, thermal, chemical and physical. Methods based on Mineral Processing principles have also been used (cyclone separation, gravimetric separation, attritioning and washing). This paper studies the possibility to remove a rich-in-contaminant soil fraction by applying flotation. The contaminated soil was artificially prepared by mixing soil with diesel fuel. Sodium dodecylsulfate (SDS) and a tertiary amine (TA) were used as collectors. The effect of collector quantity added and conditioning time was studied. The results showed that the removal of a rich-in-contaminant soil is feasible while a clean fraction is also obtained. In several cases, the removal of oil-contaminated soil approaches or reaches 100%. Introduction Human activities are intimately associated with environmental pollution to the air, water and soil, as well. Soil pollution from inorganic and organic compounds is of equal importance with that of air and water, as it results in soil degradation and its concomitant problems with respect to meeting the human needs (house building, agriculture etc). Heavy metals along with their compounds are encountered among the most usual soil inorganic pollutants. The corresponding organic soil-pollutants are oil, oil-derivatives, pesticides, various aromatic (benzene, phenols, cresols, toluene etc) and poly-aromatic (naphthalene, anthracene etc) hydrocarbons, as well as various other organic compounds (pyridine, cyclo-hexane etc). Oil remains the basic energy source sharing 38% of the global needs, being followed by coal and gas with a share of 26 and 21% correspondingly [1]. The entire activities related with oil production, transport and consumption promote pollution in the air, water and soil. Consequently, soil pollution from oil is favoured at refineries, storage sites (due to possible pipe and tank leakage, tank cleaning works, loading and transportation) and, finally, at the oil-selling sites. Oil (or diesel-oil) fuel is a mixture of hydrocarbons, with 10-20 carbon atoms in the chain of each. It is mainly composed of paraffins, naphthenes and various aromatic compounds. In general, oil is composed of 20-30% in aromatic compounds, 5-6% in naphthenes and the rest is paraffins [1]."

Jan 1, 2008

By Chao Li, Lim Hongxu, Yuyue Wang, Xie Yang

"A intergrated process of sulfuric acid leaching, sodium chloride leaching, lead chloride crystallization, lead hydroxide synthesis and roasting was carried out for the recovery of lead, cadmium, copper and zinc from lead smelting dust containing various metals. The results showed that the separation of cadmium, zinc and copper from dust could be obtained after sulfuric acid leaching and corresponding leaching rate were 92.12%, 90.25% and 88.26% respectively. The subsequent sodium chloride leaching carried out for lead with optimized parameters as: sodium chloride concentration 250g/L, leaching temperature 90°C, leaching time 2h, pulp density 33g/L, 0.12mol/L CaCh; The lead compounds crystallization rate reached 86.67%, and then transferred into mPbO·nH2O, the product of 98.66% ß-PbO could be obtained through roasting under 502°C. In the end, a feasible method of ß-PbO preparation from lead containing dust was put forward.IntroductionThe pyro-metallurgy flue dust from smelting lead activities are often stockpiled and pose potential environmental risks due to the heavy toxic elements, such as Pb, Cd, As, Cu, Cl, and Zn which are contained in the dust. With the rapid depletion of rich ore resources in recent years, it has become increasingly more financially attractive to recover valuable metals from previously stockpiled dust, while simultaneously reducing the potential of environmental pollution from toxic elements contained in the dust. Nowadays, pyro-metallurgical and hydrometallurgical routes or both combined methods can be developed to treat secondary materials such as lead ash, leach residues etc. By comparison the hydrometallurgical processes are often considered more eco-friendly.Some chloride leaching processes for recovering lead have been employed using either NaCl, or MgC2 and CaCl2, or FeCl3 [1] The effectiveness of the brine leaching method is based on the formation of complex metal chlorides in concentrated chloride solutions [2] Similarly, Other metals such as zinc, cadmium and copper contained in the flue dust are often treated with sulfuric acid, hydrochloric acid, ammonium solution and sodium hydroxide etc. as lixiviants [3] The metals from leach solutions thus has been obtained by precipitation, ion exchange, solvent extraction or replacement [4]"

