年处理60万吨原矿某锰矿选矿厂设计【含CAD图纸+文档】
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任务书学 院： 核资源工程学院 题 目： 某锰矿年处理60万吨锰矿石选矿 厂设计 论文 (设计) 内容及要求：一、毕业设计（论文）原始依据（1）工程概况：年处理60万吨锰矿石，选矿指标：原矿锰品位17.6%，精矿品位23.4%，回收率76.6%；原矿粒度480mm，中等可碎性矿石，精矿含水率小于3%。磨矿细度91%-200目。（2）收集选矿厂设计所需资料（参考类似的选矿厂）。二、毕业设计（论文）主要内容（1）确定从破碎、磨矿、选别和精矿脱水的全厂各作业合理的工艺流程和原始指标，进行流程计算和主要设备的选择与计算，编写矿物加工工艺部分设计说明书。（2）绘制选矿厂平面布置图，选矿厂设备平面布置图（示意图），工艺流程图，绘制全厂设备形象系统图和数质量、矿浆流程图。（3）图纸的绘制要求采用CAD绘制，另手工绘制一张图纸。三、毕业设计（论文）基本要求综合运用所学的基础理论与专业知识（包括以前的生产实习、毕业实习的实践知识），在老师指导下独立地、较系统地完成“某锰矿年处理60万吨锰矿石选矿 厂设计”，巩固所学的各科知识，提高综合运用所学理论知识和专业技能的能力；学会分析解决锰矿石加工过程中的实际问题，增强独立思考的能力，为以后走上工作岗位奠定良好的基础。（1）按照毕业设计任务书的要求，在指导老师的指导和帮助下，结合实际情况，按期、认真完成“某锰矿年处理60万吨锰矿石选矿设计”毕业论文的设计内容，按时提交毕业设计。（2）翻译本专业英文文献一篇（3000-5000汉字）。四、毕业设计（论文）进度安排阶段阶 段 内 容起 止 时 间1广泛查阅相关文献资料并进行分析、整理,编写开题报告。20*.12.2820*.1.152根据所掌握资料，结合实际情况，认真研究、分析，拟定设计方案。20*.2.2520*.2.283确定破碎、磨矿、选别和精矿脱水的全厂各作业合理的工艺流程和原始指标各作业合理的工艺流程和原始指标20*.3.120*.3.264流程计算和主要设备的选择与计算，绘制选矿厂平面布置图，选矿厂设备平面布置图，工艺流程图，绘制全厂设备形象系统图和数质量、矿浆流程图20*.3.2720*.5.105编写矿物加工工艺部分设计说明书20*.5.1120*.5.206检查修改，预答辩。20*.5.2120*.5.317审核、答辩。20*.6.120*.6.2五、主要参考文献1）选矿设计手册2）选矿工程设计手册3）选矿厂设计（三本书）4）有色金属硫化矿选矿5）有色金属选矿厂工艺设计规范（YSJ 01492）6）图书馆、期刊网检索相关资料。 指导老师： （签 名） 年 月 日英文原文Controlled Precipitation Control of the flow rate to the precipitation circuit is considered to be the key to the optimum metallurgical balance of the plant. In view of the closed circuit operation this flow rate affects the amount of wash solution returned for use in the filter circuit and thus the wash efficiencyAs a result, decreasing the flow rate of the pregnant solution to precipitation decreases the amount of caustic required for neutralization of bicarbonate，but at the same time，decreases the wash efficiency and results in higher uranium soluble losses to tailingsA compromise based on a ratio of 0.65 ton of pregnant solution fed to the precipitation circuit per ton of ore has been found to give maximum economy while still providing sufficient solution in the filter circuit for repulping of the filter cakes between stagesTo keep this ratio compatible with the ratio of 1.0 ton of solution per ton of ore required in the leaching circuit requires recycle of part of the pregnant solution to the mill solution tankThis results in a buildup of U3O8 in the pregnant solutions to a level between 3.0 and 3.5 grams per liter at equilibriumDeficiencies in the sodium balance due to decreased caustic addition are compensated by direct addition of soda ash to the preheat tanks preceding the leach circuit Yellow cake is precipitated from the pregnant solution by the addition of caustic soda and this product is purified in a second treatment stage by roasting and water leaching to remove vanadium In the primary precipitation stage the solution is first heated to l65and recycled yellow cake is added to increase the soluble U3O8 content from an initial value of about 3.5 grams per liter to a final value of about 7.5 grams per literThis technique has been found to improve the efficiency of the caustic precipitation from 93 percent for the normal solution to better than 98 percent for the enriched solutionHeating of the solution is accomplished by heat exchange with the autoclave discharge slurryYellow cake is recycled in an amount equivalent to 500 percent of the U3O8 in the feed solution and partially dissolved in a series of two agitated tanks providing 5 hours of residence timeUndissolved yellow cake remaining in the solution has no apparent effect on subsequent precipitationThe caustic precipitation step is accomplished in a series of eight steam-coil-heated and agitated tanks in series which provide eight hours of residence timeCaustic is normally added to the first tank in the series to obtain an excess of approximately 5.0 grams of NaOH per liter in the discharge from the final tankThe solution temperature remains at 165 Following precipitation the yellow cake is dewatered in a 40-foot diameter by 12-foot deep thickener to 35 to 40 percent solids in the underflowPart of this underflow is recycled to the dissolution tanks and the remainder further dewatered by a disk filter in preparation for treatment in the purification circuit Barren solution overflowing the yellow cake thickener is clarified by passage through Shriver filter presses and then recarbonated with boiler flue gas in packed towersThe treated solution，containing from 2 to 3 grams of NaHCO3，40 to 45 grams of Na2CO3，and about 0.07 gram of U3O8 per liter，is used to wash the second-stage filter cake in the liquid-solid separation circuit previously describedThe concentration of bicarbonate is purposely maintained at a low value during recarbonation since additional bicarbonate is formed by the oxidation of sulfides in the leaching circuitPurification The yellow cake，as precipitated with caustic soda，contains 5 to 6 percent V2O5 and 2 to 2.5 percent CO32-，both exceeding specifications and requiring a purification stepTo accomplish this，the disk filter cake is repulped with water to about 70 percent solids in an agitated tank，and 0.75 to 1.0 pound of Na2CO3 is added per pound of yellow cakeThis slurry is then fed to an 8 1/2-foot diameter，6-hearth roaster，having sufficient capacity to hold the yellow cake for approximately 1/2-hour on the last two hearths at a maximum temperature of 1,580The calcine discharges to a water-cooled screw conveyor which reduces the temperature to between 200 and 300 enroute to a water-quenching tank The high vanadium content of the initial caustic precipitate is presumed to be due to coprecipitation with the uraniumBy roasting in the presence of soda ash this vanadium is converted to the water-soluble sodium vanadate which is then dissolved by water leachingCarbonate in the initial caustic precipitate is probably present due to physical occlusion during precipitation and is either decomposed during roasting or is released from the initial crystal structure and subsequently dissolved during water leaching. The discharge from the quench tank is dewatered in a 16-foot diameter by 10-foot deep thickener followed by two 8-foot by 8-foot drum filters in seriesWater is used for washing the cake on each filter and for repulping between stages as a means of displacing vanadium-bearing solutionsThe final washed filter cake drops by gravity to feed a Proctor-Schwartz steam dryer operated at about 300The dried product is crushed through a small roll crusher to minus1/8-inch size and then packaged in 55-gallon drums for shipment The average lot contains over 85 percent U3O8 ，0.2 to 0.8 percent V2O5，about 0.2 percent CO32-；and up to 7.5 percent Na Filtrate from the drum filters returns to the thickener，and the thickener overflow is stored in a small surge tank from which part of the solution is recycled to the quench tankExcess solution in the surge tank is automatically bled to storage tanks and periodically has been sold