Jan 1, 2014

By T. Saastamoinen, N. Heino

"The Olserum deposit located in Sweden hosts an indicated mineral resource of 4.5 million tonnes over 0.6 % TREO, exhibiting a relatively high proportion of HREE. Beneficiation of a composite drill core sample was studied in the flowsheet of magnetic separation of magnetite followed by REE minerals flotation. Primary testwork was performed to select an effective REE mineral collector and then the parameters of grinding size, collector and depressant dosages, slurry pH and density, conditioning time and temperature, and flotation time were tested for optimization of REE flotation. The studies of process mineralogy showed that the REE-bearing minerals monazite, xenotime and apatite were successfully enriched from the concentrations of 0.6%, 0.31% and 2.6% in the feed to those of 17.0%, 8.9% and 65.0% in the concentrate with the recoveries of 79.0%, 81.3% and 71.0%, respectively. INTRODUCTION The Olserum deposit located in Sweden comprises lower grade placer material with higher grade apatite-magnetite veins and hosts an indicated mineral resource of 4.5 million tonnes at 0.6% TREO, with 34% of that being HREO (EURARE, 2017). The mineralization was studied by Andersson et al. (2018). At early stage, high-temperature REE is characterized by abundant monazite-(Ce) and xenotime-(Y) coexisting with fluorapatite and subordinate amounts of (Y,REE,U,Fe)-(Nb,Ta) oxides. During a subsequent stage, allanite-(Ce) and ferriallanite-(Ce) formed locally, partly resulting from the breakdown of primary monazite-(Ce). Alteration of allanite-(Ce) or ferriallanite-(Ce) to bastnäsite-(Ce) and minor synchysite-(Ce) at lower temperatures represents the latest stage of REE mineral formation. The mineralogy and beneficiation were firstly studied by GTK Mintec in the EURARE project for a composite drill core sample. The Olserum is considerably smaller than the Norra Kärr deposit. However, REE are hosted by xenotime and monazite, both of which have well established processing pathways, and meanwhile, the Olserum has a high contribution of the high value critical REE (dysprosium (Dy), yttrium (Y), neodymium (Nd) and terbium (Tb)), which provide a unique advantage for its exploration (Tasman Metals Ltd., 2013). In this paper, flotation of the REE minerals in the Olserum ore was studied."

Jan 1, 2018

By E. V. Konoplev

Russia is one of the largest manufacturers of potash fertilizers in the world. Processing of potash ores basically is carried out by a flotation way. Potash ores of the Verhnekamsky potash deposit contain sylvite (KCl), halite (NaCl) and water insoluble impurity (up to 8 %), presented silicate- carbonate minerals, easily slimed by ore crushing in rod mills. Ore of some deposit sites, developed by ?Uralkali? Co, contains a significant amount of a carnal-lite (KCl.MgCl2.6H2O). Work practice of flotation potash plants has shown, that increase of car-nallite content in ore leads to increase magnesium ion content in a circulated plant brine (saturated salt solution of potassium and sodium chlorides-sylvite brine) and to deterioration of indicators of ore processing. In the Paper we present results of researches of floatability of various size sylvite and influences silicate-carbonate slimes on sylvite flotation in brines with the various magnesium chloride content. It is established, that with increase magnesium chloride content in the brine mainly decreases the flotation of coarse sylvite and amplifies adverse effect of silicate-carbonate slimes on sylvite flotation. In this connection deterioration of potash ore flotation is caused not only decrease sylvite floatability but also strengthening adverse action of water insoluble impurity on potash flotation. It is shown that the application of frothing agents, which render the stronger dispersive action on the colloidal-micellar structures of amine in the salt solution increases the absorption of collector on the sylvite and improves the flotation of sylvite into magnesium the containing salt solutions. The application of an organic depressor to the greatest degree of decreasing the absorption amine on the slimes and to the smallest degree possessing anionic properties proves to be most effective with the flotation carnallite of the containing potassium ores. It is offered the application the frother on a basis of glycol ethers and organic depressor on a basis of urea-formaldehyde resin. Results of industrial tests developed reagent modes for flotation treatment of ores with the raised carnallite content (to 6 %) on ?Uralkali? Co plant have shown that their application has allowed to receive equivalent indicators potash flotations at change of magnesium chloride content in the brine from 0.7 % up to 2.4 % .

Jan 1, 2014

By H. Yen, D. E. Walsh, H. K. Lin

"Comminution circuits are widely used in nonferrous and precious metals milling. Steel crushers, grinding mills and grinding media are typically used in comminution circuits. Comminution circuit wear products, in the form of metallic iron (steel) fines, are unavoidably produced and report to undersize products in the classification circuit, i.e., flotation or leach feed, while tramp grinding steel (TGS) reports to classification oversize products and returns to the grinding mill, building up in the circulating load. In laboratory studies, TGS reduced grinding efficiency significantly, the impact being dependent upon the size of the TGS present in the mill. Because iron fines adsorb flotation reagents effectively, iron fines increase reagent consumption and reduce flotation recovery. Iron fines are also capable of reducing gold recovery in cyanidation circuits via a cementation reaction. When using filtered gold solution from the cyanidation circuit of an Alaskan gold mine, the gold concentration in solution was reduced by 0.2 to 2.6% over various reaction periods when iron fines were present in the solution.IntroductionComminution circuits are widely used in copper, lead, zinc, nickel, cobalt, molybdenum and precious metal mills. Flotation and leaching processes are usually preceded by comminution (crushing and grinding) circuits in order to liberate valuable minerals for effective separation, concentration or dissolution by the subsequent flotation or leaching processes.Steel crushers, grinding mills and grinding media are typically used in comminution circuits. Comminution circuit wear products, in the form of metallic iron (steel) powder, are unavoidably produced and report to undersize products in the classification circuit (flotation or leach feed), while tramp grinding steel (TGS) reports to oversize productions during classification and returns to the grinding mill, building up in the circulating load. Variables affecting the amount of metallic iron powder and TGS produced in a comminution circuit include ore hardness, grinding mill operating conditions, comminution circuit design and the metallurgy of the steel used for manufacturing crusher liners, mill liners and grinding media. As one example, a comminution circuit product of 80% passing 200 mesh at the Pogo Mine, AK is estimated to contain approximately 2 kg/t of metallic iron. A finer grind will generally produce a comminution circuit product with higher metallic iron fines content."

Jan 1, 2015