for vanadium recovery elsewhereThe vanadium concentration in the solution is held at about 8 percent V2O5 through regulation of the amount of water added for quenching and for water washing on the product filtersSodium Removal If sodium removal is required, the quenched product thickener underflow is diverted to a drum filter, and this filter cake is washed with copious amounts of water and repulped in an agitated tank in the ratio of 3 parts water to 1 part solidsThe addition of sulfuric acid to this tank to a pH of 2.5 dissolves the yellow cakeThe acid pregnant solution then flows directly to a second tank where NH3 gas is added to a pH of 7.5 to reprecipitate a purified productDue to exothermic reactions the slurry temperature in the circuit averages 140The use of vigorous agitation in the precipitation tank and the control of pH are critical in producing a crystalline，high grade product Following reprecipitation，the product is returned to the normal series of drum filters for dewatering and washingIn this case，however，a weak solution of (NH4)2SO4 is used instead of water for washing and repulping the cake so as to achieve additional sodium removal through metathesisAfter drying in the normal manner, the final product contains 1ess than 0.5 percent Na and less than 3.0 percent SO42- Filter cake from the final washing stage is repulped with water and recycled tailings solution and pumped to the disposal area where truck-mounted cyclones make a sand-slime separationThe sands are used to form the impounding dikes while the slimes flow into the pond areaThe impounded area is four-sided with up to 75-foot high dikes enclosing approximately 110 acres Clarified solution in the pond is recovered through decant towers and buried piping to sumps located outside of the impounded area，and then pumped back to the mill for use in the third filtration stageEvaporation and entrapped losses are made up by new water additions to the recycled solutions Water pumped out of the United Nuclear-Homestake Partners mines is treated to recover dissolved uranium by resin ion exchange in fixed bed columns located near the mine sitesPregnant eluate solution from the operation is transported by truck l 6 miles to the mill for final treatment Four 16-foot diameter by 8-foot high extraction columns each containing 480 cubic feet of quaternary amine-type resin are used in a series parallel arrangementApproximately 1,700 gallons per minute of mine water is passed upflow through the columns for an average cycle time of 8 daysEffluent water contains less than 1.0 ppm U3O8 which represents all extraction of about 96 percentPortions of this effluent water are pumped back to one dry mine and to two wet mines to augment the underground leaching operationsAverage resin loading is 4 pounds of U3O8 per cubic foot In an earlier phase of this operation, the loaded resin was discharged from the columns to a tank truck by gravity and then transported to the mill where pressurized water was used to force the resin from the truck to the elution column located at the mill siteLater the entire operation is completed near the mines with only the pregnant eluate being transported to the mill Elution is accomplished with four bed volumes of recycle eluate followed by four bed volumes of eluant containing 90 grams of NaCl and 4 grams of NaHCO3 per literUtilizing the split elution technique，the eluant，after passing through the columns，is saved as recycle eluate for the next elution cycleRecycle eluate，after a second passage through the columns, is recovered as pregnant eluate and contains between l2 and l6 grams of U3O8 per liter Precipitation of a uranium product from the eluate is accomplished independently of the main mill circuit by heating to l 90, acidifying to a pH of 3.0 with HCl to decompose the carbonate, and then adding NaOH to pH 7.0 to precipitate the yellow cakeThis slurry is filtered on presses and the cake then transported in drums to the main plant where it is fed into the yellow cake thickener following the caustic precipitation stageFiltrate from the presses is regenerated with NaCl and NaHCO3 additions as required and then recycled as barren eluate.Approximately 13,000 pounds of U3O8 per month are recovered from the treatment of the mine watersThis product contains approximately 1.0 percent chloride，but after dilution in the yellow cake thickeners is well within specified limits for chloride英文翻译控制沉淀控制沉淀在线路的流量被认为是使流程达到最适宜的冶金平衡的关键。鉴于闭路操控流量影响清洗溶液返回用于过滤回路的数量和洗涤效果，所以减少富液成为沉淀的流量就降低了碳酸氢钠的碱性使流量中富液的pH到中性，但与此同时，降低了清洗效率并使更多的可溶性铀丢失在尾液中。一个折中的办法已被用于获得最高的经济价值，且仍需提供充足的用于各阶段过滤回路中滤饼的再浆化，而这个办法是建立在以0.65吨富液每吨矿石的比例加入到沉降回路中的基础上的。保证这个数值能与要求在浸出回路中回收部分富液到细化溶液的容器里的每吨矿石一吨溶液的比例相容。结果导致U3O8 的含量在富液中的平均水平为3.03.5g/L。因降低碱性剂的加入使钠得不到平衡，补偿的方法是在浸出流程之前将碳酸钠加入到预先加热的容器中。从富液中加入烧碱可使黄饼沉淀，通过煅烧和水浸出除掉钒使产品在瞬间的处理过程中被纯化。在主要的沉积阶段首先加热到l65接着将回收的黄饼添加到溶液中来增加可溶性U3O8 的含量使其从初始品位约3.5g/L到最终的品位约750g/L。改进碱法沉淀的技术已经被发现，就是把93%的规定溶液换成高于98%的饱和溶液就能取得更好的效果。加热溶液通过篜压器的热传导来实现去除泥浆。黄饼被回收在大约等于500%的U3O8 给入溶液中和在一系列中的两个震动器中反应五小时后的部分溶解液中。不溶解的黄饼依然留在溶液中对后续的沉淀没有明显影响。碱法沉淀步骤完成需要在一系列的8个蒸发线圈加热器和一系列的震荡器并且持续反应8个小时。碱通常被添加在一系列容器中的第一个，而最后的容器内排出的溶液获得了大于大约5.0g/L的NaOH。溶液的温度依然是165。在接下来的沉淀中黄饼被脱水在一个40英尺的直径12英尺深的浓缩机中，达到35%-45%的固体下溢。部分溢出物被回收通过溶解到容器中，其它的由一个磁盘过滤进一步脱水准备用于净化回路处理。贫液溢出的较浓的黄饼被澄清通过Shriver压力过滤器，然后在一个密闭的碳化塔内实现再碳化。处理好的溶液。其中包含23g/L的NaHCO3，4045g/LNa2CO3，大约0.07g/L的U3O8 ，是用来洗出以前描叙的第二阶段的滤饼来实现的固液分离过程。在浸出过程中的氧化硫化物形成的额外碳酸氢盐在再碳化的过程中碳酸氢钠浓度一直维持在低品位。纯化通过加入烧碱实现沉淀的黄饼，包含5%6%V2O5、2%2.5%CO32。具备精准的技术条件和要求一个纯化阶段。为了实现这一目标，往滤饼内加水使其在固体含量大约为70%在搅拌槽内实现再浆化，0.751.0磅的Na2CO3被加入到每磅的黄饼中，然后将这种浆加入到8 1/2英尺直径、6个炉床内烘焙，能有足够的能量使黄饼在最后两个炉床上的极点温度1，580保持约1/2小时，这个烧成品排放到一个水冷螺杆式输送到一个水态骤冷槽使温度减少到200300。高钒含量包含在初始碱性沉淀物中，推测是由于与铀发生了共沉淀，通过加入苏打水煅烧这种钒使其转换为水溶性钒酸钠，这样就能在随后在水淋洗过程中溶解。最初存在碱性沉淀物中的碳酸盐的出现可能是由于在沉淀过程中的物理阻塞和烘焙过程中分解物或最初的晶体结构和随后溶解在水淋洗中的释放物。履行从骤冷槽脱水是在一个16英尺直径10英尺的浓缩机接着在一系列的两个八英尺直径由八英尺深鼓滤波器。水被用于清洗滤饼在每个过滤和再浆化阶段之间作为一种移除溶液中钒的方法。最终洗好的滤饼通过重力坠入到施瓦兹蒸汽干燥机大约300中操作，干燥好的产品被压碎通过一个小辊式破碎机到略低于1/8英寸，然后包装在55加仑的桶内出货。其中平均包含U3O8 超过85%，0.20.8%V2O5，大约0.2%的CO32;和超过7.5%的钠。从鼓式过滤器出来的滤出液返回到浓密机、浓密机底流的被储存在一个小的调浆槽，部分的溶液被回收到最终容器中。调浆槽内的额外溶液自动的流动到储藏器内，从各个流程中回收的钒被定期出售。钒集结在溶液中形成了大约8%的V2O5通过控制加入骤冷槽水的量和冲洗过滤的产品的水。钠去除如果钠去除是必须的，冷浸产品在浓密机底流中转移到一个鼓式过滤机，该种滤饼用大量的水清洗和在一个搅拌槽内实现再浆化。使水与固体的比是3/1。添加硫酸到这个容器内使pH到2.5让黄饼溶解，这种酸性富液直接流入到第二个容器，在这里加入NH3调整pH至7.5使其再沉淀成为精制产品。由于放热反应，浆的温度在整个回路中平均为140，通过在沉淀槽内剧烈搅动同时控制pH值在生产结晶的、高品位的产品临界上。接下来的再沉淀，这个产品被返回到正规级数的鼓式过滤器中脱水和洗涤。在这种情况下，贫液中(NH4)2SO4代替水来洗涤和再浆化滤饼，从而通过置换完成额外钠的去除，经过正规方式的干燥，最终产品含有低于0.5%的钠和低于3.0%的SO42。滤饼在最终的洗涤阶段是用水再浆化后再回收尾液接着抽到处理区域使用旋流器实现砂浆分离。砂石被用于作为滤网堤当矿浆落入这个池子的区域，这个过滤的区域由75英尺高堤在四面封闭组成约有110万英亩。澄清的溶液在池子里被回收通过移入塔器内，接着埋地下管在蓄水区域外边的坑内，然后抽回去研磨用于第三次过滤阶段。蒸发和滞留损耗的水通过添加新的水到回收溶液中。在联合的原子核能伙伴矿山内抽出的水被处理来回收溶解的铀通过离子交换树脂在位于接近开采位置固定床柱内，操作后富的洗出溶液通过卡车运输到6英里磨坊去做最后的处理。由八英尺高的四个16尺直径萃取柱各包含480立方米的第四纪胺类树脂用于一系列混串布置。大约 1，700加仑每分钟矿坑水向上流动通过柱子平均周期时间为8 天。抽出的96%的污水含有少于百万分之一的 U3O8 ，部分的污水被抽回到一个干的矿井和两个潮湿的矿井去增加地浸操作。平均树脂吸附
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